43-101 Dikulushi Underground

Transcription

Mawson West Limited
Technical Report on the Dikulushi Underground Project
Democratic Republic of Congo – 12 December 2013.
Democratic Republic of Congo – 12 December 2013
Perth Office
Level 4, 50 Colin Street
West Perth WA 6005
PO Box 1646
West Perth WA 6872
Australia
Doc Ref:
Tel:
Fax:
Print Date: 12 December 2013
Number of copies:
Optiro Pty Limited
ABN: 63 131 922 739
www.optiro.com
Optiro:
Mawson West Limited:
Principal Author:
Andrew
Law
MMin. MBA
FAusIMM, FIQA, MAICD, AFAIM.
Signature:
Date:
Principal Reviewer:
Contributing author:
+61 8 9215 0000
+61 8 9215 0011
12 December 2013
Ian Glacken FAusIMM(CP), CEng
Signature:
Date:
12 December 2013
Important Information:
This Report is provided in accordance with the proposal by Optiro Pty Ltd (“Optiro”) to Mawson West Limited and the
terms of Optiro’s Consulting Services Agreement (“the Agreement”). Optiro has consented to the use and publication of
this Report by Mawson West Limited for the purposes set out in Optiro’s proposal and in accordance with the Agreement.
Mawson West Limited may reproduce copies of this Report, in whole or in part, only for those purposes but may not and
must not allow any other person to publish, copy or reproduce this Report in whole or in part without Optiro’s prior
written consent.
Unless Optiro has provided its written consent to the publication of this Report by Mawson West Limited for the purposes
of a transaction, disclosure document or a product disclosure statement issued by Mawson West Limited pursuant to the
Corporations Act 2001 (Cth), Securities Act (Canada) or the rules of any relevant exchange, then Optiro accepts no
responsibility to any other person for the whole or any part of this Report and accepts no liability for any damage,
however caused, arising out of the reliance on or use of this Report by any person other than Mawson West Limited.
While Optiro has used its reasonable endeavours to verify the accuracy and completeness of information provided to it by
Mawson West Limited and on which it has relied in compiling the Report, it cannot provide any warranty as to the
accuracy or completeness of such information to any person.
P a g e | ii
Technical Report on the Dikulushi Underground Project,
Democratic Republic of Congo
A technical report on the Underground Project
Prepared for
Mawson West Limited
Andrew Law
Authors
Director –Mining, Optiro Pty Ltd
Ian Glacken
Director –Geology, Optiro Pty Ltd
MMin; MBA; FAusMM; FIQA; MAICD
BSc (Hons) (Geology);
Geology),
MSc
FAusIMM(CP), CEng
MSc (Mining
(Geostatistics),
Date of report: 12 December 2013
P a g e | iii
TABLE OF CONTENTS
1.
SUMMARY
13
1.1.
1.2.
1.3.
1.4.
1.5.
1.6.
1.7.
1.8.
1.9.
LOCATION
OWNERSHIP
MINERALISATION
MINERAL RESOURCES & RESERVES
MINING
METALLURGICAL
ECONOMIC ANALYSIS
ENVIRONMENTAL
CONCLUSIONS AND RECOMMENDATION
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14
15
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17
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2.
INTRODUCTION
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2.1.
2.2.
2.3.
2.4.
2.5.
2.6.
SCOPE OF THE REPORT
AUTHORS
PRINCIPAL SOURCES OF INFORMATION
SITE VISIT
INDEPENDENCE
ABBREVIATIONS AND TERMS
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20
21
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23
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3.
RELIANCE ON OTHER EXPERTS
30
4.
PROPERTY DESCRIPTION AND LOCATION
31
4.1.
4.2.
4.3.
4.4.
4.5.
DEMOGRAPHICS AND GEOGRAPHIC SETTING
PROJECT OWNERSHIP
PROPERTY LOCATION
THE PROPERTY TENEMENT AREA
ENVIRONMENTAL PERMITS
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31
31
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5.
ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND
PHYSIOGRAPHY
36
5.1.
5.2.
5.3.
5.4.
5.5.
5.5.1.
5.5.2.
5.5.3.
5.5.4.
5.5.5.
5.5.6.
5.5.7.
5.5.8.
ACCESS
SITE TOPOGRAPHY, ELEVATION AND VEGETATION
CLIMATE, PHYSIOGRAPHY, LOCAL RESOURCES AND INFRASTRUCTURE
SURFACE RIGHTS
SITE INFRASTRUCTURE
WATER SUPPLY
POWER SUPPLY
MINE PERSONNEL
TAILINGS STORAGE FACILITY
ADMINISTRATION AND PLANT SITE BUILDINGS
ACCOMMODATION
COMMUNICATIONS
MOBILE EQUIPMENT
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5.5.9.
SECURITY
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6.
HISTORY
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6. 1.
6. 2.
6. 3.
6. 4.
6. 5.
B E LG IA N EX PL OR A TI ON
AN VI L M INI NG L TD
M A W SO N W ES T
RE S OU RC E H IS T OR Y
PR O DU CT I ON HI S TO RY
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7.
GEOLOGICAL SETTING AND MINERALISATION
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7.1.
7.2.
REGIONAL SETTING
PROJECT GEOLOGY
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50
8.
DEPOSIT TYPES
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9.
EXPLORATION
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9.1.
9. 2.
9. 3.
BRGM
AN VI L M INI NG L TD EXP L OR AT I ON
M W L EX PL OR A TI ON
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10.
DRILLING
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10 .1 .
10 .2 .
10 .3 .
10 .4 .
10 .5 .
10 .6 .
10 .7 .
10 .8 .
10 .9 .
10 .1 0.
INT RO D UC T ION
AN VI L PR OGR AM M E 1 99 7
AN VI L PR OGR AM M E S 2 0 0 2 & 2 00 3
AN VI L PR OGR AM M E 2 00 5 / 6
M W L P RO GR AM M E 2 01 0
SU RV E Y CON TR O L
DR IL LI NG OR I EN TA T ION
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11.
SAMPLE PREPARATION, ANALYSIS AND SECURITY
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11.1.
11.1.1.
11.1.2.
11.2.
11.3.
11 .4 .
11.4.1.
11 .5 .
11 .6 .
11.6.1.
11.6.2.
11 .7 .
DIAMOND CORE SAMPLING
DIAMOND CORE RECOVERY
DIAMOND CORE LOGGING
RC SAMPLING AND LOGGING
SAMPLE QUALITY
S AM P L E PR E P AR AT I ON A N D AN A LY TI C AL PR OC E D UR E S
ANALYSES
B UL K D E NS IT Y D E T ERM I NA TI ON S
S AM P L E Q AQC
STANDARDS AND BLANKS
LABORATORY QAQC
SU M M AR Y S T AT E M ENT
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12.
DATA VERIFICATION
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13.
MINERAL PROCESSING AND METALLURGICAL TESTING
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13.1.
13.2.
13.2.1.
13.2.2.
13.3.
13.4.
INTRODUCTION
ANVIL TESTWORK
EARLY TESTWORK
LATER TESTWORK
PLANT OPERATIONAL RESULTS
METALLURGICAL PROPERTIES OF THE CUT BACK ORE AND UNDERGROUND ORE
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14.
MINERAL RESOURCE ESTIMATES
75
14.1.
14.1.1.
14.1.2.
14.1.3.
14.1.4.
14.1.5.
14.1.6.
14.1.7.
14.1.8.
14.1.9.
14.1.10.
14.1.11.
14.1.12.
14.1.13.
14.1.14.
14.1.15.
14.2.
14.2.1.
14.2.2.
14.3.
14.3.1.
14.3.2.
14.3.3.
14.3.4.
14.3.5.
14.3.6.
14.3.7.
14.3.8.
14.3.9.
14.3.10.
14.3.11.
DIKULUSHI MINERAL RESOURCE ESTIMATE
GEOLOGICAL AND MINERALISATION MODELS
DRILL DATA FOR MINERAL RESOURCE MODELLING
DATA VALIDATION
DATA PREPARATION FOR MODELLING
DATA COMPOSITING
STATISTICS
SPATIAL STATISTICS
BLOCK MODEL
DENSITY ESTIMATES IN THE BLOCK MODEL
DETERMINATION OF TOP CUTS
GRADE ESTIMATION
ORDINARY KRIGING INTERPOLATION
MODEL VALIDATION
MINERAL RESOURCE CLASSIFICATION
RESOURCE TABULATION AND INVENTORY
MINERAL RESOURCE ESTIMATE COMPARISONS
MINERAL RESOURCE STATEMENT AUGUST 2011 VERSUS OCTOBER 2007
DEPLETION OF AUGUST 2011 MINERAL RESOURCES BY AUGUST 2013 OPEN PIT CUT BACK
KAZUMBULA MINERAL RESOURCE ESTIMATE
DATA VALIDATION
STATISTICS
SPATIAL STATISTICS
BLOCK MODEL
GRADE ESTIMATION
MODEL VALIDATION
15.
MINERAL RESERVE ESTIMATES
15.1.
15.2.
DEPLETION OF THE OPEN PIT RESERVES
UNDERGROUND MINE DESIGN AND SCHEDULE BASIS
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15.2.1.
15.2.2.
15.3.
15.4.
15.5.
EXISTING WORKINGS
BASIS OF THE UNDERGROUND DESIGN AND SCHEDULE
CUT-OFF GRADE CRITERIA
MINING RECOVERY AND DILUTION
UNDERGROUND MINERAL RESERVE TABULATION
16.
MINING METHODS
16.1.
16.2.
16.2.1.
16.2.2.
16.2.3.
16.2.4.
16.3.
16.3.1.
16.3.2.
16.3.3.
16.3.4.
16.3.5.
16.4.
16.4.1.
16.4.2.
16.5.
16.6.
16.6.1.
16.6.2.
16.6.3.
16.6.4.
16.6.5.
16 .7 .
16.7.1.
16.7.2.
16.7.3.
16.7.4.
16 .8 .
16.8.1.
16.8.2.
16.8.3.
16.8.4.
16.8.5.
16.8.6.
16 .9 .
16.9.1.
16.9.2.
16.9.3.
16.9.4.
16.9.5.
HISTORICAL MINING
PROPOSED MINING METHOD – CUT AND FILL
OVERHAND CUT AND FILL
UNDERHAND CUT AND FILL
MINING OF WIDER SECTIONS OF THE OREBODY
PROPOSED MINING METHOD – EXTRACTION OF THE CROWN PILLAR
GEOTECHNICAL DESIGN PARAMETERS
STOPE LAYOUT AND SEQUENCE
DRILL AND BLAST
ORE EXTRACTION
BACKFILLING
ACCESSING THE OREBODY & REHABILITATION OF OLD WORKINGS
VENTILATION
PRIMARY VENTILATION
SECONDARY VENTILATION
DEWATERING
MINING EQUIPMENT
MINE DEVELOPMENT
MINING SCHEDULE
MINING SHIFTS
DEVELOPMENT / STOPING RATES
AIR LEG DEVELOPMENT RATES
GE O T EC HN IC A L
DATA
GEOTECHNICAL DOMAINS
POTENTIAL FAILURES
MAPPING, MONITORING AND ADDITIONAL DATA
GRO UN D S U PP OR T R E Q UI R EM EN TS
SPLIT SETS
SOLID STEEL ROCKBOLTS
CABLE BOLTS
SHOTCRETE
GEOTECHNICAL FILL REVIEW
CRF MIXING
GRO UN D S U PP OR T ST A ND AR D S
DECLINE SUPPORT STANDARD
ACCESS SUPPORT STANDARD
ORE DRIVE SUPPORT STANDARD WITH MESH
3 WAY INTERSECTION SUPPORT STANDARD
4-WAY INTERSECTION SUPPORT STANDARD
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16 .1 0.
16 .1 1.
W AS T E D UM P D ES IG N
SU RF A C E WA T ER M A N A GE M ENT
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17.
RECOVERY METHODS
17.1.
17.2.
17.3.
PLANT FLOWSHEET
TAILINGS STORAGE FACILITIES (TSF)
PROCESSING STATISTICS
18.
PROJECT INFRASTRUCTURE
18.1.
18.2.
18.3.
SURFACE FACILITIES
POWER
PROCESS WATER SUPPLY
19.
MARKET STUDIES AND CONTRACTS
19.1.
19.2.
MARKETS
CONTRACTS
20.
ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT156
21.
CAPITAL AND OPERATING COSTS
21.1.
21.2.
21.2.1.
21.2.2.
21.2.3.
21.2.4.
21.3.
CAPITAL COST ESTIMATE
OPERATING COST ESTIMATE
MINING OPERATING COST
PROCESSING OPERATING COSTS
MANAGEMENT AND ADMINISTRATION COSTS
TRANSPORT AND SMELTING COSTS
METAL PRICES
22.
ECONOMIC ANALYSIS
161
22.1.
22.1.1.
22.2.
22.3.
22.4.
OPERATIONS SUMMARY
SENSITIVITY ANALYSIS
PAYBACK
MINE LIFE
TAXATION
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23.
ADJACENT PROPERTIES
168
24.
OTHER RELEVANT DATA AND INFORMATION
169
25.
INTERPRETATION AND CONCLUSIONS
170
26.
RECOMMENDATIONS
171
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27.
REFERENCES
172
28.
CERTIFICATES
174
TABLES
Table 1.1
Table 1.2
Table 1.3
Table 1.4
Table 1.5
Table 1.6
Table 2.1
Table 4.1
Table 6.1
Table 6.2
Table 6.3
Table 6.4
Table 9.1
Table 10.1
Table 13.1
Table 13.2
Table 13.3
Table 13.4
Table 13.5
Table 13.6
Table 13.7
Table 13.8
Table 13.9
Table 13.10
Table 13.11
Table 13.12
Table 13.13
Table 14.1
Table 14.2
Table 14.3
Table 14.4
Table 14.5
Table 14.6
Table 14.7
Table 14.8
Dikulushi Mineral Resource statement as at August 2011, using a 1.0% copper cut-off grade
Depleted Dikulushi Mineral Resource statement as at August 2013, using a 1.0% copper cutoff grade
Dikulushi Mineral Reserve statement as at August 2011, at a 1.0% copper cut-off grade
Depleted Dikulushi Mineral Reserve statement as at August 2013, at a 1.0% copper cut-off
grade
Dikulushi Underground Mineral Reserve statement as at September 2013
The Kazumbula Mineral Resource statement as at November 2010
Glossary of terms
Mawson West Limited Dikulushi tenement schedule
Historical work summary at the Dikulushi Project
Mineral Resource estimate as completed by FinOre in July 2006 and published in December
2006; a cut-off grade of 1.5% copper was used
Historical Anvil production for the Dikulushi mine
Recent MWL production for the Dikulushi mine
Historical drilling summary for the Dikulushi copper silver project
MWL drilling at Kazumbula
Details of Dikulushi drill core used in Mintek metallurgical testing
Head grades of chalcocite composites
Relative abundance of significant minerals
Comminution testwork results
Effect of grind size on flotation performance (high grade chalcocite)
Effects of collector addition on flotation performance (high grade chalcocite)
Effect of grind size on flotation performance (disseminated and low grade chalcocite)
Effect of collector addition on flotation performance (disseminated and low grade chalcocite)
Effect of grind size and Eh level on flotation performance (Pb/Zn rich chalcocite)
Locked cycle flotation test results
Dikulushi processing summary (February 2007 – April 2008)
Dikulushi processing summary (June 2010 – July 2013)
Dikulushi Mineral Resource statement as at August 2011 above a 1.0% copper cut-off grade
Domain codes for Dikulushi modelling
Summary statistics for copper % and silver g/t per domain
Dikulushi variogram models with angle1 about axis 3 (Z), angle2 about axis 1 (X) and angle3
about axis 3 (Z)
Dikulushi - top cuts per domain
Mean statistics per domain comparing model estimates with data values
Dikulushi Mineral Resource statement using a 1.0% copper cut-off grade as at August 2011
Comparison of 2011 and 2007 Dikulushi Mineral Resource estimates
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Table 14.9
Table 14.10
Table 14.11
Table 14.12
Table 14.13
Table 14.14
Table 15.1
Table 15.2
Table 15.3
Table 15.4
Table 16.1
Table 16.2
Table 16.3
Table 16.4
Table 16.5
Table 16.6
Table 16.7
Table 16.8
Table 16.9
Table 16.10
Table 16.11
Table 16.12
Table 16.13
Table 16.14
Table 16.15
Table 16.16
Table 16.17
Table 16.18
Table 17.1
Table 17.2
Table 21.1
Table 21.2
Table 21.3
Table 21.4
Table 21.5
Table 22.1
Table 22.2
Comparison of August 2011 and August 2013 Dikulushi Mineral Resource estimates, showing
the Open pit cut back depletion
Summary statistics of the two metre composite data for Cu% and Ag g/t for the Kazumbula
deposit
Density estimates for the Kazumbula deposit
A table of mean statistics comparing model estimates with data values
Kazumbula Mineral Resource statement as at November 2010.
Dikulushi Mineral Reserve statement as at August 2011, using a 1.0% copper cut-off grade
Depleted Dikulushi Mineral Reserve statement as at August 2013, using a 1.0% copper cut-off
grade
Mining dilution table
Dikulushi Mineral Reserve statement as at September 2013
CRF Specifications
CAF Specifications
Dikulushi production mining equipment at site from previous mining activities
Major mining fleet and equipment required for the extraction of the Dikulushi underground
Mineral Reserves
Underground horizontal development design parameters
Underground vertical development design parameters
Mining dilution
Underground mine production physicals
Work shifts
Operational Management Labour
Technical Services labour
Support functions labour
Labour requirements: underground operations
Underground Workshop personnel
Jumbo/production drill rates by development type
Jumbo/production drill rates by individual machine
Jumbo/production drill rates by fleet
Air Leg development
Dikulushi processing summary relevant to ore to be mined in the pit cut back
Processing statistics for the LG material completed by MWL – June 2010 to May 2011
Dikulushi underground capital expenditure cost estimate.
Major mining fleet and equipment required for the extraction of the Dikulushi underground
Mineral Reserves
Mining overhead and fixed costs
Mining variable costs
Metal prices used in modelling
Dikulushi mining and financial summary
Sensitivity analysis on the cash flow forecast for underground mining and treatment at
Dikulushi
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FIGURES
Figure 1.1
Figure 4.1
Locality plan of the Dikulushi Project
Exploration Licences of the Dikulushi copper silver project
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Page |x
Figure 4.2
Figure 4.3
Figure 5.1
Figure 5.2
Figure 5.3
Figure 5.4
Figure 5.5
Figure 5.6
Figure 5.7
Figure 5.8
Figure 7.1
Figure 7.2
Figure 7.3
Figure 7.4
Figure 11.1
Figure 11.2
Figure 11.3
Figure 11.4
Figure 13.1
Figure 14.1
Figure 14.2
Figure 14.3
Figure 14.4
Figure 14.5
Figure 14.6
Figure 14.7
Figure 14.8
Figure 14.9
Figure 14.10
Figure 14.11
Figure 14.12
Figure 14.13
Figure 14.14
Figure 14.15
Figure 14.16
Figure 14.17
Dikulushi mine infrastructure within the PE 606
Dikulushi mine site aerial view
Barge on Lake Mweru
Composite Temperature and Rainfall Data recorded over the last 5 years.
Dikulushi airstrip and the G1 Charter plane provides safe staff transportation to and from site
Dikulushi Administration Centre
Dikulushi Camp Site
Dikulushi Clinic and communications centre
Dikulushi Workshop
Dikulushi Store and Fuel Farm
Regional Geology of Mawson’s convention area in the DRC
Stratigraphy of Dikulushi region with known styles of mineralisation
Local geology of the Dikulushi open pit
A typical vertical cross section through the Kazambula deposit, highlighting key geology
associated with mineralisation
GBM301-7 suggests accurate values around low value samples (~0.55% Cu)
The GBM301-8 is a high Cu value standard and suggests accurate results for high value
samples (~10% Cu)
The GBM398-4c is a low Cu value standard and suggests accurate results for low value
samples (~0.39% Cu)
Results for this blank demonstrate that contamination is well contained
Dikulushi Underground sources of ore - showing North-South section view at 50205E
An oblique southward looking 3D view of drillhole type and distribution at Dikulushi
A vertically oriented 3D view at Dikulushi, looking southwest, showing mineralisation lenses
and current drilling
A plan showing the distribution of drillhole types across Dikulushi; blasthole data from the pit
have been excluded
Quantile-Quantile (Q-Q) plot of Diamond (DD) drilled samples versus sludge drilled samples
within a common area
Cumulative distribution of sample lengths highlighting the dominant 1m sample length
Log histogram and probability plot for the main FW zone of mineralisation showing the results
of robust domaining
Variogram models for copper % across the FW zone of mineralisation
A plan view slice through the FW zone block model illustrating the good comparison between
model estimates and the nearby drillhole data
A statistical plot of estimates versus drillhole data grades for successive 30m increments in
elevation and the full strike length of the FW zone mineralisation
3D view of the Dikulushi model, looking south, and showing resource classification categories
A waterfall chart of cumulative Mineral Resource changes from 2007 to 2011
Grade tonnage curves for the combined remaining Measured and Indicated Mineral
Resources
Kazumbula vertical section, looking north, highlighting the modelled mineralisation as per the
RC and diamond drilling
Plan showing the distribution of RC and diamond drillholes across the Kazumbula deposit.
Histogram and probability plots for the Kazumbula deposit two metre sample data.
Variogram modelling for Cu % in the plane of mineralisation.
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Figure 15.1
Figure 15.2
Figure 15.3
Figure 16.1
Figure 16.2
Figure 16.3
Figure 16.4
Figure 16.5
Figure 16.6
Figure 16.7
Figure 16.8
Figure 16.9
Figure 16.10
Figure 16.11
Figure 16.12
Figure 16.13
Figure 16.14
Figure 16.15
Figure 16.16
Figure 16.17
Figure 17.1
Figure 18.1
Figure 18.2
Figure 18.3
Figure 20.1
Existing workings, showing as built underground development (grey), and the as-built pit
(green)
Underground reserve, showing as-built underground development (grey), as-built pit (green),
and measured (purple) and indicated (red) Mineral Resources
Relationship between cut-off NSR and metal grades
Overhand cut and fill mining process
Underhand cut and fill mining process
Diagrammatic representation of sequential mining in wide orebody areas
Orebody access development
Pillar ratio diagram
LHD loader with ‘rammer-jammer attachment
Underground primary ventilation circuit (full)
Underground primary ventilation circuit required for the extraction of the measured and
indicated material only
Primary ventilation fan location
Existing underground dewatering infrastructure locations
Proposed underground dewatering infrastructure locations
Ore loss due to gaps left in the backfilling process
Ore level schedule, by quarter
Dikulushi orebody rock quality, Q (Turner, 2013)
Dikulushi footwall rock quality, Q (Turner, 2013)
Dikulushi hanging wall rock quality, Q (Turner, 2013)
Dikulushi rock reinforcement chart (Turner, 2013)
Dikulushi Plant flow diagram
On-site office facilities at Dikulushi
On-site Underground change room facilities at Dikulushi
Average water balance
Community Business making work clothes for the mine.
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1.
SUMMARY
Mawson West Limited’s (MWL’s) Dikulushi Underground Project (the Project) is located in the
Katanga Province of the Democratic Republic of Congo (DRC). The Underground Mine comprises
Mineral Resources from the main Dikulushi deposit’s “Footwall” zone, which has a 230 m strike
length and true widths of up to 25 m. The open pit was recently completed as a cut back extension
of the old Dikulushi open pit mined by Anvil Mining Limited (Anvil) during its tenure of the Dikulushi
deposit. MWL has now completed an underground pre-feasibility study to re-enter and re-establish
the old underground workings and mine out the previously developed high grade Mineral Reserves
as a first stage. This study is the focus of this Technical Report, which also details information
regarding the associated Kazumbula project.
In addition to the mining of the remaining developed Mineral Reserves during the first stage, MWL
will continue to explore and evaluate depth extensions of the remaining underground Inferred
Mineral Resource. This will be done through additional underground drilling from within the reestablished Dikulushi underground workings and, once completed, will form the basis of further
underground feasibility study work based on the additional drilling and Mineral Resource evaluation
outcomes.
1.1. LOCATION
The Project is located at latitude 08°53’37.7 south and longitude 28°16’21.8 east in the south
eastern corner of the DRC, approximately 50 km north-northwest of the small town of Kilwa and
situated on the south western side of Lake Mweru (Figure 1.1).
P a g e | 13
Figure 1.1
Locality plan of the Dikulushi Project
1.2. OWNERSHIP
The Dikulushi mine is governed by the Dikulushi Mining Convention signed on January 31, 1998 with
the Government of the DRC, and ratified by Presidential Decree issued on February 27, 1998.
The Dikulushi Mining Convention is a mining concession granted to Anvil Mining Congo SARL (AMC)
which sets out the regulatory and fiscal regime applicable to the tenements owned by AMC.
Mawson West Investments Ltd, a wholly owned subsidiary of Mawson West Limited, holds 90% of
the issued capital of AMC, with the remaining 10% being held by the Dikulushi–Kapulo Foundation
NPO.
Mining operations at Dikulushi are currently conducted under the Exploitation Permit 606 (PE)
issued by Ministerial Decree under the terms of the Dikulushi Mining Convention. This guarantees
the sole and exclusive rights to the benefit of the holding company for 20 years until 2022. The
Dikulushi deposit forms part of the PE.
This report presents technical information on the Dikulushi deposit, relating to the recently
completed open pit cut back and, more particularly, to the planned re-establishment of the
underground workings and trial stoping of the previously developed levels. Additionally, further
exploration and drilling of the Inferred Mineral Resource and currently unclassified material will also
be undertaken from the re-established underground workings.
P a g e | 14
1.3. MINERALISATION
The Dikulushi copper deposit is interpreted to be a hypogene, fault-controlled orebody, comprising
disseminated, brecciated and massive chalcocite-bornite mineralisation with a supergene weathered
and oxidised zone of semi-massive malachite, azurite and nodular cuprite. Most of the oxidised
portion of the Dikulushi deposit has been mined out.
1.4. MINERAL RESOURCES & RESERVES
The current Mineral Resource of the Dikulushi orebody has been derived from a mineralisation
interpretation based upon copper drillhole grades. A block model estimate was completed in May
2009 by David Gray of Optiro and was depleted in August 2011 with updated surveyed volumes of
historical mining. The resulting Mineral Resource is stated for a 1.0% copper cut-off grade in Table
1.1.
Table 1.1
Dikulushi Mineral Resource statement as at August 2011, using a 1.0% copper cut-off grade
Category
Measured Mineral Resources
Indicated Mineral Resources
Total Measured and Indicated Mineral Resources
Category
Volume
3
(m *1,000)
184
90
274
Volume
3
(m *1,000)
Density
3
(t/m )
2.8
2.8
2.8
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
516
251
7.0
5.6
767
Tonnes
(*1,000)
Silver
(g/t)
6.6
Copper
(%)
Inferred Mineral Resources
136
2.8
380
The inferred silver grade was incorrectly reported previously at 91 g/t; the correct grade is 155g/t
211
114
179
Silver
(g/t)
6.8
91
The resulting estimates are supported by historical production and current processing grades.
The August 2011 Mineral Resource from Table 1.1 has now been depleted by the open pit cutback.
The remaining Mineral Resources are stated for a 1.0% copper cut-off grade in Table 1.2.
Table 1.2
Depleted Dikulushi Mineral Resource statement as at August 2013, using a 1.0% copper cut-off grade
Category
Category
Volume
3
(m *1,000)
74
53
127
Volume
3
(m *1,000)
Density
3
(t/m )
2.8
2.8
2.8
Density
3
(t/m )
Tonnes
(*1,000)
207
148
354
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
5.4
6.6
5.9
Copper
(%)
163
131
150
Silver
(g/t)
130
2.8
365
7.0
160
The inferred silver grade was incorrectly reported at 91 g/t in the August 2011 Mineral Resource table and should have
been 155g/t. This has now been corrected and adjusted accordingly in the depleted Mineral Resource.
The open pit Mineral Reserves, as published 16 September 2011 and revised 8 January 2013, are
shown in Table 1.3 and are stated for a 1.0% copper cut-off grade. Mineral Resources are reported
as inclusive of Mineral Reserves. The Mineral Reserve, as per the CIM definition, incorporated
mining losses and dilution material brought about by the mining operation.
P a g e | 15
Table 1.3
Dikulushi Mineral Reserve statement as at August 2011, at a 1.0% copper cut-off grade
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper (%)
Silver
(g/t)
Proven
Probable
66.6
127.8
2.8
2.8
184.7
354.3
7.27%
5.51%
207
169
Total Proven and Probable Reserves
194.4
2.8
539.0
6.12%
182
The open pit Mineral Reserves have now been depleted with the mining of the open pit cut back as
this was completed during July 2013. The open pit Mineral Reserves were based on the open pit
reaching the 810 mRL. Mining ceased at the 825 mRL following some isolated sections of the pit wall
deteriorating beyond what was predicted. Table 1.4 shows the depleted Mineral Reserves post
cessation of mining of the open pit cut back.
Table 1.4
Depleted Dikulushi Mineral Reserve statement as at August 2013, at a 1.0% copper cut-off grade
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper (%)
Silver
(g/t)
Proven
Probable
1.0
29.9
2.8
2.8
2.7
83.7
6.8
5.5
186
188
30.9
2.8
86.4
5.5
188
The above remaining Mineral Reserves have subsequently been incorporated into the Underground
Mineral Reserves, which are presented in Table 1.5 below and are now based on an NSR value cut
off value of US$329/t, using a copper price of US$3.08/lb and a Silver price of US$20 per oz. Mineral
Resources are inclusive of Mineral Reserves. The Mineral Reserve, as per the CIM definition,
incorporates mining losses and dilution material expected to be incurred through the underground
mining operation.
Table 1.5
Dikulushi Underground Mineral Reserve statement as at September 2013
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper (%)
Silver
(g/t)
Proven
Probable
0
62
0
2.8
0
173
0
5.2
0
127
62
2.8
173
5.2
127
Notes:
1) The reporting cut-off grade is based on an NSR value of US$329/t, using a copper price of
US$3.08/lb and a Silver price of US$20 per oz.
2) The above Mineral Reserve does not include any Inferred Mineral Resources.
The Mineral Reserves detailed above are derived from the depleted Measured and Indicated Mineral
Resources that remain below the open pit floor at the 825 m RL, and which can be economically
extracted based on the modifying factors as compiled in the underground pre-feasibility study.
The Kazumbula orebody was originally drilled by Anvil. MWL has developed confidence in this
deposit’s grade and geological continuity by drilling additional reverse circulation (RC) and HQ3
diamond core during 2010. A litho-structural and grade based interpretation was completed by
P a g e | 16
MWL geological staff. The Mineral Resources for Kazumbula effectively use a 0.5% copper cut-off
grade for defining the mineralised volume and are shown in Table 1.6 below.
Table 1.6
The Kazumbula Mineral Resource statement as at November 2010
Category
Volume
3
(m *1,000)
126
Density
3
(t/m )
2.5
Tonnes
(*1,000)
318
Copper
(%)
Silver
(g/t)
1.8
19
1.5. MINING
With the completion of the open pit cut back, continuation of mining activities at the Dikulushi
project will now focus on re-establishing the underground mining operation at the deposit in order
to exploit the previously-developed workings from underground mining activities conducted by
previous owner of the mine, Anvil.
Re-commencement of the underground mining activities requires the rehabilitation and reestablishment of the ventilation, electrical, air and water services. The mine is currently being
dewatered, which is nearing completion. In addition to the re-establishment of services, additional
ground support checks and repairs will be required to ensure that the existing underground
development is to the required standard in order to allow safe access to the underground mining
areas. Initial rehabilitation of the underground workings is expected to take approximately one
month, and will be tied in with the development of service infrastructure requirements for the
underground operations which includes escape way rises, return airway rises, and the installation of
the primary ventilation fans.
Production activities will commence in levels that were partially complete from previous
underground mining activities. Work on these levels will extract the remaining ore contained within.
Once the ore has been extracted from these levels, each of the drives will be backfilled using
cemented fill. Development activities will also commence in other parts of the underground mine,
establishing new levels for production.
Ore extraction from the underground will be completed using overhand and underhand cut and fill
mining practices, with the bulk of the ore being removed using the overhand mining method.
Extraction of the ore between the upper levels of the underground workings and the bottom of the
pit (the crown pillar) will be completed using a long hole stoping method.
In addition to the recommencement of underground mining activities, additional exploration drilling
is planned to be undertaken in order to upgrade the known Inferred Mineral Resource to a
Measured and/or Indicated classification and possibly extend the depth of the Mineral Resource.
The crown pillar extraction is planned to take place on a retreat method as a final operation prior to
closing the underground on completion of the extraction of the current Mineral Reserves. Should
the planned exploration drilling upgrade and extend the additional areas of the Mineral Resource
classification, and thus the Mineral Reserves, mining of the crown pillar ore tonnes will need to be
deferred.
P a g e | 17
The current underground life of mine is 19 months, including one month for the re-establishment of
the existing underground workings. The average monthly ore production rate over the life of mine is
approximately 7,200 t at an average Copper grade of 5.15% Cu and a Silver grade of 127 g/t Ag. The
production tonnes and grade quoted include the extraction of the crown pillar ore during the last 6
months of the mining schedule.
1.6. METALLURGICAL
Several metallurgical testwork programmes have been completed by Anvil on the Dikulushi ore and
are discussed in Chapter 13. These results are appropriate for deposits with similar styles of
mineralisation, such as Kazumbula, and have subsequently been compared against actual production
results during the period of operation by Anvil and more recently by MWL.
The most recent metallurgical testwork was managed by Sedgman Metals, a metallurgical consulting
company of Perth, Western Australia. Testwork was completed at AMDEL Laboratories in Perth.
Metallurgical testwork was carried out previously by Anvil on the main Dikulushi orebody.
Additional testwork was reported on in June 2004 by Independent Metallurgical Laboratories (IML),
which utilised samples provided from the mill feed and an open pit sample to perform a locked cycle
flotation test. Results indicated that from a feed grade of 8.76% copper and 306 g/t silver a recovery
of 91.1% copper and 89.7% silver could be achieved to produce a concentrate with grades of 42.1%
copper and 1,447 g/t silver. This sample contained 18% acid soluble copper in feed. Actual
production results during operations by Anvil were higher.
The float plant at Dikulushi operated from 2007 to 2008, was fed with high grade ore from the open
pit and underground mine, and yielded recoveries of 90.4% copper and 90.3% silver, producing a
concentrate with 55.5% copper and 1721 g/t silver.
The current plant, under MWL control over the past year, has been fed from the open pit cut back
and low grade stockpiles, with recoveries averaging 91.5% copper and 90.3% silver, producing a
concentrate with 56.5% copper and 1,515 g/t silver.
Plant operation under MWL over the past 6 months of production (Feb 2013 to July 2013) has seen
fresh ore feed from the open pit cut back, combined with improved operating management
practises, resulting in improved recoveries of 94.3% copper and 92.1% silver, producing a
concentrate with 61.4% copper and 1,768 g/t silver.
There has been no change in the material ore types since the previous open pit and underground
operations and it is therefore expected that the current recoveries being achieved for the fresh ore
from the open pit cut back feed will continue to be achieved with the re-establishment of the
underground operations.
The financial model uses 94% recovery for copper and 90% for silver, with a copper concentrate
grade of 60% copper.
P a g e | 18
1.7. ECONOMIC ANALYSIS
In summary, the underground operation will produce some 174, 000 tonnes of ore at a copper grade
of 5.15g/t and a silver grade of 127 g/t over a 19 month production period. This will produce some
8,100 tons of copper metal and 573,500 ounces of Silver for sale. Total operating costs are
estimated at $44.6M ($257 per ore tonne milled or $1.86 per lb of copper - net of the silver credit);
and the total net revenue is estimated at $57.7M, with net cashflow totalling $3M. The Internal
Rate of Return is 8% for the project, and the Capital cost for the project is $9.6M.
1.8. ENVIRONMENTAL
An Environmental Impact Assessment (EIA) for the Dikulushi project was lodged in 2003. In 2009, an
EIA for the underground Project was submitted to the DRC Government. Both of these reports were
compiled by African Mining Consultants of Kitwe, Zambia, an environmental company that was
licensed to work and report in the DRC. In 2011, an EIA for the cutback project was prepared by
EMIS sprl, a DRC environmental company licenced to work and report in the DRC. All three
environmental reports received DRC Government approval. A revised EIA, extending underground
mining beyond 2013, has been submitted to Government.
MWL has lodged $1.19M as an environmental bond. This financial guarantee is a contribution
towards environmental rehabilitation costs for the Dikulushi mine.
1.9. CONCLUSIONS AND RECOMMENDATION
The Project is at an advanced stage and Dikulushi may be described as a producing and developing
property. MWL has completed a pre-feasibility study in order to determine the economics of
continuing to mine the Dikulushi deposit via the previously established and developed underground
workings. Since this was previously an operating open pit and underground mine, the remaining ore
zones present the same risks as before, being somewhat mitigated for the mineralogy, metallurgical
properties and the processing aspects, which are well known. Risks associated with the mining
operations will remain, however, and constant recognition of changing conditions will need to be
ensured with appropriate changes made as mining progresses. Geotechnical knowledge will
increase with the physical mining activities and a better understanding of the underground ground
conditions will be established. There is likely to be continued resource development drilling
throughout the mining operations in order to locate and evaluate additional resources associated
with the same ore zone, either at depth or as lateral or parallel extensions. During the period
required to re-establish the underground workings and re-commence development and stoping
operations, MWL intends to continue processing the HG open pit cut back stockpile. In addition
MWL is currently in the process of defining additional deposits on the Dikulushi property and within
50 km of the Dikulushi plant.
P a g e | 19
2. INTRODUCTION
2.1. SCOPE OF THE REPORT
Mawson West Limited (MWL) commissioned Optiro Pty Ltd (Optiro) in May 2013 to review the
underground pre-feasibility study, generated by MWL, and to prepare an independent technical
report regarding the copper-silver Mineral Reserves at the Dikulushi underground deposit based on
the above study. This Technical Report has been written to comply with the reporting requirements
of the Canadian National Instrument 43-101 guidelines, “Standards of disclosure for Mineral
Properties” dated April 2011 (the Instrument) and with the “Australasian Code for Reporting of
Mineral Resources and Ore Reserves” of December 2004 (the JORC Code) as produced by the Joint
Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute
of Geoscientists and Minerals Council of Australia (JORC 2004).
The Technical Report has been written to provide the market with an update on the status of the
Mineral Resources and Reserves for the Dikulushi Open Pit cut back project (mining now complete
and processing due for completion late 2013) and to present the first stage of the underground
Project study, which is to re-enter and re-establish the old underground workings and to mine out
the previously developed high grade Mineral Reserves. This is the focus of this Technical Report.
For completeness, the Mineral Resource estimation of the related but separate Kazumbula project is
also described.
All monetary amounts expressed in this report are in United States of America dollars (US$) unless
otherwise stated.
2.2. AUTHORS
The key authors for compiling this report are:
Mr Andrew Law is the principal author and Qualified Person and takes overall responsibility
for this report. Mr Law is the Director - Mining at Optiro and is a professional Mining
Engineer. He has a HND Metalliferous Mining (1982) and an MBA from the University of
Western Australia. He has more than 30 years’ experience in the planning, development
and extraction of mineral reserves. Mr Law is a Fellow of the Australasian Institute of Mining
and Metallurgy (FAusIMM) and has the relevant qualifications, experience and
independence to be considered as a “Qualified Person” as defined in Canadian National
Instrument 43-101. Mr Law has visited the Dikulushi deposit (February 2012) and the
underground workings to the 830 m RL. Mr Law was a previous author for the Dikulushi
Open Pit Cut back NI 43-101 report generated by Optiro for MWL. Mr Law has reviewed all
sections of the “Pre-Feasibility” study generated by various other Qualified Persons, most of
whom were independent of Mawson West, and collated into a pre-feasibility study by MWL.
Mr Ian Glacken is a Qualified Person and takes responsibility for the Mineral Resources
estimation portion of this report. Mr Glacken, is a full time employee of Optiro, where he
holds the position of Geology Director, and is a professional Geologist. Mr Glacken has
P a g e | 20
degrees from Durham University (BSc (Hons) Geology, 1979), The Royal School of Mines
(MSc Mineral Exploration, 1981), Stanford University (MSc Geostatistics, 1996) and Deakin
University (Postgraduate Diploma of Computing, 1996). Mr Glacken is a Fellow of the
Australasian Institution of Mining and Metallurgy (Member Number 107194) and a
Chartered Professional Geoscientist of that Institution. He is also a Member of the Institute
of Mining, Metallurgy and Materials (UK) and a Chartered Engineer of that institution. Mr
Glacken has the relevant qualifications, experience and independence to be considered as a
“Qualified Person” as defined in Canadian National Instrument 43-101. Mr Glacken has not
visited the Dikulushi deposit but has reviewed and supervised Mineral Resource models on
the Dikulushi deposits. Optiro is an Australian-based mining and resources consulting and
advisory firm which provides a broad range of expert services and advice, locally and
internationally, to the minerals industry and financial institutions.
In September 2011 Optiro generated and supervised Mineral Resource and Reserves models for the
Dikulushi open pit cut back. In August 2013, Optiro depleted the Mineral Resource and
subsequently the open pit cut back Mineral Reserves. With the completion of the open pit cut back,
Optiro has now generated underground Mineral Reserves for the Dikulushi deposit based on the
depleted Mineral Resources as at August 2013. Optiro is an Australian based mining and resources
consulting and advisory firm which provides a broad range of expert services and advice, locally and
internationally, to the minerals industry and financial institutions.
The following authors contributed to the report:
Name
Andrew Law
Ian Glacken
Position
Director-Mining, Optiro Pty Ltd
Director – Geology, Optiro Pty Ltd
Mike Turner
Turner Mining and Geotechnical Pty Ltd
Duncan Grant-Stuart
Knight Piesold Consulting
Peter Hayward
Sedgman Ltd
NI 43-101 Contribution
Principal Qualified Person
Qualified Person and contributing
author of sections 1, 7.8, 9, 10, 11, 12
& 14.
Geotechnical, QP and author of
geotechnical submission in section 16
Engineer, QP and reviewer of tailings
storage facilities in section 17
Metallurgical, QP and input into
section 13 and 17.
2.3. PRINCIPAL SOURCES OF INFORMATION
The principal source of information used to prepare this report is the information prepared for the
development of the pre-feasibility study and the previously submitted NI 43-101 Technical Reports
covering Mineral Resources and Reserves at Dikulushi. This pre-feasibility information was provided
to Optiro by MWL. The Mineral Resource information has been sourced from the previously
submitted NI 43-101 Technical Report, by Optiro, on the Dikulushi Project, Democratic Republic of
Congo, 16 September 2011 and revised 8 January 2013. The Mineral Resource has recently
undergone a review and depletion process based on the recently completed open pit cut back.
In summary, the following are primary data sources:
P a g e | 21




