UNITED STATES
SECURITIES AND EXCHANGE COMMISSION
Washington, D.C. 20549
Form 6-K
REPORT OF FOREIGN PRIVATE ISSUER PURSUANT TO RULE 13a-16 or
15d-16 UNDER THE
SECURITIES EXCHANGE ACT OF 1934
For the month of July, 2015.
Commission File Number 001-32399
BANRO CORPORATION
(Translation of registrants name into English)
1 First Canadian Place
100 King Street West, Suite
7070
Toronto, Ontario, Canada
M5X 1E3
(Address of
principal executive offices)
Indicate by check mark whether the registrant files or will
file annual reports under cover Form 20-F or Form 40-F
Form 20-F [ X
]
Form 40-F [ ]
Indicate by check mark if the registrant is submitting the Form
6-K in paper as permitted by Regulation S-T Rule 101(b)(1): [ ]
Note: Regulation S-T Rule 101(b)(1) only permits the
submission in paper of a Form 6-K if submitted solely to provide an attached
annual report to security holders.
Indicate by check mark if the registrant is submitting the Form
6-K in paper as permitted by Regulation S-T Rule 101(b)(7): [ ]
Note: Regulation S-T Rule 101(b)(7) only permits the
submission in paper of a Form 6-K if submitted to furnish a report or other
document that the registrant foreign private issuer must furnish and make public
under the laws of the jurisdiction in which the registrant is incorporated,
domiciled or legally organized (the registrants home country), or under the
rules of the home country exchange on which the registrants securities are
traded, as long as the report or other document is not a press release, is not
required to be and has not been distributed to the registrants security
holders, and, if discussing a material event, has already been the subject of a
Form 6-K submission or other Commission filing on EDGAR.
SIGNATURE
Pursuant to the requirements of the Securities Exchange Act of
1934, the registrant has duly caused this report to be signed on its behalf by
the undersigned, thereunto duly authorized.
|
BANRO CORPORATION |
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/s/
Kevin Jennings
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Date: August 5, 2015 |
Kevin Jennings |
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Chief Financial Officer |
-2-
INDEX TO EXHIBITS
-3-
NI 43-101 TECHNICAL REPORT
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MINERAL RESOURCE AND RESERVE UPDATE
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DECEMBER 31, 2014
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TWANGIZA GOLD MINE
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DEMOCRATIC REPUBLIC OF THE CONGO
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Prepared For |
TWANGIZA MINING SA,
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A subsidiary of Banro Corporation
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Effective Date of Report: July 29, 2015 |
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Martin Pittuck, MSc., C.Eng, MIMMM |
Allan Blair, B.App.Sc, MBA, MAusIMM |
David Pattinson, PhD., BSc., C.Eng, MIMMM |
Daniel Bansah, MSc.(MinEx), MAusIMM (CP) |
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Report Prepared by |
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SRK Consulting (UK) Limited |
UK6391 |
SRK Consulting |
NI
43-101 Twangiza - Details |
COPYRIGHT AND DISCLAIMER
SRK Consulting (UK) Limited (SRK) has prepared this report in accordance
with its signed consultancy agreement dated 11 March 2015 (the
Contract) under which its services were performed for Twangiza
Mining SA (Twangiza). The Contract permits Banro to file this report on
SEDAR and EDGAR with the applicable securities regulatory authorities.
Banro assumes regulatory responsibility for the filing of this report.
Except for the purposes legislated under applicable securities law, any other
uses of this report by any third party is at that partys sole risk.
The user of this report should ensure that this is the most
recent Technical Report for the property as it is not valid if a new Technical
Report has been issued.
The quality of information, conclusions,
and estimates contained herein is consistent with the level of effort
involved in SRKs services, based on: i) information available at the
date of this report, ii) data supplied by outside sources, and iii) the
assumptions, conditions, and qualifications set forth in this report.
This report contains forward-looking statements which
address activities, events or developments which are believed, expected or
anticipated will or may occur in the future (including, without
limitation, statements regarding estimates and/or assumptions in respect
of net present values and future cash flows, Mineral Resource and Mineral
Reserve estimates, future gold production, costs, mine life, gold
recoveries, potential Mineral Resources and Mineral Reserves and
production, development and exploration plans and objectives).
These forward-looking statements reflect current expectations based on
information currently available and are subject to a number of risks
and uncertainties that may cause the actual results to differ
materially from those discussed in the forward-looking statements; these
include, among other things: uncertainty of estimates of capital and
operating costs, production estimates and estimated economic return;
the possibility that actual circumstances will differ from the estimates
and assumptions used in economic studies; failure to establish
estimated Mineral Resources and Mineral Reserves; fluctuations in gold
prices and currency exchange rates; inflation; gold recoveries being
less than those indicated by the metallurgical testwork carried out to
date; uncertainties relating to the availability and costs of
financing needed in the future; changes in equity markets; political
developments in the DRC; lack of infrastructure; failure to procure or
maintain, or delays in procuring or maintaining, permits and
approvals; lack of availability at a reasonable cost or at all, of
plants, equipment or labour; inability to attract and retain key
management and personnel; changes to regulations affecting mining
activities; the uncertainties involved in interpreting drilling
results and other geological data.
Although it is believed that the assumptions inherent
in the forward-looking statements are reasonable, forward-looking
statements are not guarantees of future performance and accordingly undue
reliance should not be put on such statements due to the inherent
uncertainty therein.
The Mineral Resource and Mineral Reserve figures
referred to in this report are estimates and no assurances can be given
that the forecast levels of gold will be produced. Such estimates are
expressions of judgment based on knowledge, mining experience, analysis of
drilling results and industry practices. Valid estimates made at a given
time may significantly change when new information becomes available.
While it is believed that the Mineral Resource and Mineral Reserve
estimates included in this report are well established, by their
nature Mineral Resource and Mineral Reserve estimates are imprecise
and depend, to a certain extent, upon statistical inferences which may
ultimately prove unreliable. Mineral Resources that are not Mineral
Reserves do not have demonstrated economic viability. There is no
certainty that Mineral Resources can be upgraded to Mineral Reserves
through continued exploration.
Due to the uncertainty attached to Inferred Mineral Resources, it
cannot be assumed that all or any part of an Inferred Mineral Resource
will be upgraded to an Indicated or Measured Mineral Resource as a
result of continued exploration. |
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43-101 Twangiza - Details |
COPYRIGHT AND DISCLAIMER (cont.)
The United States Securities and Exchange Commission (the "SEC")
permits U.S. mining companies, in their filings with the SEC, to
disclose only those mineral deposits that a company can economically
and legally extract or produce. Certain terms are used in this report,
such as "Measured", "Indicated", and "Inferred" "Resources", that the
SEC guidelines strictly prohibit U.S. registered companies from
including in their filings with the SEC. U.S. Investors are urged to
consider closely the disclosure in Banro's Form 20-F Registration
Statement, File No. 001-32399, which may be secured from Banro, or from the SEC's website at http://www.sec.gov/edgar.shtml.
© SRK Consulting (UK) Limited 2015 |
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SRK Legal Entity: |
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SRK Consulting (UK) Limited |
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SRK Address:
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5th Floor Churchill House |
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17
Churchill Way |
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Cardiff, CF10 2HH |
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Wales,
United Kingdom. |
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Date: |
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July
2015 |
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Project
Number: |
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UK6391
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Principal Consultant |
SRK Project
Director: |
Mark Campodonic |
(Resource Geology) |
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Corporate Consultant |
SRK Project
Manager: |
Martin Pittuck |
(Mining Geology) |
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Client Legal
Entity: |
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Twangiza Mining SA |
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Client
Address: |
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8
Avenue Mwanga |
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BUKAVU
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Democratic Republic of Congo.
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SRK Consulting (UK) Limited |
5th Floor Churchill House |
17 Churchill Way |
City and County of Cardiff |
CF10 2HH, Wales |
United Kingdom |
E-mail: enquiries@srk.co.uk |
URL: www.srk.co.uk |
Tel: + 44
(0) 2920 348 150 |
Fax: + 44
(0) 2920 348 199 |
NI 43-101 TECHNICAL REPORT |
MINERAL RESOURCE AND RESERVE UPDATE, DECEMBER 31
2014 |
TWANGIZA GOLD MINE, DEMOCRATIC REPUBLIC OF THE
CONGO |
SRK Consulting (UK) Limited (SRK
(UK)) was commissioned by Twangiza Mining SA (Twangiza Mining), a subsidiary of
Banro Corporation, to independently review Twangiza Minings December 31 2014
Mineral Reserve estimate which for the first time includes non-oxide material,
increasing the Mineral Reserve and extending the Mine life to 14 years.