the NI 43-101 Technical Report on the Dikulushi Project, Democratic Republic of Congo,
February 3, 2011 and subsequently revised March 7, 2011
historical and current production and processing data
the NI 43-101 Technical Report on the Dikulushi Open Pit Project, Democratic Republic of
Congo, issued 16 September, 2011 and Revised 8 January, 2013.
A pre-feasibility study for the underground prepared by Mawson West based on inputs from
various independent qualified persons.
Optiro has made all reasonable enquiries to establish the completeness and authenticity of the
information provided. In addition, a final draft of this report was provided to MWL along with a
written request to identify any material errors or omissions prior to lodgement. The following
professionals have been consulted for relevant detail contained in this report.
Name
Greg Entwistle
Chris Marissen
Gary Brabham
Mike Turner
Duncan Grant-Stuart
Peter Shephard
Peter Hayward
Andries Strauss
Glen Zamudio
Company
Mawson West Ltd
Mawson West Ltd
Mawson West Ltd
Turner Mining and Geotechnical Pty Ltd
SRK Consulting
Sedgman Ltd
Mawson West Ltd
Pre-Feasibility Contribution
Operational Management Review
Mining
Geological
Geotechnical
Tailings storage facilities
Hydrology and Water Management
Metallurgical
Tailings storage facilities
Commercial
2.4. SITE VISIT
Mr Andrew Law visited the Dikulushi Project in February 2012 and specifically visited the
underground decline and openings that were available at the time (approx. 830 mRL). He has now
reviewed all sections of the Pre-Feasibility study collated by MWL and generated by various
Qualified Persons, many of whom were independent of MWL.
Mr David Gray (a former employee of Optiro and a QP for previous Dikulushi Technical Reports)
completed a comprehensive site visit to the Dikulushi copper Project in November 2010. The
purpose of this visit was to:






verify the relative size, position and presence of copper mineralisation at the Dikulushi and
Kazumbula deposits
verify the presence and position of drillhole sampling for the respective resources and
reserves
inspect the drill core for mineralisation, geological relationships with mineralisation and
general sample quality
review the respective sampling methods and QAQC with onsite geologists
review and confirm sample and assay data as stored in the drillhole database
review historical and current production and processing data.
Mr David Gray did not take independent samples due to the operational nature of the respective
resources and the visible in-situ mineralisation which confirms drillhole sample results. Mr Ian
P a g e | 22
Glacken, Director-Geology at Optiro, has not visited the Dikulushi Operation, but has nonetheless
supervised and peer reviewed the Dikulushi Project work complied by Mr Gray since 2009, and now
accepts responsibility for the Mineral Resources estimation as stated in this report.
Site visits have been carried out by the following persons:
Name
Andrew Law
David Gray
Chris Marissen
Gary Brabham
Mike Turner
Duncan Grant-Stuart
Peter Hayward
Peter Shephard
Glen Zamudio
Company
Optiro
Optiro
Mawson West Ltd
Mawson West Ltd
Turner
Mining
and
Geotechnical Pty Ltd
Sedgman Ltd
SRK Consulting
Mawson West Ltd
Section
Mineral Reserves
Resource NI 43-101
Mining
Geology
Geotechnical
Date of Visits
February 2012
November 2010
Various as employee of MWL
December 2012
Tailings storage facility
Metallurgical
Hydrology, Water Management
Commercial
July 2010
February 2012
Once during 2007
2.5. INDEPENDENCE
Neither Mr Andrew Law or Mr Glacken, nor Optiro, have or have had any material interest in MWL
or its related entities or interests. This report has been prepared in return for fees based upon
agreed commercial rates and the payment of these fees is in no way contingent on the results of this
report.
2.6. ABBREVIATIONS AND TERMS
A listing of abbreviations and terms used in this report is provided in Table 2.1 below.
P a g e | 23
Table 2.1
Glossary of terms
/
$
%
2D
3D
A
AC
ADT
Ag
allochthonous
anticline
arenaceous
argillaceous
As
ASCu
arsenopyrite
azurite
BCM, bcm
bimodal
black copper
BOCO
bornite
boudinaged
brecciated
BMWi
°C
carbonates
carrollite
CCD
cell
chalcocite
chalcopyrite
chrysocolla
clastic
cm
CMN
Co
conglomerate
Per
Dollars
Percentage
Two dimensional
Three dimensional
Ampere(s)
Alternating Current
Articulated dump truck
The chemical symbol for the element silver
A term applied to the material forming rocks which have been transported to the
site of deposition
A description of folding of rocks which has produced a convex shape
A group of detrital sedimentary rocks, typically sandstones, in which the particles
range in size from 0.06 mm to 2 mm
A group of detrital sedimentary rocks, typically clays, shales, mudstones and
siltstones, in which the particles range in size from less than 0.06 mm
The chemical symbol for the element arsenic
Acid Soluble copper
A mineral that is made up of arsenic, iron and sulphur
A mineral that is made up of copper, up to 55% copper, with carbonate and water
Bank Cubic Metres, a measure of volume applied to unbroken rock
Statistical term for two peaks in a graph of values
An impure form of copper produced by smelting oxidised copper ores or impure
scrap, usually in a blast furnace. The copper content varies widely, usually in the
range of approximately 60 to 85% by weight
Bottom of complete oxidation
A mineral made up of copper, up to 63%, copper, iron and sulphur
A minor structure arising from tensional forces, resulting in an appearance in crosssection similar to that of a string of sausages
Describes rock made up of angularly broken or fractured rock generally indicating a
fault plane
Bond Mill Work index
Temperature measurement in degrees Celsius (also called Centigrade)
Rocks made up mainly of a metal, commonly calcium or magnesium or copper, zinc
and lead and carbon dioxide
A rare mineral that is made up of cobalt, copper and sulphur
Counter Current Decantation
A term applied to the three dimensional volume used in the mathematical
modelling by computer techniques of ore bodies
A mineral that is made up of copper, up to 80% copper and sulphur
A mineral that is made up of copper, up to 35% copper, iron and sulphur
A mineral that is made up of copper, up to 36% copper, silica and water
Rocks formed from fragments of pre-existing rocks which have been produced by
the processes of weathering and erosion, and in general transported to a point of
deposition
Centimetre
Calcaire a Minerais Noirs (limestone and dolomite with black oxides)
The chemical symbol for the element cobalt
A sedimentary rock made up of various size particles from small pebbles to large
boulders and rounded other rock fragments cemented together
P a g e | 24
Cu
CuOx
cuprous
cut-off
DC
DCF
Datamine
diagenetic
dilution
disseminated
dolomite
domain
DRC
DStrat
DTD
DTM
Dwi
E
EAF
Écaille
EGL
EIA
EMP
EW
FC
ferric
fluvial
fluvio
flotation
framboidal
g
GAC
Gécamines
geostatistics
GRAT
GST
ha
The chemical symbol for the element copper
copper in the oxide form, generally soluble in dilute sulphuric acid
copper in ionic state of one missing electron
The minimum concentration (grade) of the valuable component in a mass of rock
that will produce sufficient revenue to pay for the cost of mining, processing and
selling it
Direct Current
Discounted Cash Flow
A proprietary computer program developed to model, view, report and analyse
geological and mining data
Pertaining to the processes affecting a sediment while it is at or near the Earth’s
surface, i.e., at low temperature and pressure
A term used to describe the waste or non economic materials included when
mining ore
Ore carrying fine particles, usually sulphides scattered throughout the rock
A mineral containing calcium, magnesium and carbonate
A term used mainly in mineral resource estimation or geotechnical investigations to
describe regions of a geological model with similar physical or chemical
characteristics
Democratic Republic of Congo
Dolomies Stratifies (stratified dolomite)
Direct tailings disposal
Digital Terrain Model
Drop Weight index
Easting coordinate
Electric Arc Furnace – a smelting facility
A French term meaning ‘fragment’, used to describe the large blocks of prospective
Mines Series stratigraphy that appear to ‘float’ in a mega-breccia-type
arrangement
Effective Grinding Length
Environmental Impact Assessment
Environmental Management Plan
Electrowinning
Congolese Francs
Iron in an ionic state of three missing electrons
A geological process in, or pertaining to, rivers
A description applied to moving material by streams of water
A widely used process to concentrate valuable minerals after mining that treats
finely ground rock in a water based pulp with chemicals that allow them to float to
the surface where they are recovered in preference to waste or gangue minerals
which sink
Akin to the skin of a strawberry or raspberry
Gram
Gangue acid consumption
La Générale des Carrierés et des Mines, Parastatal copper Mining Company of the
DRC
A mathematical method based on geological spatial knowledge of grade
distributions used to estimate mineralisation grades
Grey Roches Argilo-Talcqueuse (a dolomitic and talcose argillaceous rock)
Goods and Services Tax
Hectares
P a g e | 25
HAZOP
HDPE
HG
HLS
HMS
HQ3
hrs
HT
HV
ICP
ICWi
2
ID /IDS
IT
JORC
kg
kL
km
kt
kV
kW
kWh
kriging
kurtosis
L, l
L/sec, L/s, l/sec, l/s
lacustrine
lb
LIDAR
LOB
Log
LOM
LV
m
mm
m%
m3
Ma
malachite
mamsl
massive
MAX
mbgl
Hazard and Operability Study
High Density Polyethylene
High Grade
Heavy Liquid Separation
Heavy Media Separation. A process that uses high density fluids to separate
valuable minerals from waste or gangue by exploiting differences in specific gravity
Diamond drill core with a diameter of 63.5 mm
Hours
High tension
High voltage
Inductively Coupled Plasma Mass Spectrometry
Impact Crushing Work index
Inverse Distance Squared (method of estimating grades by mathematically
weighting samples based on their distance away from the estimation point)
Information technology
An acronym for Joint Ore Reserve Committee, an Australian committee formed by
the Australian Stock Exchange and Australasian Institute of Mining and Metallurgy,
the purpose of which is to set the regulatory enforceable standards for the Code of
Practice for the reporting of Mineral Resources and Ore Reserves
Kilogram
Kilolitre
Kilometre
Kilotonne
Kilovolt
Kilowatt
Kilowatt hour
A geostatistical method (named after the South African, D. G. Krige) of estimating
the unknown grade of resource blocks from the grades of samples, taking
cognizance of the sample distribution
Statistical term for peaked graph shape (peakedness)
Litres
Litres per second
Sediment deposition in lakes
Pounds
Light Detection and Ranging – a remote sensing system used to collect topographic
data
Lower Orebody
Natural logarithm to the base 10
Life of Mine
Low voltage
Metre
Millimetre
Metre percentage (obtained by multiplying metres by % of assay value)
Cubic metre
Mega annum (Million years)
A mineral containing copper, up to 57% Cu, carbonate and water
Metres above mean sea level
A term used to describe a large occurrence of a pure mineral species, often with no
structure
Maximum
Metres below ground level
P a g e | 26
mbs
MCC
MCK
Mg
MIR
MIN
MINDIL
mineralisation
mm
ML
MN
MODFLOW
MPa
Mt
MVa
MW
N
Neo-Proterozoic
NI
OC
ore
orogeny
P80
Pb
PBC
PDT
PE
PFDs
PFS
P&IDs
pH
PLC
PLS
ppm
pseudomalachite
PVC
QAQC
raffinate
RAT
RC
recovery
Metres below surface
Motor Control Centre
Mining Company of Katanga
Milligrams
Milling in raffinate
Minimum
A Whittle Four-X mine planning software term for mining dilution
The presence of minerals of possible economic value or the description of the
process by which the concentration of valuable minerals occurs
Millimetre.
Millions of litres
Magnetic North.
A groundwater modelling program used to assess the impact on the regional
groundwater table of pumping and abstraction, and also contaminant flow
Millions of Pascals
Millions of tonnes
Millions of Volt Amps
Millions of Watts
Northing Coordinate
The term used in the geological time scale for the period from 545 million years
ago to 1000 million years ago
National Instrument
Organic Continuous
A natural aggregate of one or more minerals which, at a specified time and place,
may be mined and sold at a profit or from which some part may be profitably
separated
Greek for ‘mountain generating’ - the process of mountain building. Orogenic
events occur as a result of plate tectonic processes
80% of product passes
The chemical symbol for the element lead
Pinned Bed Clarifier
Phase Disengagement Time
Permis d’Exploitation (Exploitation Permit or Licence)
Process Flow Diagrams
Pre-feasibility Study
Piping and Instrumentation Drawings
Concentration of hydrogen ion
Programmable Logic Controller
Pregnant Liquor Solution
Parts per million (same as grams per tonne)
Pseudomalachite or ‘false malachite’ – named because it is visually similar in
appearance to malachite
Polyvinyl chloride
Quality Assurance and Quality Control
A liquid stream that remains after the extraction with the immisciable liquid to
remove solutes from the original liquor. From French: raffinere, to refine.
Roches Argilo-Talcqueuse (a dolomitic/talcose argillaceous rock)
Reverse circulation (as in drilling)
A measure in percentage terms of the efficiency of a process, usually metallurgical,
in gathering the valuable minerals. The measure is made against the total amount
of valuable mineral present in the ore
P a g e | 27
reserve (Ore Reserve)
resource (Mineral Resource)
RL
Roan Supergroup
RMWi
ROM
RSA
RSC
RSF
S
s, sec
SAG
sandstone
SCADA
SD
SEM
SG
siltstone
silica
SMC
SNEL
SPLP
S/S, SS
storativity
stratiform
strings
supergene
SURPAC
SX
SX-EW
t
TCu
termitaria
TN
TOFR
tpa
The term for the economic quantities and grade of valuable materials as strictly
applied in compliance with the definition in the Australian JORC Code and in the
Canadian National Instrument (NI) 43-101
The term for the estimate of the quantities and grade of valuable materials but
with no economic considerations as strictly applied in compliance with the
definition in the Australian JORC Code and in the Canadian National Instrument (NI)
43-101
Reduced Level (same as elevation coordinate)
Describes the stratigraphic succession of sedimentary rocks of Neo-Proterozoic age,
in the Katanga Province of the Democratic Republic of Congo
Rod Mill Work index
Run-of-Mine (ore)
Republic of South Africa
Roches Silicieuses Cellulaires (siliceous rocks with cavities)
Roches Siliceuses Feuilletees (foliated and silicified dolomitic shales)
South Coordinate.
Second
Semi-autogenous Grinding
A sedimentary rock consisting of sand size grains, generally the mineral quartz,
which is in a consolidated mass
Supervisory Control and Data Acquisition System
Shales Dolomitiques (dolomitic shales)
Scanning Electron Microscopy
Specific Gravity
A sedimentary rock consisting of grains from 0.063 to 0.25 mm, generally the
mineral quartz and clay, which is in a consolidated mass
A compound of silicon and oxygen, generally occurring in the form of a mineral
called quartz
SAG mill comminution
Société Nationale d’Electricité – the provider of electrical power in the DRC
Simulated Precipitation Leach Procedure
Stainless steel
The volume of water an aquifer releases from or takes into storage per unit surface
area of the aquifer per unit change in head
Describes a layered or tabular shaped body of mineralized rock within a
sedimentary rock and implies that the layering of the mineralisation is parallel to
the bedding planes in that sedimentary rock
A term used to a digital line drawn within a computer program that outlines or
describes a shape of an object or interpretation
Pertaining to that part of an ore deposit in which the mineralisation has been
increased as a result of the downward percolation of fluids carrying metal in
solution
A proprietary computer program developed to model, view, analyse and report on
geological and mining data
Solvent Extraction
Solvent Extraction and Electrowinning
Metric tonne
Total copper
Termite mounds
True North
Top of fresh rock
Tonnes per annum
P a g e | 28
tpd
tph
transmissivity
TSF
TSS
UCS
UTM
V
VAT
VESDA
VSD
%v/v
W
Whittle Four-X
WNW
WRD
%w/w
Zn
μm
Tonnes per day
Tonnes per hour
The volume of water flowing through a defined cross-sectional area of an aquifer
Tailings Storage Facility
Total Suspended Solids
Unconfined Compressive Strength
Universal Transverse Mercator grid
Volts
Value Added Tax
Very Early Smoke Detection and Alarm
Variable Speed Drive
Percent by volume
Westing Coordinate
A mine planning software program used to optimise resource models, based on
economic and mining/processing parameters
West North West
Waste Rock Dump
Percent by weight
The chemical symbol for the element zinc
Microns, micrometers
P a g e | 29
3. RELIANCE ON OTHER EXPERTS
This Technical Report has been prepared and approved under the supervision of Mr Andrew Law,
Director Mining, Optiro Pty Ltd. Mr Andrew Law, who is the principal author of the report, is an
independent Qualified Person as defined in National Instrument 43‐101.
In preparing this report, the Qualified Persons have relied upon and taken responsibility for the
information provided by MWL relating to mining, legal, environmental and financial information as
noted below:






Legal title to the tenements held by MWL in the DRC and MWL’s permits to mine, which is
relevant to Sections 4 and 20 of this report.
Environmental permit and bond information which is relevant to Sections 4 and 20 of this
report.
The nature and validity of any off-take agreements for concentrate held by MWL, which is
relevant to Section 19 of this report.
Financial and cash flow models were provided to Optiro by MWL which is relevant to Section
22 of this report.
Metallurgical balance and current production information leading to the assessed head
grade of the copper-silver concentrate produced from treatment of the mined ore, which is
relevant to Sections 13 and 17 of this report.
Mine design, geotechnical, hydrology, planning, scheduling and costing which is relevant to
Sections 15, 16, 21, and 22 of the report.
The Qualified Persons have made all reasonable inquiries to establish the completeness and
authenticity of the information provided. Drafts of this report were provided to MWL with a request
to identify any material errors or omissions prior to filing. Notwithstanding the reliance of the
Qualified Persons on MWL for the financial and cash flow models, metallurgical balance information
and mine design noted above, the Qualified Persons accept responsibility for all of the
scientific/technical information related to these matters.
P a g e | 30
4. PROPERTY DESCRIPTION AND LOCATION
4.1. DEMOGRAPHICS AND GEOGRAPHIC SETTING
The Democratic Republic of the Congo (DRC) is located in central Africa and straddles the equator.
The DRC has an east-west lateral extent of approximately 1,500 km and extends over a north-south
distance of some 1,800 km. The DRC is Africa’s second largest country, covering an area of
approximately 2.3 million km2 and shares land borders with Angola, Zambia, Rwanda, Tanzania,
Uganda, the Republic of the Congo, Sudan, Burundi and the Central Africa Republic. The capital city
is Kinshasa, which is located in the western portion of the country. The DRC’s main port is Matandi,
approximately 115km from the coast on the Congo River.
The DRC has a population in excess of 75 million of which approximately 50% are aged between 15
and 64 years old. There are over 200 African ethnic groups within the country’s borders, although
the Bantu and Hamitic groups account for approximately 45% of the population. The majority of the
population reside in rural areas with one-third living in urban centres.
Christianity is the dominant religion in the DRC, with approximately half of the population being of
the Roman Catholic faith, with a further 20% Protestant. The remaining population follow the
Kimbanguist (10%), Muslim (10%) and other (10%) faiths.
The national language is French, although Lingala, Kingwana, Kikongo and Tshiluba are widely
spoken.
4.2. PROJECT OWNERSHIP
The Dikulushi mine is governed by the “Dikulushi Mining Convention”, signed on the January 31,
1998 with the Government of the DRC, and ratified by Presidential Decree issued on February 27,
1998.
The Dikulushi Mining Convention is a mining concession granted to AMC. Mawson West
Investments Ltd a wholly owned subsidiary of MWL, holds 90% of the issued capital of AMC, the
remaining 10% is held by the Dikulushi – Kapulo Foundation (NPO).
For the purposes of this report, the Mawson West Limited ownership structure, referred to as MWL,
is used in this report for ease of reference.
4.3. PROPERTY LOCATION
The Project is located within the Katanga Province in the south-eastern DRC, some 400 km north of
Lubumbashi and 50 km north of the regional town of Kilwa. The Project is centred at approximately
S 08° 53’ E 28° 16’, some 25 km west of Lake Mweru near the DRC border with Zambia.
Figure 4.1 shows the property location of MWL’s holding within the DRC, which are effectively two
distinct properties – the Dikulushi property (shown in green in figure 4.1) and the Kapulo property
(shown in blue in figure 4.1). The focus of this report and the projects discussed herein, relate
specifically to the Dikulushi property only.
P a g e | 31
Figure 4.1
Exploration Licences of the Dikulushi copper silver project
4.4. THE PROPERTY TENEMENT AREA
MWL holds title to the Dikulushi mine and surrounding exploration tenements, as governed by the
Dikulushi Mining Convention. Under the Dikulushi Mining Convention the exploration tenements
known as “PR’s” were issued for an initial five year period and are renewable a further three times,
each time for a period of five years; that is a total of 20 years. The Dikulushi PR’s shown in Table 4.1
below and Figure 4.1 above, and these were first granted on the 22 May 2001 and currently have
“renewed” expiry dates April 2016. A further 5 year renewal period is available post this date. MWL
currently holds 18 Exploration Permits and three Exploitation Permits under the Dikulushi Mining
Convention, covering 7,283km².
Under the Dikulushi Mining Convention, MWL is guaranteed sole and exclusive rights for exploitation
for a period totalling 20 years from the date of the issue of the permit.
P a g e | 32
Mining operations at the Dikulushi mine are conducted under an Exploitation Permit PE 606, issued
on 29 December 2003 by Ministerial Decree. The exploitation permit recognised that AMC had
commenced mining operations form 31 January 2002. The PE covers an area of 40.77 km2 over the
Dikulushi mine area (Figure 4.1 and Figure 4.2).
Table 4.1
Mawson West Limited Dikulushi tenement schedule
Tenement Schedule
Project
Group Entity Permit No.
Type
Granted
Expiry
Dikulushi
AMC
PE606
Area km²
40.77
Mining
29-Dec-03
30-Jan-22
Dikulushi
Dikulushi
AMC
AMC
PR546
PR1693
283.8
398.6
Exploration
Exploration
23-May-11
12-Apr-11
22-May-16
11-Apr-16
Dikulushi
Dikulushi
AMC
AMC
PR1694
PR1700
398.5
398.4
Exploration
Exploration
12-Apr-11
12-Apr-11
11-Apr-16
11-Apr-16
Dikulushi
Dikulushi
AMC
AMC
PR1703
PR1705
398.3
237.0
Exploration
Exploration
22-May-11
22-May-11
21-May-16
21-May-16
Dikulushi
Dikulushi
AMC
AMC
PR1706
PR1707
398.0
397.7
Exploration
Exploration
22-May-11
23-May-11
21-May-16
22-May-16
Dikulushi
Dikulushi
Dikulushi
AMC
AMC
AMC
PR1708
PR1709
PR1710
405.1
345.0
397.0
Exploration
Exploration
Exploration
22-May-11
22-May-11
22-May-11
21-May-16
21-May-16
21-May-16
Dikulushi
AMC
PR1711
396.9
Exploration
22-May-11
21-May-16
Total Area
Figure 4.2
4,495.1
Dikulushi mine infrastructure within the PE 606
P a g e | 33
4.5. ENVIRONMENTAL PERMITS
An EIA for the Dikulushi project was lodged in 2003. In 2009, an EIA for the underground project was
submitted to the DRC Government. Both of these reports were compiled by African Mining
Consultants of Kitwe, Zambia, an environmental company that was licensed to work and report in
the DRC. An EIA was lodged for the cutback project, prepared by EMIS sprl, a DRC environmental
company licences to work and report in the DRC. All three environmental reports received DRC
Government approval. A revised EIA, extending underground mining beyond 2013, has being
submitted to Government.
Each EIA includes commitments relating to mine decommissioning. Annual reporting of
environmental issues and measurements to relevant government bodies is a condition of the
operating license and EMP.
MWL have lodged $1.19M as an Environment Bond. The financial guarantee is a contribution
towards an estimate of the total costs of closure, rehabilitation and re-vegetation of the Dikulushi
mine. The development of the financial guarantee is conducted in compliance with:



Articles 410 of the Mining Regulations
Articles 124 and 125 of Appendix XI of the DRC Mining Regulations 2003; and
Appendix II of the Mining Regulations 2003
Regular environmental audits are carried to determine the mine’s compliance with its Environmental
Management Plan.
An environmental monitoring database is maintained at the mine, comprising the following:








wet/dry, min/max temperatures
rainfall
dust exposure
noise levels
ground and surface water quality
groundwater levels
Tailings Dam piezometer water levels
light levels.
A study into the acid rock drainage potential of the process plant tailings was conducted in 2005 and
they were classified as low risk. Ongoing testwork and monitoring continues to support this
conclusion.
P a g e | 34
Figure 4.3
Dikulushi mine site aerial view
P a g e | 35
5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,
INFRASTRUCTURE AND PHYSIOGRAPHY
5.1. ACCESS
Access to the Dikulushi Mine is by sealed road from Lubumbashi to Kasenga, along the Luapula River
by boat to Kilwa and then approximately 54 km by gravel road from Kilwa to Dikulushi. The total
travelling distance is approximately 500 km. The closest international airport is at Lubumbashi,
approximately 450 km to the south. A gravel airstrip is located at the Dikulushi mine and charter
flights using a G1 plane (as shown in Figure 5.3) from Lubumbashi can land directly at site. Supplies
for the project are typically trucked on sealed roads from South Africa via Botswana to Nchelenge
port on the Zambian side of Lake Mweru. Supplies are then transferred from Nchelenge to Kilwa on
the Congo side of Lake Mweru on a 340 t capacity barge (Figure 5.1) owned by AMC; the water
journey takes 5 hours. Access from Kilwa port to the mine is via a 54 km gravel road and takes
approximately 1 hour by light vehicle.
Figure 5.1
Barge on Lake Mweru
5.2. SITE TOPOGRAPHY, ELEVATION AND VEGETATION
The Dikulushi deposit is located on a plateau approximately 1000 m above sea level. The area
surrounding the Dikulushi site is almost entirely covered with woodland and forest, with some
swamps or wetland areas. The plateau rises into the Kundelungu ranges 60 km to the west of
Dikulushi and forms an escarpment 25 km to the east along the fault-bounded edge of Lake Mweru.
A minor ephemeral stream is located near the Dikulushi mine site. The Luapula River is the main
drainage into Lake Mweru and both form the international boundary between Zambia and the DRC.
P a g e | 36
5.3. CLIMATE, PHYSIOGRAPHY, LOCAL RESOURCES AND
INFRASTRUCTURE
The average annual rainfall, as indicated by mission records, is 1,260 mm, with a range of 800 mm to
2,200mm. An Oregon Scientific weather station was installed at Dikulushi in 2006. A composite
graph of the weather data collected at Dikulushi over the past 5 years is shown in Figure 5.2. The
wet season begins towards the end of October and finishes at the end of April, with 90% of the
annual rainfall occurring during this period. The average minimum recorded temperature is 15°C
and the average maximum temp is 29°C during the year.
Figure 5.2
Composite Temperature and Rainfall Data recorded over the last 5 years.
35
200
180
30
160
140
120
20
100
15
80
Rainfall, mm
Temperature, °C
25
Rainfall
Temperature (High)
Temperature (Low)
60
10
40
5
20
0
0
Jan
Feb
Mar
Apr
May
Jun
Jul
Aug
Sep
Oct
Nov
Dec
The wet season generally has minimal effect on mining or processing operations at Dikulushi.
5.4. SURFACE RIGHTS
The Dikulushi mine is based on Exploitation Licence (PE606) granted on 29 December 2003. The
lease is valid for 20 years and can be renewed for up to a further 20 years.
There are no competing mining rights (for example, small artisanal mining licenses) in the project
area.
5.5. SITE INFRASTRUCTURE
The development of the Dikulushi mine has required development of seven major locations:
1. the treatment plant area, which includes the mine administration building
2. the mine services area, including workshops, fuel farm and powerhouse
3. the explosives storage area
P a g e | 37
4.
5.
6.
7.
the staff village
the airstrip
the process water dam
tailings storage facility.
These items of infrastructure are depicted in Figure 4.2 and Figure 4.3. This infrastructure was in
place for the previous operations under Anvil and has been used for the open pit cut back operations
with it being well established and maintained, as well as being of sufficient size for the current and
future underground requirements.
Figure 5.3
Dikulushi airstrip and the G1 Charter plane provides safe staff transportation to and from site
5.5.1.
WATER SUPPLY
Mine water is sourced from a raw water dam located adjacent to the Tailings Dam. Supernatant
tailings water is reclaimed via penstock arrangements for use in the processing plant. A water
supply flowchart and site wide water balance is provided in section 18 of this report.
Potable water is supplied from various bores on the property which are tested regularly.
5.5.2.
POWER SUPPLY
The project is located in a remote area where there is no electrical utility grid. The mine power is
supplied by diesel generators. There is sufficient back-up capacity.
The existing power station at Dikulushi comprises the following generators: 4 x 1.2 MW FG Wilson (
being new units and installed during the 3rd quarter 2013 ) , 1 x 2.0 MW Caterpillar, 1 x 1.6 MW
Caterpillar, 1 x 0.8 MW Mirrlees for a current total capacity of 9.2 MW. The current power demand
for the plant and infrastructure is in the order of 2.0 MW. The 2.0 MW Caterpillar, 1.6 MW
Caterpillar currently require major overhauls which will be completed during 2014. The 1 x 0.8 MW
Mirrlees will be decommissioned during the 4th quarter of 2013. The new FG Wilson generating sets
were installed to supply power to the operations as well as dewatering of the underground and
P a g e | 38
normal underground operations. The 2.0 MW Caterpillar and the 1.6 MW Caterpillar will be used as
backup standby power.
MWL recognises that a consistent reliable fuel supply is crucial to the success of the Dikulushi
operation. The operation currently uses approximately 450,000l of diesel per month. This fuel is
supplied by three DRC based companies, two receive supplies from the port of Beira and the other
receives supplies from the port of Dar es Salaam. MWL has contacted a further supplier from Dar es
Salaam whom would be able to supply fuel to Dikulushi. MWL is regularly speaking to suppliers to
guarantee no interruptions in fuel supply. MWL believes that it has mitigated the risk of fuel supply
by having a number of suppliers whom source fuel from different ports and transport routes.
5.5.3.
MINE PERSONNEL
As at June 2013, the Dikulushi mine employed 515 people, of which 39 were expatriates. The
requirements for the underground operations and other associated activities will require a total
workforce of 500-550 employees. Contractors will be used as required. This a change from the open
pit operations where the workforce was mainly contractor supplied.
5.5.4.
TAILINGS STORAGE FACILITY
There are currently three tailing storage facilities (TSF) on site. The initial TSF designed for HMS
tailings, dormant since 2004, has had a section of the coarse portion reclaimed and retreated in
early start up operations by MWL. The second TSF is dormant whilst the third is in use to
accommodate the tailings resulting from the treatment of the current open pit operations. The third
TSF has been reviewed for extended use beyond its current life. This will be raised to accommodate
tailings resulting from the final open pit cut back mining operations and the planned underground
operations.
More detail on the TSF is covered in Section 17.
5.5.5.
ADMINISTRATION AND PLANT SITE BUILDINGS
The infrastructure on site includes administration offices (Figure 5.4), a warehouse, mining
equipment and maintenance workshops, mechanical workshops and a service area with access pit
for inspection and repair of vehicles.
There is a fully equipped clinic on site (Figure 5.6) and a hospital at Kilwa, approximately 50 km from
the mine. An assay laboratory on site facilitates metallurgical, exploration and grade control
sampling assaying requirements.
P a g e | 39
Figure 5.4
Dikulushi Administration Centre
5.5.6.
ACCOMMODATION
A staff village has been constructed 1.8 km from the process plant. A mess hall, fully equipped
kitchen, food storage and laundry facilities serve all employees. Recreational facilities are also
available to employees. Figure 5.5
Figure 5.5
Dikulushi Camp Site
P a g e | 40
5.5.7.
COMMUNICATIONS
Mobile phone coverage is available through a dedicated mast located on top of the waste dump.
There are satellite systems for data transmission and VOIP telephone coverage. There is a base
station radio system, along with vehicle and hand-held radios. Figure 5.6
Figure 5.6
Dikulushi Clinic and communications centre
5.5.8.
MOBILE EQUIPMENT
Sufficient mobile equipment for the efficient running of the operations is in place, comprising light
vehicles (including an ambulance), light trucks, forklifts, buses and generators. Figure 5.7
P a g e | 41
Figure 5.7
Dikulushi Workshop
5.5.9.
SECURITY
Security is provided by a contractor. Appropriate secure facilities are provided for the storage of fuel
and explosives. Figure 5.8
P a g e | 42
Figure 5.8
Dikulushi Store and Fuel Farm
P a g e | 43
6. HISTORY
The development of exploration and mining at Dikulushi and surrounds can be broken down into the
following periods:
•
Early History – Copper mineralisation was first reported in the early 20th Century by
Simkat. Other assessments were made in the 1950s and by the French Bureau de
Recherches Geologiques et Minieres (BRGM) during the 1970s.
•
Recent History – Anvil, from 1996 to April 2008.
•
Current – Mawson West Limited from June 2010 to present.
The history of the project is summarised in Table 6.1.
Table 6.1
Historical work summary at the Dikulushi Project
Year
1900s
1974-1981
1996-1997
1998-2002
2002-2006
2006-2008
2008
Supervision
Belgian explorers
BRGM
Anvil
Anvil
Anvil
Anvil
Anvil
2010
onwards
MWL
6.1.
Work Completed
Rock chips
48 Diamond drillholes
Dikulushi Mining Convention signed and some drilling
Modified convention signed, more drilling and metallurgical testwork, feasibility study
Mining of open cut and regional soils and termite sampling and drilling
Underground development at Dikulushi and further drilling of targets defined from above
Anvil closes down Dikulushi Mine.
Restarted plant on LG stockpiles and then proceeded to process ore from the pit cut back
operation until July 2013 when the pit concluded operations due to reaching its design
limits.
BELGIAN EXPLORATION
Copper mineralisation in the area was initially evaluated by Belgian explorers (Simikat) from 1910
until 1923.
BRGM purchased an interest in the Dikulushi Deposits during the early 1970s and completed adit
sampling, diamond drilling, metallurgical testwork, soil geochemistry and geophysics. The projects
lay dormant until Anvil pegged the ground in the late 1990s and subsequently signed the JV
agreement with Mawson West.
6.2.
ANVIL MINING LTD
An open pit mine was commissioned at Dikulushi in October 2002 by Anvil, with run-of-mine ore
delivered to an on-site heavy media separation (HMS) concentrator at the rate of 250,000 tonnes
per year. The copper-silver concentrate was subsequently transported by barge across nearby Lake
Mweru into Zambia and then by road to smelters in South Africa and Namibia.
During the first 15 months of operation, the geology within the open pit was extensively mapped
and, with results of the drilling, resulted in a re-interpretation of the mineralised envelope at
Dikulushi.
The DevMin consulting group was approached by Anvil to undertake the Open Pit Mine Plan study to
estimate the remaining open pit reserves and prepare an open pit life-of-mine schedule for the
Dikulushi Mine. This study was initiated in August 2003 when a preliminary pit optimisation was
P a g e | 44
undertaken using Anvil’s current in-house resource model. The initial study was largely an update of
previous work carried out by DevMin and showed that with current economic parameters, the
feasible depth for open pit mining at Dikulushi could potentially extend down to 150 m below
surface. An earlier work phase, carried out in 2002 prior to the mine being commissioned, had
stopped the pit at 130 m depth with remaining mineralisation proposed to be exploited by an
underground mine.
Further metallurgical work was carried out by Anvil in 2003, resulting in the installation of a ball mill
and flotation plant in mid-2004 increasing the throughput to 350,000 tonnes per year. A second ball
mill was installed in June 2005 with a further increase in throughput to 520,000 tonnes. This is the
current capacity of the treatment plant and the mining fleet was increased to provide the
appropriate plant feed tonnage.
With the completion of the planned open cut operations in 2006, the focus moved to developing an
underground operation. The ROM stockpile at the end of 2006 was considerable and it was
expected that this would see the mine through to the commencement of production from the
underground in early 2008.
Upon commencement of underground mining, and due to time constraints, Anvil adopted a sublevel caving method for the underground. However with the realisation that such a method was
inappropriate for the type of orebody geometry at Dikulushi, the decision was made to move to an
Avoca method of mining. This necessitated a hiatus in ROM feed to the plant in April 2008 due to
the intensive nature of development required for a bottom-up mining method.
By the end of 2008, with the World financial markets in turmoil and the subsequent plunge in the
copper price, the decision was made to place the Dikulushi Mine on care and maintenance.
6.3.
MAWSON WEST
MWL acquired the Dikulushi project from Anvil in April 2010. Plant refurbishment was started
immediately and completed in July 2010 at which point MWL started processing the LG stockpile
which continued into 2012.
MWL has continued with the cut back of the open pit until July 2013, where practicable completion
of the pit was reached. Ore is stock piled on the ROM pad to continue the milling operations until
late 2013. The next phase is the re-establishment of the underground operations below the current
open pit (825 mRL). Previously, the underground was developed by Anvil down to the 750 mRL.
6.4.
RESOURCE HISTORY
During the BRGM tenure of the deposit (1974 – 1981) a resource of 1.65 Mt at a copper grade of
10.46% and a silver grade of 310g/t to a vertical depth of 220 m was estimated for the Dikulushi
deposit. Anvil published a Mineral Resource estimate for the Dikulushi deposit in December 2006.
The estimate was completed by FinOre Mining Consultants (FinOre) in July 2006 and used 3D
wireframe volumes to define the mineralisation. A 0.5% copper mineralisation cut-off was used to
guide the wireframe volume. Estimates were completed using Datamine software and the resulting
Mineral Resource is detailed in Table 6.2 below.
P a g e | 45
Table 6.2
Mineral Resource estimate as completed by FinOre in July 2006 and published in December 2006; a cut-off grade of
1.5% copper was used
Tonnes
Cu(%)
Cu Metal
(Tonnes)
Ag g/t
Ag Metal
Mozs
Measured
410,000
9.10
37,300
288
3.8
Indicated
Measured and
Indicated
650,000
7.90
51,100
184
3.84
1,060,000
8.30
88,400
224
7.64
Inferred
1,380,000
5.80
80,600
141
6.26
Category
There are no other historically published Mineral Resources available for Kazumbula or any of the
surrounding exploration targets.
6.5.
PRODUCTION HISTORY
Anvil mined the Dikulushi deposit between 2002 and 2008. Production statistics are presented in
Table 6.3 below.
Table 6.3
Historical Anvil production for the Dikulushi mine
Year
Tonnes
processed
Grade Cu%
Grade Ag g/t
2002 (HMS)
2003 (HMS)
2004 (HMS/Float)
2005 (Float)
2006/07 (Float)
2008 (No Dec)
36,010
273,500
245,000
410,000
1,059,950
471,590
7.26
7.68
6.39
5.07
5.12
3.16
144
195
177
149
151
84
Average
recovery
(%)
66.7
66.1
69
86
81
75
Cu
produced
(tonnes)
1340
13,613
10,840
16,900
46,507
11,177
Ag produced
Mozs
0.08
1.16
0.89
1.63
4.45
0.97
From 2010 MWL recommenced processing operations of low grade stockpiles whilst the cut back of
the open pit commenced mining in Jan 2012 and ceased July 2013, with ore processing due to be
completed late 2013. Production statistics for this period are presented in Table 6.4 below.
Table 6.4
Recent MWL production for the Dikulushi mine
Year
Tonnes
processed
Grade Cu%
Grade Ag g/t
2010/11
2011/12
2012/13*
467,958
347,863
355,043
1.46
1.80
5.21
35.31
34.2
141.5
Average
recovery
(%)
64.0
63.0
91.6
Cu
produced
(tonnes)
4251
3,948
16,925
Ag produced
Mozs
0.36
0.27
1.46
*Note: 2013 Production is to the end of July 2013 (13months).
[Note: These figures are available in table 17.2]
P a g e | 46
7. GEOLOGICAL SETTING AND MINERALISATION
7.1. REGIONAL SETTING
The Dikulushi and Kazumbula copper-silver deposits are located west of Lake Mweru (08°53’37”S;
28°16’21”E), near the eastern margin of the Katanga sedimentary basin, in an area known as the
Kundelungu Plateau. Deformation of the Katanga Supergroup sedimentary rocks is mild, and
comprises open upright folds developed in association with major north-northwest and subordinate
north-northeast-trending faults. An angular unconformity separates the undeformed, uppermost
subdivision of the Kundelungu Group (Plateau Series) from the remainder. Regionally, copper
mineralisation is known from at least three stratigraphic levels in the Kundelungu Group, but all
occur beneath this unconformity. These deposits are hosted by the Kalule Formation, and ore is best
developed in red sandstones and shales of the Mongwe member, above a conspicuous
reduction-oxidation (redox) and pH boundary with the grey carbonates of the underlying Kiaka
member. Mineralisation at Dikulushi was strongly fault-controlled (Haest et al., 2007), but across
the district many copper occurrences, including the nearby Kazumbula deposit, are associated with
the same stratigraphic position. Stratabound mineralisation is known from higher (Mwitapile; Sonta
member, Kiubo Formation) and lower in the stratigraphy (Lufukwe; Monwesi Formation; El Desouky
et al., 2007). In the greater Dikulushi district, prospective parts of the stratigraphy are exposed in
areas of low terrane from which the upper, un-mineralised sequences have been removed by
erosion. Mineralisation occurred during the waning stages of the Lufilian Orogeny (560 Ma) as
compression and ductile deformation gave way to extension and brittle deformation. Several
hundreds of millions of years later the Dikulushi deposit was chemically reworked and upgraded by
circulating groundwaters (Haest, 2009; Haest et al., 2010). Figure 7.1 depicts the regional geology of
the Dikulushi district and Figure 7.2 summarises the regional stratigraphic sequence.
P a g e | 47
Figure 7.1
Regional Geology of Mawson’s convention area in the DRC
P a g e | 48
Figure 7.2
Super Group
Stratigraphy of Dikulushi region with known styles of mineralisation
Group
Formation
Plateaux
Ks-3
Upper Kundelungu - Ks
Kiubo
Ks - 2
Lithology
m-Approx
thickness
Kilunga Lupili
Arkose
Pink Arkose
<200
Kapenga Schists
Interbedded red argillaceous
sandstone and red argillates
350
Sonta Sandstone
Red sandstone and quartzite
100
Sampwe schists
300
Kyafwama Kombo Sandstone
Red sandstone
50
Lufila schists
150
Kiubo Sandstone
Red sandstone
50
Mongwa Schists
Argillaceous red sandstones
interbedded with argillates.
Minor Sandstone
<100
Lubudi
Dolomites
Interbedded pink, cross
bedded dol-arenites,with olitic
caps
<50
Kanianga
Sandstone
Sandstone - argillaceous and
weakly calcareous
<50
Pink intramicrite flakestones
<50
Interbedded white/pink
carbonate muds and arenites
<50
Basal dolomite (BD)
4
Le Petit
Conglomerate
Diamictite
<100
Monwesi
Fluvio glacial sandstones
<50m
Le Grand
Conglomerate
Diamictite
500
Dolomitic shales and
sandstones
?
Kiaka
Carbonates Ks 1.2
Lusele Pink
Dolomites
Lower
Kundelungu Ki
Ks - 1.1
Roan R
Mineralization
Disseminated chalcocite and
malachite at base of Sonta
Sandstone
Ks - 2.2
Ks - 1.3
Kalule
Ks - 1
Unit
Monwesi
Likasi Ki-1
Ki -0 1.1
Mwashya
R-4
R - 4.2
Fault controlled Cu-Ag
mineralisation at the
dolomite/ sandstone contact
Zn-Pb mineralization at
contact with diamictite
Stratiform copper Kinkumbi,
Lufukwe anticline
P a g e | 49
7.2. PROJECT GEOLOGY
The Dikulushi and Kazumbula deposits are principally fault-hosted lodes of massive to semi-massive
copper sulphides with minor disseminated and stockwork mineralisation in the wallrocks. The
deposits are mostly hosted by red siltstones, with lesser ore hosted by underlying grey carbonate
rocks and fragmental rocks that mark the contact between the two. The mineralised Dikulushi Fault
trends east-northeast and breaches the northeastern nose of a north-northeast plunging elongate
anticline. This orebody therefore is suborthogonal to stratigraphy. Prior to mining at Dikulushi, it
was approximately 400 m long, 10 metres wide and extended from surface to a depth of at least 450
metres (Figure 7.3).
Figure 7.3
Local geology of the Dikulushi open pit
P a g e | 50
The Kazumbula orebody (Figure 7.4) comprises a fault-controlled lode of disseminated and minor
fracture hosted copper sulphides and oxides. The orebody is hosted by siltstones and intercalated
granular lithic sandstones. These have pervasive grey-green hydrothermal illite alteration when they
occur in the immediate wallrocks to the orebody, but are red-brown elsewhere. The mineralised
Kazumbula Fault trends east-northeast, but mineralisation is located at the intersection of the fault
with other structures trending north-northeast and north-northwest. These structures are
considered to have been active in concert as parts of a conjugate strike-slip shear array. Gently
folded stratigraphy on the eastern limb of the Kabangu Antiform abuts the fault plane. The orebody
is sub-orthogonal to stratigraphy and appears to be restricted to preferred brittle and permeable
stratigraphy. As it is presently known, the deposit is approximately 180 m long, 12 m wide and
extends from surface to a depth of approximately 80 m (Zukowski et al., 2010) as confirmed by
drillhole intercepts.
Figure 7.4
A typical vertical cross section through the Kazambula deposit, highlighting key geology associated with mineralisation
P a g e | 51
8. DEPOSIT TYPES
The Dikulushi and Kazumbula copper deposits are interpreted to be hypogene, fault-controlled
deposits, containing semi-massive and disseminated chalcopyrite-bornite mineralisation with
azurite, malachite and cuprite developed in the supergene zone. The mineralisation is lithostructurally controlled and is hosted in shales and sandstones of the Kundelungu Group. The host
sedimentary rocks show varying degrees of brecciation, with the highest grade zones comprising
semi-massive bornite-chalcopyrite fault fill. The Dikulushi and Kazumbula deposits are not typical of
the stratiform copper deposits which are common within the central African Copperbelt.
Dikulushi ore comprises massive to semi-massive and fracture-disseminated copper sulphide
minerals that filled open spaces and cemented breccias along the fault zone. Early mineralisation
was polymetallic and contained Cu-Fe-Pb-Zn and Ag as chalcopyrite, bornite, galena and sphalerite.
Subsequent remobilisation dissolved most of the Fe, Pb and Zn and led to upgrading of the coppersilver content of the ore. The central parts of the deposits commonly contain >10% Cu and >200
ppm Ag. As a result of this process, the deposits are now composed largely of the copper sulphide
mineral chalcocite. Silver occurs as atomic-level substitutions and as very fine grained inclusions in
chalcocite grains (Dewaele et al., 2006; Haest et al., 2009; Haest et al., 2010). Gangue minerals
typically associated with the mineralisation at Dikulushi are quartz, feldspar, pyrite and clays. Figure
7.3 depicts the mapped geology of the Dikulushi open pit.
P a g e | 52
9. EXPLORATION
The Dikulushi areas copper deposits have been explored by three main groups over the last 40 years.
The three main periods of exploration are:



Simkat - BRGM (1910-1981)
Anvil (1997 to 2008)
MWL (2010 - current)
9.1. BRGM
The Dikulushi deposit was discovered by Simkat around 1910 but no records of work completed are
available. In the period from 1910 to 1936 the deposit was investigated by Simkat and Societe
Miniere du Lac Moero. BRGM then appraised the deposit from 1974 to 1981, during which time 48
diamond holes for 5,223 m were drilled and a resource of 1.65 Mt @ 10.46% Cu and 310g/t Ag to a
vertical depth of 220 m was estimated. In 1988 the BRGM withdrew from Zaire (previous name for
the DRC) and the deposit was left dormant until Anvil took control in 1996.
9.2.
ANVIL MINING LTD EXPLORATION
In 1996 Anvil submitted a Mining Convention application to the Zaire government for the Dikulushi
deposit. The documents were signed in January 1997 and this document was then re-negotiated
with the new Kabila government in 1998, with a presidential decree issued the month after signing
ratifying the convention.
During the latter half of 1997, Anvil completed 26 reverse circulation drillholes and 18 diamond
drillholes which, together with the BRGM drilling, formed the basis for a pre-feasibility study
completed by Signet Engineering of Perth, Western Australia. Anvil subsequently carried out
additional drilling in 2000, 2003, 2004, 2005-6, 2007 and 2008, which is detailed in Table 9.1.
Table 9.1
Company
BRGM
Historical drilling summary for the Dikulushi copper silver project
Period
1974-1981
1997
2000
2003
Anvil
2004
2005-2006
2007
2008
Type
DDH
DDH
RC
RC
DDH
RC
RC/DDH
DDH
DDH
DDH
No. Holes
48
18
26
22
4
21
14
14
9
78
2
Metres
5226
2115
2305
786
885
1768
414/3811
5779
2061
4130.9
1251
Sequence
DIK1-47
DDH1-14, DR15, 19, 20, 38
DRC15-40
DRC043-064
DDH16-19
DRC065-085
DDH020-035 (no 032)
DDD38-052
DDD053-061
UGD001 to UGD97
SUR001&005
P a g e | 53
9.3.
MWL EXPLORATION
At Dikulushi MWL has completed some geotechnical drilling for pit cut back and underground
development studies, but this has not been sampled.
MWL has also carried out drilling at the Kazumbula deposit. Details of this are provided in Section
10.8.
P a g e | 54
10. DRILLING
10.1.
INTRODUCTION
All the drilling prior to 2010 was completed by Anvil. This section describes the various Anvil
programmes and the programmes carried out by MWL in 2010 which were relevant to the
Kazumbula deposit.
10.2.
ANVIL PROGRAMME 1997
In 1997 Anvil carried out an exploration programme at Dikulushi which aimed at confirming the
BRGM results and allowing estimation of Mineral Resources to the standard of the Australian JORC
Code. A total of 40 drillholes were completed using two drill rigs between September and October
1997. All holes were oriented at 60° towards an azimuth of 340° (grid north), and all holes were in
the area previously drilled by the BRGM. These resources and reserves were subsequently reviewed
in the light of the requirements of NI43-101, since Anvil was a listed company in Canada.
A total of 18 HQ and NQ diamond drill (DDH) holes (DDH1-14 and DDH15, 19, 20, 38), including 4
with reverse circulation (RC) pre-collars, were completed in the 1997 programme. Mineralisation
was intersected in 17 of the 18 holes. Downhole camera surveys were completed at least every 12
metres in the DDH holes and the results showed the DDH remained essentially straight with a
maximum deflection of 2° in declination and 4° in azimuth. A core orientation spear was also used
after every core run (usually 3 metres). Upon recovering the core it was oriented when possible, and
was then logged for geotechnical defects, core recovery and geology. Core recovery averaged about
90%, except for minor soil, sandy and cavernous zones. Recoveries in mineralised zones are
reported to have been about 90%.
A total of 22 RC drillholes (DRC16-18, 21-37, 39, and 40) were completed during the 1997
programme using a booster compressor, which was essential due to large water inflows and broken
ground. Mineralisation was intersected in 19 of the 22 RC holes. None of the wholly RC drilled holes
was downhole surveyed. Four RC pre-collared holes were surveyed throughout their length but due
to the presence of steel casing, only the cored section returned valid azimuth readings. The dip
variations in these holes were minor; however, the cored sections showed consistent anticlockwise
azimuth rotations of between 9 and 18°. Consequently, an average azimuth correction factor of 3.2°
anti-clockwise deviation per 20 metre downhole has been entered into the RC drillhole database to
compensate for this interpreted drillhole rotation.
Of the holes drilled during the 1997 programme, 11 DDH holes and 6 RC drillholes were specifically
collared to twin earlier BRGM DDH holes.
Anvil also cleaned out and re-sampled six trenches and four test pits that were originally dug and
sampled by the BRGM. A total of 90 channel samples and 18 rock-chip samples (pits) were taken
from 191 metres of trenching as well as 18 rock chip samples from pits. Orientation soil and stream
sampling and other regional reconnaissance sampling and exploration were also completed during
the 1997 programme.
P a g e | 55
10.3.
ANVIL PROGRAMMES 2002 & 2003
The drilling programmes carried out in 2002 and 2003 were partly undertaken in response to the
observation in Munro (1998) that as the “upper 30 metres of the deposit is poorly defined, infill
pattern diamond drilling is required”, along with the need to clarify the extent of the mineralisation
and to sterilise certain areas prior to the commencement of mining.
The diamond core drilling procedures used during the 2002 and 2003 programs were compatible
with those of the 1997 Anvil drilling program. A Stanley Drilling Longyear 38 Rig was used in the
2002-03 programmes. To avoid the need to reduce core size at depth to NQ, the first 60 m was
drilled using PQ diameter, followed by HQ. This practice ensures that a good volume of mineralised
intercept core is available for both assay and archive.
Core orientation procedures were routinely conducted, although drilling conditions at times limited
successful achievement of orientation results. Core recoveries across the mineralised horizons
typically exceeded 95%.
Regrettably, all archived core from the 1997 and 2002 campaigns was rendered useless after core
trays were overturned by army units during the latter stages of the civil war. Fortunately,
photographs of this core, taken for geotechnical purposes, are still available.
10.4.
The objective of the 2004 programme was to extend the resource down to 300 m below surface and
to provide data for a pre-feasibility study on an underground mine. 14 holes were pre-collared with
RC (414 m) and drilled to a maximum ore intercept depth of 280 m below surface (3,811 m).
The diamond core drilling procedures used during the 2004 programme were largely compatible
with those of the previous two programmes, although core recoveries were not as good due to
technical problems with the rig. Stanley Drilling was the contractor for both the RC and diamond
drilling.
Downhole surveys were carried out at 50 m intervals using an Eastman camera. Core orientation
was attempted using the spear method, but poor ground conditions rendered the data to be of little
practical use.
10.5.
ANVIL PROGRAMME 2005/6
The objective of the 2005/6 programme was to increase the confidence in the geological model and
upgrade the resource classification to a depth of 400 m below surface for a possible future
underground mining operation. Diamond drillholes from this programme were identified with the
prefix “DDD”.
The programme was drilled with one of Anvil’s own Boart Longyear LF90 rigs, managed by Wallis
Drilling. The holes were all drilled to 45 m with HQ (to which depth the holes were cased) and
drilling continued with NQ. Drilling procedures were upgraded following recommendations made by
Arnold (2004b), and included reducing the downhole survey interval (to approximately 30 m) and
P a g e | 56
establishing the daily maintenance of an up-to-date digital database of geological and geotechnical
logs, survey data, and QA/QC data.
Each hole was surveyed every 30-40 m, using a single shot Tropari tool. Drilling procedures were
similar to those of the 2004 programme.
10.6.
A surface programme was drilled from the pit in 2007 to increase the confidence in the underground
resource. A total of 9 diamond holes were completed for 2061 m to better define the mineralisation
beneath the current pit at the time. The drillholes were all diamond HQ near surface and reduced to
NQ at depth for the deeper holes. A total of 78 UGD series underground holes were completed as
grade control diamond holes.
10.7.
In late 2008 there was the recognition that another surface drilling programme was needed as there
was significant and increasing capital cost being sunk into the Dikulushi Mine. The objective of the
2008 programme was to better define the resource below the 630 mRL to assist with the planned
underground development. The orebody below the 630 mRL was at the bottom of the resource
model at the time.
A decision was made to mobilize a Titan drill rig to site in late October 2008 to commence drilling.
By the time the mine was placed on care and maintenance in early December 2008, only two out of
the 5 proposed holes had been completed. Both holes intersected the mineralisation but were not
sampled.
10.8.
MWL PROGRAMME 2010
The Kazumbula deposit was drilled by Anvil during 2008. The Anvil drilling data assisted MWL with
drillhole planning and targeting of the Kazumbula deposit. MWL drilled RC and diamond holes
(Table 10.1) to define the near surface copper mineralisation during August and September 2010.
The drillhole spacing was approximately 15 m along drill lines spaced 20 m apart. Drillholes were
drilled at approximately 60 degrees to the south-southeast to maximise the angle of intersection
with the orebody.
Table 10.1
Prospect
Kazumbula
Kazumbula
Type
RC
DDH/tail
No Holes
17
10
Metres
1676
674.4
Samples
1676
674
The RC drilling was completed by Titan Drilling of Lubumbashi, utilizing a truck mounted RC rig. A
supervising geologist was on site at all times during the drilling and industry standard procedures
were followed during the RC drilling programme. The diamond drilling was contracted and
completed by Chantete Emerald, who completed five diamond holes (HQ3) from surface and four
diamond tails from RC pre-collars (HQ3).
P a g e | 57
10.9.
SURVEY CONTROL
Drill hole collar locations have been located using a Leica total station by mine surveyors. Collar
surveys were stored with both local mine grid and UTM coordinates. A local Dikulushi Grid was
established and collar locations stored in both local mine grid and UTM co-ordinates. The
relationships between Dikulushi Mine Grid and magnetic (MN) and true north (TN) orientations are
as follows:
•
•
10.10.
Grid North = MN - 20.2˚
Grid North = TN - 22.0˚
DRILLING ORIENTATION
Downhole surveys of drillholes were completed every 30 m to 50 m of advance in order to ensure
that each hole was not deviating too much from the planned dip and azimuth. The surveys were
measured using Eastman downhole electronic single-shot cameras that record the dip and azimuth
and results were then tabulated for each hole as a report, which was checked by the site geologists
for accuracy. The camera has a stated accuracy of 1 degree in dip and azimuth. No magnetic
minerals have been noted in the logging of the drill core at Dikulushi and thus the recorded azimuths
are regarded as reliable.
Holes at both Dikulushi and Kazumbula were generally drilled orthogonal to the mineralisation, and
thus significant true width conversions were not required. The average mineralisation true thickness
is significantly greater than the average (1 m) downhole drilling increment, so distorted intersections
of the mineralisation at Dikulushi and Kazumbula were not obtained.
P a g e | 58
11. SAMPLE PREPARATION, ANALYSIS AND SECURITY
All of the drilling data used in the Dikulushi resource estimate was collected by Anvil. A review of
Anvil’s procedures by Mawson personnel concluded that they were of an acceptable industry
standard. The sampling procedures described below have been used for all diamond core drilled at
the Dikulushi deposits. All sampling and logging data are stored in secure database systems and
have been subject to routine validations during capture and storing.
11.1. DIAMOND CORE SAMPLING
Drill core (HQ and NQ size) is sampled by splitting it in half with a water lubricated diamond blade
saw. MWL has ensured the diamond blade is cleaned frequently using a brick to prevent across
metre contamination (especially in zones of massive sulphide). All drill core was sampled on a metre
basis from the start to end of hole. Minor residual sample lengths (less than one metre) may occur
along mineralisation contacts or at the end of holes. Each metre sample of half drill core is collected
from a consistent side of the drill core tray and placed into a sequentially numbered sample bag.
Sample log books are used to record the drillhole number, sample number and the “from” and “to”
sample depths. MWL has additionally incorporated electronic data capture of the sampling into a
toughbook laptop computer.
Calico/sample bags were tied up and placed into larger labelled plastic bags for transportation.
Submission forms for the laboratory were completed and placed into a small plastic bag within the
large labelled plastic bag. The samples were appropriately packaged for transport to the respective
international or minesite laboratories. Transport documentation and customs clearance were
completed by the company representatives. Laboratory turn-around times varied from a few days
to typically three to four weeks for the international laboratories.
11.1.1.
DIAMOND CORE RECOVERY
At the end of a core run, the drillers attach a water hose and pump the HQ/NQ core out of the barrel
into an angle iron ensuring minimal disturbance. The driller records the total depth and core run
length on a core block, also noting any core loss, and places this into the core tray at the end of the
run. The site geologist regularly checks the depths provided by the drillers with the core in the trays
during site visits. Diamond core recovery is good and was noted to be above 95%. Any handling
core breaks are marked with a cross. Once a core tray is full, the tray was labelled with from and to
depths and the hole number. The labelled core trays were moved to the core logging/storage area
at the Dikulushi mine site.
11.1.2.
DIAMOND CORE LOGGING
The drill core was washed to remove any residual cuttings. Downhole metre marks were made by
the geologist on the consistent half of the core. Wet core was photographed for the more recent
drillholes, using a digital camera before logging. Labels showing hole number, tray number and from
and to depths were placed in the photo frame for each core tray photograph. Core recovery, RQD,
geology, alteration and mineralisation were logged onto standard paper logs and more recently by
MWL into a toughbook laptop computer using LogChief software. The logs were electronically
P a g e | 59
captured into an Access database at Dikulushi or electronic logs were emailed to Mawson West’s
Perth office and loaded into the central Datashed database.
11.2. RC SAMPLING AND LOGGING
RC drilling by Anvil was generally limited to depths of less than 120 metres below the pre-mining
surface. The portion of the Mineral Resource and Ore Reserve estimates that are the subject of this
report are informed entirely by diamond core drill holes.
11.3. SAMPLE QUALITY
Drill core samples that inform the portion of the Mineral Resource and Ore Reserve estimates that
are the subject of this report are of good quality with no major risks identified for use in the resource
estimate. Open pit blast hole grade control samples and underground sludge hole samples have not
been used to inform the Dikulushi resource and reserve estimates.
With respect to sample security, no sample preparation other than diamond core cutting was carried
out by MWL or Anvil employees.
Reference materials, including chips, core, pulps and residues are retained and stored at the
Dikulushi mine site. Assessment of the data indicates that the assay results are generally consistent
with the logged alteration and mineralisation tenor.
11.4.
SAMPLE PREPARATION AND ANALYTICAL PROCEDURES
Preparation and assaying of samples from the Dikulushi Project has been carried out at three
independent laboratories:



Genalysis (RSA) and Genalysis (Western Australia)
ALS-Chemex (RSA)
(from Jan 2008)
SGS (Dikulushi).
11.4.1.
ANALYSES
Samples sent to Genalysis in Johannesburg, South Africa, were processed and analysed using the
following methods. All samples were weighed, then dried at 110° for 8 hours and then crushed to a
nominal 10 mm crush size in a conventional jaw crusher. The entire sample was then pulverised to a
nominal 85% passing 75µm in an LM-5 mixer-mill. A scoop of the pulverised sample was then
digested by the AX method which was a modified (higher precision) 4 acid digest for base metals.
Analysis technique was by AAS for Copper (0.01% detection limit) and Inductively Coupled Plasma
Mass Spectrometry (ICP-MS) for Ag (1ppm detection limit), As (10ppm detection limit), Co (1ppm
detection limit) and U (0.1ppm detection limit). Results were reported electronically via email and a
hard copy report was mailed to MWL and Anvil staff.
Samples sent to ALS Chemex in Johannesburg, South Africa, were prepared and analysed by the
following procedures. Samples were weighed and then dried for 8 hrs at 110° and then fine crushed
to 2 mm with a 250 g split of the sample taken for pulverising to 85% passing -75µm. The sample
was then digested in a four acid mixture (HF, HNO3, HClO4) and a HCL leach with analysis by AAS
P a g e | 60
(method AA62) for Copper (0.01 to 40% reporting range). Other elements were analysed by the ICPAES (optical emission method) with detection limits of Ag (0.5ppm), As (5ppm), Co (1ppm), U
(10ppm) and Co (1ppm).
Samples sent to the SGS laboratory at Dikulushi were prepared and analysed by the following
procedures. Received samples were sorted and dried at 105 degrees for a minimum of 8 hours and
then crushed to a nominal 10 mm crush size using a jaw crusher. Samples were split to 250g and
then pulverised to 90% passing -75µm. The sample was then digested in triple acid digest (A103
method;0.4g, Hydrochloric acid, Nitric acid and Perchloric acid) and finally analysed by AAS machine
with low detection limit (DL) for Copper of 0.001% and an upper detection limit (UL) of 5%, Silver (DL
was 5ppm and UL was 500ppm), Cobalt (DL was 20ppm and UL was 5%), Lead (DL was 10ppm and UL
2.5%), Arsenic (DL was 0.01% and UL was 5%), Zinc (DL was 10ppm and UL was 5%) and Fe (DL was
0.01% and UL was 100%). The sample batches included 2 standards, 2 blanks, 2 repeats and 1
replicate per 43 samples.
11.5.
BULK DENSITY DETERMINATIONS
Samples were collected every metre from the massive sulphide zones and from a representative
selection from the transitional and primary zones and un-mineralised zones. Diamond core samples
were prepared by ‘squaring off’ the ends of approximately 10-20 cm billets of half core. A total of
1,294 specific gravity (SG) measurements were made of dried half core to obtain the dry weight at
Kazumbula. The same piece of core was then measured in water on a suspension cage below the
same electronic scale. The conventional formula for SG determination was used, i.e.
SG = Dry Sample Weight / (Dry Sample Weight – Wet Sample Weight)
11.6.
SAMPLE QAQC
11.6.1.
STANDARDS AND BLANKS
QAQC for exploration drilling samples includes use of standards, blanks and duplicates, together
with internal/laboratory batch control information. Results from submitted standards are shown in
Figure 11.1, Figure 11.2 and Figure 11.3. GBM398-4c (Figure 11.3) is a low copper value standard
and suggests accurate results for low value samples (~0.39% Cu).
P a g e | 61
Figure 11.1
GBM301-7 suggests accurate values around low value samples (~0.55% Cu)
Figure 11.2
The GBM301-8 is a high Cu value standard and suggests accurate results for high value samples (~10% Cu)
Figure 11.3
The GBM398-4c is a low Cu value standard and suggests accurate results for low value samples (~0.39% Cu)
Analytical results from sample blanks, Figure 11.4, suggest that contamination was kept to a
minimum.
P a g e | 62
Figure 11.4
Results for this blank demonstrate that contamination is well contained
Both blank and standard sample results indicate that minor sample mislabelling occurred.
11.6.2.
LABORATORY QAQC
The respective laboratories perform internal QAQC checks as per the following description from ALS
Chemex (Johannesburg):
“The Laboratory Information Management System (LIMS) inserts quality control samples (reference
materials, blanks and duplicates) on each analytical run, based on the rack sizes associated with the
method. The rack size is the number of samples, including QC samples, included in a batch. The blank
is inserted at the beginning, standards are inserted at random intervals, and duplicates are analysed
at the end of the batch. Quality control samples are inserted based on the following basis:
Sample Count
40
Methods
Regular AAS, ICP-AES and ICP-MS methods
QAQC Sample Allocation
2 standards, 1 duplicate, 1 blank
The laboratory staff analyses quality control samples at least at the frequency specified above. If
necessary, laboratory staff may include additional quality control samples above the minimum
specifications.”
Failed batches are automatically repeated until acceptable results are achieved.
11.7.
SUMMARY STATEMENT
Sampling of drillhole material and QAQC is comprehensive in its coverage of the mineralisation and
does not favour or misrepresent in-situ mineralisation. Sampling and sub-sampling procedures are
of good standard industry practice and have occurred in a safe and secure manner, with minimal
time lags between drillhole sampling and analysis. Sufficient drillhole material has been retained
should additional verification of results be required. Sample security, preparation and analytical
procedures are believed to be able to support representative sample assay results for estimation.
Submitted blanks did not raise any risks with regard to contamination.
P a g e | 63
12. DATA VERIFICATION
The comprehensive program of multiple standard and blank inserted at regular but random intervals
has highlighted that the Dikulushi sampling is both accurate and precise. The accredited and
independent laboratories have evidence of good internal QAQC practices. These results, combined
with the good spatial distribution of QAQC sampling, support accurate, precise and uncontaminated
sample assay results and have been verified by the principal author and Qualified Person. According
to these results and the number of samples available for estimation, the Dikulushi and Kazumbula
drillhole databases provide satisfactory sample support and quality for estimating in situ
mineralisation.
P a g e | 64
13. MINERAL PROCESSING AND METALLURGICAL TESTING
13.1. INTRODUCTION
Historically Anvil has completed a significant amount of testwork for Dikulushi, and a summary of
this work is presented below. Relevant operational data from the Dikulushi processing plant is also
tabulated.
As the underground ore will be mined from the same or close to the same areas as the ore
previously treated at or below the current open pit floor, it is not unreasonable to expect that it will
exhibit similar metallurgical characteristics during processing through the existing Dikulushi
Processing plant. It should be remembered that the underground was previously developed, mined
and ore processed through the current operating plant pre MWL ownership.
13.2. ANVIL TESTWORK
13.2.1.
EARLY TESTWORK
The following information was supplied by Mawson West as background to the original design for
the process plant that was built at Dikulushi. Sedgman has not been able to review the original
testwork reports and as such cannot verify the information in this sub-section.
A significant amount of metallurgical testwork was undertaken by Anvil for the pre-feasibility phase
of their Dikulushi Project between February 1998 and April 1998 by the Minerals Engineering Group
of Mintek at their laboratories in Randburg, South Africa. Resource Management Group (RMG)
established and supervised the testwork on behalf of Anvil. Local coordination and support in South
Africa were provided by Fluor Daniel, Southern Africa. The Mintek data were used as the process
design basis for the pre-feasibility study completed by Signet Engineering in Perth in April 1998.
A previous testwork program was carried out by the Bureau de Recherches Géologiques et Minières
(BRGM), the results of which were available in Report no. 80 SGN 260 MIN, issued in April 1980. A
limited amount of preliminary testwork was initiated by Anvil and undertaken by Goldfields in
Johannesburg and was detailed in their report no. FL04\ks dated 4 November, 1996.
The metallurgical testwork program carried out by Mintek in 1998 was on various sulphide, oxide
and host rock samples from Dikulushi. The locations of these samples, their average grades and the
rock type classification are listed below in Table 13.1. Each composite comprised material from one
to three drillholes.
P a g e | 65
Table 13.1
Details of Dikulushi drill core used in Mintek metallurgical testing
Composite No.
1
2
3
4
5
6
Drillholes
DIK 15, 22
DIK 28, 31
DIK 6, 11, 14
DIK 26
DIK 5, 14
DIK 12, 13, 23
Classifications
East-oxidised
East-deeper
West-main
West-disseminated
West-complex
East-transition
% Copper
Total
Oxide
9.5
1.5-2
15.2
0.8
10.1
1.0
2.8
0.3
9.0
1.3
7.9
0.6
Silver
g/t
360
525
150
60
50
260
The sample nomenclature indicates that compositing was based upon special and oxidation
properties of the ore. Sedgman cannot comment on the representivity of these samples with
respect to the current study.
Physical tests were undertaken for typical composites of massive sulphide and light grey sandstone.
Flotation tests were carried out on primary, transition, oxidised and highly oxidised composites from
the east zone, and primary and complex sulphide composites from the west. These composites
represented an arbitrary sub-division of the ore body.
Head analyses revealed a relatively high total copper grade of 15.2% for the East Primary composite,
while the others were in the range of 8.2 - 11.4%, which was reasonably close to the target grade of
10% copper. Silver assays were variable, with a range of 138 - 562 g/t, the highest being for the East
Primary. Iron and sulphur levels were relatively low. Potential penalty elements identified were
lead and zinc in the West Complex, arsenic in the West Primary and West Complex, and fluorine in
all composites.
The previous testwork by BRGM in the 1980s indicated good flotation characteristics, with
recoveries ranging from 84 - 96% for copper, and 79 - 96% for silver. High grade concentrate grades
of 63 - 72% copper and 950 - 2,600 g/t silver were produced. BRGM found that sulphidation with
Na2S was required for oxidised material, though highly oxidised near surface ore was not tested.
Mineralogical examination revealed that the dominant copper sulphide mineral was chalcocite, in
both massive and disseminated forms. Some of the massive chalcocite was crystalline, and may
tend to slime during grinding. Complex sulphides in the west zone contained chalcopyrite, bornite
and sphalerite. Sphalerite is also common in other areas associated with chalcopyrite. Near surface
oxide contained malachite, azurite and chrysocolla. The latter did not float even when sulphidised.
Silver was assumed to be present mostly in solid solution in chalcocite, and occasionally as selenide.
Arsenic occurred as arsenopyrite and tennanite. Sandstone was the dominant host rock.
The physical tests revealed that the Dikulushi ore was of moderate hardness, with figures of 14.1 17.4 for the Rod Mill Work Index (RMWI), 10.5 - 12.5 for Ball Mill Work Index (BMWI) and 0.21 - 0.39
for Abrasion Index (AI) being reported. The higher indices generally related to the massive ore.
Flotation results at a grind size of 80% passing 75 microns were comparable to those in the BRGM
data, with recoveries of 71 - 97% for copper and 63 - 95% for silver. The lower figures were for near
surface highly oxidised material. The predicted concentrate grades were 48 - 70% copper, and 661 P a g e | 66
2,300 g/t silver. Detailed concentrate analyses revealed that fluorine was the only impurity over the
penalty threshold. Reagent usage appeared modest, except for the Na2S required for the oxidised
material, which required up to 3.2 kg per tonne of ore
13.2.2.
LATER TESTWORK
Additional testwork was performed by Independent Metallurgical Laboratories (IML) in Perth during
2003. The related testwork reports have been reviewed and Sedgman has been able to verify the
information detailed in this sub-section.
Five separate copper ore composites from Dikulushi were used for the testwork:





high grade chalcocite
disseminated and low grade chalcocite
lead and zinc rich chalcocite
bornite
stockpiled dense media separation tailings.
The various chalcocite composite assays are detailed in Table 13.2.
Table 13.2
Element
Unit
Cu (Total) – Assay
Cu (Total) – Calc.
Cu (Total – Sequential.) – Calc.
Cu (Acid Soluble)
Cu (Cyanide Soluble)
Cu (Residual)
Ag
Pb
Zn
%
%
%
%
%
%
ppm
High
grade chalcocite
21.9
20.1
20.4
3.52
16.8
0.14
624
39 ppm
189 ppm
Disseminated
& low grade
chalcocite
3.05
2.99
3.05
1.27
1.73
0.04
75
21 ppm
115 ppm
Pb/Zn rich
chalcocite
6.30
5.53
5.67
0.26
3.83
1.58
23
1.58%
10.88%
The Sequential Diagnostic Leach Analysis identifies the oxide component as Acid Soluble copper, the
Secondary Sulphides (including Chalcocite and Covellite) report as Cyanide Soluble species and the
residual fraction relates to primary copper sulphides such as chalcopyrite.
Mineralogical examinations identified the abundance of various minerals as illustrated in Table 13.3.
P a g e | 67
Table 13.3
Relative abundance of significant minerals
Disseminated
Pb/Zn rich
& low grade
chalcocite
Mineral
chalcocite
+0.1mm
-0.1mm
+0.1mm
-0.1mm
+0.1mm
-0.1mm
Chalcocite
Dominant Dominant Dominant Dominant
Minor
Accessory
Malachite
Major
Major
Major
Major
Bornite
Accessory
Accessory
Trace
Trace
Accessory
Chalcopyrite
Trace
Major
Minor
Pyrite
Trace
Major
Minor
Sphalerite
Accessory
Dominant Dominant
High
grade chalcocite
Note. Dominant: >50%, Major: 20 - 50%, Minor: 10 – 20%, Accessory: 1 – 10%, Trace: <1%.
The comminution data derived for these composites relating to the Bond Ball Mill, Bond Rod Mill
and Abrasion Indices are summarised in Table 13.4.
Table 13.4
Comminution testwork results
Composite
High Grade Chalcocite
Disseminated & Low Grade Chalcocite
Pb/Zn Rich Chalcocite
BRMWi (kWH/t)
15.7
17.3
17.7
BBMWi (kWh/t)
12.4
13.8
-
BAi
0.1472
0.4224
0.2360
A series of flotation tests was performed on the composites.
HIGH GRADE CHALCOCITE
There was minimal difference in rougher flotation performance between grind P 80s of 75, 106 and
150 microns using a stainless steel mill. See Table 13.5.
Table 13.5
Effect of grind size on flotation performance (high grade chalcocite)
Grind P80 - mic
75
106
150
Cumulative Rougher Concentrates
Copper
Silver
Assay (%) Distribution (%) Assay (ppm) Distribution (%)
52.0
97.8
1567
97.3
53.2
97.6
1632
97.3
54.9
97.7
1543
97.0
Using a grind P80 of 150 microns in each case, rougher flotation tests at potassium amyl xanthate
(collector) additions of 70, 105 and 140 g/t resulted in high copper grades and recoveries in each
case although flotation kinetics were significantly slower at the lower addition rate, see Table 13.6.
P a g e | 68
Table 13.6
Effects of collector addition on flotation performance (high grade chalcocite)
Collector Addition (PAX) – g/t
70
105
140
Copper
Silver
55.3
95.8
1818
96.8
51.0
96.2
1644
96.8
54.9
97.7
1543
97.0
DISSEMINATED AND LOW GRADE CHALCOCITE
A set of flotation tests was conducted at various grind sizes. A grind P80 of 150 microns produced
similar results to the finer grind sizes, see Table 13.7.
Table 13.7
Effect of grind size on flotation performance (disseminated and low grade chalcocite)
Grind P80 - mic
75
106
150
Copper
Silver
13.4
82.3
303
80.6
14.3
81.5
326
79.4
13.1
81.5
303
78.8
The effect of variation in collector dosing was investigated. Although a higher collector addition
produced better results, these tests were performed at the fine grind P 80 of 75 microns and before
an optimised pulp Eh had been established. Consequently the testing was inconclusive, see Table
13.8.
Table 13.8
Effect of collector addition on flotation performance (disseminated and low grade chalcocite)
Collector Addition (PAX) – g/t
100
165
Copper
Silver
22.4
75.8
521
75.3
25.5
78.0
615
79.5
PB/ZN RICH CHALCOCITE
Two sets of tests were performed to investigate the effect of grind size at different pulp Eh levels.
The results are shown in Table 13.9.
Table 13.9
Effect of grind size and Eh level on flotation performance (Pb/Zn rich chalcocite)
Grind P80 - mic
Eh – mV
(Ag/AgCl/Sat KCl)
75
106
150
150
106
150
70
70
Copper
Zinc
Assay (%) Distribution (%) Assay (%) Distribution (%)
11.9
98.0
23.8
88.6
13.1
94.8
27.3
82.0
12.0
12.6
97.6
98.1
23.2
25.3
90.7
90.4
P a g e | 69
The tests showed high copper recoveries but the copper grades were diluted by the amount of zinc
also reporting to concentrate.
A series of tests were performed to determine the effect of a range of Zinc Depressants – Sodium
Cyanide, Zinc Sulphate and Sodium Meta-bisulphite. The results were disappointing with only
sodium meta-bisulphite demonstrating any depression of zinc, but unfortunately it also depressed
copper.
A mineralogical examination of a first rougher concentrate showed that approximately 50% of the
sphalerite was locked with chalcopyrite and another 10-20% of the sphalerite was associated with
other sulphides.
MIXED CHALCOCITE COMPOSITE
A locked cycle flotation test was performed on a composite comprising 41.6% High Grade Massive
Chalcocite and 58.4% Disseminated and Low Grade Chalcocite which produced a calculated head
grade of 9.41% copper.
A combined rougher/cleaner copper concentrate grade of 54.6% was produced at an overall
recovery of 86.9%. The combined silver concentrate grade was 1,683 ppm at a recovery of 91.9%.
ROM LOCKED CYCLE TEST
In 2004 a locked cycle test was performed on a plant feed sample dated 25/11/2003 producing a
unit flash flotation cell, rougher and cleaner concentrate.
The head feed sequential analysis is shown in Table 13.10 and the test results in Table 13.11.
Table 13.10
Element
Cu (Total) - Assay
Cu (Total) – Calc.
Cu (Total – Sequential.) – Calc.
Cu (Acid Soluble)
Cu (Cyanide Soluble)
Cu (Residual)
Ag
Pb
Zn
Table 13.11
Unit
%
%
%
%
%
%
ppm
ppm
ppm
25/11/2003 Feed Sample
9.18
9.23
9.31
1.98
7.32
0.01
Not Assayed
41
857
Locked cycle flotation test results
Product
Wt%
Unit Cell Conc.
Rougher Conc.
Cleaner Conc.
Scavenger Tail
Calculated Head
4.99
8.62
5.32
81.07
100.00
Assay (%)
61.32
47.80
14.25
1.03
8.79
Copper
Distribution (%)
34.92
47.72
8.47
8.89
100.00
Assay (%)
2400
1600
305
39
306
Silver
Distribution (%)
39.21
45.14
5.31
10.34
100.00
P a g e | 70
The results indicate a combined concentrate grade of 42.1% copper at an overall recovery of 91.1%.
The combined silver concentrate grade was 1,447ppm at an overall recovery of 89.7%.
TESTWORK SUMMARY
Of the three chalcocite composites tested at IML in 2003 the high grade chalcocite composite was
the most relevant to the Dikulushi Open Pit Project. However it cannot be considered truly
representative as the head grade was far higher than the planned feed grade and operational data at
Dikulushi showed that there was a positive correlation between copper head grade and recovery.
Overall the testwork did demonstrate that provided the flotation conditions, including Redox
potential, was carefully controlled, chalcocite ore could be effectively recovered by flotation
producing fast kinetics, high concentrate grades and good recoveries.
13.3. PLANT OPERATIONAL RESULTS
MWL has indicated that the flotation plant at Dikulushi previously operated from 2004 to 2008 and
processed high grade ore from both the open pit and underground mine. According to Anvil
production data between September 2004 and April 2008, it achieved recoveries of 88.3% copper
and 88.5% silver, producing a concentrate containing 54.7% copper and 1,659 g/t silver. The plant
was shut down in November 2008 after treating low grade stockpile material during the last months
of operation.
In May 2010 the plant was refurbished and commenced production in June 2010 by treating low
grade stockpile ore and HMS tails. Over the past 3 years, recoveries vary between 60 - 70% for the
low grade stockpile material and the open pit cut back ROM feed grade recoveries have varied
between 75% - 95%. Concentrate grades in the last 12 months have averaged 56% copper and 1,515
g/t silver over the past year. Table 13.13 in the next section, shows the last 3 years production on a
month by month basis.
Sedgman has reviewed the production data as supplied by MWL for the periods Feb 2007 to Apr
2008 and June 2010 to May 2011, however, Sedgman has not reviewed the production data for the
period June 2011 to July 2013.
13.4. METALLURGICAL PROPERTIES OF THE CUT BACK ORE AND
UNDERGROUND ORE
The Dikulushi deposit was mined and processed by Anvil for several years and the high grade
chalcocite ore below the current pit floor has previously been processed in the mill during
underground mining operations. Anvil monthly production reports, for the previously mined
underground are tabled in table 13.12, where the mill feed was from the old open pit and the then
underground mining operations. In reviewing the historic operating data, it can be seen that the
copper recovery was approximately 90.4% over the period, with underground recoveries averaging
86.9 to 92.8%.
P a g e | 71
Table 13.12
Month
Feb07*
Mar-07
Apr-07
May-07
Jun-07
July
07*
Aug-07
Sep-07
Oct-07
Nov-07
Dec-07
Jan-08
Feb 08*
Dikulushi processing summary (February 2007 – April 2008)
Blend %
ROM
RL mined (Ore Only)
60
Cu
(%)
Ag
(ppm)
Cu
(%)
Ag
(%)
Concentrate
Grade
Cu
Ag
(%)
(ppm)
27,779
5.93
181
85.7
86.9
56.0
1730
Flotation
Plant
Tonnes
Plant Feed
Recovery
100
100
100
100
860 pit stockpile
860 pit stockpile
850 pit stockpile
850 pit stockpile
28,508
28,487
26,188
30,805
8.44
7.68
7.61
7.74
264
240
231
233
91.6
90.7
90.2
91.0
90.4
90.5
90.1
90.5
56.2
55.1
55.1
55.3
1734
1722
1670
1654
91.4
870 Dev
31,838
7.28
214
89.5
89.8
56.6
1668
30,802
25,934
31,193
30,286
30,641
30,746
30,789
7.96
7.97
8.18
7.81
8.45
6.00
5.09
245
258
272
250
266
187
154
91.2
91.3
92.4
92.2
92.8
90.6
87.2
90.5
91.4
92.5
91.8
92.0
89.4
87.7
56.2
54.9
54.6
56.3
56.8
55.2
55.7
1717
1777
1821
1793
1772
1694
1687
37,998
5.50
170
88.2
88.5
54.5
1691
33,400
455,395
4.76
7.04
139
218
86.9
90.4
88.4
90.3
54.0
55.5
1601
1721
100
100
100
100
100
100
81.4
Mar-08
100
Apr-08
90
stockpile
850 Dev
850 Dev & 890 Stoping
870 Dev & 890 stoping
830 Dev & 890/870
Stoping
Total
* Low grade ore blended in with the development or stoping ore.
Table 13.3 below shows the current processing statistics for the MWL operations from June 2010
through to July 2013. During this period a combination of LG stockpile material from the old Anvil
open pit was processed in the early months, to satellite orebodies such as Boom Gate etc, through
to the current mill feed being exclusively from the open pit cut-back material. This process feed
material is set to continue until December 2013.
In reviewing the recent operating data, it can be seen that the copper recoveries realised over the
past 7 months, reflecting the fresh open pit cut-back material feed, was approximately 94.3% over
the period.
Table 13.13
Ore Processed
Mill Feed Grade
Mill Feed Grade
Tails Grade Cu
Tails Grade Ag
Conc Tonnes
Conc Grade Cu
Conc Grade Ag
Cu metal in Conc
Ag metal in Conc
Recovery Cu
Recovery Ag
Dikulushi processing summary (June 2010 – July 2013)
tonnes
Cu %
Ag g/t
Cu %
Ag g/t
dmt
Cu %
Ag g/t
dmt
oz
%
%
Jun10
5,387
1.28
35.87
0.34
10.1
128
38.7
1,067
51.45
4,384
74.62
70.57
Jul-10
36,157
1.45
40.4
0.39
11.1
896
43.5
1,138
389.6
32,778
74.31
69.88
Aug10
43,882
1.04
27.63
0.35
10.5
719
42.7
1,107
306.9
25,581
67.25
65.62
Sep10
40,839
1.27
31.72
0.46
10.9
783
43.0
1,119
336.7
28,177
64.92
67.65
Oct-10
27,450
3.78
77.17
1.64
23.0
1,380
44.1
1,139
608.5
50,534
58.64
74.20
Nov10
49,029
1.52
41.2
0.63
13.6
1,066
41.5
1,188
442.7
40,726
59.41
62.68
Dec10
41,111
1.17
28.5
0.52
10.70
684
39.74
1139
272
25,057
56.54
66.45
Jan-11
49,650
1.33
32.6
0.46
11.3
1001
40.1
1070
400
32,737
64.13
62.70
Feb11
42,839
1.32
29.2
0.43
7.95
890
41.6
1033
366
29,385
66.91
73.09
Mar11
46,054
1.28
27.8
0.46
7.85
893
39.35
941
351
27,279
62.66
63.34
Apr11
40,855
1.40
34.6
0.44
8.7
906
40.2
1092
365
31,904
67.46
71.25
May11
44,705
1.32
33.31
0.52
9.1
865
41.7
993
361
27,559
61.37
61.3
YTD
467,958
1.46
35.31
0.54
10.95
10,211
41.66
1089
4,251
356,101
64.05
66.68
P a g e | 72
Jun-11
Jul-11
Aug-11
Sep-11
Oct-11
Nov-11
Dec-11
Jan-12
Feb-12
Mar-12
Apr-12
May-12
Jun-12
YTD-12
Ore Processed
tonnes
41,684
44,113
31,913
35,566
32,799
34,417
30,084
14,008
23,756
15,009
13,061
14,096
17,355
Reconciliated Mill Feed Grade
Cu %
1.55
1.44
2.12
1.70
1.77
1.36
1.43
1.96
1.95
2.06
2.24
2.73
2.90
1.80
Ag g/t
28.14
29.65
37.08
33.20
32.12
27.30
29.86
41.37
37.87
38.07
40.57
49.30
50.06
34.15
Tails Grade Cu
Cu %
0.62
0.59
0.92
0.68
0.73
0.54
0.53
0.71
0.93
0.74
0.69
0.76
0.74
0.69
Tails Grade Ag
Ag g/t
7.40
8.28
9.02
7.77
8.90
7.70
7.85
14.25
16.50
13.10
12.60
14.76
13.84
10.75
10,437
Concentrate Tonnes Produced dmt
347,863
991
924
944
926
863
723
699
495
780
633
621
804
1,032
Concentrate Grade Cu
Cu %
39.54
41.36
41.36
39.96
40.13
39.62
39.28
36.00
32.02
32.07
33.46
35.24
37.15
37.83
Concentrate Grade Ag
Ag g/t
803.56
997.08
911.84
883.23
865.56
900.22
936.86
781.60
667.08
604.74
600.72
620.20
622.99
790.95
Cu metal in Concentrate
dmt
392
382
391
370
346
287
275
178
250
203
208
283
383
3,948
Ag metal in Concentrate
oz
25,608
29,629
27,683
26,290
24,007
20,934
21,064
12,440
16,737
12,314
11,998
16,036
20,667
265,405
Recovery Cu
%
60.7
60.2
57.7
61.0
59.8
61.3
63.6
65.0
53.9
65.6
70.9
73.7
76.0
63.0
Recovery Ag
%
67.9
70.5
72.8
69.3
70.9
69.3
72.9
66.8
57.9
67.0
70.4
71.8
74.0
69.5
Ore Processed
tonnes
Cu %
2.65
2.28
2.33
2.04
5.09
5.41
5.48
6.32
5.46
6.50
10.27
10.88
7.02
5.21
Ag g/t
47.22
42.24
39.35
34.53
152.94
139.04
145.47
183.09
146.87
164.68
316.98
354.49
226.97
141.47
Tails Grade Cu
Cu %
0.65
0.59
0.55
0.45
0.42
0.44
0.40
0.43
0.36
0.44
0.51
0.61
0.46
0.48
Tails Grade Ag
Ag g/t
11.59
10.92
9.94
6.03
21.39
14.95
15.08
11.72
15.47
20.70
22.94
25.50
20.14
14.93
1,448
1,334
1,476
1,200
1,138
2,711
2,685
2,939
2,936
3,087
2,470
2,798
3,723
29,944
Cu %
34.86
36.51
37.57
40.49
61.16
58.42
59.42
61.40
62.29
62.70
62.85
60.79
58.97
56.52
Ag g/t
620
677
622
725
1,733
1,462
1,532
1,787
1,620
1,499
1,903
1,952
1,864
1,515
dmt
505
487
554
486
696
1,584
1,596
1,804
1,829
1,936
1,552
1,701
2,195
16,925
oz
28,876
29,014
29,496
27,950
63,410
127,429
132,235
168,837
152,974
148,824 151,060
175,621
223,058
1,458,783
Recovery Cu
%
76.9
75.2
77.4
78.7
92.4
92.6
93.3
93.8
93.4
93.9
95.8
95.4
94.2
91.5
Recovery Ag
%
76.9
75.4
76.0
83.2
87.1
90.2
90.5
94.2
90.3
88.7
93.9
94.0
92.1
90.3
Jul-12
24,739
Aug-12
Sep-12
Oct-12
Nov-12
Dec-12
Jan-13
Feb-13
Mar-13
Apr-13
May-13
Jun-13
28,351
30,690
30,242
14,808
31,614
31,232
30,442
35,865
31,706
15,787
16,387
Jul-13
33,181
YTD-13
355,043
Figure 13.1shows a cross section of the Mineral Resource as it relates to the previously mined parts
of the orebody via the Anvil open pit, Anvil underground and MWL open pit cut-back.
Figure 13.1
Dikulushi Underground sources of ore - showing North-South section view at 50205E
The planned underground ore production is to be from the previously developed and mined levels at
the 810 mRL down to the fully developed 770 mRL and a minor amount of ore from the partially
P a g e | 73
developed 750 mRL. Thus large portions of the underground ore reserve ore tons have been
previously treated in the processing plant, as can be seen in the production processing summary
table 13.12; and the more recent production processing data in table 13.13, and hence significant
variations in ore quality is not expected from the mining of the underground Mineral Reserves.
P a g e | 74
14. MINERAL RESOURCE ESTIMATES
14.1.
DIKULUSHI MINERAL RESOURCE ESTIMATE
The Dikulushi Mineral Resource estimate was prepared in May 2009 by Mr. David Gray, Qualified
Person and principal author of the technical report which was originally submitted in February 2011.
The May 2009 Mineral Resource was subsequently updated in August 2011 using the latest available
survey data of the historical volumes mined by Anvil and updated pre-feasibility study cut-off grades
for the proposed cut back of the open pit.
A previous (October 2007) Mineral Resource estimate for Dikulushi was generated for the purposes
of evaluating underground Mineral Resources. The geological interpretation of copper-silver
mineralisation beneath the open pit was largely based on the diamond drillhole database and
enabled the main Footwall zone of mineralisation to be extended to the Kiaka Carbonates.
Between the October 2007 estimate and the May 2009 estimate, an additional 23,610 m of
underground, infill and extensional drilling was completed (Figure 14.1) across the Dikulushi ore
body and can be broken down by sampling type:




802 m were derived from underground channel sampling
3,747 m from underground grade control diamond drilling
4,789 m from RC drilling
14,272 m from surface diamond drilling.
The May 2009 estimate was based on all available data as at the end of November 2008, with no
outstanding core logging, sampling or assay results remaining. Since that time there has been no
additional data that impacts on the estimate of resources remaining below the base of the open pit
cut back that was completed in July 2013.
Dikulushi mineralisation (Figure 14.2, showing footwall mineralisation in green and hanging wall
mineralisation in orange) is characterised by a hydrothermal copper-silver vein system hosted by
Proterozoic sediments of the Upper Kundelungu Group, and has two distinct ore zones. A dominant
“Footwall” zone is intersected over a 230 m strike length with thicknesses of up to 25 m, which
decreases with increasing depth. This zone comprises semi-massive chalcocite and/or bornite veins,
strikes east-northeast and dips southeast at approximately 65°. Exhibiting good strike continuity, it
can be traced to depths of approximately 500 m below surface. A secondary “Hanging Wall” zone is
observed within 50 m of the Footwall zone, and comprises discontinuous, steeply dipping, chalcocite
veins, veinlets and disseminations. These dip at varying angles to the Footwall zone and may
occasionally intersect it. Apart from minor other occurrences, the Hanging Wall zone is largely
absent below the base of the open pit.
Grade interpolation was undertaken for total copper (%) and silver grade (g/t). Wireframes were
created for the domains and defined zones of similar weathering, faulting, stratigraphy and copper
grade. Sample copper and silver analytical results were composited to one metre interval lengths
per domain. Variography displayed reasonable continuity with low nugget values.
P a g e | 75
Figure 14.1
An oblique southward looking 3D view of drillhole type and distribution at Dikulushi
The resulting Mineral Resource statement was depleted for open pit and underground material
mined as surveyed from mined volumes and since the previous October 2007 estimate through to
November 2008. The estimate is representative of all data acquired. Mineral Resources have been
classified into Measured, Indicated and Inferred categories for the fresh sulphide mineralisation
located below the November 2008 pit surface, as per Table 14.1.
Table 14.1
Dikulushi Mineral Resource statement as at August 2011 above a 1.0% copper cut-off grade
Volume
3
(m *1,000)
184
90
Density
3
(t/m )
2.8
2.8
Tonnes
(*1,000)
516
251
Copper
(%)
7.0
5.6
Silver
(g/t)
211
114
Measured & Indicated
274
2.8
767
6.6
179
Inferred
136
2.8
380
6.8
91
Category
Measured
Indicated
It should be noted that this model has now been further depleted with the recently completed open
pit cut back; the depletion tables are presented and discussed in Section 14.2.
P a g e | 76
14.1.1.
Lithology and lode profiles were developed using five metre spaced north-south cross sections. The
ore body was modelled as a Footwall fault zone with sporadic mineralisation intersected within 50 m
of the overlying hanging wall. Two Hangingwall domains, as observed in the pit, were delineated
and modelled. The open pit has mined most of the weathered material and has exposed weathering
to depths of 35 m; the impacts of weathering were therefore not considered in the 2009 estimate.
Wireframes representing the boundaries relevant to the mineralisation were constructed in three
dimensions (3D) using north-south vertical cross sections. Mineralisation outlines were guided by
geological continuity between drillholes and a mineralisation threshold between 0.3% and 0.7%
copper.
Both blasthole and underground channel data (Figure 14.2) supported depth extensions of the
Footwall Fault zone.
Figure 14.2
A vertically oriented 3D view at Dikulushi, looking southwest, showing mineralisation lenses and current drilling
P a g e | 77
14.1.2.
Drill data was stored in Dikulushi’s on-site Access database. While some risk exists regarding the
reliability of manually handled data in an Access database, the drillhole de-surveying process
revealed only minor location errors, which were immediately corrected.
A plan view of drillhole data by type is presented in Figure 14.3. A total of 567 holes were available
for geological modelling, comprising 22,129 m surface diamond, 4,951 m surface reverse circulation,
1,285 m channel, 4,131 m underground grade control diamond and 1,369 m sludge samples. This
translates as a net increase of 23,610 m from the previous resource estimate.
Diamond drilling was undertaken along north-south oriented lines spaced 20 - 25m apart, with holes
at 25 m intervals along each line. To maximise the true widths of the intersections, most drilling was
angled at 50 to 60 degrees to the south. As the risk of undetected changes to orebody orientation
increases with depth, additional infill drilling will naturally assist in improving the confidence in
deposit geometry. In 2008, a total of 4 surface exploration drillholes were drilled to both infill and
extend Footwall zone mineralisation. While the deposit remains open at depth, this recent drilling
has led to only minor east-west extension.
Figure 14.3
A plan showing the distribution of drillhole types across Dikulushi; blasthole data from the pit have been excluded
Since twin-hole drilling was not completed, drilling and sampling methods were compared for
potential bias across a similar volume of the FW zone mineralisation using quantile-quantile (Q-Q)
plots. Diamond core was accepted as generally providing the most representative sample. This
comparison emphasises the difference in copper values between diamond and sludge hole samples
(Figure 14.4), with the latter decreasing as the former increases. As a direct result, sludge hole data
was not used in this estimate.
P a g e | 78
Figure 14.4
Quantile-Quantile (Q-Q) plot of Diamond (DD) drilled samples versus sludge drilled samples within a common area
14.1.3.
DATA VALIDATION
A series of data validations were completed prior to de-surveying the drillhole data into a three
dimensional format. These included:






verification of collar coordinates with existing topography and underground development
wireframes, with virtually no problems observed
visualisation of downhole survey data to identify improperly recorded downhole survey
values, with all minor discrepancies corrected
dataset examination for sample overlaps and/or gaps in downhole survey, sampling and
geological logging data, with none observed
database interrogation for negative values representing codes such as ‘insufficient sample’,
with all such samples set to absent
examination for negative assays reflecting ‘below detection’ range; these values were all reset to 0.01%
testing for absent or duplicate samples, with none recorded.
14.1.4.
The de-surveyed 3D assay drillhole file was coded and selected within the mineralisation and
lithological 3D wireframes. Each sample interval was coded with a mineralisation zone and
weathering profile, providing mineralised domain codes for estimation (Table 14.2). The coded
drillhole data was exported for subsequent geostatistical analysis and grade interpolation.
P a g e | 79
Table 14.2
Domain codes for Dikulushi modelling
Field name
OREZONE
WEATH
MINED
14.1.5.
Domain
Oxidised FW zone
Fresh FW zone
Shallow HW zone A
Shallow HW zone B
Internal FW zone waste
Soil to 5m
Oxidised to 35m
Transitional to 75m depth
Fresh rock
Air
Not mined
Open pit mined
Mined underground
Open pit reserves
Code
50
100
200
300
400
0.1
0.2
0.3
0.4
0
0
1
2
3
DATA COMPOSITING
To determine the most common sample length, the distribution of raw sample lengths was plotted.
Approximately 45% of the data had a sample length within a few centimetres of 1 m (Figure 14.5).
All data was composited to 1 m sample lengths, ensuring that intervals provided good resolution
across domain boundaries. The total raw sample length is identical to the composited total sample
length.
Figure 14.5
Cumulative distribution of sample lengths highlighting the dominant 1m sample length
P a g e | 80
14.1.6.
STATISTICS
Statistical analyses of the data, including spatial statistics, were carried out using Supervisor
software. The statistical analysis of composite copper grades was undertaken within each of the
final domains and the summary results are presented in Table 14.3.
Statistics for copper and silver were investigated by domain with histograms and probability plots.
The objective of the domain selections was to reduce internal variability and domain mixing, thereby
assisting with spatial analysis and providing a more robust estimate.
The selected domains appear to be well defined, with a minimal degree of mixing as depicted in
Figure 14.6 for Dikulushi’s principal Footwall zone.
Table 14.3
Summary statistics for copper % and silver g/t per domain
Samples
Min
Max
Mean
Std Dev
CV
Variance
Skewness
Log variance
Geometric mean
Waste
domain (0)
Oxide FW zone
domain (50)
Fresh FW zone
domain (100)
HW zone A
(200)
HW zone B
(300)
Cu (%)
Cu (%)
Cu (%)
Cu (%)
Cu (%)
10429
0.01
5.00
0.20
0.48
2.37
0.23
5.59
1.61
0.07
Ag (g/t)
Samples
Min
Max
Mean
Std Dev
CV
Variance
Skewness
Log variance
Geometric mean
14.1.7.
5456
1.00
325.00
14.22
31.09
2.19
966.27
6.19
1.43
6.13
1284
0.01
63.80
7.14
9.69
1.36
93.92
2.42
2.08
3.08
Ag (g/t)
1108
1.00
2615.00
214.27
340.51
1.59
115947.00
2.70
2.64
69.46
17145
0.01
74.34
6.06
8.47
1.40
71.81
2.84
2.75
2.32
Ag (g/t)
16221
1.00
1800.00
251.69
305.00
1.21
93023.50
2.02
2.10
111.94
956
0.01
11.00
2.29
1.76
0.77
3.09
1.60
1.21
1.54
Ag (g/t)
849
1.00
325.00
58.70
53.95
0.92
2910.94
2.11
0.96
39.20
204
0.02
23.00
3.03
4.27
1.41
18.23
2.68
1.87
1.36
Ag (g/t)
179
4.00
730.00
101.73
145.82
1.43
21262.30
2.70
1.37
50.07
Internal
waste
(400)
FW
zone
Cu (%)
1629
0.01
17.00
0.50
1.69
3.41
2.84
6.62
2.02
0.12
Ag (g/t)
709
1.00
470.00
27.19
59.91
2.20
3588.70
4.98
1.24
11.82
SPATIAL STATISTICS
For Dikulushi, variography was analysed using composite data located within the mineralised
envelopes of each domain, based on the following methodology:



data was declustered prior to variogram modelling so as to remove the effect of closely
spaced blast hole and underground channel data
the principal axes of anisotropy were determined using semi-variogram (variogram) fans
based on normal scores variograms
normal scores variograms were calculated for each of the principal axes of anisotropy
P a g e | 81



downhole normal scores variograms were modelled for each domain and adjusted to
determine the normal scores nugget effect
variogram models were then determined for each of the principal axes of anisotropy using
the nugget effect from the downhole variogram
the variogram models were back-transformed to the original distribution and used to guide
search parameters and complete ordinary kriging estimation.
Figure 14.6
Log histogram and probability plot for the main FW zone of mineralisation showing the results of robust domaining
P a g e | 82
Variogram orientations were largely controlled by the strike of mineralisation and downhole
variography. Variogram models for silver and copper were similar, with silver tending to have a
slightly longer range of influence. Variogram models for the Footwall zone of mineralisation were
robust with a clearly-defined nugget value and well-defined structure (Table 14.4). Omni-directional
variogram models were derived for both HW zones and the upper oxidised FW zone. These domains
were not critical to this Mineral Resource estimate as this ore has already been mined. They were
included to ensure continuity with the deeper domains. Key variogram models for the main FW
zone are depicted in Figure 14.7.
Table 14.4
No.
1
2
3
4
5
6
7
8
9
10
11
12
No.
1
2
3
4
5
6
7
8
9
10
11
12
Assay
CU
AG
CU
AG
CU
AG
CU
AG
CU
AG
CU
AG
Assay
CU
AG
CU
AG
CU
AG
CU
AG
CU
AG
CU
AG
Dikulushi variogram models with angle1 about axis 3 (Z), angle2 about axis 1 (X) and angle3 about axis 3 (Z)
Domain
0
0
50
50
100
100
200
200
300
300
400
400
Domain
0
0
50
50
100
100
200
200
300
300
400
400
Angle1
-5
-5
0
0
-10
-10
0
0
0
0
140
140
St2 par1
11.5
20
34.5
57
26.5
29
15.5
25.5
16
23
40
31
Angle2
130
130
0
0
100
100
0
0
0
0
80
80
St2par2
15
10
34.5
57
18.5
16
15.5
25.5
16
23
33
31
Angle3
10
10
0
0
-80
-80
0
0
0
0
-100
-110
St2 par3
9.5
11.5
34.5
57
8
6.5
15.5
25.5
16
23
14.5
31
Nugget
St1 par1
0.06
0.06
0.04
0.04
0.21
0.2
0.11
0.12
0.27
0.28
0.07
0.06
St2 par4
0.29
0.25
0.3
0.33
0.25
0.25
0.32
0.35
0.15
0.26
0.23
0.15
St1 par2
4
10.5
5
5
9
11
5
4.5
3
4
5.5
3
St3 par2
191
399
5
6
5
5
3
1.5
5
4.5
3
4
5
3
0.54
0.51
0.66
0.63
0.3
0.29
0.4
0.33
0.58
0.46
0.7
0.79
St3 par3
St3 par4
10
89
0.12
0.18
39
118.5
-
84
121.5
38.5
48
-
St1 par4
6
5
5
5
5
4.5
5
4.5
3
4
5.5
3
St3 par1
-
St1 par3
49.5
84.5
38.5
48
-
15
15.5
38.5
48
-
0.25
0.27
0.17
0.2
-
P a g e | 83
Figure 14.7
Variogram models for copper % across the FW zone of mineralisation
14.1.8.
BLOCK MODEL
The block model dimensions and parameters were based on the geological boundaries and average
drill grid spacing. Sub-blocks were used to ensure that the block model honoured the domain
geometries and volume. Block estimates were controlled by the original parent block dimension.
The individual parent block dimensions were 15 mE by 4 mN by 15 mRL, with sub-blocking allowed.
This dimension was supported by a kriging neighbourhood study which demonstrated little change
in the kriging efficiency or slope of regression (a measure of bias) from this block size to larger block
sizes.
P a g e | 84
14.1.9.
Density estimates were based on approximately 61 samples from the Footwall mineralisation and
1,236 samples from the surrounding waste material. These values have been tested and confirmed
via two mill feed samples. The assigned density of the Footwall ore zone was 2.8 t/m 3 and the
surrounding waste material 2.6 t/m3.
14.1.10.
DETERMINATION OF TOP CUTS
Top cut analysis was used to describe the maximum reasonable metal grade for a composite sample
value within a given domain. If the grade of a sample exceeded this value, the grade was reset to
the top cut value. The objective of applying top cuts is to minimise the risk of uniquely high metal
concentrations biasing individual block estimates, especially those located within areas of low
sample support.
Top cuts for Dikulushi were established by investigating univariate statistics and histograms of
sample values by domain. A top cut level was selected if it reduced the sample variance and did not
materially change the mean value. The following top cuts were applied to the data for resource
estimation (Table 14.5).
Table 14.5
Dikulushi - top cuts per domain
Domain
0
50
100
200
300
400
14.1.11.
Copper%
5
56
11
23
17
Silver g/t
325
2000
1800
325
730
470
GRADE ESTIMATION
Grades for copper and silver were estimated into parent blocks of an empty domain coded block
model using ordinary kriging (OK). OK was deemed an appropriate interpolation technique owing to
near normal data distributions and differentiable grade ranges particular to the lode style
mineralisation. Estimation into parent blocks used a discretisation of 8 (X points) by 3 (Y points) by 8
(Z points) to better represent estimated block volumes.
14.1.12.
ORDINARY KRIGING INTERPOLATION
Estimation parameters for kriging were based on variography, geological continuity and the average
spatial distribution of data. The first pass search radius was set within half to two thirds of the
variogram range to improve the quality of the local block grade estimate for areas of close spaced
drilling and to ensure that grade was not smeared laterally. Most blocks (75%) were estimated
within the first search radius. Subsequent search radii were set to ensure that remaining blocks
within the mineralised domain were interpolated with a copper grade.
For the ore domains, a minimum of 8 samples were required for a single block estimate and a
maximum of 40 samples to limit grade smoothing. Due to the long drillhole intercepts within the
orebody estimates were limited to a maximum of 10 samples per drillhole.
P a g e | 85
Soft boundaries were created between the oxidised and fresh weathering domains in order to
represent the variable nature of this boundary and the transition in values. All other domain
boundaries were hard and data between domains was not included for estimation.
14.1.13.
MODEL VALIDATION
The first pass of model validation included:



visual comparisons (Figure 14.8) of drillholes and estimated block grades
checks for negative grade estimates; if there were any, they were reset to a minimum 0.01 %
grade
checks to ensure that only blocks significantly distal to the drillholes remained without grade
estimates.
The model was further validated by statistical comparison of mean composite grades and model
grades, in addition to visual comparisons with drillholes. A table comparing the mean values for the
estimate with those of the data (Table 14.6) illustrates acceptable correlation.
Table 14.6
Mean statistics per domain comparing model estimates with data values
Domain
100
100
50
50
400
400
Field
Ag g/t
Cu%
Ag g/t
Cu%
Ag g/t
Cu%
Data
219.01
7.48
172.11
6.07
17.87
0.42
Model
201.22
7.44
169.34
6.18
15.82
0.42
% Variance
8.12
0.44
1.61
-1.81
11.46
0.19
Spatial statistical plots by domain are used to compare the mean model and drill grades data by
relative elevation slices (Figure 14.9). Model estimates respond well to changes in the composite
grade data, but local estimates are likely to be improved with additional drillhole intersections.
Based upon the summary statistics, visual validations and graphical plots, the OK estimates are
consistent with the drillhole composites, and are believed to constitute a reasonable representation
of the Footwall mineralisation.
P a g e | 86
Figure 14.8
A plan view slice through the FW zone block model illustrating the good comparison between model estimates and the
nearby drillhole data
Figure 14.9
A statistical plot of estimates versus drillhole data grades for successive 30m increments in elevation and the full strike
length of the FW zone mineralisation
P a g e | 87
14.1.14.
Classification of the Mineral Resource was primarily based on confidence in assayed grade,
geological continuity, and the quality of the resulting kriged estimates.
Geological confidence is supported by extensive open pit exposures and underground geological
mapping and channel data, which in turn reinforces drillhole logging and domain volumes.
Confidence in the kriged estimate is associated with drillhole coverage, analytical data integrity,
kriging variance and efficiency and regression slope values. Specifically, kriging variances below 0.2,
kriging efficiencies above 80% and regression slope values above 0.8 were considered appropriate
for a Measured Mineral Resource category of classification. Whereas the use of mean domain
density values is appropriate, subsequent models should make use of increased density data for
more robust estimates.
Regarding drillhole spacing, a Measured Mineral Resource category was considered appropriate with
a 20 m separation between drill holes and drill line spacing between 25 m to 50 m. An Indicated
Mineral Resource category was considered appropriate where there was a drill spacing of about 50
m to 75 m along drill lines and a line spacing of approximately 50 m. An Inferred Mineral Resource
category was considered where there was a drill spacing of about 75 m to 100 m along drill lines and
where the line spacing was around 100 m.
The Measured Mineral Resources are located below the pit and where underground sampling and
drilling is closely spaced. Indicated Resources extend as a consistent rim below the Measured
Resources. Confidence in the estimates deteriorates rapidly into Inferred Resources with the
increase in grid spacing and the short ranges of influence/grade continuity.
Figure 14.10
3D view of the Dikulushi model, looking south, and showing resource classification categories
P a g e | 88
The Mineral Resource has been classified and reported using the guidelines of the JORC Code (JORC,
2004), which in turnalign with the Standards on Mineral Resources and Reserves of the Canadian
Institute of Mining, Metallurgy and Petroleum (CIM, 2000).
14.1.15.
RESOURCE TABULATION AND INVENTORY
The Mineral Resource at Dikulushi is derived from that portion of the block model which occurs
below the current pit surface. Mineralisation appears to be open at depth, but is restricted to the
west by the Kiaka carbonates and is observed to pinch out to the east. Resources were depleted for
production and development from the underground mine, according to surveyed volumes. 112,000
tonnes of Mineral Resource was mined underground at an average of 8.5% copper.
The Measured and Indicated Resources for Dikulushi (Table 14.7) total 0.77 million tonnes at 6.6%
copper, and were determined above an economic cut-off grade of 1.0% copper. This is composed
of:


0.52 million tonnes at 7.0% copper in the Measured Resource category
0.25 million tonnes at 5.6% copper in the Indicated Resource category.
Table 14.7
Dikulushi Mineral Resource statement using a 1.0% copper cut-off grade as at August 2011
Category
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
184
90
2.8
2.8
516
251
7.0
5.6
211
114
274
2.8
767
6.6
179
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
136
2.8
380
6.8
91
14.2. MINERAL RESOURCE ESTIMATE COMPARISONS
14.2.1.
MINERAL RESOURCE STATEMENT AUGUST 2011 VERSUS OCTOBER 2007
The August 2011 Mineral Resource estimates were compared to those of October 2007. These
results (Table 14.8) reflect an overall tonnage decrease of 23%, together with a 7% increase in
copper% and a 6% decrease in silver grade. Variance is against all resource categories: Measured,
Indicated and Inferred.
Notable category changes include a 124% increase in Measured Resource category tonnes and an 8%
increase in Inferred Resource category tonnes, associated with the presence of additional data from
underground exposures and from drilling. Most of these resources represent conversion from
Indicated Resource material.
There is a significant decrease in the Measured Resource copper % grades associated with
extensional drilling within deeper, lower grade areas. In contrast the deeper infill and extensional
drilling has supported an increase in the Inferred Resource copper grades.
P a g e | 89
These comparisons were carried out using the 1.5% copper cut-off resource as the 2007 resources
were only available at that cut-off grade.
Table 14.8
Comparison of 2011 and 2007 Dikulushi Mineral Resource estimates
Dikulushi Mineral Resource statement as at August 2011, using a 1.5% copper cut-off
grade
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
Measured
Indicated
176
86
2.80
2.80
493
241
7.32
5.79
219
118
Total Measured & Indicated
262
2.80
733
6.82
186
Inferred
129
2.80
361
7.11
94
Dikulushi Mineral Resource statement as at October 2007, using a 1.5% copper cut-off
grade
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
Measured
Indicated
78
307
2.83
2.83
220
869
9.63
6.50
289
155
385
2.83
1,089
7.13
182
Inferred
119
2.83
336
4.30
112
Comparison by percentage variation between the August 2011 and October 2007 results.
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
-24%
-11%
Silver
(g/t)
Measured
Indicated
126%
-72%
-1%
-1%
124%
-72%
-32%
-1%
-33%
-4%
2%
9%
-1%
8%
65%
-16%
Inferred
-24%
-24%
The 2011 Mineral Resource estimates have been guided by additional drillholes, underground
sampling, density, geological and in-pit blasthole data available as of November 2008. The
additional data has enabled an increase of some 21,000 copper tonnes from previous Indicated and
Inferred Mineral Resources to be upgraded to a Measured category.
Figure 14.11 illustrates the relative and cumulative change in copper tonnes between the 2007 to
2011 estimates. The 2011 Mineral Resource estimate has dropped by 14%, or a total of 13,000
tonnes of copper. Some of this is associated with mining depletion and significant changes to the
volumes of mineralisation. Grade reductions for the Measured and Indicated categories are offset
by increases in the Inferred category.
P a g e | 90
Figure 14.11
14.2.2.
DEPLETION OF AUGUST 2011 MINERAL RESOURCES BY AUGUST 2013
OPEN PIT CUT BACK
The August 2011 Mineral Resource estimates were compared to those of August 2013, which
features the depleted Mineral Resources after the mining of the open pit cut back. These results, as
presented in Table 14.9 below, reflect an overall tonnage decrease of 37.3%, together with a 3.2%
decrease in copper% and a 9.3% decrease in silver grade (corrected for a previous reporting error).
The variance has been calculated for all resource categories, i.e., Measured, Indicated and Inferred.
Notable category changes include a 60% decrease in Measured Resource tonnes, a 41% decrease in
Indicated Resource tonnes, and a 3.9% decrease in the Inferred Resource category tonnes. All of
these changes are wholly due to depletion as a result of mining of the open pit cut back, which was
completed in July 2013.
P a g e | 91
Table 14.9
Comparison of August 2011 and August 2013 Dikulushi Mineral Resource estimates, showing the Open pit cut back
depletion
grade*
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
Measured
Indicated
184
90
2.80
2.80
516
251
7.0
5.6
211
114
274
2.80
767
6.6
179
Inferred
136
2.80
380
6.8
91+
+
grade
Category
Measured
Indicated
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
74
53
2.80
2.80
207
148
5.4
6.6
163
131
127
2.80
354
5.9
150
Inferred
130
2.80
365
7.0
160
Comparison by percentage variation between the August 2011 and August 2013 results.
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper
(%)
Silver
(g/t)
Measured
Indicated
-51.1%
-41.1%
0%
0%
-59.9%
-41.0%
-22.7%
+17.9%
-22.7%
+14.9%
-53.7%
0%
-53.9%
-10.6%
-16.2%
-4.4%
0%
-3.9%
+2.9%
+3.2*
Inferred
Note: + - The inferred silver grade was incorrectly reported at 91 g/t in the August 2011 Mineral Resource table and
should have been 155 g/t.
Note: * - This % has been corrected to show the “Real” comparison % based on the correction note above.
Figure 14.12 is a waterfall chart which illustrates the relative and cumulative changes in Mineral
Resource tonnes between the 2011 and the 2013 estimates. Figure 14.13 shows tonnage-grade and
metal-grade curves for the depleted remaining Mineral Resource as at August 2013 as tabulated in
Table 14.9.
P a g e | 92
Figure 14.12
Figure 14.13
Grade tonnage curves for the combined remaining Measured and Indicated Mineral Resources
Total Measured and Indicated tonnage and grade,
remaining Dikulushi Mineral Resource
25.0
0.35
20.0
0.30
0.25
15.0
Cu%
Resource tonnage and copper metal tonnage
0.40
0.20
10.0
0.15
0.10
5.0
0.05
0.00
0.0
0
2
4
6
8
10
12
14
16
18
Copper cut-off grade (Cu%)
Resource tonnes (1,000,000)
Cu tonnes (100,000)
Cu %
P a g e | 93
14.3. KAZUMBULA MINERAL RESOURCE ESTIMATE
The Kazumbula Mineral Resource estimates were prepared in November 2010 by Optiro in
conjunction with MWL geological staff, who worked on the Kazumbula Mineral Resources between
March 2009 and December 2010. The content of this technical report is guided by the reporting
requirements of the Canadian National Instrument 43-101 ‘Standards of Disclosure for Mineral
Projects’ and the JORC Code.
The Kazumbula deposit is located some 15 km north-northeast of the Dikulushi plant and has good
drillhole coverage from both Anvil and Mawson West. The Kazumbula deposit is relatively small
when compared to Dikulushi, and has a strike length of approximately 200 m.
RC and diamond drilling define the Kazumbula deposit with a grid spacing of approximately 20 m to
25 m. Only the MWL drilling data have been used for estimating the Kazumbula Mineral Resource
due to reliability issues with other drill campaigns.
14.3.1.
The Kazumbula mineralised volume (Figure 14.14) was delineated on vertical sections per drill line.
A 0.5% copper cut-off was used as a guideline for defining the mineralised volume. The delineated
string envelopes per section were linked with wireframe surfaces to define the mineralised volume
of the Kazumbula ore body. The Kazumbula mineralisation exhibits good downhole and betweenhole continuity. Mineralisation was also guided by the position of the interpreted fault surface. The
boundary between oxidised and sulphide mineralisation was modelled according to the logged
geology, oxidation and mineralogy; however, due to the relatively small size of the Kazumbula
orebody and the limited number of intersections within the oxide and sulphide domains, it was not
deemed appropriate to subdivide the mineralised domain.
P a g e | 94
Figure 14.14
Kazumbula vertical section, looking north, highlighting the modelled mineralisation as per the RC and diamond drilling
14.3.2.
The deposit was drilled by Anvil during 2008, but MWL was not able to completely validate these
drillhole data (i.e. logging, sampling, assay, drillhole collars and downhole surveys), and as a result
the Anvil data were not used in this resource estimate. The Anvil drilling data has, however, assisted
MWL with drillhole planning and targeting of the Kazumbula deposit. MWL drilled RC and diamond
holes (Table 14.10) to define the near surface copper mineralisation during August and September
2010. The drillhole spacing was approximately 15 m on drill lines spaced 20 m apart (Figure 14.15).
Drillholes were drilled at approximately 60 degrees to the south-southeast to optimise the angle of
intersection with the orebody. No twin holes were completed for this programme, but significant
mineralised intersections are comparable with those completed by Anvil.
P a g e | 95
Table 14.10
Prospect
Kazumbula
Kazumbula
Type
RC
DDH/tail
No Holes
17
10
Metres
1676
674.4
Samples
1676
674
The RC drilling was completed by Titan Drilling of Lubumbashi, utilising a truck mounted RC rig. A
supervising geologist was on site at all times during the drilling and industry standard procedures
were followed during the RC drilling programme. The diamond drilling was contracted and
completed by Chantete Emerald, who completed six diamond holes (HQ3) from surface and four
diamond tails from RC pre-collars (HQ3).
Figure 14.15
Plan showing the distribution of RC and diamond drillholes across the Kazumbula deposit.
14.3.3.
DATA VALIDATION
Drillhole collar coordinates were surveyed by the qualified Dikulushi mine surveyor. Collars were
verified against the topography. Visual inspection of the downhole survey measurements was
completed in order to identify any anomalously different bearings and dips. Micromine software
was used to validate the drillhole logging and sampling data for any gaps or overlaps, with only
minor errors identified. These errors were associated with typing and were corrected immediately
in the MWL database.
P a g e | 96
14.3.4.
The de-surveyed 3D assay drillhole file was composited to two metre lengths. As discussed, a single
mineralised domain was used to model the Kazumbula orebody. As a result, the two metre
composites were deemed appropriate for defining the extents of mineralisation and reduced some
of the variability associated with the one metre samples.
14.3.5.
STATISTICS
Statistical analyses of the two metre sample data, including spatial statistics, were carried out using
Supervisor software. The summary results are tabulated in Table 14.11.
Statistics for copper and silver grade data were investigated with histograms and probability plots
(Figure 14.16). The statistics highlight only minor internal variability and domain mixing, thereby
assisting with spatial analysis and supporting a reasonable estimate.
Table 14.11
Summary statistics of the two metre composite data for Cu% and Ag g/t for the Kazumbula deposit
Cu (%)
14.3.6.
Ag (g/t)
Samples
Min
Max
Mean
Std Dev
CV
Variance
Skewness
Log variance
Geometric mean
137
0.03
8.41
2.00
1.59
0.79
2.53
1.53
0.83
1.44
137
2.50
142.00
22.48
27.20
1.21
739.86
2.09
1.42
11.56
Log mean
137
137
SPATIAL STATISTICS
For the copper and silver variography, no definitive anisotropy was evident from spatial analysis. A
standardised nugget value of 0.37 was clearly defined from the downhole variography (Figure
14.17). Variography was oriented according to the plane of the fault and mineralisation, which dips
at 75 degrees towards 340. An isotropic variogram model (Figure 14.17) in this plane of
mineralisation was used to define the horizontal continuity of 60 m for copper and silver grades.
The true width variogram range was set to 20 m.
14.3.7.
BLOCK MODEL
The block model dimensions were guided by the mineralised wireframe shape, orientation and
volume, together with the drill grid spacing. Sub-blocks were used to ensure that the block model
honoured the wireframe volume. The block model volume was within 1% of the mineralised
wireframe volume. The individual parent block dimensions were 15 mE by 10 mN by 2 mRL, with
sub-blocking allowed down to 5, 4 and 1 m respectively. Block estimates were controlled by the
original parent block dimension.
P a g e | 97
Figure 14.16
Histogram and probability plots for the Kazumbula deposit two metre sample data.
P a g e | 98
Figure 14.17
Variogram modelling for Cu % in the plane of mineralisation.
14.3.8.
Diamond core samples were prepared by ‘squaring off’ the ends of approximately 10 cm to 20 cm
billets of half core. A total of 118 bulk density (BD) measurements were made of dried half core.
The same piece of core was then measured in water on a suspension cage below the same electronic
scale. The conventional formula for BD was then applied, viz.
BD = Dry Sample Weight / (Dry Sample Weight – Wet Sample Weight)
The average BD measurements and statistics for Kazumbula for mineralised and un-mineralised core
are shown in Table 14.12. The lower density for the mineralised oxide samples is explained by the
lack of hematite alteration in the mineralised samples, being replaced by an illite clay assemblage
and lowering the overall density of the rock.
Table 14.12
Density estimates for the Kazumbula deposit
Type
Mineralised Oxide
Unmineralised Oxide
Mineralised Sulphide
Unmineralised Sulphide
14.3.9.
SG
2.41
2.47
2.65
2.61
GRADE ESTIMATION
Grades for copper and silver were estimated into the parent blocks of an empty block model using
ordinary kriging (OK). OK is believed to be an appropriate estimation method due to the pseudonormal data distributions of the mineralisation. Estimation into parent blocks uses a discretisation
of 6 (X points) by 6 (Y points) by 2 (Z points) to better represent estimated block volumes, in addition
to applying an octant sample selection strategy of four sectors and a minimum of 4 samples and
maximum of 20 samples per sector.
P a g e | 99
14.3.10.
MODEL VALIDATION
The model was validated by visual techniques. Block estimates, while smoothed, do reflect the
average higher and lower grades of the drillhole samples. In addition, the model was also validated
with an inverse distance to the power of two estimate (ID2), using the same sample selection
parameters and search parameters. The mean grade estimate compared to the mean drillhole value
and the mean ID2 estimate are all close to within 10% of each other. The declustered mean of the
drillhole data has a value of 1.8% copper (Table 14.13).
Table 14.13
A table of mean statistics comparing model estimates with data values
Field
Copper %
Data
2.0
2
ID
2.0
Declustered data
1.8
OK estimate
1.8
From the summary statistics and visual validations, the OK estimates are consistent with the drillhole
composites, and while smoothed, are believed to constitute a reasonable representation of the
Kazumbula grade.
14.3.11.
Classification of the Kazumbula Mineral Resource was based on quality of sample assays, grid
spacing, the assigned density and the resulting kriged estimates. The Kazumbula deposit has been
classified in its entirety as an Indicated Mineral Resource. The mineralised volume is adequately
supported by a regular 20 m grid of drillhole intercepts and has been defined using an effective 0.5%
copper cut-off.
The Mineral Resource has been classified and reported using the guidelines of the JORC Code (JORC,
2004), which in turn comply with the Standards on Mineral Resources and Reserves of the Canadian
Institute of Mining, Metallurgy and Petroleum (CIM, 2005). A resource summary is given in Table
14.14.
Table 14.14
Kazumbula Mineral Resource statement as at November 2010.
Category
Indicated Oxide Mineral Resources
Indicated Sulphide Mineral Resources
Total Indicated Mineral Resources
Volume
3
(m *1,000)
66
60
126
Density
3
(t/m )
2.41
2.65
2.52
Tonnes
(*1,000)
159
160
318
Copper
(%)
1.75
1.89
1.82
Silver
(g/t)
14.7
22.9
18.8
P a g e | 100
15. MINERAL RESERVE ESTIMATES
15.1. DEPLETION OF THE OPEN PIT RESERVES
The open pit Mineral Reserves have been depleted with the mining of the Dikulushi open pit cut
back, which commenced in August 2011 and was completed during July 2013. Processing of the
stockpiled cut back ore is continuing, and it is estimated that this ore will be processed by midDecember 2013. Table 15.1 shows the Dikulushi Mineral Reserves statement as at August 2011.
Table 15.1
Dikulushi Mineral Reserve statement as at August 2011, using a 1.0% copper cut-off grade
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper (%)
Silver
(g/t)
Proven
Probable
66.6
127.8
2.8
2.8
184.7
354.3
7.27%
5.51%
207
169
194.4
2.8
539.0
6.12%
182
The Open Pit Mineral Reserve estimate was based on the Open Pit reaching the 810 mRL. Mining
ceased at the 825 mRL due to safety concerns, with some isolated sections of the pit wall
deteriorating beyond what was predicted. Table 15.2 shows the depleted Mineral Reserves post the
cessation of mining of the open pit cut back.
Table 15.2
Depleted Dikulushi Mineral Reserve statement as at August 2013, using a 1.0% copper cut-off grade
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper (%)
Silver
(g/t)
Proven
Probable
1.0
29.9
2.8
2.8
2.7
83.7
6.8
5.5
186
188
30.9
2.8
86.4
5.5
188
The above remaining Mineral Reserves have been incorporated into the Underground Mineral
Reserves, which are now presented in the following sections and discussions below.
15.2. UNDERGROUND MINE DESIGN AND SCHEDULE BASIS
15.2.1.
EXISTING WORKINGS
Mine design for the Dikulushi underground mining operations has utilised the existing decline level
development completed by Anvil, which was completed prior to abandoning the underground. The
mining operations consisted of a decline and level development down to the 750 mRL. The capital
and development design criteria now used by MWL for the underground Mineral Reserve estimation
are the same as the original underground workings completed by Anvil, with the following design
cross-sectional areas:


decline and decline stockpiles; 5.5 m high by 5.5 m wide arched profile
level access / level stockpiles; 5.0 m high by 5.0 m wide arched profile
P a g e | 101




return airway level drives; 5.0 m high by 5.5 m wide arched profile
cut and fill ramp access; 5.0 m high by 4.5 m wide square profile
ore drives / cut and fill; 5.0 m high by 4.5 m wide square profile
return airway rise; 5.0 m diameter.
Prior to the abandonment of the underground mining operations, Anvil completed stoping activities
in the upper levels of the mining operations between the 910 mRL and 850 mRL. The stopes were
mined using long-hole stoping. The success of stopes was limited due to ore dilution from the
surrounding waste material located around the hanging wall. Handheld stoping was also trialled in
the latter stages in some selected stopes, with better success.
Figure 15.1 shows the open pit cut-back and the original underground workings, as developed
previously by Anvil, which will now be utilised by MWL to form the basis of the underground Mineral
Reserve development and production. Additional development will be undertaken to access the undeveloped Mineral Reserves.
Figure 15.1
Existing workings, showing as built underground development (grey), and the as-built pit (green)
P a g e | 102
15.2.2.
BASIS OF THE UNDERGROUND DESIGN AND SCHEDULE
The Mining schedule and designs focussed on the Measured and Indicated Mineral Resources, and
were used in the development of the Mineral Reserves. No Inferred Mineral Resource Material was
used in the schedule or for economic evaluation of the Mineral Reserves. Figure 15.2
Underground reserve, showing as-built underground development (grey), as-built pit
(green), and measured (purple) and indicated (red) Mineral Resources and additionally shows the
relationship between the current development (as built), and the Measured and Indicated Mineral
Resources.
Figure 15.2
Underground reserve, showing as-built underground development (grey), as-built pit (green), and measured (purple)
and indicated (red) Mineral Resources
15.3. CUT-OFF GRADE CRITERIA
Cut-off grade evaluation for the deposit has been completed using a Net Smelter Return (NSR)
calculation method due to the polymetallic nature of the deposit, as both the silver and copper
metals provide significant value to the final revenue of the mine.
The first part of the calculation process was to identify the NSR of one tonne of material with an
average grade of 5.38% copper and 128 g/t silver using Equation 15.1. The average grades for the
copper (x1) and silver (x2) were taken from the Mineral Reserve schedule.
Equation 15.1
Polymetallic NSR using average metal grades
(
)
(
)
(
(
)
(
)
P a g e | 103













x1 - is the average copper grade expressed as a percentage
x2 - is the silver grade expressed as a percentage
r1 - is the percentage of copper recovered when producing the concentrate
r2 - is the percentage of silver recovered when producing the concentrate
p1 - is the copper concentrate percentage received after smelting charges
p2 - is the copper concentrate percentage received after smelting charges
V1 - is the market value of copper per tonne
V2 - is the market value of silver per tonne
R1 - is the refining cost of the copper per tonne of copper produced
R2 - is the refining cost of the silver per tonne of silver produced
Cs - is the smelting and refining cost per tonne of copper concentrate
Ct - is the smelting and refining cost per tonne of copper concentrate
K - is the number of ore tonnes required to be mined to produce one tonne of copper
concentrate.
The values used in the calculation were derived from both the mining schedule created for the
Mineral Reserves Schedule and MWL’s financial model. The K value used was calculated using the
tonnes and grade from the mining schedule. From the application of Equation 15.1 a NSR value of
one tonne of ore averaging 5.38% copper and 0.0128% silver was calculated. The NSR value of one
tonne using the average copper and silver grades from the schedule was $329.
To allow the calculation of the NSR cut-off value for one tonne of rock with variable copper and
silver grades, Equation 15.2 has been derived from Equation 15.1. This equation provides a
mathematical relationship that allows the calculation of one variable component, if the other two
variable components are known, providing a breakeven revenue position. For example, if the
copper grade and an NSR cut-off value are known for a tonne of material, the silver content can then
be calculated. This allows the metal content of each tonne mined to be economically evaluated with
the aim of determining whether the material should go to the waste dump or be sent to the mill for
processing.
Equation 15.2
Variable grade NSR equation
A stoping
(cut-off) value for the underground mining operations was determined using
Equation 15.3. The stope cut-off value relates to the drive/drift development used to extract ore
from the level development.
Equation 15.3
NSRc stope Stope cut-off equation
(




)
- is the copper processing cost per tonne of ore mined and milled
- is the incremental silver processing cost per tonne of mined and milled
- is the waste processing cost per tonne mined
- is the ore mining cost per ore tonne mined
P a g e | 104
The mining costs used in the evaluation of the
were taken from the mining schedule
prepared for the Mineral Reserves, with the processing costs from MW’s financial model. As the
processing cost provided did not differentiate between the individual metals contained, the variable
for incremental cost of silver processing was set to $0 per ore tonne. When applying the mining and
processing costs, a NSRc Stope value of $143.44 per tonne is required to determine whether the
material should be mined or not.
The development cut-off
was calculated using the Equation 15.4. The NSR value from this
calculation is used to determine whether material that must be mined from the underground
operation should be sent to the mill or waste dump. This calculation determines the NSR value
based on the difference between the mining and processing cost of ore and waste.
Equation 15.4 – NSRc dev Development Cut-off equation
(