This report presents a compilation of
information prepared by Twangiza Mining and presents findings from a
reconciliation study completed as part of the review process and Mineral Reserve
estimate.
This report provides:
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an update on the expansion and upgrade of the
original processing plant to allow processing of non-oxide material at an
annual throughput of 1.7 million tonne per annum (Mtpa); |
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a summary of changes implemented as a result of
production reconciliation with historical estimates; |
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historical and forecast mining and processing
operating costs based on the ability to process non-oxide material at the
increased annual throughput; |
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mine plan optimisation for the 1.7Mtpa
processing operating costs, at a range of gold prices and mining operating
costs; |
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practical pit designs based on an optimal pit
shell; |
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a production schedule based on an economic
cut-off grade; |
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the tailings management facility plan to
accommodate the increased Mineral Reserve tonnage and throughput rate;
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a financial model that incorporates the
additional capital to be expensed during the mine life and assesses the
sensitivity of the project to gold price fluctuation; and |
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SRK (UK)s review comments.
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The Twangiza project is located at
approximately 2º52 South and 28º45 East in the South Kivu Province of the
Democratic Republic of the Congo (DRC), some 35 kilometres west of the Burundi
Border and 45 kilometres to the south-southwest of Bukavu, the provincial
capital.
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Registered Address: 21 Gold Tops,
City and County of Newport, NP20 4PG, |
Group
Offices: Africa |
Wales, United Kingdom. |
Asia |
SRK Consulting (UK) Limited Reg No
01575403 (England and Wales) |
Australia |
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Europe |
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North America |
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South America |
SRK Consulting |
NI
43-101 Twangiza Summary |
Banros properties all lie within the
Kibara Belt, a Proterozoic intracontinental mobile belt situated between the
Congo Craton in the west and the Tanzanian Craton in the east. Gold
mineralisation at Twangiza is hosted by a folded package of mudstone and
siltstone sediments and porphyry sills, confined by a doubly plunging anticlinal
structure. Mineralisation is found along a 3.5 kilometre long, north trending
corridor which hosts the two principal deposits of Twangiza Main and Twangiza
North.
The Twangiza property consists of six
exploitation permits totalling 1,156 square kilometres which are wholly-owned by
Twangiza Mining, a subsidiary of Banro Corporation. These exploitation permits
will expire in 2016 and are subject to renewal for consecutive 15 year
periods.
The Twangiza Project poured its first
gold in October 2011. The mine commenced with a refurbished plant designed
primarily to process oxide material at 1.3Mtpa. Since commissioning, the plant
has been improved and expanded to 1.7Mtpa and has been shown to be able to
process harder non-oxide material.
1.2 |
Mineral Resource Statement |
The Mineral Resources at Twangiza are
based on a model originally prepared by SRK (UK) in 2009. The model has
subsequently been updated under the supervision of Daniel Bansah of Twangiza
Mining and Banro and has been reviewed by Martin Pittuck of SRK (UK). The
revised model includes modifications to density values to account for a variance
between the original model and tonnages mined to date, and modifying factors
based on an analysis of historical production records.
The Mineral Resource estimate is
reported according to the definitions and guidelines given in the Canadian
Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral
Resources and Mineral Reserves. The Mineral Resource Statement uses a cut-off
grade of 0.4 g/t gold; it has been restricted to a pit shell which uses a
USD1,600/oz gold price which is considered therefore to have reasonable
prospects for economic extraction by open pit mining.
Table ES 1 below details the Oxide
and Non-Oxide components of the Twangiza Mineral Resource estimate split by
confidence category, at a cut-off grade of 0.4 g/t gold. The Mineral Resources
are inclusive of the Mineral Reserves.
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Table ES 1: |
Twangiza Mineral Resource
Estimate By Confidence Category
(December 31, 2014) |
OXIDE MINERAL RESOURCE
CATEGORY |
TONNES
(Mt) |
GRADE(g/t
Au ) |
GOLD
OUNCES (Moz) |
MEASURED |
3.72 |
2.30 |
0.28 |
INDICATED |
8.76 |
1.88 |
0.53 |
MEASURED AND INDICATED |
12.48 |
2.02 |
0.81 |
INFERRED |
1.34 |
1.32 |
0.06 |
NON-OXIDE MINERAL RESOURCE
CATEGORY |
TONNES
(Mt) |
GRADE(g/t
Au ) |
GOLD
OUNCES (Moz) |
MEASURED |
3.80 |
2.23 |
0.27 |
INDICATED |
93.00 |
1.40 |
4.18 |
MEASURED AND INDICATED |
96.80 |
1.43 |
4.45 |
INFERRED |
11.65 |
1.12 |
0.42 |
NB: Any apparent errors are due to
rounding and are therefore not considered material to the estimate
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43-101 Twangiza Summary |
1.3 |
Mineral Reserve Statement |
Mineral Reserves were estimated by
Twangiza Mining under the supervision of Daniel Bansah and reviewed by a team
from SRK (UK) led by Martin Pittuck who is a Qualified Person as such term is
defined in National Instrument 43-101. The Mineral Reserve Statement is reported
in accordance with National Instrument 43-101 requirements.
The Mineral Reserves stated in 2009
were restricted to oxide material; oxide mining commenced in 2011 and in recent
years some non-oxide material has been increasingly blended into the plant feed
averaging 20% in the second half of 2014.
The Mineral Reserves given in Table ES
2 below have now increased due to the addition of non-oxide material; they are
contained in a practical pit design and they include the Valley Fill material.
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Table ES 2: |
Twangiza Mineral Reserve Estimate (December
31, 2014) |
CATEGORY |
TONNES (Mt) |
GRADE(g/tAu ) |
GOLD (Moz) |
PROVEN |
7.47 |
2.41 |
0.58 |
PROBABLE |
14.91 |
2.22 |
1.06 |
PROVEN + PROBABLE |
22.38 |
2.28 |
1.64 |
Twangiza Mining will need to invest in
expanding its waste containment capability. Metago Environmental Engineers, who
designed the current tailings management facility (TMF), have evaluated
accelerated wall raising costs on the existing TMF and are designing and costing
an additional TMF to supplement the existing designed tailings capacity.
Table ES 3 below summarizes the
estimated capital costs associated mainly with mining operations and process
plant upgrade to meet the requirements of the blend feed at a throughput of
1.7Mtpa.
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Table ES 3: |
Capital Cost Summary
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ITEM |
COST (USD million) |
CAPITALISED EXPENDITURE |
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MINING SUSTAINING CAPITAL |
25.5 |
PROCESSING SUSTAINING CAPITAL |
6.4 |
TMF CONSTRUCTION |
46.0 |
TAILINGS SUSTAINING CAPITAL |
30.5 |
GENERAL & ADMINISTRATION - SUSTAINING CAPITAL |
20.6 |
BANRO FOUNDATION |
1.2 |
TOTAL CAPITALISED EXPENDITURE |
130.2 |
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43-101 Twangiza Summary |
1.5 |
Operating Cost Summary |
The operating costs in Table ES 4 below
were estimated and incorporated into the financial analysis. Estimates have been
based on a zero-based cost analysis following review of 2014 historical costs
and implementation of several cost saving measures.
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Table ES 4: |
Summary of LoM Operating Costs
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ITEM |
USD million
/ annum |
USD /
t processed |
USD /
oz poured |
MINING |
9.53 |
5.96 |
107 |
PROCESSING |
32.1 |
20 |
361 |
G&A |
20.6 |
12.9 |
232 |
TOTAL OPERATING COSTS |
62 |
39 |
699 |
The cash flow model for the Twangiza
project summarised in Table ES 5 below is based on the 31 December 2014 Mineral
Reserve. It assumes a base case gold price of USD1,200 per ounce and a 5%
discount rate. The financial model also reflects the fiscal aspects of the
mining convention governing the Twangiza project, which includes a 100% equity
interest and a 10 year tax holiday from the start of production. An
administrative tax of 5% for the importation of plant, machinery and consumables
has been included in the projected capital and operating costs.
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Table ES 5: |
Financial Analysis Summary
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ITEM |
UNIT |
AMOUNT |
LIFE OF MINE GOLD PRODUCTION |
koz |
1,246 |
PRODUCTION PERIOD |
years |
14 |
ANNUAL GOLD PRODUCTION FOR FIRST 5 YEARS |
koz |
109 |
TOTAL CAPITAL COSTS |
USD/oz |
104 |
ALL IN COSTS |
USD/oz |
888 |
POST-TAX NET PRESENT VALUE |
USD million |
285 |
NET CASH FLOW AFTER TAX AND CAPEX |
USD million |
395 |
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43-101 Twangiza Summary |
1.6.1 |
Sensitivity analysis |
A sensitivity analysis was performed on
the after tax profits by varying the gold price between USD1,000 and USD1,600
per ounce. The results are given in Table ES 6.
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Table ES 6: |
Cash Flow Sensitivity
|
GOLD PRICE (USD/oz) |
NET PRESENT VALUE (USD
million) |
1600 |
1500 |
1400 |
1300 |
1200 |
1100 |
1000 |
NPV 0.0% |
847 |
734 |
621 |
508 |
395 |
282 |
169 |
NPV 5.0% |
612 |
530 |
449 |
367 |
285 |
203 |
122 |
NPV 8.0% |
516 |
447 |
378 |
310 |
241 |
172 |
103 |
NPV 9.5% |
476 |
413 |
350 |
286 |
223 |
159 |
96 |
NPV 10.0% |
464 |
402 |
341 |
279 |
217 |
156 |
94 |
NPV 12.5% |
411 |
356 |
302 |
247 |
193 |
139 |
84 |
NPV 15.0% |
366 |
318 |
270 |
222 |
173 |
125 |
77 |
1.7 |
Conclusions and
Recommendations |
SRK (UK) has reviewed the technical and
economic work presented by Twangiza Mining and has assisted in the technical
review of historical production records at the mine.
Overall the Mineral Resource base is
considered to be well known; the oxide and non-oxide resources are modelled to
an appropriate level of confidence for estimation of Mineral Reserves.
Adjustment of the resource model densities has been implemented following a
reconciliation of the block model with historical production which also gave
support to the modifying factors used for estimating Mineral Reserves.
Mineral Reserves are contained within
open pits which have been appropriately designed, however a redesign is
recommended to bring old designs closer to more recent optimisations. Of
particular note is the addition of non-oxide material to Mineral Reserve
following the completion of a plant upgrade and capacity increase; this has
allowed some 12.8 Mt of non-oxide ore to be added the reserve. The non-oxide
material has very variable metallurgical recoveries, the detailed lithological
and weathering models ensure accurate estimation of recoverable gold grade
within the deposit and a cut-off grade is applied on this basis. The various ore
types currently appear in different proportions in the mining schedule on an
annual basis and SRK (UK) recommends further smoothing of the mining schedule to
even this out.
The process plant has not worked at
design capacity historically, mainly due to a shortage of funding. The plant has
now been refurbished and SRK (UK) considers that it will be capable of operating
at 1.7 Mtpa to fulfil the mine plan presented in this report as long as
reagents, consumables and spares are adequately funded and supplied in good
time.
The increased Mineral Reserve requires
an increased tailings storage capacity. The existing facility is not cost
effective and plans to supplement this capacity with a new facility are being
developed.
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Compared with 2014 costs, the annual
and unit operating costs in the mine plan are lower primarily as a result of
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plant upgrade investment coming to an end,
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diesel fuel price reductions, |
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a number of cost saving measures with suppliers
and contractors that have been or will be implemented, and |
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switching to hydroelectric power
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The cost savings are based on plans
which are at an early stage of implementation and require confirmation in
practice. SRK (UK) is confident that if cost savings are made, then the project
NPV will be as presented in this report, however if the changes are not realised
then the NPV may be considerably lower.
Regarding the lower power costs that
are anticipated from a hydroelectric scheme; these savings will only be realised
after the hydroelectric facility has been built, however investors and
contractors have yet to be identified.
There are a number of capital items to
be incurred in the mine pan, the most significant of which is for tailing
management facility construction; SRK considers the cost estimate to be in line
with tailings construction cost metrics currently incurred on site and to be
broadly in line with some design studies completed some years ago. These
designs, costs and construction schedules need to be updated to address the
specific requirement of the current mine plan. The capital costs are substantial
and SRK considers that the final realised cost will be within the means of the
projects cash flow.
The Twangiza North pit area and the
area identified for the new tailings dam both require resettlement of dwellings
and relocation of artisanal mining activity. Whilst Twangiza Mining has a good
track record of achieving this at Twangiza Main pit, the new areas may yet
present challenges; SRK (UK) recommends that these activities are planned in
good time in order to avoid significant delays to the mine plan.
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43-101 Twangiza Table of Contents |
Table of Contents
1 |
SUMMARY |
4 |
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1.1 |
Project Overview |
4 |
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1.2 |
Mineral Resource Statement |
5 |
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1.3 |
Mineral Reserve Statement |
6 |
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1.4 |
Capital Cost Summary |
6 |
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1.5 |
Operating Cost Summary |
7 |
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1.6 |
Financial Analysis |
7 |
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1.6.1 Sensitivity analysis |
8 |
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1.7 |
Conclusions and Recommendations
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8 |
2 |
INTRODUCTION |
20 |
3 |
RELIANCE ON OTHER
EXPERTS |
20 |
4 |
PROPERTY DESCRIPTION AND
LOCATION |
20 |
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4.1 |
Mineral Tenure |
21 |
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4.2 |
Natural and Existing Features |
25 |
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4.3 |
Royalties and Other Payments
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25 |
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4.4 |
Environmental Liabilities |
25 |
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4.5 |
Required Permits and Approvals
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25 |
5 |
ACCESSIBILITY, CLIMATE, LOCAL
RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY |
26 |
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5.1 |
Topography, Elevation,
Vegetation and Seismic Activity |
26 |
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5.2 |
Means of Access |
26 |
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5.3 |
Climate |
26 |
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5.4 |
Surface Rights and Available Local
Infrastructure |
27 |
6 |
HISTORY |
28 |
7 |
GEOLOGICAL SETTING AND
MINERALISATION |
29 |
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7.1 |
Regional Geology |
29 |
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7.2 |
Property Geology |
30 |
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7.3 |
Weathering |
31 |
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7.4 |
Mineralisation |
31 |
8 |
DEPOSIT TYPES
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32 |
9 |
EXPLORATION |
33 |
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9.1 |
Historical Exploration |
33 |
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9.2 |
Recent Exploration (October 2005 December
2006) |
33 |