)
(
)
(
)
- is the overhead costs associated with the mining and processing 1 tonne of ore
- is the overhead costs associated with the mining and processing of 1 tonne of waste
- is the waste mining cost per waste tonne mined
When applying Equation 15.4, a
value of $71.38 per tonne is required to determine if a
tonne of material that has to be mined should be processed or not.
Due to the polymetallic nature of the orebody, the cut-off grades are variable. Figure 15.3 shows a
relationship between the metal grades and the NSR cut-off for both the development and stope cutoffs. Each of the lines shown on the graph has been calculated using the variable grade NSR
Equation 15.2, with the NSR value in the equation substituted with either the NSRc Stope or
.
The silver content in g/t was then calculated for a variety of copper grades to produce the graph
lines.
P a g e | 105
Figure 15.3
Relationship between cut-off NSR and metal grades
15.4. MINING RECOVERY AND DILUTION
Mining ore loss has been estimated at 5%. Ore loss is estimated to occur as a result of small gaps
left in the backfill process, caused by the inability to properly level off the top of backfill. This means
that when using the backfill as the floor for the next level, the fill will generally have a rounded
profile, and this will result in some material being lost in these corners as the next level is developed.
There will also be a small amount of ore loss resulting from general activities, such as rehandle and
material movement.
Mining dilution for the project has been estimated at 8%. This has been calculated by allowing a
200mm dilution skin on the walls of development drives, which has then been used to determine the
percentage dilution using the cross-sectional area of the drives. The figures use the average
dimensions for development drives, and also assume a square cross-section for calculation purposes.
Table 15.3 outlines the specific numbers for each development type.
Table 15.3
Mining dilution table
Development
Width
(m)
Height
(m)
Drive
Area (m2)
Dilution
width (m)
Dilution skin
area (m2)
Dilution
(%)
4.50
5.00
22.50
0.20
2.00
8.9
5.50
5.50
5.00
5.00
5.00
4.50
5.50
5.50
5.00
5.00
5.00
5.00
30.25
30.25
25.00
25.00
25.00
22.50
0.20
0.20
0.20
0.20
0.20
0.20
2.20
2.20
2.00
2.00
2.00
2.00
7.3
7.3
8.0
8.0
8.0
8.9
Ore
Ore Drive
Waste
Decline
Decline Stockpile
Level Access
Level Stockpile
Return Airway Drive
Cut and Fill Ramp
P a g e | 106
Average
5.08
5.17
0.20
26.26
2.07
7.9
15.5. UNDERGROUND MINERAL RESERVE TABULATION
A financial model has been developed and analysis indicates that a positive return is expected.
There are several areas where it is considered by Optiro that conservative estimates of mining costs
have been made. It should be noted that the waste dilution rate of the ore is considered to be
conservative for this type of operation, mining style and deposit.
The resulting Mineral Reserves for the underground mine are only based upon Measured and
Indicated Mineral Resources from the depleted Mineral Resources post mining of the open pit cut
back. The Mineral Resources have been classified as Indicated due to the risks associated with
underground mining of this deposit.
The resulting Mineral Reserves are supported by historical production and current processing data
and are tabulated in Table 15.4, using a cut-off grade based on an NSR value of US$329/t, at a
copper price of US$3.08/lb and a Silver price of US$20 per oz. All stated Mineral Resources are
inclusive of Mineral Reserves. The Mineral Reserve, as per the CIM definition, incorporates mining
losses and diluting materials brought about by the mining operation.
Table 15.4
Dikulushi Mineral Reserve statement as at September 2013
Category
Volume
3
(m *1,000)
Density
3
(t/m )
Tonnes
(*1,000)
Copper (%)
Silver
(g/t)
Proven
Probable
0
62
0
2.8
0
173
0
5.2
0
127
62
2.8
173
5.2
127
Note:
1) Cut-off grade is based on a NSR value of US$329/t, at a copper price of US$3.08/lb and a
Silver price of US$20 per oz.
2) The above ore reserve does not include any Inferred category Mineral Resource material
P a g e | 107
16. MINING METHODS
16.1. HISTORICAL MINING
Anvil commenced Underground Mining operations at the Dikulushi deposit in 2006 and continued
through until 2008, when falling copper prices and challenging mining conditions led to the
underground operations being abandoned. During the period of underground mining, mine
development was completed from the surface down to the 750 mRL, with extensive level
development completed from the 790 RL to the surface. Longhole open stoping activities were also
completed between the 900 and 850 mRL’s at the eastern end of the deposit.
MWL acquired the Dikulushi mining operations through its purchase of Anvil’s subsidiary AMC in
2010, and recommenced the open pit mining operations through the implementation of a cut back
of the original pit. The cut back was designed to reach the 805 mRL, and in doing so, would mine out
parts of the “old” underground workings that had been previously developed by Anvil. Mining of the
cut back commenced in late 2011 and continued through to July 2013, when mining was stopped at
the 825 mRL. This was 20 m short of the cut back design depth, due to safety concerns with some
isolated sections of the pit wall deteriorating beyond what was predicted.
16.2. PROPOSED MINING METHOD – CUT AND FILL
The planned mining method to be used at the Dikulushi underground operations is cut and fill, using
a combination of overhand and underhand variants. Initial ore production will be mined from the
800 mRL down to 755 mRL. Sections of the ore body between these two levels have good and poor
rock mass characteristics and contain the bulk of the Measured and Inferred material contained
within the Mineral Resource. Due to the variability of the rock mass in the zone, the backfilling of
drives will be completed using one of three backfill methods



Cemented rock fill (CRF),
Cemented aggregate fill (CAF)
Un-cemented rock fill (RF).
Use of the different backfill types as planned in the mining schedule has been determined by the
mining sequence and related mining activities that will be conducted after the fill has been placed.
16.2.1.
OVERHAND CUT AND FILL
Overhand cut and fill is a variant of cut and fill which moves from the bottom of the orebody
upwards, using the fill from previous levels as the floor for the next level. Levels accessed are
excavated from the decline through to the orebody, perpendicular to the strike of the orebody.
From these accesses, ore drives are mined out along the strike of the orebody (shown in Figure
16.1), and once a level is fully extracted it is backfilled with CRF. Once the cement is cured, mining
recommences directly on top of the backfilled level, using the backfill as the floor for the next ore
drive. Overhand is the more productive of the two cut and fill variants suggested, and can be used
where the ore (which forms the backs of the mining levels) is competent enough to work under
P a g e | 108
safely. Overhand cut and fill is also initially more capital-intensive than underhand, as it requires
development to the bottom of the ore block before commencing mining in an upwards direction.
Figure 16.1
Overhand cut and fill mining process
16.2.2.
UNDERHAND CUT AND FILL
Underhand cut and fill mining works in a similar way to the overhand variety; however, mining
advances downwards underneath previously mined and backfilled levels, with backfill forming the
backs of each successive level. This is done by driving an access from the footwall of the deposit
through to the orebody. In the narrow sections of the orebody stoping is done by developing a
single strike drive to extract the ore on the level. The strike drives are developed and designed to
minimise openings created for the extraction of ore, with the openings kept as small as possible thus
minimising the amount of waste dilution taken with the ore being extracted. Once the orebody on
the level has been fully extracted using the drive, cemented fill (CAF) is then used to fill up the void
created by mining. After the backfill has been given sufficient time to cure and obtain the
appropriate strength requirements, the next level is then developed directly below the level above
with the cemented backfill forming the backs (roof) of the stoping drives developed below. Figure
16.2 shows the development sequence of this mining method in a horizontal long section view.
Underhand cut and fill is the less productive and a higher cost method; however, it has the
advantage of being able to control the condition of the backs. This means that it is suitable for any
areas where the ore is not competent enough for the overhand method, or anywhere where there is
a high propensity for rockbursts.
P a g e | 109
Figure 16.2
Underhand cut and fill mining process
The primary mining method that has been selected and scheduled is the overhand cut and fill
method, which accounts for 70% of the ore extraction. The underhand mining method is primarily
used where the ore is extracted around previously developed levels and accounts for approximately
20% of the ore. The remaining 10% of the ore is extracted by Longhole stoping methods, which are
confined to stope areas within the crown pillar below the pit floor.
16.2.3.
MINING OF WIDER SECTIONS OF THE OREBODY
In the sections of the orebody where the mineralisation exceeds the drive development width
(4.5m), multiple drives will be developed to extract the ore in these zones. The development
sequence of all drives will involve developing the initial drive along the hanging wall contact to the
end of the orebody. This drive will then be filled with either CRF or CAF. Once the fill has been
allowed to cure in the hanging wall drive, a second drive in the footwall will developed. Once the
footwall drive development has been completed, CRF or CAF will be used to backfill the drive.
Additional drives will then be developed to extract the remaining parts of the orebody between the
hanging wall and footwall contact drives. This process should allow for complete extraction of the
orebody where it exceeds the maximum drive width specified under the geotechnical requirements
(4.5 m). Figure 16.3 is an example of the extraction sequence for sections where the orebody width
exceeds the maximum stoping drive width.
P a g e | 110
Figure 16.3
Diagrammatic representation of sequential mining in wide orebody areas
5
3
5
15m
2
4
1
3
6
7
8m
1
Plan view
The current design work produced for the underground has access of the decline to the orebody at
20 m vertical spacing. Each of the level take-off points from the decline will provide access to four
stoping levels, with the levels developed from top to bottom (Figure 16.4 shows the access
development of the stoping levels from the decline). The bottom stoping drives from each level
access will sit directly above the top stoping drive from the decline access, 20 m below.
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Figure 16.4
Orebody access development
As the mining sequence will require multiple levels to be mined in the same period to provide ore
feed to the processing plant, controls will need to be in place to maintain a minimum horizontal
distance between mining levels. The vertical spacing between all drives should be a small proportion
of the maximum void space created from the level above, in order to maintain a 1 to 1 void to pillar
ratio (Figure 16.5) between active mining levels. This has been employed to minimise the risk of
rock mass failure between the mining levels.
Figure 16.5
Pillar ratio diagram
16.2.4.
PROPOSED MINING METHOD – EXTRACTION OF THE CROWN PILLAR
Extraction of the crown pillar from the 815 mRL to the base of the pit at the 825 mRL is to be
extracted using a longhole stoping method. The extraction of ore via this method will involve mining
through back filled sections of the 810 mRL to create a drive for the drilling of the ore from
underground, with the blasted ore transported to the mill via the underground decline. Extraction
and transportation of the ore via the open pit workings was not considered due to the possible
undercutting of the existing pit walls.
Removal of the crown pillar, as set out in the schedule, is undertaken at the end of the underground
mine’s operational life as it stands. The reason for this is to minimise the risk of water inrushes into
the underground and the potential destabilising of the open pit walls. MWL plans to continue
P a g e | 112
underground exploration drilling of the orebody, and will review the success of those programs prior
to undertaking to remove the crown pillar.
16.3. GEOTECHNICAL DESIGN PARAMETERS
16.3.1.
STOPE LAYOUT AND SEQUENCE
The mining sequence and layout developed for the mineral reserves schedule has been created in
consultation with both Mike Turner from Turner Mining and Geotechnical Pty Ltd and John Keogh
from Peter O’Bryan & Associates. For the purpose of the stope layout and designs, drives have been
limited to 4.5 m wide by 5 m high in the ore body. This is due poor ground conditions experienced
when mining was originally conducted.
16.3.2.
DRILL AND BLAST
Rock breakage for the development of the underground mining operations will be conducted using
drill and blast. There are currently three development drills (Jumbos) located at the Dikulushi
operations from previous mining activities. These three jumbos will be used to complete all drilling
activities for horizontal mine development and stoping within the mining operations for ore and
waste. The three jumbos on site include two twin boom jumbos and one single boom jumbo, with
all three rigs able to drill holes using a 3.7 m steel. For the purposes of scheduling the underground
mining operations it was decided that each jumbo would be able to take a mining cut to a depth of
3.3 m.
Blasting of the rock will be done primarily using ANFO as the main explosive. In development cuts,
reduced charging will be used in the perimeter holes of each cut to minimise the impact of the
explosives on the drive walls and backs.
On completion of drill and blast activities, the installation of weld mesh and steel bolts or split sets
will be used to provide ground control for the drive walls and backs. Ground support work will be
predominantly done using the single boom jumbo, with the twin boom jumbos bolting and meshing
as required.
16.3.3.
ORE EXTRACTION
Extraction of the ore and waste material after blasting will be completed using standard
underground loading and hauling practices. Ore material will be removed from the face of the drive
by an LHD (load –haul-dump unit) to stockpiles located on the level. The ore is then loaded onto
haul trucks and transported to the mill processing area on the surface. Waste material will be
removed from the underground workings in the same manner as the ore. The waste material used
for the production of CAF or CRF will be transported to surface where it will be screened or crushed
to provide the appropriately sized material for each of the filler types. Where RF is to be used in the
underground workings, this material will be transported to stockpiles in strategic locations
underground for use as backfill on levels where ore mining has been completed.
P a g e | 113
16.3.4.
BACKFILLING
Cut and fill mining methods rely heavily on the use of backfill, with the mining process using backfill
material to provide stable material for either the floor (overhand cut and fill) or the backs
(underhand cut and fill) or walls for the extraction of adjacent stoping drives. Backfilling operations
for the Dikulushi underground will use three different types of back fill for the safe extraction of the
orebody.
ROCK FILL (RF)
The use of RF within the mining schedule has occurred where all of the mining activities beside or
below the location backfill have ceased, and the RF is used to provide the floor for extraction of the
level directly above. This fill method is used extensively in certain sections of the orebody where a
single drive is required to extract the ore from the level, and mining activities below the drive have
been completed. Some levels use a combination of CRF and RF. RF is significantly cheaper than both
the CRF and CAF backfilling methods, as no secondary screening or crushing is required and the RF
can be directly hauled between locations with material dumped in a stockpile close to the backfill
location.
RF DESIGN AND PLACEMENT
Placement of the RF will be done using underground LHD units, with the rock placed in the each of
the drives working from the furthest end of the drive back towards the access. In addition to moving
the RF to the fill location the LHD unit can be used to push up and pack in the RF (commonly by using
a ‘rammer-jammer’ attachment on the bucket, as shown in Figure 16.6.
Figure 16.6
LHD loader with ‘rammer-jammer attachment
CEMENTED ROCK FILL (CRF)
Selection of CRF as the primary method for backfilling stopes was done on the basis that materials
required for the process were easily available, with waste development providing the rock fill and
cement being easily imported to site with minimal technical or specialised gear required to produce
the fill. The process for creating CRF involves the mixing of cement, water and rock from waste
headings using a LHD unit in a stockpile located close to the area to be backfilled.
P a g e | 114
Once each of the ingredients has been mixed the LHD unit then transports the mix from the
stockpile and deposits it in the area being backfilled in a layered approach. A half to one meter layer
of CRF material is deposited onto the floor of the drive. The LHD unit then mixes more CRF, which is
deposited on top of the initial layer. As the LHD unit deposits the next layer of CRF it compacts the
layer below, helping to improve the strength of the CRF by removing any voids.
To maximise the fill in the drive and minimise any voids between levels, the CRF in the upper
sections of drive should be pushed up and packed in with the loader (commonly by using a ‘rammerjammer’ attachment on the bucket, Figure 16.6).
To ensure there is sufficient supply of rock material, additional waste will be sourced from the open
pit stockpiles. On completion of open pit mining operations, material in the pit was blasted, but not
removed; this has been identified as an alternate source of rock fill, should the ground waste
development not provide the volumes required to achieve the underground mining production.
CRF DESIGN AND PLACEMENT
CRF has been designed conservatively at a minimum of 10% cement, and rock fragments of 300 mm
diameter or less. A water to cement ratio is designed at 0.45, i.e. for a mix of 1 tonne cement and 9
tonnes rock (10 tonnes total), 450 litres of water is required (Table 16.1).
Table 16.1
CRF Specifications
CRF Specifications
Cement content
Rock fragments
Water:Cement ratio
Curing time until full strength is obtained
≥10%
≤ 300mm
0.45
28 days
The floor is to be covered in steel mesh, to be anchored into the sidewalls using resin encapsulated
rebar bolts (post-tensioned) at a spacing of 1.2 – 1.5 m and where mesh sheets overlap. The mesh
sheets are also to be shackled together along the overlaps.
The minimum curing time of the CRF prior to the development of adjacent development has been
set at 2 days within the schedule. This has been deemed as a significant time for the CRF to obtain
ample strength to allow mining activities to occur alongside them.
CEMENTED AGGREGATE FILL (CAF)
Cemented Aggregate Fill (CAF) is similar to CRF in design; however, the rock has been sized (and can
include sand) so as to achieve maximum strength. CAF is more expensive to produce than CRF due
to the additional requirement of crushing and screening the rock prior to use (as CAF) as backfill. It
is intended that CAF will be used on any levels that will subsequently form the backs for a level
directly below.
CAF DESIGN AND PLACEMENT
CAF has been designed conservatively at a minimum of 10% cement and rock fragments of 25 mm
diameter or less. The water to cement ratio is designed at 0.45, i.e. the CRF can be mixed at the
P a g e | 115
surface using AGI trucks or by LHD using a mix of 1 tonne cement and 9 tonnes rock (10 tonnes
total), 450 litres of water is required (Table 16.2).
Table 16.2
CAF Specifications
CRF Specifications
Cement content
Rock fragments
Water : cement ratio
Curing time until full strength is obtained
≥10%
≤ 25mm
0.45
28 days
The floor is to be covered in steel mesh, to be anchored into the sidewalls using resin encapsulated
rebar bolts (post-tensioned) at a spacing of 1.2 – 1.5 m and where mesh sheets overlap. The mesh
sheets are also to be shackled together along the overlaps.
The minimum curing time of the CRF prior to the development of adjacent development has been
set at 2 days within the schedule. This is being deemed as significant time for the CRF to obtain
ample strength to allow mining activities to occur alongside them.
CEMENTED PASTE FILL
Testwork is being conducted by MWL to identify whether the tails material produced from the
production of the copper concentrate is suitable for use as paste backfill in the underground
operations. Initial results of the testwork have indicated that there is a high likelihood of this
material being suitable for backfilling operations. Should the testwork prove successful paste fill
could be used as an alternative to either CRF or CAF. The advantage of paste fill is its ability to
provide a better filling ratio in the backfilled drives, with minimal void space remaining after the
completion of the backfilling process. The use of paste fill is highly unlikely in the current reserve
mining plan due to the short mine life based on the Measured and Indicated Reserves. Should
Inferred material within the geological reserve model be able to be upgraded to higher confidence
categories, paste fill would provide a viable option for future mining activities.
16.3.5.
ACCESSING THE OREBODY & REHABILITATION OF OLD WORKINGS
Initial access to the ore body will be provided by the existing workings left behind by Anvil. This
development includes extensive level development on the 810, 790 and 770 mRLs. The decline has
been developed down to the 750 mRL and includes the initial take-off drive development for level
access. Development of the decline below the 750 mRL will be completed using twin boom
underground development drills, with ground support implemented as per the geotechnical
recommendations from Mike Turner’s report ‘Geotechnical Input for re-opening Dikulushi
Underground’, August 2013.
Dewatering activities have made it easy to access the main decline. Prior to closure of the
underground mine in 2008, M. Turner (Geotechnical Consultant) undertook a detailed inspection of
the underground working areas. He highlighted a number of areas which required remedial support,
and the need for ongoing inspections for mining activities to proceed. In late 2010, MWL asked
Australian Mining Consultants (AMC) to undertake a geotechnical assessment of underground
P a g e | 116
mining options proposed by the previous owners Anvil. As part of this work AMC provided the
following comment: “Rehabilitation requirements below 900mRL are unknown, but for budgeting
purposes, it should be assumed that all of the (ground) support will require replacement.”
In December 2012 an inspection was performed by M. Turner to assess the decline down to the
groundwater level (830 mRL). This concluded that the rock mass and support in the decline was in
good condition, limiting the number of areas requiring support rehabilitation. Some areas would
require bleeding of scats from behind damaged and sub-standard mesh, but most areas could be
made safe by installing additional sheets of mesh over the existing support. Additionally, it was
observed that some areas below 900 mRL had corroded mesh. After testing the mesh strength,
MWL was satisfied that this would still perform as required. As dewatering proceeds, ongoing mesh
testing is planned.
As open pit mining has proceeded, the previously mined crosscuts and ore drives have been
intersected. The installed ground support, which includes galvanised Split Set bolds and mesh, is
exposed on the pit floor. This has allowed close inspection of installed support, and very little
corrosion was consequently noted. This has been consistent when mining down through the various
ore drives. Again, this lends weight to the premise that there has been only minor deterioration of
the installed ground support.
16.4. VENTILATION
16.4.1.
PRIMARY VENTILATION
A series of primary ventilation rises will be developed in conjunction with the main decline. These
rises will be mined using hand-held mining equipment, with a rise driven between the return air
drives at a 1.5 m width and then stripped out from the top down to a diameter of 5 m. The rises will
be connected to the decline through a series of small return airway drives strategically positioned at
20 vertical metre intervals. Ventilation of the decline between return airway levels will be provided
through secondary fans in the development phase. A new section of the primary ventilation system
will be developed, and ventilation bulkheads will be used to seal off the return airway drives located
above the lowest return airway drive. This process will then provide primary ventilation to the
bottom of the decline, allowing for further development. Figure 16.7 shows the proposed
underground primary ventilation circuit down to the 520 mRL created by Red Rock Engineering.
Figure 16.8 shows the ventilation development required for mining the existing Measured and
Indicated material in the Mineral Reserve.
P a g e | 117
Figure 16.7
Underground primary ventilation circuit (full)
Primary Ventilation
Return Air Rises
Figure 16.8
Underground primary ventilation circuit required for the extraction of the measured and indicated material only
Primary Ventilation
Return Air Rises
The ventilation system sets up the return airways from each of the levels up to the 825 mRL, where
the primary fans will be located. These fans will draw air out of the return airway system and
exhaust it into the current open pit, with fresh air being drawn down the decline to complete the
ventilation circuit. Figure 16.9 shows the proposed primary fan location.
P a g e | 118
Figure 16.9
Primary ventilation fan location
Underground as built
design information
16.4.2.
Primary Ventilation
Fan Location 825 mRL
SECONDARY VENTILATION
Secondary ventilation will be provided using smaller fans located in the decline, with ventilation bags
extending into the levels and secondary ventilated areas. MWL plans to use the secondary
ventilation fans and the electrical installations that were kept from the original underground
workings completed by Anvil. All this equipment will be refurbished and made fit for purpose prior
to re-installation in the underground workings.
16.5. DEWATERING
When mining the Dikulushi open pit cut back, MWL decided against dewatering from within the
open pit, and instead decided to access the existing decline and use it as a pumping platform. A
110mm diameter line was connected to a small floating pontoon within the decline, and as the
water level receded, the pontoon was shifted down the decline. This strategy successfully
minimised the impact of flood water on open pit mining. In the main decline, dewatering was rapid
until established levels were encountered at 20 m vertical intervals, below 900 mRL. The extensive
development encountered on these levels stored a significant amount of water and consequently
reduced dewatering rates. As the open pit was mined, depressurisation holes were drilled at 10m
vertical intervals, and 20 m horizontally apart within the walls. Significant volumes of groundwater
were encountered in the eastern and western walls of the pit, which was allowed to flow onto the
pit floor, eventually making its way into the underground workings. Prior to the 2012-13 wet
season, (October 2012), the groundwater level had been lowered to the 820 mRL. During the wet
season, a number of heavy downpours dramatically increased water flow into the underground,
such that by the end of April 2013, the water level had risen to 849.5 mRL. Dewatering activities at
the time were limited by the equipment available at the time of the rain events.
The commencement underground mining will depend upon the success in reducing the water level
to below the planned working areas. For initial mining, the water level will need to be below the 770
mRL level. The near term mining will need to effectively dewater all underground development
P a g e | 119
down to the 750 mRL. For long term mining, ground water inflow will have to be carefully managed
to minimise its impact on mining activities.
Long-term dewatering of the underground mining operation would use the existing dewatering
sumps and pumping locations originally set up. Dewatering stations will be set up as the water level
is lowered throughout the existing mine workings. It is anticipated that dewatering equipment left
behind by Anvil, once refurbished, will be suitable to control the water entering underground
workings. In addition to the existing pumping locations and sumps, new sumps and pumping
locations will be developed as the decline is advanced down to lower regions of the orebody. Figure
16.10 and Figure 16.11 show the existing underground dewatering infrastructure locations and
proposed future dewatering locations respectively. The existing dewatering infrastructure
implemented by Anvil had a sump located along the decline, with vertical distances between 60 and
75 m between stations. All dewatering pipelines were run between pumping stations using the main
decline.
Figure 16.10
Existing underground dewatering infrastructure locations
Figure 16.11
Proposed underground dewatering infrastructure locations
New Sump
Location
P a g e | 120
16.6. MINING EQUIPMENT
With the purchase of Anvil, MWL acquired all the mining equipment associated with the mining
operation. This included all of the underground equipment previously used by Anvil in the
underground operations closed in 2008. MWL plans to refurbish this equipment and use it as the
underground mining fleet for extraction of the underground Mineral Reserves. Table 16.3 is a list of
the major pieces of mining equipment available. Additional mining equipment has also been
retained from the previous underground mining activities, including dewatering pumps, ventilation
fans, starter boxes, electrical supply and distribution equipment that was recovered as part of the
abandonment process.
Table 16.3
Dikulushi production mining equipment at site from previous mining activities
Equipment type
Make
Model
Description
Development Drill Rig
Haulage Truck
Haulage Truck
Haulage Truck
LHD
Diamond Drill
Diamond drill
Integrated tool carrier
Sandvik
Sandvik
Sandvik
Sandvik
Sandvik
Atlas Copco
Sandvik
Boart Longyear
Kempe
Caterpillar
Axera 6
Axera 6
Axera 5
EJC533
EJC533
MT440
Toro1400
LM75
Twin Boom Jumbo
Twin Boom Jumbo
Single Boom Jumbo
30 t dump truck
30 t dump truck
30 t dump truck
Underground Loader
Diamond Drill Rig
Diamond drill rig
Integrated tool corner
924
The composition of the mining fleet required for extraction of the underground Mineral Reserve is
shown in Table 16.4. In the equipment listed, the following items are still required to be hired or
purchased





one LHD
two light vehicles - will be taken from existing LVs on site
one charge up vehicle
six air leg drills and
a wire line scraper.
Due to the short mine life based on the Measured and Indicated Mineral Resource, the purchase of
major equipment such as the LHD has not been included as part of the capital cost. It is expected
that these items will be obtained on a hire arrangement, with the cost included as part of the main
operating costs.
It is expected that parts of the fleet will have low utilisation due to the small tonnages being
produced from the underground mining operations, providing ample coverage for breakdowns. The
additional pieces will still be required, as they will provide the operation with the flexibility to
increase mining rates as required.
P a g e | 121
Table 16.4
Major mining fleet and equipment required for the extraction of the Dikulushi underground Mineral Reserves
Mining fleet
Number
2
3
2
1
1
1
1
2
6
1
LHD
Truck
Jumbo twin
Jumbo single
Integrated Tool Carrier
Charge up vehicle
PC
Ute
Air leg drills
Wire Line Scraper
16.6.1.
MINE DEVELOPMENT
The development of the underground operations will be completed using traditional drill and blast
methods utilising two twin-boom jumbos. A single-boom jumbo is to provide rock support and
rehabilitation capabilities, using weldmesh and rock bolts for the bulk of the mining development, as
per the geotechnical guidelines provided by Mike Turner. In addition to the standard ground
support outlined in the report, cable bolts will be installed at intersections and any areas of poor
ground encountered. Table 16.5 is a list of the horizontal development design parameters, including
the development location/type, width, height, back profile, ground support method and gradient.
Table 16.5
Underground horizontal development design parameters
Development type
Width
(m)
Height
(m)
Decline
Decline Stockpile
Level Access
Level Stockpile
Return Airway Drive
Cut and Fill Ramps
Ore Drives
5.5
5.5
5.0
5.0
5.0
4.5
4.5
5.5
5.5
5.0
5.0
5.0
5.0
5.0
Profile
Arched
Arched
Arched
Arched
Arched
Arched
Square
Ground
support
method
Mesh & Bolt
Mesh & Bolt
Mesh & Bolt
Mesh & Bolt
Mesh & Bolt
Mesh & Bolt
Mesh & Bolt
Gradient
1 in 7
1 in 50
1 in 100
1 in 100
1 in 50
1 in 6
1 in100
Vertical development to be completed in the underground will be conducted using hand-held rising
methods. Escape way rises will be developed in a single pass from bottom to top. Other vertical
development, such as vent rises, will be developed initially with a single rise of 1.5 m in diameter
developed from bottom to top. Stripping of the rise will then take place working from top to
bottom, with all blast material scraped into the centre of the rise where it will fall down to the lower
level for removal. Installation of split sets and mesh will be done as required in the larger
development openings. Table 16.6 is a list of the expected vertical development required in the
mining operation.
P a g e | 122
Table 16.6
Underground vertical development design parameters
Diameter
Development type
(m)
Length
(m)
Profile
Vent Rise
Escape Way Rise
15 to 20
15 to 20
Round
Round
16.6.2.
5.0
1.5
Ground
support
method
N/A
As Required
Angle
(deg)
60
60
MINING SCHEDULE
For the purposes of determining reserves for the Dikulushi underground mining operation several
schedules were created. Each of the schedules created was produced in Enhanced Production
Scheduler (EPS) using information from the reserve block model and underground designs provided
by MWL.
Initial scheduling of the underground operations involved creating a Measured, Indicated and
Inferred schedule that would be the basis of identifying the economic prospects of the underground
mining operation. This schedule was created to identify the probability of the mining operation
having a longer life than would have been indicated by just evaluating the Measured and Indicated
ore reserve only. This approach was taken due to the limited amount of Measured and Indicated ore
reserve remaining after the completion of the open pit mining activities.
The initial Measured Indicated and Inferred schedule was evaluated by MWL using an existing
financial model to identify the project’s potential viability and to understand the impact of the
operation within the company’s portfolio of mining operations. From the information output from
MWL’s financial model, the following costs were used as a benchmark for evaluating the Measured
and Indicated tonnes contained within the Initial schedule with



an operational mining cost of $100 per ore tonne mined,
a processing cost of $55 per ore tonne and
an operational overhead costs of $68.20 per tonne.
For the process of identifying the viability of the Measured and Indicated ore tonnes contained
within the underground design, a total mining and processing cost per ore tonne was then estimated
at $223.20. This cost was estimated for an underground operation with an ore mining and
processing production rate of 183,291t per annum. In addition to the ore mining cost, a copper
price of $6,800 per tonne and silver price of $20 per ounce were then used to determine the
revenue from the contained Measured and indicated copper tonnes and silver ounces.
The tonnes and grade for each of the levels were exported from EPS into a spreadsheet where a high
level economic evaluation was performed. For the economic evaluation of the Measured and
Indicated ore tonnes, each level was split into two sections - east and west of the level access.
The revenue for each section was calculated by taking the Measured and Indicated copper tonnes
and silver ounces contained in the mined material multiplied by the processing recoveries and
product prices. Each section was then evaluated by subtracting the mining cost from the section
from the revenue obtained from the sale of the measured and indicated copper tonnes and silver
P a g e | 123
ounces to determine if it had a positive revenue. Each level that had a positive revenue was
included as part of the Mineral Reserve schedule.
From the evaluation process nine levels were identified as having positive revenues. Seven of the
levels included both the drives to the East and West of the access, with two levels having
development east of the access only. A second EPS schedule was set up to provide a practical
mining schedule for the basis of the underground reserves.
A second mining schedule was created to provide a new mining sequence, incorporating the
development of capital mining development to provide