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9.2.1 Soil geochemical
programme |
33 |
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9.2.2 Trenching programme |
34 |
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9.2.3 Prospect scale mapping
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34 |
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9.3 |
Recent Exploration (January 2007 November
2008) |
34 |
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9.3.1 Geophysical exploration
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34 |
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9.3.2 |
LIDAR survey |
34 |
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9.3.3 |
Additional regional work |
34 |
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9.4 |
Recent Exploration (January 2009
December 2010) |
35 |
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9.4.1 |
Twangiza West Zone |
35 |
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9.4.2 |
Twangiza East Zone |
35 |
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9.4.3 |
Valley Fill |
35 |
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9.4.4 |
Geochemical surveys |
35 |
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9.4.5 |
Regional work |
36 |
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9.5 |
Recent Exploration (January 2011
December 2012) |
36 |
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9.6 |
Recent Exploration
(January 2013 December 2014) |
36 |
10 |
DRILLING |
37 |
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10.1 |
Drilling (September
1997 March 1998) |
37 |
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10.2 |
Drilling (February 2006 May 2008)
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37 |
|
10.3 |
Drilling (May 2008
November 2008) |
39 |
|
10.4 |
Drilling (December 2008 December
2012) |
39 |
|
10.5 |
Drilling (January
2013 December 2014) |
40 |
|
10.6 |
Results |
40 |
11 |
SAMPLE
PREPARATION, ANALYSES AND SECURITY |
41 |
|
11.1 |
SAMPLING METHOD AND APPROACH |
41 |
|
|
11.1.1 |
Soil Geochemistry |
41 |
|
|
11.1.2 |
Trench Samples |
41 |
|
|
11.1.3 |
Stream Sediment Samples |
41 |
|
|
11.1.4 |
Drill Core Samples |
41 |
|
|
11.1.5 |
Auger Samples |
41 |
|
11.2 |
Sample Preparation and Analysis |
42 |
|
|
11.2.1 |
Sample Preparation |
42 |
|
11.3 |
Laboratory Analysis |
43 |
|
11.4 |
Quality Control
Procedures |
43 |
|
11.5 |
Assessment of Quality Control Data
|
43 |
|
|
11.5.1 |
Certified Reference Material
|
43 |
|
|
11.5.2 |
Inter-laboratory check assays |
46 |
|
|
11.5.3 |
Duplicate coarse split |
46 |
|
|
11.5.4 |
Blank samples |
48 |
12 |
DATA
VERIFICATION |
49 |
|
12.1 |
Database and Data Quality |
49 |
|
12.2 |
Adit Check Sampling
|
49 |
|
12.3 |
Data Validation |
50 |
13 |
MINERAL PROCESSING
AND METALLURGICAL TESTING |
52 |
|
13.1 |
Background |
52 |
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13.2 |
Sample Selection |
55 |
|
13.3 |
Review of Scoping Study Metallurgical
Testwork |
55 |
|
13.4 |
Review of
Pre-Feasibility Study Metallurgical Testwork |
56 |
|
|
13.4.1 |
Mineralogy |
56 |
|
|
13.4.2 |
Comminution results of
composite samples |
57 |
|
|
13.4.3 |
Comminution results of variability samples |
57 |
|
|
13.4.4 |
Gold recovery testwork |
57 |
|
|
13.4.5 |
Gravity and intensive cyanidation testwork
(oxides only): |
57 |
|
|
13.4.6 |
Variability leach testwork
(oxides only) |
58 |
|
|
13.4.7 |
Cyanide Detoxification (Oxides Only) |
58 |
|
|
13.4.8 |
Transition and fresh (main and
north) optimisation recovery testwork |
59 |
|
|
13.4.9 |
Settling and viscosity |
59 |
|
13.5 |
Review of Feasibility
Study Metallurgical Testwork |
60 |
|
|
13.5.1 |
Gravity recovery main and north transition
and fresh samples |
60 |
|
|
13.5.2 |
Cyanide destruction |
61 |
|
|
13.5.3 |
Comminution appraisal |
62 |
|
|
13.5.4 |
Recovery variability testing
|
64 |
|
|
13.5.5 |
Transition CMS and fresh CMS refractory ore
testwork |
67 |
|
13.6 |
Review of Comminution
Circuit Parameters to Process 1.7 Mtpa |
72 |
|
|
13.6.1 |
Background |
72 |
|
|
13.6.2 |
Summary of investigation by
Orway Mineral Consultants |
72 |
14 |
MINERAL RESOURCE ESTIMATE |
74 |
|
14.1 |
Introduction |
74 |
|
14.2 |
Approach |
74 |
|
14.3 |
Density
determinations |
74 |
|
|
14.3.1 |
Feasibility Study Data collection and analysis
|
74 |
|
|
14.3.2 |
Pit Density Determinations |
75 |
|
|
14.3.3 |
Bulk Sample Density Determinations |
76 |
|
|
14.3.4 |
Revised Density Model |
76 |
|
14.4 |
Descriptive statistics of assay data
|
79 |
|
14.5 |
Geological Modelling
|
79 |
|
|
14.5.1 |
Geological wireframes |
79 |
|
|
14.5.2 |
Mineralisation wireframe |
82 |
|
|
14.5.3 |
Geological block model |
83 |
|
14.6 |
Topography,
oxide/transition sub-models |
85 |
|
14.7 |
Statistical analysis of the
mineralised data |
85 |
|
|
14.7.1 |
Selection of composite lengths
for statistics |
85 |
|
|
14.7.2 |
Summary statistics and histograms |
85 |
|
|
14.7.3 |
High grade capping |
87 |
|
14.8 |
Geostatistical analysis |
88 |
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|
14.8.1 |
Semi-variograms |
88 |
|
|
14.8.2 |
Block estimation |
91 |
|
14.9 |
Mineral Resource
Classification |
92 |
|
|
14.9.1 |
Geological complexity |
92 |
|
|
14.9.2 |
Quality of data used in the
estimation |
93 |
|
|
14.9.3 |
Results of the geostatistical analysis |
93 |
|
|
14.9.4 |
Classification Method |
93 |
|
14.10 |
Valley Fill model |
94 |
|
14.11 |
Mineral Resource
Statement |
96 |
15 |
MINERAL RESERVE ESTIMATE |
98 |
|
15.1 |
Geotechnical |
98 |
|
15.2 |
Open pit optimization |
98 |
|
|
15.2.1 |
Introduction |
98 |
|
|
15.2.2 |
Cost inputs |
99 |
|
|
15.2.3 |
Mining factors |
100 |
|
|
15.2.4 |
Cut-off grade |
100 |
|
|
15.2.5 |
Whittle results |
100 |
|
15.3 |
Practical pit design |
107 |
|
15.4 |
Mine production
schedule |
112 |
|
15.5 |
Modifying Factors |
118 |
|
15.6 |
Mineral Reserve
Estimate |
118 |
16 |
MINING METHODS |
121 |
|
16.1 |
Mining Method and
Site Layout |
121 |
|
16.2 |
Site Preparation |
122 |
|
16.3 |
Drill and Blast |
122 |
|
16.4 |
Loading and Hauling |
122 |
|
16.5 |
Manpower |
124 |
|
16.6 |
Mining Operations |
126 |
17 |
RECOVERY
METHODS |
128 |
|
17.1 |
Process Flow Sheet |
128 |
|
17.2 |
Process Plant Design
|
128 |
|
|
17.2.1 |
Front End / Crushing and Screening |
129 |
|
|
17.2.2 |
Secondary Plant Feed Point |
130 |
|
|
17.2.3 |
Upgrade of feed weightometers |
130 |
|
|
17.2.4 |
Milling |
130 |
|
|
17.2.5 |
CIL |
131 |
|
|
17.2.6 |
Acid Wash, Elution and
Regeneration |
131 |
|
|
17.2.7 |
Electro-winning and Smelting |
132 |
|
|
17.2.8 |
Tailings |
132 |
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|
17.2.9 |
Water Supply Systems |
132 |
|
|
17.2.10 |
Summary and Conclusions |
132 |
|
17.3 |
General Process Plant Description |
133 |
|
|
17.3.1 |
Introduction |
133 |
|
|
17.3.2 |
Mobile Crushing and Screening |
133 |
|
|
17.3.3 |
ROM Pad Storage Area Primary
Feed Point |
133 |
|
|
17.3.4 |
In-Plant Crushing and Scrubbing |
134 |
|
|
17.3.5 |
Secondary Feed Point |
134 |
|
|
17.3.6 |
Milling and Classification |
134 |
|
|
17.3.7 |
Gravity and Intensive
Cyanidation |
135 |
|
|
17.3.8 |
Trash Removal |
136 |
|
|
17.3.9 |
Leach and Adsorption Circuit
|
136 |
|
|
17.3.10 |
Carbon Safety and Detoxification |
137 |
|
|
17.3.11 |
Tailings Dam Storage and Return
|
138 |
|
|
17.3.12 |
Acid Wash |
138 |
|
|
17.3.13 |
Elution |
139 |
|
|
17.3.14 |
Electrowinning |
139 |
|
|
17.3.15 |
Regeneration |
140 |
|
|
17.3.16 |
Gold Room |
140 |
|
17.4 |
Consumables |
141 |
|
|
17.4.1 |
Mill Balls |
141 |
|
|
17.4.2 |
Cyanide |
141 |
|
|
17.4.3 |
Caustic |
141 |
|
|
17.4.4 |
Lime |
142 |
|
|
17.4.5 |
Sodium Metabisulphite |
142 |
|
|
17.4.6 |
Copper Sulphate |
143 |
|
|
17.4.7 |
Plant Diesel |
143 |
|
|
17.4.8 |
Carbon |
143 |
|
|
17.4.9 |
Hydrochloric Acid |
144 |
|
|
17.4.10 |
Air Services |
144 |
|
|
17.4.11 |
Water services |
144 |
18 |
PROJECT
INFRASTRUCTURE |
146 |
|
18.1 |
Infrastructure |
146 |
|
|
18.1.1 |
Roads |
146 |
|
|
18.1.2 |
Process Plant Buildings |
146 |
|
|
18.1.3 |
Process Plant Warehousing and
Workshop |
146 |
|
|
18.1.4 |
Process Plant Ancillary Infrastructure |
146 |
|
18.2 |
Accommodation |
147 |
|
18.3 |
Security |
147 |
|
18.4 |
Power requirements
|
148 |
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18.5 |
Communications |
148 |
|
18.6 |
Sewage collection and
treatment |
148 |
|
18.7 |
Fuel and lubricant storage and
distribution |
149 |
|
18.8 |
Tailings Management
Facility |
149 |
|
|
18.8.1 |
Introduction |
149 |
|
|
18.8.2 |
Rainfall |
149 |
|
|
18.8.3 |
Production rates and design life |
150 |
|
|
18.8.4 |
Particle size distribution and
specific gravity |
150 |
|
|
18.8.5 |
Slurry characteristics |
151 |
|
|
18.8.6 |
Survey information |
152 |
|
|
18.8.7 |
Climate data summary |
152 |
|
|
18.8.8 |
Rate of rise of tailings |
152 |
|
|
18.8.9 |
TMF construction works |
153 |
|
|
18.8.10 |
Conclusions |
154 |
19 |
MARKET STUDIES AND CONTRACTS
|
155 |
|
19.1 |
Gold Price Trend |
155 |
20 |
ENVIRONMENTAL STUDIES,
PERMITTING AND SOCIAL OR COMMUNITY IMPACT |
156 |
|
20.1 |
Environmental and
Social Impact Assessment status |
156 |
|
20.2 |
Key risks to the project |
157 |
|
|
20.2.1 |
Social conflict arising from
resettlement |
157 |
|
|
20.2.2 |
Twangiza North Resettlement |
157 |
|
|
20.2.3 |
New TMF Resettlement |
158 |
|
|
20.2.4 |
General Resettlement Considerations |
158 |
|
20.3 |
Local political
instability |
159 |
|
20.4 |
Biophysical aspects |
159 |
|
20.5 |
Environmental
monitoring |
159 |
21 |
CAPITAL AND OPERATING COST |
161 |
|
21.1 |
Operating Cost
Estimate |
161 |
|
|
21.1.1 |
Summary |
161 |
|
|
21.1.2 |
Diesel and Power |
161 |
|
|
21.1.3 |
Sundry and Expenses |
161 |
|
|
21.1.4 |
Payroll |
162 |
|
|
21.1.5 |
Processing Materials and Contracts |
162 |
|
|
21.1.6 |
Mining Materials and Contracts
|
162 |
|
|
21.1.7 |
Basis of Estimates |
163 |
|
21.2 |
Capital Cost Estimate
|
164 |
|
21.3 |
Introduction |
164 |
|
|
21.3.1 |
Basis of Estimates |
165 |
|
21.4 |
Escalation |
166 |
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21.5 |
Taxes |
166 |
22 |
ECONOMIC ANALYSIS |
167 |
|
22.1 |
Economic Analysis |
167 |
|
22.2 |
Model Assumptions |
168 |
|
22.3 |
Sensitivity Analysis
|
168 |
|
|
22.3.1 |
Project life cash-flow |
170 |
23 |
ADJACENT
PROPERTIES |
172 |
|
23.1 |
Adjacent Properties |
172 |
24 |
OTHER RELEVANT
DATA AND INFORMATION |
173 |
25 |
INTERPRETATION AND CONCLUSIONS
|
174 |
|
25.1 |
Geology |
174 |
|
25.2 |
Resource Model |
174 |
|
25.3 |
Metallurgy and
Processing |
174 |
|
|
25.3.1 |
Historical Plant Performance |
174 |
|
|
25.3.2 |
Recent Plant Modifications |
175 |
|
25.4 |
Mine Plan |
175 |
|
25.5 |
Operating &
Capital Costs |
175 |
|
25.6 |
Infrastructure |
177 |
|
25.7 |
Environmental |
177 |
|
25.8 |
Financials |
177 |
26 |
RECOMMENDATIONS |
178 |
|
26.1 |
Geology |
178 |
|
26.2 |
Resource model |
178 |
|
|
26.2.1 |
Density |
178 |
|
|
26.2.2 |
Reconciliation |
178 |
|
|
26.2.3 |
Grade Control Modelling |
178 |
|
26.3 |
Metallurgy and
Processing |
178 |
|
|
26.3.1 |
Measuring and Reporting Production |
178 |
|
|
26.3.2 |
Technical Improvements |
179 |
|
|
26.3.3 |
Fund Fully to Maximise Performance |
180 |
|
|
26.3.4 |
Additional Testwork |
180 |
|
|
26.3.5 |
Ore Type Definition and Management |
181 |
|
26.4 |
Mine Plan |
181 |
|
26.5 |
Operating & Capital Costs |
181 |
|
26.6 |
Infrastructure |
182 |
|
26.7 |
Environmental |
182 |
27 |
REFERENCES |
183 |
28 |
DATE AND SIGNATURE PAGE |
184 |
29 |
CERTIFICATES OF
QUALIFIED PERSONS |
185
|
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List of Tables
Table 5-1: |
Monthly Rainfall (mm) For
Twangiza (October 2006 to December 2014) |
27 |
Table 10-1: |
Drilling Summary |
37 |
Table 11-1: |
Statistics of Results of
Standard Certified Sample Submissions |
45 |
Table 11-2: |
Coarse Split Sample Pairs pre FS |
46 |
Table 11-3: |
Coarse Split Sample Pairs Post
FS |
46 |
Table 12-1: |
Data Type and Amount of Data Employed In
Current Geological Modelling |
49 |
Table 12-2: |
Adit sampling - statistics of
composites |
50 |
Table 13-1: |
Ore Characteristics |
54 |
Table 13-2 |
Predicted Plant GRG Recoveries
|
60 |
Table 13-3: |
Summary of Cyanide Destruction Tests |
62 |
Table 13-4: |
Summary of Bond Work Index
Tests |
62 |
Table 13-5: |
Summary of AMCT And JKTech Drop Weight Tests
|
62 |
Table 13-6: |
Liner and Ball Consumptions |
64 |
Table 13-7: |
Main Oxide Variability Results |
65 |
Table 13-8: |
North Oxide Variability Tests
|
65 |
Table 13-9: |
Main Transitional and Fresh Variability Results
|
66 |
Table 13-10: |
North Transitional and Fresh
Variability Results |
66 |
Table 13-11: |
Head Assay Carbon and Sulphur Speciation |
67 |
Table 13-12: |
Full Elemental Analysis |
68 |
Table 13-13: |
Grind Optimisation Results Summary - Fresh |
69 |
Table 13-14: |
Grind Optimisation Results
Summary - Transition |
69 |
Table 13-15: |
Reagent Optimisation at Natural pH |
70 |
Table 13-16: |
Reagent Optimisation at Acidic
pH |
70 |
Table 13-17: |
Reagent Optimization by Varying Collector Type
|
70 |
Table 13-18: |
Rougher Rate Tests - Fresh |
71 |
Table 13-19: |
Rougher Rate Tests - Transitional |
71 |
Table 13-20: |
Bulk Concentration Results |
71 |
Table 14-1: |
Feasibility Study Dry Solid Densities |
75 |
Table 14-2: |
Pit Density Determinations |
76 |
Table 14-3: |
Summary of 2015 Revised Dry Density Values |
78 |
Table 14-4: |
Summary of Raw Statistics per
Sampling Phase |
79 |
Table 14-5: |
Summary of Kriging Zones Used in the Latest
Block Model |
83 |
Table 14-6: |
Details of Block Model
Dimensions For Geological Model |
83 |
Table 14-7: |
Summary of Fields Used For Flagging Different
Geological Properties |
84 |
Table 14-8: |
Rock Codes Used in Twangiza
Main |
84 |
Table 14-9: |
Rock Codes Used in Twangiza North |
84 |
Table 14-10: |
Summary Statistics of 2m
Composites |
86 |
Table 14-11: |
High-Grade Capping |
88 |
Table 14-12: |
Back-Transformed Gaussian
Variogram Parameters |
91 |
Table 14-13: |
Search Radius For Pass 1 |
92 |
Table 14-14: |
Valley Fill Block Model
Parameters |
95 |
Table 14-15: |
Mineral Resource Estimate as at December 31,
2014 |
96 |
Table 14-16: |
Mineral Resource Estimate by
Pit, Material Type and Confidence Category as at December 31, 2014 |
97 |
Table 15-1: |
Mining Cost Summary |
99 |
Table 15-2: |
Government Royalties, Refining
and Selling Costs Summary |
100 |
Table 15-3: |
Recoverable Gold Cut-Off Grade by Deposit |
100 |
Table 15-4: |
Whittle Parameters for Open Pit
Optimization |
101 |
Table 15-5: |
Comparison of Inventory Between EOY2014 Optimum
Pit and 2014 Practical Pit Design |
106 |
Table 15-6: |
Stockpile Closing Balance as at
December 31, 2014 |
113 |
Table 15-7: |
Annual Mine Production Schedule by Material
Type |
114 |
Table 15-8: |
Mine Production Schedule
Summary |
117 |
Table 15-9: |
Resource to Reserve Modifying Factors |
118 |
Table 15-10: |
Summary of Twangiza Mineral
Reserves as at December 31, 2014 |
119 |
Table 15-11: |
Mineral Reserve by Pit as at December 31, 2014