primary ventilation of the mining operation
access to the design ore drives from the existing underground workings
development of the ore drives
the use of backfill.
This schedule excluded the higher risk ore contained in the crown pillar zone between the 805 and
825 mRL’s. The Measured and Indicated mining sequence created in EPS mined 121,000 ore tonnes
at 5.4% copper over a period of 19 months. This timeframe allows for one month lead for
rehabilitation of the existing underground workings and development of rises to establish primary
ventilation along with one extra month at the end of the schedule to complete ore backfilling
operations. The mining costs used in the schedule were a combination of time variables such as
labour and administration costs which are independent of production activities, and direct
production costs which are driven by the day-to-day production of the mining and processing
operations. All costs used in the schedule were provided by MWL, with direct production costs
created from first principle methods and the overhead costs calculated using historic site costs.
The final reserves schedule included the extraction of the crown pillar tonnes from the 805 to the
825 mRL’s. The economic selection of the drives to be mined as part of the Mineral Reserve was
completed using the same process as the second schedule. In addition an economic evaluation
process for the ore drives was evaluated, based on their interaction with the existing Dikulushi open
pit. Where the underground development could have a significant impact on the stability of the
open pit shell, these ore tonnes were removed.
The Measured and Indicated resource mining sequence created in EPS mined 173,000 ore tonnes at
5.2% copper over a period of 20 months. The extraction of the additional tonnes in the crown pillar
area was scheduled in parallel with the ore tonnes produced from a second mining schedule being
created.
CONSTRAINTS
The Measured and Indicated Mineral Reserve mining schedule produced in this report has tried to
replicate the same parameters as outlined in the Measured Indicated and Inferred schedule
produced for MWL. Mining activities were restricted to levels that provided a positive economic
evaluation as described, with the mining sequence altered to minimise the mining timeframe
required for the extraction of the ore. All drives were reduced in length where the copper grade fell
P a g e | 124
below 1.5% towards the end of the drive. Ore drive sections where the grade fell below 1.5% but
then rose above 1.5% further down have been retained in the schedule, and were treated as
marginal ore, requiring that only the processing costs were covered by putting the ore tonnes
through the processing plant as a tonnes would need to be mined regardless of grade to access the
high-grade sections of the orebody.
To ensure that the full cost of mining for each ore drive has been included as part of the economic
evaluation, the Inferred and unclassified ore tonnes mined have not been included as part of the
cost calculation process. The grade component from the Inferred and unclassified tonnes was not
included as part of the revenue from the drive, diluting the grade from the Measured and Indicated
tonnes mined.
The production drilling rates that are used, are the maximum scheduled advance per month rate, as
Jumbo development is the critical limiting factor in the schedule. Twin-boom Jumbo lineal advance
rates have been scheduled at 193m per month per jumbo, or a total of 386 m per month for both
twin-boom jumbos. Waste development has been scheduled at 100 m per month. Development
rates for ore drives have been scheduled at the following rates:
 52 m per month in the western end of the orebody
 65 m per month in the central and eastern areas of the orebody
 91 m per month for capital development
Development rates used are outlined in greater detail in section 16.6.4.
DEVELOPMENT AND LEVEL DESIGN
The designs used in the schedule were in addition to work already completed by Red Rock
Engineering, which consisted of designs for capital development and an ore body wireframe shape
(stope6.dtm). Optiro has produced level designs in Datamine along the width of this stope shape
using 5m vertical intervals and drive dimensions of 4.5 mW x 5.0 mH. Levels have then been broken
into approximately 10m sections to allow more granular scheduling.
The level development design sections were then imported into Mine 5D Planner, where they were
evaluated against the block model and each section was assigned the corresponding average grade,
tonnes and material properties of the model cells contained within it. Additional properties were
also added such as development length and segment identification fields.
ORE LOSS AND DILUTION
For scheduling purposes, mining ore loss has been estimated at 5%. Ore loss will primarily occur as a
result of gaps left in the backfill process, caused by the inability to properly level off the top of
backfill as it is being placed in the drive. This means that when using the backfill as the floor for the
next level the fill will generally have a rounded profile, and this will result in some material being lost
to fill these corners as the next level is developed (see Figure 16.12). There will also be a small
amount of ore loss from general activities such as rehandle and material movement.
P a g e | 125
Figure 16.12
Ore loss due to gaps left in the backfilling process
Mining dilution for the project has been estimated at 8%, which is the average dilution calculated for
the various drive dimensions. Dilution was estimated by allowing a 200mm dilution skin on the walls
of development drives, which has then been used to determine the percentage dilution using the
cross-sectional area of the drives. Drive profile calculations are outlined in Table 16.7. These
estimations assume a square cross-section for calculation purposes.
Table 16.7
Mining dilution
Width
(m)
Height
(m)
Drive
Area (m2)
Dilution
width (m)
Dilution
skin area
(m2)
Dilution
(%)
Ore
Ore Drive
4.50
5.00
22.50
0.20
2.00
8.9
Waste
Decline
Decline Stockpile
Level Access
Level Stockpile
Return Airway Drive
Cut and Fill Ramp
Average
5.50
5.50
5.00
5.00
5.00
4.50
5.08
5.50
5.50
5.00
5.00
5.00
5.00
5.17
30.25
30.25
25.00
25.00
25.00
22.50
26.26
0.20
0.20
0.20
0.20
0.20
0.20
0.20
2.20
2.20
2.00
2.00
2.00
2.00
2.07
7.3
7.3
8.0
8.0
8.0
8.9
7.9
Development
Ore loss has been applied to the schedule by reducing the contained metal for each segment by 5%.
Dilution has been applied to the in situ tonnes of the blocks, multiplying them by 1.08 and then
subtracting the ore loss material. This only occurs in areas producing ore, and is limited to the
Measured and Indicated tonnes only. Updated metal grades (including both ore loss and dilution)
are then calculated from the ore-loss applied contained metal and the new diluted tonnes. For this
schedule ore loss has been assumed for both copper and silver. Waste tonnes have been calculated
for the purpose of financial evaluation of the schedule, and have been determined by simply taking
the total tonnes for a segment and subtracting the ore tonnes.
P a g e | 126
BACKFILL
Backfilling has been scheduled by calculating the mined volume of the drives and allowing for an
80% fill factor due to the difficulty of pushing material all the way up to the backs. Backfill has been
scheduled at a rate of 1,008 m3 per week per machine for all fill types, and minimum curing times
have been scheduled at 28 days for mining underneath fill and 2 days for mining alongside fill. A
density of 2.5 t/m3 has been assumed for backfill, which is conservatively high. Backfill sequencing
has allowed for 5 m pillars between active levels.
HAULAGE CALCULATIONS
The schedule has included production tonnes-km figures, which are a metric used to represent
haulage in terms of both the total tonnes and the total required haul distance for a given schedule
block. Surface haul distances were calculated by measuring the haul distance from the waste dumps
and ROM pad from the portal. Underground haul distances were calculated using differences in
vertical distance between the portal (1015 mRL) and the schedule level RL, and multiplying it by 7 to
estimate a decline distance. Total haul distances (in km) were then multiplied by the total tonnes
mined in each schedule block to calculate the haulage tonnes-km value.
SCHEDULE SEQUENCE
The Mineral Reserve schedule utilises the existing capital development completed in previous
underground mining activities. The mining activities that are required before a stoping operation
can re-commence are the development of ventilation and escape way rises in the upper levels of the
mine that were not previously completed. The development of the ventilation and escape way rise
system is to be completed within the first 4 months of the mining operation’s re-commencement.
This opens the opportunity to develop the maximum number of all drives from month 5 in the
schedule onwards.
Level development in the first 3 months of mining production is concentrated in the 800, 790, 765
levels. Towards the end of the quarter the 755 mRL opens up for mining activities with backfilling
operations. Backfilling operations are initially concentrated in the eastern section on the 790 mRL
for the first half of the quarter, with the 765, 770, and 800 RL backfilling activities commencing
towards the end.
Mining activities in the second quarter are scheduled to continue in the 790, 765 mRL’s. New mining
activity then commences on the 770, 780 and 750 mRL’s. Backfilling activities continue in the 790
and 800 mRL’s with backfilling activities starting in the 770 and 755 mRL’s. Mining activities in both
the 755 and 790 mRL’s is completed during this period.
Quarter three mining activities concentrate around the 770, 780, 795 mRL’s with backfilling activities
occurring on the same three levels. Mining of the 770 and 780 mRL’s is completed during this
period.
Quarter four mining activities are concentrated around the 795, 785 and 775 mRL’s and backfilling
activities are concentrated around the 780 and 795 levels, with backfilling activities also occurring in
the 785 and 775 levels towards the end of the period.
P a g e | 127
The remaining mining activities for quarters five and six are concentrated in the 785 and 775 mRL’s,
extracting the remainder of the current Measured and Indicated material. Due to the requirement
for backfill to minimise the open stoping widths, mining and production significantly reduces.
Backfilling activities during this period include the 785 and 755 mRL’s along with the 795 mRL.
Mining activities are completed within the first month of quarter six, with backfilling activities
finishing midway through the third month of the quarter.
Figure 16.13 is a pictorial representation of the Measured and Indicated Mineral Reserve extraction
sequence as described.
Figure 16.13
Ore level schedule, by quarter
Table 16.8 details of the planned mining production physicals on a quarterly basis.
Table 16.8
Underground mine production physicals
16.6.3.
MINING SHIFTS
Mining personnel shifts will be split up into three 8 hour panels; a day shift, afternoon shift and night
shift (outlined in Table 16.9). Day shift commences at 07H00 and shift changeover for the day shift
to afternoon shift will occur at 15H00. This is timed to coincide with the primary blasting time, and
afternoon shift start time will be dependent on re-entry periods following blasting. Shift changeover
P a g e | 128
from the afternoon to the nightshift will occur at 23H00. Work rosters will be a 2 day shift / 2
afternoon shift / 2 night shift / 2 off basis.
Table 16.9
Work shifts
Start time
Finish time
Duration
(hrs)
Day
07H00
15H00
8
Afternoon
15H00
23H00
8
Night
23H00
07H00
8
Shift
Project Management, Technical Services, and Support Personnel will generally work day shift only,
covering the day and part of the afternoon mining shifts. Table 16.10 to Table 16.12 shows the
estimated total number of personnel required for the site’s management, technical services and
operational support functions.
Table 16.10
Operational Management Labour
Operational Management
Project Manager
Maintenance Manager
Electrical Supervisor
Senior Auto Electrician
Maintenance Crew Leader/ Trainer
Op. Crew Leader / Trainer
Table 16.11
Quota
1
1
1
2
3
3
Technical Services labour
Technical Services
Technical Manager
Mine Survey
Mine Geology
Geotechnical Engineer
Mine Engineering
Quota
1
2
2
2
2
P a g e | 129
Table 16.12
Support functions labour
Labour
HSE Manager
Onsite Medical Doctor
Senior Environmental Officer
Onsite Medical Nurse
Environmental Technician
General Clerk
Ambulance Driver
General Driver
General Assistant
Quota
1
2
2
6
3
3
3
3
6
Table 16.13 and Table 16.14 detail the shift personnel numbers required for the operation of each
shift. The total numbers of personnel required to maintain the mining operations will be the
number of personnel per shift multiplied by the number of shifts. Where the manning number is
accompanied by an asterix, this means that this role is only filled during the day and afternoon shift,
with no personnel rostered into this position during the nightshift.
Table 16.13
Labour requirements: underground operations
Underground Operations
Mine Foreman
Shift Supervisor
Develop Drill Crew
Rehab Crew
General Service Crew
Material Haul
Material Load
Charge Crew
Air leg Drillers
Backfill Crew
Quota
1*
1
2
1
3
3
2
2
2*
2
P a g e | 130
Table 16.14
Underground Workshop personnel
Labour – Underground Workshop
Workshop Foreman
Crew Supervisor
Develop Drill Fitter
Workshop Fitter
UG Fitter
Trade Assistants
Senior Electrician
Mine Electrician
16.6.4.
Quota
1*
1
1
2
2
2
1
1
DEVELOPMENT / STOPING RATES
Development rates are outlined in Table 16.15, showing the estimated development turnaround,
length and maximum number of cuts available to be taken in each type of heading per week. All the
drive profiles can be found previously in Table 16.5 and Table 16.6.
Table 16.15
Jumbo/production drill rates by development type
Drive type
Cut Length
(m)
Turn Around
Maximum
cuts per week
Decline
Decline Stockpile
Level Access
Level Stockpile
Return Airway Drive
Cut and Fill Ramps
Ore Drives
3.3
3.3
3.3
3.3
3.3
3.3
3.3
1.5 shifts
1.5 shifts
1.5 shifts
1.5 shifts
1.5 shifts
1.5 shifts
1.5 shifts
7
7
7
7
7
5
5
Table 16.16 details the individual Jumbo production rates applied in the scheduling process; these
rates have been created by applying basic mining principles on a conservative production rate. The
twin-boom development rate (193.1 m per month) has been applied to the schedule, and the singleboom jumbo rates provided are intended as back-up rates only. Monthly development for 2 twinboom jumbos has been scheduled at a lineal rate of advance of 386.1 m per month, or 117 cuts
(shown in Table 16.17). The single-boom jumbo is assumed to be used primarily for cable bolting
and rehabilitation work.
P a g e | 131
Table 16.16
Jumbo/production drill rates by individual machine
Cut
length
(m)
Cuts
per
shift
No. of cuts
per week
Dist.
advanced per
week (m)
No. of cuts
per month
Dev. adv. per
month (m)
Twin-boom
3.3
0.64
13.5
44.6
59
193.1
Primary
Single-boom
3.3
0.31
6.5
21.5
42
139.4
Back up
Jumbo
advance
Table 16.17
Dev.
status
Jumbo/production drill rates by fleet
Units
Cut
length
(m)
Cuts
per
Shift
No. of
cuts per
wk.
Dist.
advanced
per wk.(m)
No. of
cuts per
month
Dev. adv.
per month
(m)
Twin-boom
2
3.3
1.29
27
89.1
117
386.1
Primary
Single-boom
1
3.3
0.31
6.5
21.5
42
139.4
Back up
Monthly
Development
16.6.5.
Development
Status
AIR LEG DEVELOPMENT RATES
Development work using air leg miners will occur on an eight hour day shift only, as the work
conducted by these workers is regarded as high risk. Table 16.18 outlines the production rates that
have been used in the schedule. Air leg/rise mining is a critical part of the mining plan, as it provides
the return airways for the primary ventilation circuit and the alternate secondary escape a path for
an emergency.
Development of the air leg rises for escape ways has been split up into production rates for single
and double rises. The development rate of rises varies with the air leg miners’ access to multiple
headings. For a single rise being developed by itself the advance is 1.5 m per shift, if there are two
rises close together this rate is able to be doubled due to the availability of equipment.
This scheme was applied for the development of the ventilation rises in the schedule and involves air
leg miners initially developing a rise from the lower level of the ventilation rise to the upper level.
Once the rises can be completed, the air leg miner then proceeds to strip the surrounding parameter
of the final rise diameter into the hole created by the initial rise. The process used by the air leg
miner involves drilling and firing the material to be stripped into the hole during one shift, with the
following shift required to clean out the fired material from the stripping, and installation of
appropriate ground support around the rise as it is developed down. It is expected that the
development of a 20 m rise will take one air leg miner approximately 42 days to complete the
development from start to finish.
P a g e | 132
Table 16.18
Air Leg development
Length
advance
(m)
Turn around
Advance per
week (m)
1.5
per shift
7
3
per shift
7
Drill and Fire
1.5
per shift
3.5
Bog and Ground support
1.5
per shift
3.5
Development type
Rise Development
Single
Double (side by side)
Return Air Way stripping
16.7. GEOTECHNICAL
16.7.1.
DATA
A Geotechnical study into re-opening the underground operation has been performed by Turner
Mining and Geotechnical (Turner, 2013). The study investigated geotechnical aspects of mining from
the pit bottom around 825mRL to 500mRL, and was based on the following data:






Diamond drill core logging databases (31 holes, Anvil and MWL).
Underground mapping data (7 sites, personal data, undertaken for Anvil).
Surpac files of current open pit and underground excavations (MWL).
Surpac files of planned open pit and underground excavations (MWL and Red Rock
Engineering Pty Ltd).
Previous geotechnical reports (Turner Mining and Geotechnical (2008(a), 2008(b), 2012 and
AMC Consultants (2004, 2011)).
Observations made during multiple site visits (2003 to December 2012).
16.7.2.
GEOTECHNICAL DOMAINS
Rock Quality classification has been conducted by Turner Mining and Geotechnical, with the
following geotechnical domains identified:



The orebody ranges from Extremely Poor to Fair, with the majority classed as Poor.
The footwall rockmass averages Fair.
Hangingwall ranges from Extremely Poor to Fair, with the majority classified as Poor.
The contours of Q for the orebody (Figure 16.14) show the very poor ground above the750mRL and
west of 50300mE. This poor ground zone in the orebody is critically important to manage as it
indicates severe ground control problems could be encountered when trying to extract ore out of
this zone
The contours of Q (rock quality) for the footwall (Figure 16.15) show 2 zones of very poor ground,
centred around the 725mRL at 50125mE; and 525mRL at 50280mE. These two zones indicate that
even though split sets and mesh might still be appropriate, there could be a need for shorter
cuts/round lengths and secondary installation of grouted bolts.
P a g e | 133
The contours of Q for the hanging wall (Figure 16.16) show a zone of very poor ground centred
around the 725mRL at 50130mE. This poor hanging wall zone eliminated the use of any longhole
stoping or benching in this zone.
Figure 16.14
Dikulushi orebody rock quality, Q (Turner, 2013)
Figure 16.15
Dikulushi footwall rock quality, Q (Turner, 2013)
P a g e | 134
Figure 16.16
Dikulushi hanging wall rock quality, Q (Turner, 2013)
16.7.3.
POTENTIAL FAILURES
From the previous mapping data gathered during mining activities in the 810 mRL in 2008 following
the fall of ground in 810 mRL W2 drive. Unwedge analysis was performed to evaluate the potential
for wedge instability in the walls and backs of excavations of the underground excavations. From
this analysis it was identified that the use of solid steel bolts compared to standard spit sets in
ground support activities provided an increase the factor of safety, reducing the likelihood of an
wedge failure occurring. Recommendations from the report by Turner ‘geotechnical input for reopening of the Dikulushi underground’ suggest the use of solid steel rock bolts to be used during
development of levels in the orebody where poor ground conditions have been identified (above the
755 mRL and West of the 50130 m E).
16.7.4.
MAPPING, MONITORING AND ADDITIONAL DATA
Mapping and monitoring of the underground drive development will form a significant part of the
ground support risk mitigation activities for the underground development, especially where ore
drives are located in areas of poor rock mass quality.
16.8. GROUND SUPPORT REQUIREMENTS
The orebody rock mass quality in the poor ground section from 810 to 780 mRL west of 50130 mE
ranges from Very Poor to Extremely Poor and the hanging wall is also Very Poor. Only small voids
will remain stable in this zone, and will require very intensive support, including solid steel rockbolts,
P a g e | 135
solid steel spiling bars and weldmesh or fibrecrete (Figure 16.17). Even with such support there will
still be a risk of collapse that will need to be managed by operations.
Figure 16.17
Dikulushi rock reinforcement chart (Turner, 2013)
Detailed pre-scheduling of drives is also essential to ensure sufficient working places are available,
taking into account the fill curing constraint. The strength and curing time need to correlate with
the scheduled extraction sequence to ensure sufficient ore is produced.
Testing of the different fill mixtures with varying cement percentages is another essential function
prior to the introduction of this method. The fill will need to have a guaranteed strength of at least
10 MPa before excavation can proceed under the fill. It is also common practice to increase the
cement content above that indicated from tests by up to 50% to cater for poor mix control.
P a g e | 136



Stope drifts should be maintained at no greater than 4.5 m width and split sets and mesh
used as support
quality control on cement percentage and mixing is essential
the extraction sequencing of adjacent drifts in wide areas will be critical.
2.4 m solid steel bolts should be inserted into 1.2 m holes 0.3 m above the floor at 1.2 m spacing
along each wall of the drive prior to fill being introduced in order to improve the tensile strength and
wall adhesion in the critical lower corners of the filled drifts. Mesh should also be left at the same
height across the width of the drifts if paste fill is used.
16.8.1.
SPLIT SETS
Most ore drives may be supported with split sets and mesh. Split sets (friction stabilisers) should be
galvanised, 2.4 m long, 46 mm diameter and of the type installed by jumbos. Hole diameter control
is an essential part of split set installation and bit sizes must be checked regularly to ensure that the
hole diameter is 44 to 45 mm. 0.9 m, SS39 stubby split sets may be used for mesh overlaps.
16.8.2.
SOLID STEEL ROCKBOLTS
Solid steel rockbolts that are immediately active (no delay for grout curing) will be required for any
development in the extremely poor ground in the orebody (above 775 mRL and West of 50130 mE).
Suitable solid steel bolts include:




20 mm Posimix bolts
20 mm CT-Bolts
20 mm Gemini Bolts (South African version of the CT-Bolt)
20 mm MD Bolt (combination split set and mechanical wedge).
These are either anchored using a mechanical shell/wedge or with resin, and can be used to install
mesh to the face. Plain solid steel bars are also suitable for use as spiling bars in extremely weak
ground to stabilise the backs ahead of the drive.
16.8.3.
CABLE BOLTS
All intersections should be cable bolted unless a geotechnical engineer is on site to map the
intersection and determine if it is stable, without a potential for wedge failure. Cable bolts should
consist of fully grouted, twin-strand 15.2 mm, plated and tensioned units on a 2.5 m spacing.
Historically cable bolts have never been installed correctly at Dikulushi and suitable equipment
should be purchased. Effective training and supervision will also be essential.
16.8.4.
SHOTCRETE
Fibrecrete may be used instead of mesh but the logistics of maintaining an operational shotcrete
fleet at Dikulushi would probably preclude this option. Shotcrete with fibres would be useful for
intersections of extremely weak ground, but these are only expected in the orebody, and if the
ground is that weak there will be subsequent stope stability issues.
P a g e | 137
16.8.5.
GEOTECHNICAL FILL REVIEW
Brief comments on fill strength and sizing have been recommended by Turner Mining and
Geotechnical Pty Ltd in the 2013 report with additional comments provided by Peter O’Brian and
Associates Pty Ltd.
The 4 main types of fill considered for the operations in both reports were;




rock fill (RF)
cemented rock fill (CRF)
cemented aggregate fill (CAF) and
cemented paste fill (Paste)
Rock fill was recommended for overhand cut and fill stoping .It was recommended that the use of
standard underground development waste was sufficient, with the main function of the rock fill
being to provide a working floor for the next mining levels/lift.
Cemented rock fill (CRF) was recommended to be use in open stoping, bench stoping and cut and fill
mining activities. The percentage of cement recommended was 10% and this is significant more
than stand a cut and fill operations, which generally use 5 to 6% cement. In addition to the use of
rock fill it was also recommended that material no greater than 300 mm be used.
CAF was also recommended to be used in open stoping, bench stoping, underhanded cut and fill or
standard cut and fill mining activities.
In addition to the uses of CRF and CAF a review of the tailings from the Dikulushi processing plant
was also conducted to identify there suitability for their uses in cemented paste fill. This review
identified that the tailings from the plant would be suitable for the creation of paste fill for the
underground mining operations, the tailings from the processing plant had an even size distribution
which will lead to an increase in strength and fast curing times.
The use of paste fill requires the construction of both a paste fill plant and underground delivery
system either through pipework located in the decline or a series of boreholes to deliver fill to each
of the levels. Additionally barricades need to be constructed at the end of the voids being filled with
paste fill to contain the fill wallet sets. Construction of barricades could be undertaken using waste
material and hand packed cemented fill bags sealing the opening to the backs of the drive. Due to
the limited height of ore drives/Stopes in the underground working it is unlikely there will be a need
for engineered barricades to contain the paste fill.
16.8.6.
CRF MIXING
Cemented Rock Fill (CRF) consists of waste rock mixed with cement. The source of the waste rock
should be unweathered and without an excess of large rocks or fine material. Underground
development waste is suitable but waste rock from the open pit waste dumps is not suitable due to
the much larger fragment size.
The method of mixing cement into the rock is critical and has a major impact on the cement content.
P a g e | 138
Mixing methods can include:




Spray mixing with cement slurry at the entry to the stope
Spray mixing at the tipping point into trucks.
Batch mixing at the entry for trucks
Batch mixing by loader in stockpile bays (reasonably common in small mines in Australia)
A cement percentage of 5 to 6% is normally required for the scale of the mining methods proposed
at Dikulushi, but where there is a risk of poor mixing and excess water content this should increase
to 10%. A 10% composition is typically used in small-scale filling where the fill is mixed by loaders in
stockpiles bays, calculated by adding the required number of cement bags to each loader bucket of
waste placed in the stockpile bay.
16.9. GROUND SUPPORT STANDARDS
Ground support standards have been developed for the Dikulushi underground mining operation by
Turner mining and geotechnical Pty Ltd. The standards include ground support for the declines,
access drives, ore drives and intersections and are shown in sections to 16.9.1 to 16.9.5
P a g e | 139
16.9.1.
DECLINE SUPPORT STANDARD
P a g e | 140
16.9.2.
ACCESS SUPPORT STANDARD
P a g e | 141
16.9.3.
ORE DRIVE SUPPORT STANDARD WITH MESH
P a g e | 142
16.9.4.
3 WAY INTERSECTION SUPPORT STANDARD
P a g e | 143
16.9.5.
4-WAY INTERSECTION SUPPORT STANDARD
P a g e | 144
16.10.
WASTE DUMP DESIGN
Waste material from the underground mining activities is expected to be minimal due to the use of
waste material as a major component of the backfill required for the ore mining activities. In
addition to the underground waste produced it is expected that suitable waste rock material will
need to be sourced from the base of the open pit and surrounding waste dumps to make up any
shortfall in waste material required for the production of backfill.
Should the underground mining activities change to mining methods which require significantly less
backfill material, it is expected that the existing waste dump surrounding the Dikulushi open pit can
be modified to accept additional waste material.
16.11.
SURFACE WATER MANAGEMENT
The existing surface water management arrangements for the Dikulushi open pit mining operations
are in conjunction with the underground dewatering system and it is expected to be capable of
dealing with any surface water inflow into the underground mining operations.
P a g e | 145
17. RECOVERY METHODS
17.1. PLANT FLOWSHEET
The processing plant and associated infrastructure was refurbished prior to start up in June, 2010,
and has been in continuous operation since.
The crushing plant consists of 3 stages; primary jaw crushing, followed by 2 stages of cone crushing
in a closed circuit with a double deck vibrating screen, producing a minus 20 mm product for the
grinding circuit feed. The grinding circuit consists of two overflow ball mills in parallel configuration
in closed circuit with a 250 mm hydrocyclone. Each ball mill is powered by a 750 kW motor. The
grind sizing parameter is 70% passing 106 microns. The mill is capable of treating in excess of
520,000 tonnes of ore per annum.
Both ball mills discharge to a common sump, and the slurry is pumped to a single 250 mm diameter
cyclone. The cyclone underflow gravitates to an Outokumpu SK240 Unit Flotation Cell to recover
coarse liberated copper sulphides, which report directly to the final concentrate. The cyclone
overflow reports to conditioning and conventional flotation at 35% solids.
A relatively simple flotation circuit is in place; the circuit consists of two sections, a primary sulphide
flotation and a secondary sulphide/oxide flotation. See figure 17.1.
Collector and frother addition is conventional when processing low grade ore. The splitting of the
circuit is due to the presence of oxide minerals in some of the ore blends which require activation
using sodium hydrosulphide (Na2S) to enable them to be recovered. As sodium hydrosulphide can
depress some sulphide minerals, the majority of the sulphide minerals are recovered in the primary
sulphide flotation circuit.
The tailings from the primary sulphide flotation circuit are sulphidised and the liberated oxides and
additional sulphides are recovered. In the event of the ore blend containing little or no oxides and
thus not requiring sulphidising, the secondary sulphide circuit acts as a sulphide scavenger. The
primary rougher circuit has provision for bypassing initial rougher concentrates directly to final
concentrate. Lower grade rougher concentrates report to the cleaning flotation cells for upgrading.
Final tailings from the secondary rougher circuit are pumped to the tailings storage facility.
Supernatant water is recovered from the tailings dam and recycled to the processing plant. The
circuit is based on a nominal flotation time of 20 minutes in each of the rougher flotation stages and
a minimum 15 minutes in each of the cleaner stages.
Final concentrate is pumped to a thickener and the underflow is pumped to a concentrate storage
tank. The storage tank has sufficient capacity for 8 hours of concentrate production. A filter press
with a capacity of 194t per day is operated in batch mode. Filter cake discharges directly onto a
concrete floor below the filter where it is recovered and transported to a simple hopper/bagging
arrangement with a skid steer loader. Concentrate is loaded into two tonne capacity bulk bags.
Moisture content is near 10%.
P a g e | 146
Each bag is weighed ready for despatch by truck to the Kilwa port as outlined in (Figure 17.1)
Figure 17.1
Dikulushi Plant flow diagram
17.2. TAILINGS STORAGE FACILITIES (TSF)
The first TSF for HMS tailings, designed by Knight Piésold Consulting, South Africa, covers 1.8
hectares and has been dormant since September, 2004. A particularly coarse portion of the HMS
tailings was recovered and processed through the flotation plant by Anvil. MWL has recovered and
processed approximately 15,000t of coarse sand and slime material from this dump
A second TSF (TD2), designed by D.E. Cooper and Associates, Australia, was built during 3Q, 2004 to
receive flotation tailings. This facility is located ~100m North of the HMS TSF, covers about 12
hectares and is 12 m high on the eastern embankment. This facility has also reached capacity.
A third TSF (TD3), designed by Knight Piésold, is located adjacent to and north of the second TSF and
covers a 21 hectare area. This dam is a typical hillside impoundment and provides the needed area
to limit the rise rate of tailings at acceptable norms.
Supernatant tailings water is reclaimed via penstock arrangements for use in the processing plant.
The third TSF (TD3) was utilized until December, 2008 and lay dormant until it was recommissioned
in July 2010. At this juncture Knight Piésold was employed to carry out a volumetric assessment
study to determine the storage capacity of the dam to accommodate 840,000t of tailings resulting
P a g e | 147
from the processing of the low grade ore stockpile. The study concluded that a 2 m embankment
raise would be required during 2011.
Deposition continued until October 2010 when Knight Piésold was commissioned to further assess
TD3 expansion capabilities to make provision for an additional 1,500,000t of deposition. The study
concluded that the walls would require to be raised by 6m to accommodate this quantity of tailings.
The raise will be carried out in 2 stages of 3m each, with the first raise completed in mid-2012.
Sub aerial tailings deposition to TD 3 has continued since completion of this first stage 3m raise and
as of June 2013, the TD 3 survey pick up indicated a residual storage volume capacity of 425,000 m3.
Based on operational beach densities and freeboard management allowance, the residual storage
capacity allowance is ensured until the TD 3 stage 2 raise, which is planned for during the dry season
in 2014.
In preparation for uninterrupted TD 3 works during this construction period, an interim storage
capacity lift of 2m is proposed in TD 2 before TD 3 works commence. This will also allow for ongoing
back-up operational capacity.
This will provide a tailings storage facility capable of supporting the underground mining operation.
The information on the upgrading of TD 3 has been supplied by MWL and Sedgman has not reviewed
this data.
17.3. PROCESSING STATISTICS
ANVIL PROCESSING
Anvil processed 137,256 tonnes of low grade between May 2008 and December 2008 when the
open cut run of mine ore ran out prior to full production from underground. Some production
results from the February 2007 to April 2008 can be seen in Table 17.1.
MWL PROCESSING
MWL blended material from surface stockpiles and the HMS Tails through the plant to maximise
copper output between June 2010 and February 2012. This was followed by the treatment of ore
from satellite orebodies such as Boom Gate. The recoveries from this activity were much lower than
from the fresh ore material from either the open pit or underground. It is reasonable to say that
process recoveries and values are associated more with those from the previous open pit and
underground mining operations carried out by Anvil. Processing statistics for the LG material
completed by MWL are shown in Table 17.2.
Commercial production from mining the Dikulushi open pit cutback commenced in November 2012
and was completed in July 2013. Processing of the cut back material is incomplete and open pit cut
back ROM feed material will be processed till the end of 2013. Underground ore processing of ore
will commence for approximately 18 months to Mineral Reserve Completion.
P a g e | 148
Table 17.1
Dikulushi processing summary relevant to ore to be mined in the pit cut back
Grade
Grade
RL mined (Ore Only)
Conc
copper%
Conc silver
g/t
86.9
860 pit stockpile
56.0
1730
91.5
87.5
860 pit stockpile
56.0
1745
Month
Blend
%
ROM
Plant
Feed
copper%
Silver
g/t
Copper
Rec%
Silver
Rec %
Feb-07*
60
5.93
181
85.7
Mar-07
100
8.41
273
Apr-07
100
7.65
233
90.7
92.2
850 pit stockpile
55.0
1696
May-07
100
7.61
231
90.2
90.1
850 pit stockpile
55.0
1670
Jun-07
100
7.74
233
91.0
90.5
870 Dev
55.0
1654
July 07*
91.4
7.28
214
89.5
89.8
stockpile
56.5
1668
Aug-07
100
7.92
245
91.1
89.4
850 Dev
56.0
1695
Sep-07
100
7.98
262
91.3
90.2
54.0
1890
Oct-07
100
8.18
272
92.4
92.5
55.0
1821
Nov-07
100
7.81
250
92.2
91.8
56.0
1793
Dec-07
100
8.45
266
92.8
92.0
57.0
1772
Jan-08
100
6.00
187
90.1
89.4
55.0
1694
Feb 08*
81.4
5.09
154
87.2
87.7
56.0
1687
Mar-08
100
5.45
188
88.1
79.2
830 Dev & 890/870
Stoping
54.0
1668
54.0
1601
Apr-08
90
4.76
139
87.0
88.4
830/810 Dev & 870
Stoping
* Low grade ore blended in with the development or stoping ore.
Table 17.2
Ore Processed
Mill Feed Grade
Mill Feed Grade
Tails Grade Cu
Tails Grade Ag
Conc Tonnes
Conc Grade Cu
Conc Grade Ag
Cu metal in Conc
Ag metal in Conc
Recovery Cu
Recovery Ag
Processing statistics for the LG material completed by MWL – June 2010 to May 2011
tonnes
Cu %
Ag g/t
Cu %
Ag g/t
dmt
Cu %
Ag g/t
dmt
oz
%
%
Jun10
5,387
1.28
35.87
0.34
10.1
128
38.7
1,067
51.45
4,384
74.62
70.57
Jul-10
36,157
1.45
40.4
0.39
11.1
896
43.5
1,138
389.6
32,778
74.31
69.88
Aug10
43,882
1.04
27.63
0.35
10.5
719
42.7
1,107
306.9
25,581
67.25
65.62
Sep10
40,839
1.27
31.72
0.46
10.9
783
43.0
1,119
336.7
28,177
64.92
67.65
Oct-10
27,450
3.78
77.17
1.64
23.0
1,380
44.1
1,139
608.5
50,534
58.64
74.20
Nov10
49,029
1.52
41.2
0.63
13.6
1,066
41.5
1,188
442.7
40,726
59.41
62.68
Dec10
41,111
1.17
28.5
0.52
10.70
684
39.74
1139
272
25,057
56.54
66.45
Jan-11
49,650
1.33
32.6
0.46
11.3
1001
40.1
1070
400
32,737
64.13
62.70
Feb11
42,839
1.32
29.2
0.43
7.95
890
41.6
1033
366
29,385
66.91
73.09
Mar11
46,054
1.28
27.8
0.46
7.85
893
39.35
941
351
27,279
62.66
63.34
Apr11
40,855
1.40
34.6
0.44
8.7
906
40.2
1092
365
31,904
67.46
71.25
May11
44,705
1.32
33.31
0.52
9.1
865
41.7
993
361
27,559
61.37
61.3
YTD
467,958
1.46
35.31
0.54
10.95
10,211
41.66
1089
4,251
356,101
64.05
66.68
P a g e | 149
Table 17.3
Processing statistics for the LG material completed by MWL – June 2011 to June 2012
Table 17.4
Processing statistics for the Open pit cut back ROM completed by MWL – Jul 2012 to July 2013
Jul-12
Aug-12
Sep-12
Oct-12
Nov-12
Dec-12
Jan-13
Feb-13
Mar-13
Apr-13
May-13
Jun-13
Jul-13
YTD-13
24,739
28,351
30,690
30,242
14,808
31,614
31,232
30,442
35,865
31,706
15,787
16,387
33,181
355,043
Ore Processed
tonnes
Cu %
2.65
2.28
2.33
2.04
5.09
5.41
5.48
6.32
5.46
6.50
10.27
10.88
7.02
5.21
Ag g/t
47.22
42.24
39.35
34.53
152.94
139.04
145.47
183.09
146.87
164.68
316.98
354.49
226.97
141.47
Tails Grade Cu
Cu %
0.65
0.59
0.55
0.45
0.42
0.44
0.40
0.43
0.36
0.44
0.51
0.61
0.46
0.48
Tails Grade Ag
Ag g/t
11.59
10.92
9.94
6.03
21.39
14.95
15.08
11.72
15.47
20.70
22.94
25.50
20.14
14.93
1,448
1,334
1,476
1,200
1,138
2,711
2,685
2,939
2,936
3,087
2,470
2,798
3,723
29,944
Cu %
34.86
36.51
37.57
40.49
61.16
58.42
59.42
61.40
62.29
62.70
62.85
60.79
58.97
56.52
Ag g/t
620
677
622
725
1,733
1,462
1,532
1,787
1,620
1,499
1,903
1,952
1,864
1,515
dmt
505
487
554
486
696
1,584
1,596
1,804
1,829
1,936
1,552
1,701
2,195
16,925
oz
28,876
29,014
29,496
27,950
63,410 127,429 132,235 168,837 152,974 148,824 151,060 175,621 223,058
1,458,783
Recovery Cu
%
76.9
75.2
77.4
78.7
92.4
92.6
93.3
93.8
93.4
93.9
95.8
95.4
94.2
91.5
Recovery Ag
%
76.9
75.4
76.0
83.2
87.1
90.2
90.5
94.2
90.3
88.7
93.9
94.0
92.1
90.3
As stated in Section 13.3 the production data for the Period June 2011 to July 2013 was supplied by
MWL and has not been reviewed by Sedgman.
P a g e | 150
18. PROJECT INFRASTRUCTURE
The Dikulushi operation is an operating mine and the infrastructure remains in place. It has been
used and maintained by MWL since it took over the project site. The infrastructure is considered
adequate for the continuation of the operations with the resumption of underground mining
activities.
18.1. SURFACE FACILITIES
The existing surface facilities (Figure 18.1) remaining from the previous underground operations and
open pit cut back operations will be suitable for use by the underground mining personnel (Figure
18.1 and Figure 18.2).
Figure 18.1
On-site office facilities at Dikulushi
Figure 18.2
On-site Underground change room facilities at Dikulushi
P a g e | 151
18.2. POWER
The project is located in a remote area where there is no electrical utility grid. The mine power is
supplied by diesel generators. Power for the Dikulushi operation will be provided by the existing
diesel powered electricity generation installation. This installation has previously supplied power to
the camp and the processing plant. Current production plans will not exceed previous levels and the
installed capacity is expected to be sufficient for future activities. There is sufficient back-up
capacity.
The existing power station at Dikulushi comprises the following generators: 4 x 1.2 MW FG Wilson (
new units and installed during the 3rd quarter 2013 ), 1 x 2.0 MW Caterpillar, 1 x 1.6 MW Caterpillar
and 1 x 0.8 MW Mirrlees, for a total capacity of 9.2 MW. The current power demand for the plant
and infrastructure is in the order of 1.8 MW. The 2.0 MW Caterpillar and 1.6 MW Caterpillar
generators currently require major overhauls, which will be completed during 2014. The 1 x 0.8 MW
Mirrlees will be decommissioned during the 4th quarter of 2013. The new FG Wilson generating sets
were installed to supply power to the operations as well as dewatering of the underground and
normal underground operations.
18.3. PROCESS WATER SUPPLY
Lake Newton on the perennial Dikulushi stream provides storage for dewatering and serves as a
reservoir for the supply of process water.
Meteorological data has been collected between August 2005 and August 2013, with the exception
of 2009 and 2010, where little or no data was collected due to reduced activities on site. The mine
water flow regime has changed over the past 3 years and the current water supply and balance
system is shown in Figure 18.3.
There are several sources of water on site:




The Dikulushi stream, which traverses adjacent to the mine, has two abstraction points. The
first abstraction point feeds water to the Process Plant, with the flow being measured. The
second abstraction point is at Lake Newton which stores the water before routing it to the
Return Water Dam (RWD) as make-up water. The flow between Lake Newton and the RWD
is measured.
The second source of water is from the Stream Borehole which supplies the Process Plant.
The borehole is not currently in operation.
The current main source of water is from the Open Pit which has a single supply pipeline to
the Process Plant which is metered.
Water from Tailings Dam 3 (TD3) is captured at the RWD where it is routed to the Process
Plant, This flow is also metered.
Other flow metered points are the Admin Building and the Power House which are internal plant
meters. The Truck Feed receives water from Lake Newton and is used for dust suppression around
the mine. A recent review and update was carried out on the full water balance during June and July
P a g e | 152
which are amongst the driest months in the year and this has been used to update the current
average water balance model, which is presented in Figure 18.3
Figure 18.3
Average water balance
Process Plant Water Sources
First option, must keep the water level down.
Last option.
Tailing
Dam
Lake
Newton
A 150m3/Day
Returned
Water Pond
1715m3/Day B
To the Plant
F
G
Main saurce for Process and Raw water.
3387m3/Day
The river pond feed also the Process only
when required.
E
2205m3/Day
Process Water
U/G
D 515m3/Day
DIKULUSHI
RIVER
A
B
C
D
E
F
G
H
I
J
KEY
Lake Newton
Return Dam Water
Boom Gate water
River
U/G Water 1
U/G Water 2
Process Water
Admin/gardens
Power House
Gland service/lube cooling water
Raw water
Flowmeter
32m3/Day
H
To admin/garden
117m3/Day
Pump
J
1956m3/Day
To Power House
32m3/Day I
C 764m3/Day
Emergency backup for Raw water.(Gland service)
Camp
Bore
Hole
Boom gate is the backup for Raw water and also Process water.
Can feed also the river pond and overflow to Lake Newton.
BG pit
Based on the water balance review the following were noted:








The total inflow onto the Tailings dam under present conditions is 3,607 m3/d of which
1,856 m3/d is returned back to the return water pond for reuse in the plant;
The RWD gets about 2,433 m3/d from Lake Newton via the river and 1,856 m3/d from the
tailings dam, while a total of 1,940 m3/d is sent to the plant for reuse;
The main losses from the Tailings dam are seepage, evaporation and interstitial storage;
The rainfall onto the open pit that is collected in the sumps below is reused in the plant and
only when water cannot be reused in the plant is the water discharged into Lake Newton
after the water is settled;
The extended waste footprint means that there will be runoff from the dump that will need
to be settled in paddocks and evaporated where possible;
Borehole water is currently not being used but will be used as potable water and make-up
water when it is in operation;
Approximately 1,589 m3/d will need to be supplied from Lake Newton or from boreholes to
sustain the mine during the dry months;
During the wet season there will be times where the water will discharge from the RWD into
the perennial stream as the plant will not be able to use all the water in the process.
P a g e | 153
19.
MARKET STUDIES AND CONTRACTS
19.1. MARKETS
The Dikulushi plant is currently configured to produce a copper/silver concentrate which contains
approximately 50% copper.
MWL has not yet committed any of the concentrate that will be produced from the underground.
This will be done via a tender process.
In July 2013, the Company entered into a copper hedging transaction, forward selling 3,500 tonnes
of copper at $6,875 per tonne for financial settlement at intervals commencing in November 2013
and ending in March 2014.
19.2. CONTRACTS
MWL currently has a contract to sell the copper concentrate produced from the open pit project to
Transamine Trading. This contract was awarded after a tender process.
There are various contracts either already in place or required to be entered into for the following
major areas:





Explosives
Diesel supply
Transport
Reagents
Spares
MWL recognises that a consistent reliable fuel supply is crucial to the success of the Dikulushi
operation. The operation currently uses approximately 500,000 litres of diesel per month and will
need additional fuel for the underground operations. This fuel is supplied by four DRC based
companies; two receive supplies from the port of Beira in Mozambique and the other two receive
supplies from the port of Dar Es Salaam in Tanzania. MWL had no interruptions during the open pit
project when it was receiving up to 1,200,000 litres per month. Thus MWL believes that it has
mitigated the risk of fuel supply by having a number of suppliers whom source fuel from different
ports.
MWL has a number of supply contracts for various inputs required for operations. MWL also has
contracts to transport of concentrate from Dikulushi to Kilwa.
The current revenue estimates include the concentrate being sold to the Concentrate trader who
will on sell the concentrate to smelters, where it will be converted to metal and sold to the market.
MWL receives 90% provisional payment for material delivered to Nchelenge in Zambia. A further
10% is received after finalisation of QP pricing and assay exchange.
P a g e | 154
Costs for the transport, as well as the treatment and refining charges for the copper and silver
concentrates, along with the final net smelter returns have been sighted and used in calculations for
the Mineral Reserves, but due to the commercial in confidence agreements with the smelters are
not shown here.
The study has used a copper price of $6,800/tonne copper ($3.08/lb. copper) and a silver price of
$20 /oz silver.
No formal off-take agreements have been confirmed to support these assumptions, but the
expected revenue parameters are based on assessments completed by Mawson West of likely
conditions and forward price curves
The average cost per tonne of copper product for transport, treatment, refining and clearing is
estimated to be $1,153 per tonne of copper metal sold.
Long term commodity price projections have not been evaluated due to the short life of the
underground and processing operations being less than 24 months, at this stage.
P a g e | 155
20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL
OR COMMUNITY IMPACT
An Environmental Impact Assessment (EIA) for the Dikulushi project was lodged in 2003. In 2009,
an EIA for the underground Project was submitted to the DRC Government. Both of these reports
were compiled by African Mining Consultants of Kitwe, Zambia, an environmental company that was
licensed to work and report in the DRC. In 2011 an EIA for the cutback project was prepared by EMIS
sprl, a DRC environmental company licences to work and report in the DRC. All three environmental
reports received DRC Government approval. A revised EIA, extending underground mining beyond
2013, has been submitted to the Government for approval.
MWL is required to provide annual environmental reports and demonstrate that it is in compliance
with the EIA. Mine remediation is one of the compliance items in the EMP.
MWL has lodged an environmental bond of $1.19M. The financial guarantee is a contribution
towards an estimate of the total costs of closure, rehabilitation and re-vegetation of the Dikulushi
mine. The development of the financial guarantee is conducted in compliance with:



Articles 410 of the Mining Regulations
Articles 124 and 125 of Appendix XI of the DRC Mining Regulations 2003; and
Appendix II of the Mining Regulations 2003.
The company recently had completed an annual review of the EIA which has been lodged. This
review did not find any non-compliant items, or any breaches of the permitted conditions and
requirements.
MWL has number of corporate social responsibility programs that are run on the Dikulushi Property.
The key programs are –
A) The Dikulushi-Kapulo foundation – a community foundation to initiate, develop and support
development projects for the benefit of local communities in health, education,
infrastructure and reinforcement capabilities. The foundation acts as a catalyst to support
community initiatives and development projects.
B) Employment and training – MWL employs approximately 900 local employees. MWL has
introduced various training programs that are also available to the local community. A
teacher has been employed to assist women with language education in French, English and
Swahili.
C) Community Health – MWL has joined with the Australian Government’s DAP program and
the program has contributed funds to improve the available health care facilities present in
the Dikulushi community clinic.
D) Education – MWL has contributed funds towards the upgrade of the Dikulushi School, to
enclose classrooms and provide classroom equipment and resources. This project benefits
1,100 local school students.
E) Kilwa Electrification – this project is working with the Kilwa community to provide power to
the hospital and surrounding buildings in the village.
P a g e | 156
F) Kipeto Community garden project – This project is to assist the community to establish a
vegetable nursery in the community.
G) MWL is committed to supporting local business by sourcing certain supplies from local
villages surrounding Dikulushi and the Kapulo projects. Figure 20.1 is an example of this
commitment.
Figure 20.1
Community Business making work clothes for the mine.
P a g e | 157
21. CAPITAL AND OPERATING COSTS
21.1. CAPITAL COST ESTIMATE
Capital expenditure for the underground mining operations is estimated at around US$9M. The bulk
of the Capital expenditure for the project is focussed on the mining equipment related to the
underground operations. All other capital for the processing plant, infrastructure and administration
is already in place and ongoing from the Open pit cut back operations.
The list of items included in capital expenditure cost is shown in Table 21.1. Due to the short time
period that the Mineral Reserves will be mined over, it has been planned that any additional major
mining equipment required will be obtained on a hire arrangement, with costs covered as part of the
annual operating cost. This has provided a significant saving in the capital spend for the reestablishment of the underground operations. Additional capital savings have been achieved
through the refurbishment of the previous underground mining fleet and this is to provide the bulk
of the mining fleet required to mine the underground Mineral Reserves.
Table 21.1
Dikulushi underground capital expenditure cost estimate.
Item
Refurbishment of mobile equipment
2 x UG Loader & 2 x Truck (new)
UG dewatering system & bores
Vehicles & ancillary units
Ventilation fans
UG compressors
Safety equipment and systems
Ancillary support equipment
Total
Expenditure US$
000’s
1,109
3,362
2,570
560
400
150
590
865
9,606
The total composition of the mining fleet is shown in Table 21.2. The majority of this fleet is already
located on-site and was used as part of the previous underground mining activities conducted by
Anvil.
It is expected that parts of the fleet will have low utilisations due to the small tonnages being
produced from the underground mining operations, and this will provide ample coverage for
breakdowns. The additional pieces of equipment required will provide the operation with the
flexibility to increase mining rates as required.
P a g e | 158
Table 21.2
Major mining fleet and equipment required for the extraction of the Dikulushi underground Mineral Reserves
Mining fleet
LHD
Truck
Jumbo twin
Jumbo single
IT
Charge up vehicle
PC
Ute
Air leg drills
Wire Line Scraper
Number
2
3
2
1
1
1
1
2
6
1
The majority of the above capital is spent over the first 6 months of the project.
21.2. OPERATING COST ESTIMATE
The underground mining operations at the Dikulushi project will be run under an owner operator
model. MWL will provide all the equipment and personnel to complete the mining operations. For
the purpose of estimating the operating costs, the costs have been broken up into mining variable
costs and overhead/fixed costs. This approach has been chosen so that the costs that are directly
involved with development and ore production from the underground, will vary with the advance of
the physical mining operation and are separated from other costs such as labour, management and
supervision that are generally fixed operating costs on a month by month basis.
21.2.1.
MINING OPERATING COST
Table 21.3 shows the fixed mining and processing costs associated with the Dikulushi underground
operation. The costs included in this table were provided by MWL and are based upon cost
estimates in financial modelling already completed to evaluate underground mining at the site. G&A
costs are included in the Site Costs item, along with camp costs, OH&S, site wide power and water
supply for the camp.
Variable costs are listed in Table 21.4, including development, haulage backfill and rehabilitation.
Capital development, vertical development and production development costs are all inclusive of
jumbo drilling, blasting, loading, ground support and services, including all the equipment operating
costs involved. Costs have been calculated on a lineal metre basis.




Decline costs have been calculated assuming drive dimensions of 5.5 mH by 5.5 mW.
Level access and return airway drives have been calculated using dimensions of 5.0 mH by
5.0 mW.
Production access and ore drives assume dimensions of 5.0 mH by 4.5 mW.
Vertical development rises assume dimensions of 1.5 mH by 1.5 mW for escape ways and
5m diameter for vent rises.
P a g e | 159

Haulage costs are inclusive of loading, hauling to the waste dump or mill, and equipment
operating costs.
Table 21.3
Mining overhead and fixed costs
Mining overhead and
fixed costs
Units
Rate
(million)
Mining Personnel
$/yr
$0.75
Workshop Personnel
$/yr
$0.33
Technical Support Roles
$/yr
$2.60
Fixed Plant Costs
$/yr
$0.66
Power Consumption
$/yr
$1.76
Table 21.4
Engineering, Geology, survey etc.
Capital Development
Decline
Level access / Stock piles
Return Air Way Drives
Vertical Development
Air leg Rises Escape way
Vent Rise
Production Development
Cut and Fill Ramp Access
Ore Drives / Cut and fill
Haulage
Ore
Waste
Backfill
CRF
CAF
RF
Rehabilitation
Drive rehab
21.2.2.
Comments
Units
Rate
$/m
$/m
$/m
1,739
1,677
1,739
$/m
$/m
83
568
$/m
$/m
1,330
1,330
5.0mH by 4.5mW
5.0mH by 4.5mW
$/tkm
$/tkm
2.21
3.19
Loading and haulage to waste dump or mill –
incl. equipment operating costs
$/m
3
$/m
3
$/m
44.57
55.70
3.23
-
$/m
814
3
Comments
5.5mH by 5.5mW
5.0mH by 5.0mW
5.0mH by 5.5mW
1.5mH by 1.5mW - including equipping
5.0m diameter, air leg pilot + airleg strip
Assumes all ground support is replaced
PROCESSING OPERATING COSTS
A $55.00 per ore tonne processing cost has been applied to all tonnes processed from the
underground mining operations. This was provided by Mawson West and represents the overhead
and variable processing costs associated with the operation of the processing plant for the
production of the copper concentrate.
21.2.3.
MANAGEMENT AND ADMINISTRATION COSTS
The management and administration costs are based on the existing cost structure of the current
operations and includes management and administration personnel, OH&S costs, logistics costs,
camp costs and sustaining capital required for non-production related activities. The total annual
cost is estimated at $10.6M per year.
P a g e | 160
21.2.4.
TRANSPORT AND SMELTING COSTS
Costs for the transport, as well as the treatment and refining charges for the copper and silver
concentrates, along with the final net smelter returns have been sighted and used in calculations for
the Mineral Reserves, but due to the commercial in confidence agreements with the smelters are
not shown here.
21.3. METAL PRICES
Financial modelling has used a copper price of US$6,800/t and a silver price of US$20/oz. Details are
provided in Table 21.5.
Table 21.5
Metal prices used in modelling
Product
Copper
Sensitivity Low
Sensitivity High
Silver
Sensitivity Low
Sensitivity High
Units
$/t
$/t
$/t
$/oz
$/oz
$/oz
Rate
6,800
6,120
7,480
20
18
22
22. ECONOMIC ANALYSIS
22.1. OPERATIONS SUMMARY
Table 22.1 below provides a summary of the mining and cashflow performance of the Project mine
including the extraction of the Crown pillar ore. The total material mined from the underground
operations is 221 kt of material of which 173 kt is ore at a diluted and recovered grade of 5.15%
copper.
The average mining cost $83 per total tonne and $106 per ore tonne. The processing cost is $55 per
ore of tonne processed. Management and administration costs are $97 per tonne of ore processed.
Due to the long history of mining operations at the Dikulushi project, capital expenditure is minimal
as the site already has the processing and major infrastructure in place from previous open pit and
underground mining activities. In addition to the facilities the majority of the mining equipment
required for the underground operations exists on site. The total capital costs allowed for as part of
the underground mining project is $9M. Sustaining capital for the operations has been included in
the above capital cost.
The mine life is 19 months with the majority of the ore mined in the first 12 months of the mining
operation, and ore production reduces towards the end of the mine life, as the number of
production ore headings available reduces. Processing, copper production and sales have been
scheduled in the month of extraction for the purpose of this economic evaluation.
The processing recovery for copper used for this estimate was 94%, with silver recovery at 90% to
produce a copper concentrate grading approximately 60% copper.
P a g e | 161
The total cash cost of the operation is $44.6M. A total of 8.1 Kt of copper is sold along with 573 Koz
of silver to produce net revenue of $57.7M, and net cashflow of $3.0M. Due to the short life of the
existing underground workings no discount rate has been applied to the revenues in Table 22.1
P a g e | 162
Table 22.1
Dikulushi mining and financial summary
Pre-mining
Qtr 1
Qtr 2
Qtr 3
Qtr 4
Qtr 5
Qtr 6
Total
Physical Schedule
Total material mined
Waste Mined
tonnes
tonnes
3,564
3,564
50,672
17,346
30,517
4,271
25,044
6,084
38,069
5,576
38,713
11,044
34,548
0
221,128
47,885
Ore Mined
Copper mined grade
Silver mined grade
tonnes
copper%
silver g/t
0
0.00
0.0
33,327
5.08
142.7
26,246
4.97
102.9
18,961
6.16
126.9
32,493
5.78
151.1
27,669
4.38
94.3
34,548
4.83
133.7
173,243
5.15
127.0
Copper mined
Silver mined
t
oz
0
0
1,693
152,913
1,304
86,871
1,168
77,338
1,877
157,804
1,212
83,929
1,667
148,518
8,921
707,373
$
$
$
$
$
$
$
$
$
$
134,790
42,538
3,811
0
0
24,091
0
88,110
55,414
146,792
0
1,186,964
28,678
0
140,714
119,642
307,128
264,331
166,243
440,376
0
829,203
0
0
114,775
31,200
994,495
264,331
166,243
440,376
0
349,937
0
0
40,237
37,816
306,330
264,331
166,243
440,376
0
702,380
0
0
94,240
37,977
586,851
264,331
166,243
440,376
0
923,050
0
0
115,531
68,728
472,853
264,331
166,243
440,376
0
324,052
0
427,327
59,772
0
172,134
264,331
166,243
440,376
134,790
4,358,124
32,488
427,327
565,269
319,454
2,839,791
1,674,093
1,052,874
2,789,046
Technical Support $
Total Mining Opex $
216,313
711,859
648,938
3,303,014
648,938
3,489,560
648,938
2,254,206
648,938
2,941,334
648,938
3,100,050
648,938
2,503,172
4,109,938
18,303,195
Costs
Mining
Decline Rehabilitation
Jumbo development
Airleg Development
Longhole Stoping
Ore Haulage
Waste Haulage
Backfill
Mining & Workshop Labour
Fixed Plant
Power
P a g e | 163
Total Mining Opex $/ore tonne mined
$/total tonne
mined
Pre-mining
Qtr 1
Qtr 2
Qtr 3
Qtr 4
Qtr 5
Qtr 6
Total
0.00
199.72
99.11
65.18
132.95
114.35
118.89
90.01
90.52
77.26
112.04
80.08
72.45
72.45
105.65
82.77
0
0.00
1,832,958
55.00
1,443,546
55.00
1,042,835
55.00
1,787,108
55.00
1,521,768
55.00
1,900,139
55.00
9,528,355
55.0
883,333
0.00
2,650,000
79.52
2,650,000
100.97
2,650,000
139.76
2,650,000
81.56
2,650,000
95.78
2,650,000
76.70
16,783,333
96.88
7,378,442
7,271,818
7,053,311
44,614,883
Processing
Processing $
$/ore tonne milled
Management & Admin
Administration $
$/ore tonne milled
Total Operating Costs
$
1,595,192
7,785,972
7,583,106
5,947,041
Total Capital Costs
Sustaining Capital
$
$
4,990,000
26,316
3,921,000
78,947
495,000
78,947
200,000
78,947
78,947
78,947
78,947
9,597,000
500,000
Metal in concentrate copper t
silver oz
0
0
1,592
140,680
1,225
79,922
1,097
71,151
1,765
145,180
1,139
77,215
1,567
136,636
8,385
650,783
Metal sold copper t
silver oz
0
0
1,537
124,309
1,184
70,157
1,060
62,448
1,705
128,109
1,101
67,845
1,514
120,705
8,100
573,573
Sales & Transport Costs $
0
1,394,211
1,059,225
948,462
1,540,389
986,344
1,371,896
7,300,528
Duties and Taxes $
0
297,711
223,249
199,859
327,811
208,192
292,751
1,549,574
Revenue
P a g e | 164
Copper NSR $
$/t mined
Pre-mining
0
0
Qtr 1
10,454,484
206.31
Qtr 2
8,049,030
263.75
Qtr 3
7,208,970
287.85
Qtr 4
11,591,067
304.48
Qtr 5
7,484,181
193.33
Qtr 6
10,294,229
297.97
Total
55,081,959
0
0
2,486,180
49.06
1,403,131
45.98
1,248,965
49.87
2,562,176
67.30
1,356,902
35.05
2,414,105
69.88
11,471,459
0
0
0
11,546,452
297,711
11,248,741
8,392,935
223,249
8,169,686
7,509,473
199,859
7,309,614
12,612,854
327,811
12,285,043
7,854,738
208,192
7,646,546
11,336,438
292,751
11,043,687
59,252,891
1,549,574
57,703,317
Silver NSR $
$/oz mined
Total Revenue
Revenue from Sales $
Royalties/taxes $
Net Revenue $
Cashflow from Operations
NPV
$
-6,611,508
NPV
-537,178
$2,982,434
12,632
8%
1,083,626
4,827,653
295,781
3,911,429
2,982,434
IRR
P a g e | 165
22.1.1.
SENSITIVITY ANALYSIS
MWL has carried out a sensitivity analysis on the cash flow forecasts, and this is provided in Table
22.2.
Table 22.2
Sensitivity analysis on the cash flow forecast for underground mining and treatment at Dikulushi
Table
Dikulushi Copper Project
Project Sensitivity to a Change in copper Price
NPV (US$ million)
Change in copper Price
-10%
-5%
0%
5%
10%
-9.8
0.7
3.0
5.7
8.4
Table
Project Sensitivity to a Change in Silver Price
NPV (US$ million)
Change in Silver Price
-10%
-5%
0%
5%
10%
1.8
2.4
3.0
3.6
4.1
Table
Project Sensitivity to a Change in Operating Costs
NPV (US$ million)
Change in Operating Costs
-10%
-5%
0%
5%
10%
7.4
5.2
3.0
0.8
-1.5
22.2. PAYBACK
As discussed the refurbishment cost of the mill has already been covered by the revenues from the
LG stockpile treatment and the previous open pit mining activities, thus there is no formal capital
payback period. The development of the underground mining is to be fully funded out of MWL’s
current existing cash reserves. The maximum negative cashflow (including capital costs) is -$1.7M
(end of the pre-mining period) and cash flow moves back into positive territory during the third
quarter of operation.
P a g e | 166
22.3. MINE LIFE
Mine life is based on the Probable mining Mineral Reserve schedule and is approximately 19 months.
This allows a one month lead for rehabilitation of the existing underground workings, and the
development of ventilation rises to establish primary ventilation along with one extra month at the
end of the schedule, to complete ore backfilling operations.
22.4. TAXATION
The Dikulushi mine operates under the Dikulushi Mining Convention, which provides for
concessionary rates of taxation for each new mine. The first five years of production were tax free,
the effective tax rate from the sixth through tenth years of production is 16% and for the eleventh
through fifteenth years of production 18%, thereafter 40%. Dikulushi has been producing for
approximately eleven years.
In addition to the usual deductions of expenses and accruals, the Dikulushi Mining Convention
provides that taxable income is adjusted by allowances for:



depreciation of moveable and immoveable fixed assets,
a “depletion allowance” equal to 15% of gross sales up to 50% of net profit, and
all exploration and evaluation expenses.
AMC also receives the benefit of concessionary import duty rates. During the construction phase,
2% import duties are applied and then during production import duties are applied at the rate of 3%
for fuel, lubricants and mining consumables and 5% of all other supplies.
P a g e | 167
23. ADJACENT PROPERTIES
There are no significant mining properties adjacent to the Dikulushi Property.
MWL’s Kapulo copper exploration and development project is also part of the Dikulushi Mining
Convention that includes the Dikulushi Property. However, MWL considers the projects to be
separate and non-contiguous. The Dikulushi Mining Convention applies to an area of approximately
7,300 km2 and the two projects are located 124 kilometres apart, are on distinctly separate leases,
and are separated across this distance by Lake Mweru and the Luapula River. Road access to the
Dikulushi Mine is from the DRC side of Lake Mweru, while road access to the Kapulo Project is from
the Zambian side of Lake Mweru. There is only rudimentary road access between the two projects.
Development of the Kapulo Project is not dependent on, and will not share infrastructure with, the
Dikulushi Mine. Each of the projects will have their own separate mills, facilities, equipment and
administration, and will conduct independent processing operations. A definitive feasibility study on
the development of the Kapulo Project is the subject of a NI 43-101 technical report dated June 30,
2011 entitled “Kapulo Copper Project, DRC, National Instrument 43-101 Technical Report”.
P a g e | 168
24. OTHER RELEVANT DATA AND INFORMATION
Historically, Dikulushi was a producing open pit operation from 2002 until 2006. It continued for a
period of time supplying ore from underground operations until closure in November 2008.
The Dikulushi mine was acquired from Anvil by Mawson West Limited in April 2010 and work started
immediately on refurbishment of the plant, which was completed in June 2010. Since June 2010,
MWL has produced copper-silver concentrate from a feed of blended HMS tails and reclamation of
the LG stockpile, as well as the successful mining and processing of the Open pit cut back.
The next stage of the operations is to now re-establish the underground workings and re-commence
underground production as outlined in this report. It is also MWL’s intention to continue
exploration drilling of the Dikulushi orebody from accessible locations within the underground.
P a g e | 169
25. INTERPRETATION AND CONCLUSIONS
The Dikulushi Property is a producing and developing property. Current processing of the Dikulushi
LG stockpile reserves and open pit cut back ore has provided MWL with a robust cash flow, and
production results demonstrate reliable grades of remaining ore stocks when compared with
Mineral Reserve estimates.
The Dikulushi deposit has a history of exploration and successful mining. Data quality across the
unmined volume of the deposit is of good quality and has representative sample values for reliable
Mineral Resource estimates. Mineral Resource classification supports both Proven and Probable
Reserve categories within the underground Project. The pre-feasibility study and resulting Mineral
Reserves from the underground further extends MWL’s production life from the Dikulushi Project.
MWL has an opportunity, with the re-establishment of the underground, to actively pursue
exploration drilling from selected locations in the underground to be able to upgrade Inferred
Mineral Resource material and extend the total Mineral Resource. MWL intends to continue
processing the open pit cut back ROM material during the re-establishment and build up phase to
production from the underground.
MWL’s strategy is to continue to develop satellite deposits around Dikulushi, such as Kazumbula, in
addition to extending the remaining Dikulushi Mineral Resources located below the underground.
The recent exploration drilling at Kazumbula has provided geological and sample information to
support a robust Mineral Resource estimate. Upon completion of the mine design, scheduling and
financial analysis, the Kazumbula deposit is most likely to be of reasonable size and grade to be able
to contribute feed to the Dikulushi plant. Additional satellite deposits within 50 km of Dikulushi are
currently being drilled by MWL.
P a g e | 170
26. RECOMMENDATIONS
It is recommended that MWL continues with the planned underground Project. The key aspects for
the success of the underground Project are the establishment and maintenance of sound
underground mining practices, including a key focus on the drilling and blasting operations and the
post blasting ground support regimes. The backfilling operations and the selected fill support
method for the ground conditions, will require on-going attention and review, as well as close
attention to the fill specifications as recommended. Ongoing test and study work is recommended
on the investigation of using the tails dam material as underground paste fill. The initial prefeasibility tests showed positive results.
With the recommencement of undergrounds operations, it is recommended that exploration drilling
be continued from selected underground positions to test the orebody at depth and to assist with
the re-classification of the remaining Mineral Resource up to the Indicated Category and thus
Probable Reserves.
Ongoing annual reviews are required for the environmental approvals and permitting and a key
component is to ensure the integrity of the tailings dam continues to be maintained.
The development of additional targets within the 50 km radius of Dikulushi has good synergies with
the overall MWL strategy.
P a g e | 171
27. REFERENCES
DevMin Pty Ltd (Feb 2004): Anvil Mining Ltd “Dikulushi Copper-Silver Deposit, NI34-101 Technical
Report. February 16, 2004.
Franey, N., Hillbeck, M. and Fahey, G. (2006): Technical Report, Dikulushi Copper – Silver Deposit.
February 21, 2006
JORC (2004): Australasian Code for Reporting of Mineral Resources and Ore Reserves, Effective
December 2004. Prepared by the Joint Ore Reserves Committee of The Australasian Institute of
Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia (JORC).
National Instrument 43-101, Standards of Disclosure for Mineral Projects, Supplement to the OSC
Bulletin, April 8, 2011
Form 43-101F1 Technical Report, Supplement to the OSC Bulletin, April 8, 2011
Munro, K.D. & Associates (1998): Dikulushi Copper-Silver Project. Geological Review and Mineral
Resource Estimate for Dikulushi Copper-Silver Project.
Lemmon, T., Boutwood, A., Turner, B., (2003) The Dikulushi copper-silver deposit, Katanga, DRC. In,
Proterozoic Sediment-hosted base metal deposits of Western Gondwana, ed., J. Cailteux, Abstract of
the IGCP 450 conference and field workshop, July 14-24. Lubumbashi, DRC.
Dewaele, S., Muches, P., Heijlen, W., Lemmon, T., Boutwood, A., (in press), Reconstruction of the
hydrothermal history of the CU-Ag vein-type mineralisation of Dikulushi, Kundelunga foreland,
Katanga, DRC.
Fahey,G.,Franey,N., Anvil Mining Limited Dikulushi Copper-Silver Mine Katanga Region Democratic
Republic of Congo technical Report (NI43-101), December 22nd, 2006
Mawson West Ltd Pre-Feasibility study, July 2011
Independent Metallurgical Laboratories (IML): Metallurgical Ore Characterisation of Dikulushi
Copper Ores for Anvil Mining NL, August 2003
Independent Metallurgical Laboratories (IML): Confirmatory Metallurgical Testwork on ROM
Dikulushi Copper Ore for Anvil Mining NL, June 2004
Metallurgical Design and Management Pty Ltd; Dikulushi Copper Silver Project, Stage 2 Flotation
Project Interim Metallurgical Rreport, July 11, 2003
F Chikosha, Dikulushi Copper Mine Tailings Disposal Facility TD3 Expansion Study, June 2011
A J Strauss, Dikulushi Copper Mine Tailings TD3 Volumetric Assessment, July 2010
M.Turner, Indpendent geotechnical consultant: Dikulushi north wall cable bolts 270711, July 2011
M.Turner, Indpendent geotechnical consultant: MHTurner Project stability 260711, July 2011.
SRK Consulting: Project No: 436159 Water Balance for Dikulushi Mine – 2011 Update
P a g e | 172
Mawson West, Dikulushi Underground Mining Study, June 2013
Peter Wade, Capital Mine Consulting; Dikulushi Mining Operating Cost Review, June 2013
M.Turner, Dikulushi Underground Re-opening Geotechnical Study Report No. 0713, August 2013.
J.Keogh, Dikulushi Underground Underhanded Cut & Fill Cemented Rockfill Preliminary Design
Technical Note 13032, August 2013
P a g e | 173
28. CERTIFICATES
OPTIRO PTY LTD
CERTIFICATE OF QUALIFIED PERSON – ANDREW LAW
As the lead author and a Qualified Person of the report entitled “Technical Report on the Dikulushi
Underground Project, Democratic Republic of Congo” (the Study) dated 12 December 2013, on the
Underground Project of Mawson West Limited, I hereby state:
1. My name is Andrew Law and I am a full time employee of the firm Optiro Pty Ltd of Level 4,
50 Colin Street, West Perth, WA, 6005, Australia.
2. I am a practising Mining Engineer and a Fellow of the AusIMM (107318), also a Fellow of the
Institute of Quarrying Australia (991004), and a Member of the Australian Institute of Company
Directors (0044149).
3. I am a graduate of the Witwatersrand Technikon, Johannesburg, South Africa, with a HND
Metalliferous Mining, in 1982.
4. I have practiced my profession continuously since 1983.
5. I am an “independent” and “qualified person” as the terms are defined in National Instrument
43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
6. I have performed consulting services and reviewed files and data associated with the Dikulushi
Project from August 2011 to the present.
7. I visited the Dikulushi Project property and the underground as far as the 850Mrl (est water
level) in February 2012. I have performed consulting services and reviewed files and data
associated with Dikulushi between August 2011 and the present time
8. Based on the information provided by Mawson West Ltd and reviewed by myself, I contributed
to Sections 1,4,5,6,15, 16, 19, 20, 21, 22, 24, 25, and 26.
9. As of December 12, 2013, the effective date of the Study, to the best of my knowledge,
information and belief, the Study contains all scientific and technical information that is required
to be disclosed to make the Study not misleading.
10. I have read the National Instrument and Form 43-101F1 (the “Form”) and the Study has been
prepared in compliance with the Instrument and the Form.
11. I do not have nor do I expect to receive a direct or indirect interest in the Dikulushi property of
Mawson West Ltd, and I do not beneficially own, directly or indirectly, any securities of Mawson
West Ltd or any associate or affiliate of such company.
Dated at Perth, Western Australia, on the 20 December 2013.
Andrew Law
FAusIMM
Director - Mining (Optiro Pty Ltd)
P a g e | 174
OPTIRO PTY LTD
CERTIFICATE OF QUALIFIED PERSON – IAN GLACKEN.
As one of the authors of the report entitled “Technical Report on the Dikulushi Underground Project,
Democratic Republic of Congo” (the Study) dated 12 December 2013, on the Underground Project of
Mawson West Limited, I hereby state:
1. My name is Ian Glacken and I am a full-time employee of the firm Optiro Pty Ltd of Level 4,
50 Colin Street, West Perth, WA, 6005, Australia.
2. I am a practising geologist and a Fellow of the AusIMM (107194) and a Chartered Professional
Geologist. I am also a Member of the Institution of Metals Mining and Materials (IMMM, 46394)
and a Chartered Engineer of this Institution.
3. I am a graduate of Durham University in the United Kingdom with a BSc (Hons) in Geology in
1979, the Royal School of Mines in the United Kingdom with MSc in Mineral Exploration in 1981
and Stanford University in the USA with an MSc in Geostatistics in 1996.
5. I am an “Independent” and “qualified person” as the terms are defined in National Instrument
6. I have not visited the Dikulushi Project property. I have performed consulting services and
reviewed files and data associated with the Dikulushi and Kazumbula Projects between May
2009 and the present.
7. I take responsibility for Sections 1 (in part), 7, 8, 9, 10, 11, 12 and 14 of the Study and have
contributed to Sections 17.1 and 17.3 and the associated text in the summary, conclusions and
recommendations.
9. I am independent of Mawson West Ltd pursuant to section 1.4 of the Instrument.
Dated at Perth, Western Australia, on the 20 December, 2013.
Ian Glacken
BSc (Hons) (Geology), FAusIMM(CP), MIMMM, CEng
Principal Consultant (Optiro Pty Ltd)
P a g e | 175
TURNER MINING AND GEOTECHNICAL PTY LTD
CERTIFICATE OF QUALIFIED PERSON – MIKE TURNER
1. My name is Mike Turner and I am a full-time employee of Turner Mining and Geotechnical Pty
Ltd of 3B Valley Road, Wembley Downs, WA, 6019, Australia.
2. I am a practising geotechnical and mining engineer and a Chartered Professional Fellow of the
AUSIMM (205399).
3. I am a graduate of Imperial College, London University with a BSc (Eng) (Hons) in Mining in 1979.
I also obtained a Master of Science in Mineral Production Management at the Royal School of
Miners in 1984 and a Chamber of Mines Certificate in Rock Mechanics in South Africa in 1987.
4. I have practiced my profession continuously since 1979, apart from a 12 month period from
1983/1984 during which I completed the MSc course at Imperial College.
6. I visited the Dikulushi Project from 10th to 13th December 2012, the most recent visit prior to
completion of the Study. I have performed consulting services and reviewed files and data
associated with Dikulushi between July 2003 and the present.
7. I am responsible for the geotechnical section in Section 16.
Dated at Perth, Western Australia, on 20 December 2013
Michael Harry Turner, MSC, DIC, BSc (Eng) (Hons) (Mining), ARSM, FAusIMM (CP), RPEQ
Director, Turner Mining and Geotechnical Pty Ltd
P a g e | 176
KNIGHT PIESOLD
CERTIFICATE OF QUALIFIED PERSON – DUNCAN JOHN GRANT-STUART
As a reviewer of the report entitled “Technical Report on the Dikulushi Underground Project,
1. My name is Duncan John Grant-Stuart and I am a full time Engineer with the firm of Knight Piesold
(Pty) Limited of PO Box 221, Rivonia, 2128, South Africa.
2. I am a practising Civil Engineer and member of the Institution of Civil Engineers (UK)(MICE) and am
registered with the Engineering Council (UK) (C.Eng) and the Engineering Council of South Africa
(PR.Eng).
3. I am a graduate of the University of the Witwatersrand with a BSC (Eng) degree completed in 1976.
5. I am an “Independent” and “qualified person” as the terms are defined in National Instrument 43101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
6. I visited the Dikulushi Project property. I have performed consulting services and reviewed files and
data supplied by Mawson West Ltd in 2011.
7. I reviewed Section 17 of the Study, as well as the associated text in the summary, conclusions and
recommendations.
8. I am responsible for the geotechnical section and the associated text in the summary conclusions
and recommendations.
9. I am not aware of any limitations imposed upon my access to persons, information, data or
documents that I consider relevant to the subject matter of the study.
10. I am not aware as at 12 December 2013, the effective date of the Study, of any material fact or
material change with respect to the subject matter of the Study, which is not reflected in the Study,
the omission of which would make the Study misleading.
11. I am independent of Mawson West Ltd and AMC SARL pursuant to section 1.4 of the Instrument.
Dated at Rivonia, South Africa, on 20 December 2013
Duncan John Grant-Stuart
PR Eng 900014
C.Eng
Technical Consultant
P a g e | 177
SEDGMAN
SEDGMAN
CERTIFICATE OF QUALIFIED PERSON – PETER HAYWARD
1. My name is Peter George Hayward and I am Senior Process Engineer with the firm Sedgman
Limited, Suite 3, 3 Craig Street, Burswood, 6100.
2. I am a practicing Metallurgist and a Fellow of the Australian Institute of Mining and Metallurgy.
3. I am a graduate of the Ballarat Institute of Advanced Engineering and hold a Diploma of
Metallurgy.
4. I have practiced my profession continuously since February 1974.
6. I have personally visited the Dikulushi property in February 2012. I have reviewed files and data
supplied by Mawson West Ltd in September 2011 and in January 2013.
7. I have contributed to the Sections 13 and 17.
8. I am not aware of any limitations imposed upon my access to persons, information, data or
documents that I consider relevant to the subject matter for the Study (apart from as indicated
in the text).
9. I am not aware as at 12 December 2013, the effective date of the Study, of any material fact or
material change with respect to the subject matter of the Study, which is not reflected in the
Study, the omission of which would make the Study misleading.
Dated at Perth, Western Australia, on 20 December 2013
Peter Hayward
P a g e | 178

43-101 Dikulushi Underground

Transcription

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