|
120 |
Table 16-1: |
Mining Work Schedule |
124 |
Table 16-2: |
Labour Schedule |
125 |
Table 16-3: |
Mining Equipment Asset Table
|
126
|
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Table 16-4: |
Mining Equipment Schedule Matched to Production
Requirements |
127 |
Table 16-5: |
Replacement Mining Equipment
Schedule |
127 |
Table 18-1: |
24 Hour Design Storm Event Estimates |
149 |
Table 18-2: |
Particle Size Distribution and
Sieve Analysis |
150 |
Table 18-3: |
Tailings Parameters |
151 |
Table 20-1: |
Summary of Key Legislation and
Relevant Compliance |
157 |
Table 21-1: |
Summary of Unit Operating Costs |
163 |
Table 21-2: |
Twangiza Capital Cost Summary
|
165 |
Table 22-1: |
Financial Analysis Summary |
167 |
Table 22-2: |
Financial Model Assumptions |
168 |
Table 22-3: |
Sensitivity Analysis on Gold Price |
168 |
Table 22-4: |
Cash Flow Summary |
170 |
List of Figures
Figure 4-1: |
Africa Locality Plan |
22 |
Figure 4-2: |
Location of the Property in Democratic Republic
of the Congo |
23 |
Figure 4-3: |
Map of the Twangiza Property
|
24 |
Figure 7-1: |
Regional Setting of The Kibara Belt |
29 |
Figure 7-2: |
Magnetic Image Showing
Litho-Structural Domains Of The Twangiza Concession |
30 |
Figure 8-1: |
Deposit Geology |
32 |
Figure 10-1: |
Plan of Resource Drillhole
Locations |
38 |
Figure 11-1: |
Summary of Returned Assays per Standard |
44 |
Figure 11-2: |
Original vs. Coarse duplicate
splits - all samples |
47 |
Figure 11-3: |
Summary of All Blank Submissions to SGS Mwanza
|
47 |
Figure 12-1: |
Comparison of Acme Vancouver
vs. MGL Kamituga Laboratories |
50 |
Figure 13-1: |
Effect of Grind on Recovery |
69 |
Figure 14-1: |
Cross-Section Showing Digitised
Interpretation of Oxide Contacts & Lithology |
80 |
Figure 14-2: |
3D Screenshot of Porphyry Wireframe at Twangiza
Main |
81 |
Figure 14-3: |
3D Screenshot showing
cross-section of porphyry wireframe |
81 |
Figure 14-4: |
3D Screenshot 0.3 g/t Au Leapfrog Iso-Surface
|
82 |
Figure 14-5: |
Histogram of Sample Lengths In
Twangiza Database |
85 |
Figure 14-6: |
Comparative Histograms Per Oxidation Domain |
87 |
Figure 14-7: |
Example Plot of Cumulative Mean
and Covariance |
88 |
Figure 14-8: |
Summary of Gaussian Transformed Variograms For
Lower Oxide Zone |
90 |
Figure 14-9: |
Typical Section Displaying
Classification |
94 |
Figure 14-10: |
Plan Showing Valley Fill and Twangiza Main |
95 |
Figure 15-1: |
Whittle Optimization Results
Twangiza Central Deposit December 2014 |
102 |
Figure 15-2: |
Twangiza Central Practical Pit Design Versus
Blocks Within Whittle Pit Number 38 |
102 |
Figure 15-3: |
Whittle Optimization Results
Twangiza Main Deposit December 2014 |
103 |
Figure 15-4: |
Twangiza Main Interim Practical Pit Design
Versus Blocks Within Whittle Pit Number 12 |
103 |
Figure 15-5: |
Whittle Optimization Results
Twangiza North Deposit December 2014 |
104 |
Figure 15-6: |
Twangiza North Practical Pit Design Versus
Blocks Within Whittle Pit Number 29 |
104 |
Figure 15-7: |
Whittle Optimization Results
Twangiza West Deposit December 2014 |
105 |
Figure 15-8: |
Twangiza West Practical Pit Design Versus
Blocks Within Whittle Pit Number 41 |
105 |
Figure 15-9: |
Optimized Pit Shells Selected
For Practical Design |
107 |
Figure 15-10: |
Twangiza Main Intermediate Pit (Cut 1) Plan
View |
109 |
Figure 15-11: |
Twangiza Main Intermediate Pit
(Cut 2) Plan View |
109 |
Figure 15-12: |
Twangiza North Pit Plan View |
110 |
Figure 15-13: |
Twangiza Central Pit Plan
View |
110 |
Figure 15-14: |
Twangiza West Final Pit Plan View |
111 |
Figure 15-15: |
Twangiza Valley-Fill Final Pit
Plan View |
111 |
Figure 15-16: |
Twangiza Final Open Pits Plan View |
112 |
Figure 16-1: |
Site Layout Plan |
121 |
Figure 18-1: |
TMF Stage Capacity Curves |
153 |
Figure 19-1: |
Gold Price In USD Trend 2000
November 2014 |
155 |
Figure 20-1: |
Plan Showing Twangiza Water Monitoring
Positions |
160 |
Figure 22-1: |
Sensitivity Analysis of Gold
Price Versus Net Present Value |
169 |
Figure 22-2: |
Sensitivity Analysis of Total
Operating and Capital Cost Versus Net Present Value 169
|
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SRK Consulting (UK)
Limited |
5th Floor Churchill House
|
17 Churchill Way |
City and County of Cardiff
|
CF10 2HH, Wales |
United Kingdom |
E-mail:
enquiries@srk.co.uk |
URL: www.srk.co.uk |
Tel: +
44 (0) 2920 348 150 |
Fax: +
44 (0) 2920 348 199 |
NI 43-101 TECHNICAL REPORT, MINERAL RESOURCE AND
|
RESERVE UPDATE, DECEMBER 31 2014, TWANGIZA GOLD
MINE, |
DEMOCRATIC REPUBLIC OF THE CONGO
|
SRK Consulting (UK) Limited, (SRK
(UK)) was commissioned by Twangiza Mining SA (Twangiza Mining) to independently
review its 31 December 2014 Mineral Reserve estimate. SRK helped review over
three years of production data in a detailed reconciliation study which has been
used to inform the forecast production schedule and to assess the ability of the
plant to process non-oxide material.
This report presents a compilation of
information prepared by Twangiza Mining and its consultants and SRKs opinions
on core aspects of the Mineral Reserve. Additional details for background
information can be found in the "Updated Feasibility Study NI43-101 Technical
Report, Twangiza Gold Project, South Kivu Province, Democratic Republic of Congo
(2009 Feasibility Study); a copy of which can be obtained from SEDAR at
www.sedar.com.
The Qualified Persons (within the
meaning of National Instrument 43-101 (NI43-101)) for the purposes of this
report are Martin Pittuck and David Pattinson of SRK (UK) and who are
independent of the project, and Daniel Bansah of Twangiza Mining and Banro. All
the Qualified Persons visited the Twangiza mine site between 12th and
15th March 2015.
Twangiza Mining has warranted in
writing that it has openly provided all material information to SRK (UK), which,
to the best of its knowledge and understanding, is complete, accurate and true,
having made due enquiry. SRK (UK) is not aware of any current or pending
litigation or liabilities attached to Twangiza Mining.
3 |
RELIANCE ON OTHER EXPERTS |
SRK UK has reviewed production data and
mine planning information provided by Twangiza Mining. The environmental, fiscal
and legal data has been provided directly by Twangiza Mining and SRK UK has
received verbal assurance from Twangiza Mining that none of these aspects
presents an impediment to achieving the production outlined in this report.
The modified Tailings Management
Facility (TMF) plan was provided by Daniel Bansah of Twangiza Mining; the plan
is derived from previous work completed by Metago (now SLR Consulting) on behalf
of Twangiza Mining.
4 |
PROPERTY DESCRIPTION AND
LOCATION |
The property is located in the South
Kivu Province of the Democratic Republic of the Congo (DRC), centred at
approximately 2º52 South and 28º45 East, roughly 35 kilometres west of the
Burundi Border and 45 kilometres south-south-west of Bukavu.
|
Registered Address: 21 Gold Tops,
City and County of Newport, NP20 4PG, |
Group
Offices: Africa |
Wales, United Kingdom. |
Asia |
SRK Consulting (UK) Limited Reg No
01575403 (England and Wales) |
Australia |
|
Europe |
|
North America |
|
South America |
SRK Consulting |
NI 43-101 Twangiza Main Report |
In April 2002, the Government of the
DRC formally signed an agreement which entitled Twangiza Mining to hold a 100%
interest in the Twangiza Property under a revived mining convention which
expires in March 2027 (subject to extension under the new DRC Mining Code).
The exploitation permits give Twangiza
Mining exclusive rights to carry out exploration, development, construction and
exploitation works within the perimeter over which they have been granted.
The six exploitation permits, or
Certificat / Permis dExploitation (PE), covering a total area of 1,156 square
kilometres define the Twangiza Property for which Twangiza Mining has exclusive
mining rights; these are listed below and shown in Figure 4-3. The exploitation
permits are 100% owned by Twangiza Mining, a subsidiary of Banro Corporation.
|
PE40 Concession No 92 |
|
PE41 Concession No 91 |
|
PE42 Concession No 90 |
|
PE43 Concession No 89 |
|
PE44 Concession No 88 |
|
PE68 Concession No 66 |
The property boundaries are located by
co-ordinates provided with each exploitation permit. Twangiza Mining has had no
need to physically beacon the boundaries as the mineralized zone and areas
affected by the mining activities are located well within the boundaries defined
by PE42.
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4.2 |
Natural and Existing
Features |
Details of the known mineralized zones,
mineral resources and mineral reserves are included within Section 7 of this
report.
During the colonial era there was small
scale hard-rock mining for tin, alluvial gold and tungsten diggings. However,
these sites (all outside Twangiza Minings mining area) have now become
over-run by artisanal mining activity and there is little evidence of the
earlier activity.
With the exception of informal
artisanal activity there are no current or historical formal mine workings,
tailings ponds or waste dumps located on the Twangiza Property other than those
belonging to the operation described in this report
4.3 |
Royalties and Other
Payments |
A royalty calculated at 1.0% of revenue
is paid to the DRC government.
4.4 |
Environmental Liabilities |
Twangiza Mining will be liable to the
DRC government for any damage to the environment resulting from a breach of the
requirements of the DRC Mining Code, approved environmental impact statement
(EIS) or associated environmental management plan of the project (EMPP).
More information regarding the status
of the EIS and EMPP is included in Section 20 of this report.
On the basis of limited monitoring data
and observations in the field, the following potential liabilities accruing from
past mining (artisanal mining) and prospecting activities on site may occur:
Landslides due to destabilization of
slopes;
Twangiza river: denuded deposits of
tailings along 3-4 km stretch of riverbed, may be entrained by floodwaters,
leading to high suspended sediment loads in downstream river reaches and
resulting in negative impacts on aquatic biota and human consumption;
Mercury used by artisanal miners may be
adsorbed in sediments and may be mobilized and transported in river water should
acid rock drainage occur. However, this proposition has not been tested and
proven. Acid rock drainage potential is indicated in preliminary tests, but has
not been confirmed.
4.5 |
Required Permits and
Approvals |
The permits and approvals required to
conduct the work at the Twangiza property are specified in the DRC Mining Code.
Such permits and approvals have been obtained.
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5 |
ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,
INFRASTRUCTURE AND PHYSIOGRAPHY |
5.1 |
Topography, Elevation, Vegetation and Seismic
Activity |
Twangiza is situated in a mountainous
area with deeply incised valleys with slopes typically greater than
30o, forming a dendritic drainage pattern. The mining area occupies a
steep ridge running north/south between two fast-flowing rivers, which join just
to the north of the mine. Elevation in the area ranges from 1500 m to 2400 m
above sea level.
Vegetation on the Twangiza property is
a mosaic of transformed, agricultural plots and woodlots of cypress and
eucalyptus, and montane grassland. One small, 2.18 ha patch of indigenous forest
remains to the east of the mine area, the Lusirwe sacred forest.
Due to its location within the western
arm of the Rift Valley system, the property is subject to seismic activity.
Detailed discussion of the seismic hazard potential within the Twangiza area is
included the 2009 Feasibility Study.
Twangiza Minings offices are in
Bukavu, the capital city of South Kivu Province some 45 kilometres
north-northeast of the Twangiza Property. Bukavu has an airport, Kavumu, however
access by road from Rwanda is the current preferred route for international
access.
Road access from Bukavu to the Twangiza
Property is possible by travelling some 55km on the recently upgraded N2
National Road and then 30km on the recently widened and upgraded Twangiza access
road. The journey time is 2.5 hours during the dry season and extends to 4 hours
under wet conditions. The property is also serviced by a helicopter and the
journey between Bukavu and Twangiza is some fourteen minutes.
The climate at Twangiza can be
classified as tropical to sub-tropical; the wet season falls between September
and April and the main dry season is from May to August. Due to its close
proximity to the equator, Twangiza experiences daylight and night hours that are
almost equal, with daylight lasting between 6am and 6pm. The relative humidity
generally exceeds 85 % and the mine is often in cloud.
Twangiza has an average annual rainfall
of 1,796 mm. Rain generally occurs as soft, lengthy rainfall in the mid to late
afternoons, but violent thunderstorms are also frequent. For the period October
2006 to December 2014 the highest monthly rainfall recorded for Twangiza was
357.5 mm in December 2011 and the lowest total monthly rainfall was in March
2008 (see Table 5-1).
The on-site weather station at Twangiza
recorded data for the period October 2006 to December, 2014. The average
temperature measured on-site is 18°C with a maximum temperature of 27.7°C,
measured during March 2007, and a minimum temperature of 8.6°C, measured during
July 2008.
The prevailing north-easterly wind
direction for Twangiza is relatively consistent throughout the year. The average
annual wind speed for all hours is 4.33 m/s.
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Table 5-1: |
Monthly Rainfall (mm) For Twangiza (October
2006 to December 2014) |
YEAR |
J |
F |
M |
A |
M |
J |
J |
A |
S |
O |
N |
D |
2006 |
- |
- |
- |
- |
- |
- |
- |
- |
- |
99* |
238 |
238 |
2007 |
275 |
118 |
109 |
144 |
40 |
134.6 |
22.4 |
24.4 |
111 |
37.2* |
177.4 |
203.2 |
2008 |
164.2 |
140 |
4.6* |
86.2 |
61.8 |
53.8 |
27.6 |
12.2 |
29.2 |
103* |
- |
- |
2009 |
- |
- |
- |
- |
63.6 |
14.8 |
0.4 |
12.4 |
112 |
73.4 |
162.2 |
- |
2010 |
- |
202.6 |
207.1 |
73 |
122.4 |
18.8 |
5.2 |
7.2 |
98 |
124.4 |
154.8 |
112.8 |
2011 |
299.4 |
87.1 |
262.9 |
48.5 |
59.7 |
92.8 |
36.3 |
21.6 |
82.7 |
107.4 |
264.5 |
357.5 |
2012 |
151.5 |
177 |
137.6 |
141 |
81.3 |
13.7 |
34 |
130 |
93 |
161 |
79.4 |
151 |
2013 |
197.9 |
123.6 |
282.4 |
141.6 |
11.8 |
0.2 |
0 |
39.5 |
213.8 |
88.4 |
248.7 |
215 |
2014 |
257.3 |
208.7 |
223.6 |
84.9 |
10.9 |
50.2 |
0 |
28.2 |
85.7 |
196.7 |
187.6 |
154 |
* Incomplete data set
-
No data
5.4 |
Surface Rights and Available Local
Infrastructure |
Twangiza Mining has mining rights to
the Twangiza mine. The Twangiza Property is remotely located and there is no
existing supply of power suitable for the Project requirements. A
diesel-generator power plant has been established to provide the power required
to the Twangiza Project. Twangiza Mining plans to eventually take power from a
hydroelectric scheme due to come on line in the next few years.
The topography in the area is
challenging. Water containment dams and a tailings disposal facilities can be
placed in nearby valleys of limited dimensions or in wider river plains within
several km of the mine.
The local workforce consists primarily
of subsistence farmers and artisanal miners. In addition the communities contain
a number of skilled workers (i.e. artisans - builders, carpenters, electricians,
plumbers etc).
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The earliest recorded work programs
within the Twangiza Property consisted of alluvial mining for tin and gold as
early as the 1930s along prominent rivers and creeks by Mines des Grandes Lacs
(MGL). MGL began exploration for in-situ resources in 1957 and followed alluvial
gold deposits upstream from the Mwana River to the present day Twangiza deposit.
MGL tested the Twangiza deposit through
8,200 metres of trenching and 12,100 metres of adits (20 metre by 20 metre grid)
on seven levels (Levels 2100 to Level 2220).
In 1974 to 1976, Charter Consolidated
Limited (Charter) undertook an evaluation program of the Twangiza area in
order to verify the results obtained by MGL and to look for possible extensions
of the deposit. Work also included metallurgical studies.
From 1982 to 1984, SOMINKI undertook a
feasibility study which was completed by ABAY, a Belgian consulting company and
in 1988 the Northern Queensland Company assessed the deposit and generated some
financial models. A report was prepared by Billiton in 1989 for SOMINKI and
submitted to the Ministry of Planning for tax exoneration purposes.
In January 1996, Banro Resource
Corporations (now Banro Corporation) wholly owned subsidiary, African Mineral
Resources Inc. (AMRI), in conjunction with its joint venture partner Mines
D'Or du Zaire (MDDZ), completed the purchase of the outstanding privately held
shares of SOMINKI. The joint venture partners controlled 72% (AMRI - 36%, MDDZ -
36%) of SOMINKI, with the remaining 28% held by the Government of Zaire (DRC).
Banro subsequently acquired MDDZs 36% interest in SOMINKI in December 1996.
In early 1997, Banro, SOMINKI and the
government of the DRC ratified a new 30 year mining convention that provided for
SOMINKI to transfer its gold assets to a newly created company. Société Aurifère
du Kivu et du Maniema, SARL (SAKIMA) was incorporated to acquire the assets of
SOMINKI as stipulated in the new mining convention. Banro consolidated the
information and from August 15, 1997, to April 15, 1998, undertook a field
exploration program managed by CME and Company (CME).
In addition to this asset transfer, the
new mining convention included a ten year tax moratorium from the start of
commercial production, the ability to export all gold production, the ability to
operate in US currency, the elimination of import duties and title confirmation
for all of the concessions. The new mining convention provided for Banro to
control 93% of SAKIMA with the remaining 7% held by the Government of the DRC as
a net carried interest.
In July 1998, President Laurent D.
Kabila issued presidential decrees which, amongst other things, effectively
expropriated SAKIMAs gold assets. Banro initiated arbitration proceedings
against the Government of the DRC seeking compensation for the expropriation of
the assets.
In April 2002, the Government of the
DRC formally signed a settlement agreement with Banro. The agreement called for,
among other things, Banro to hold a 100% interest in the Twangiza Property under
a revived mining convention which expires in March 2027 (subject to extension
under the new DRC Mining Code).
Additional information regarding the
Twangiza Property with respect to history is set out in the 2009 Feasibility
Study.
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7 |
GEOLOGICAL SETTING AND MINERALISATION |
|
|
7.1 |
Regional Geology |
The Twangiza Property is located in the
northern half of the Great Lakes sub-province of High Africa, one of the worlds
principal Precambrian orogenic-metallogenic provinces, see Figure 7-1. Twangiza
lies within the Kibara Belt, a Proterozoic intracontinental mobile belt situated
between the Congo Craton in the west and the Tanzanian Craton in the east. The
belt trends in a NNE-SSW direction for over 2,000 km from Katanga to Lake
Victoria, and attains its maximum width of about 500 km slightly to the north of
the Twangiza-Namoya area.
The belt has a long and complex
evolution, stretching from the Palaeoproterozoic prior to the Eburnean orogeny,
through to the Neoproterozoic and the Pan African event. The belt is dominated
by clastic sedimentary rocks with minor carbonates and volcanics, which have
been intruded by granitoids, mafics and alkaline complexes.
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The concessions-scale geology can be
divided into three distinct litho-structural terrains as illustrated in Figure
7-2. The eastern terrain is characterised by folded, broadly N-S trending
Neoproterozoic sediments, which are part of the Itombwe synclinorium, a
regional-scale fold which extends southwards from the Twangiza area for about
150 km. The western domain has a distinct NW-SE tectonic grain, and is believed
to be Palaeoproterozoic in age. The third domain occurs in the north, where
recent basalts blanket the Proterozoic rocks.
The sediments in the Neoproterozoic
terrain are weakly metamorphosed. The dominant lithology is mudstone, often with
a significant amount of carbonaceous material; units of siltstone are commonly
interbedded with the mudstone. Quartz wacke and sandstone occur as thin beds or
lenses. A characteristic feature of the Twangiza area is the presence of a
conglomerate consisting of clasts of granite, mudstone and siltstone supported
by a matrix of dark grey silty mud. It frequently contains a significant amount
of detrital magnetite, and forms the relatively highly magnetic unit that
clearly defines the geometry of the concession-scale folds in the magnetic
images.
In the vicinity of the Twangiza Mine,
the Neoproterozoic sediments have been intruded by porphyritic sills, ranging in
thickness from less than 1 m to over 50 m. The sills have undergone extensive
hydrothermal alteration and are thought to be part of a suite of alkaline
intrusives emplaced at about 750 Ma.
The Neoproterozoic terrain at Twangiza
is characterised by a series of N-S trending, concession-scale folds, which
plunge to the north. These folds vary from being open to almost isoclinal,
although the average limb dips are usually between 50° and 80°. Smaller-scale
folds, probably parasitic to the larger structures, are commonly seen on a
prospect scale; they display plunges to the north and south, or are
doubly-plunging like the fold hosting the Twangiza deposits. The folding is
considered to have developed in response to E-W compression in the Pan African
orogeny at about 550 Ma.
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Faulting in the Neoproterozoic terrain
is common, the main trends being NE-SW to E-W. In addition, zones of shearing
and/or brecciation have been mapped sub-parallel to the fold axes at several
prospects, and may have had a control on the mineralisation.
The rocks in the area have been
affected by tropical weathering resulting in degradation of the original rock
fabric down to a depth 100m; in the deposit area where most data is available
the weathering is particularly intense in the top 30m.
The Twangiza Main ore body consists of
a wide (up to 200 m) zone of pervasively altered mudstone, siltstone and
porphyry sills, with abundant sulphidic veins. The veins form a complex
irregular network, although veining parallel to bedding is relatively common.
Hydrothermal fluids have exploited both the fracture system which developed
during folding due to competency contrasts between the lithologies, and
dilational zones between bedding planes to form saddle reefs.
The style of mineralisation in the
sediments and sills varies, but can be sub-divided into two main types as
discussed below:
Mineralisation in the sills is
characterised by the presence of pyrite and arsenopyrite (approximately 65%
pyrite: 35% arsenopyrite). The total abundance of sulphide is variable,
averaging about 3% of the rock, but locally comprising up to about 30% of a 1 m
sample. There is a positive correlation between grade and sulphide content. The
sulphides occur in a variety of habits: (a) disseminated crystals, (b)
stringers, (c) coarsely crystalline veins up to 10 cm in width, but usually 1
3 cm across, and often with intergrown quartz, and (d) irregular massive
patches. Further details can be found in Section 13.4.1.
The quantity of sulphides disseminated
in the mineralised sediments is generally lower, they are generally finer
grained and are more common in the relatively porous siltstone units. The
sulphide veins in the sediments generally contain more quartz, either intergrown
with the pyrite and arsenopyrite, or forming borders to the veins.
In the oxidised zone, the veins in both
the porphyry and sediments have weathered to limonite-silica intergrowths. This
limonite-silica veining is a common feature of the mineralisation in outcrop.
Limonite-filled boxworks, and irregular limonite patches and coated vugs have
formed due to oxidation of the disseminated sulphides and patches.
Hydrothermal alteration associated with
the gold mineralisation has formed three broad assemblages.
Additional information regarding the
Twangiza property with respect to geological setting and mineralization is set
out in the 2009 Feasibility Study.
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The spatial association of gold
deposits with Sn-W mineralisation in the Kibara belt has led to the suggestion
that the gold bearing fluids were also related to the intrusion of the
tin-bearing G4 granites. However, the rocks which host the Twangiza deposit are
considered to be Neoproterozoic in age, and therefore post-date the G4 granites
(c. 975 Ma).
It is proposed that the Twangiza
mineralisation (and possibly the other gold deposits in the Kibara belt) are
rather related to fluids derived from the devolatisation of the lower crust
during the Pan African orogeny at about 550 520 Ma.
The Pan African orogeny at about 550
Ma, involved E-W compression leading to deformation of the Neoproterozoic
sediments and sills into N-S trending folds. Auriferous fluids were focused into
these structural traps; at Twangiza Main, the most important traps were the
low-pressure hinge zones of anticlinal folds, see Figure 8-1.
The feeder structures were probably
sub-vertical, limb-parallel structures or limb shears. Mineralisation would be
expected to deteriorate down the fold limbs away from the fold closure and away
from the sill zone where the stratigraphy is relatively homogeneous. Twangiza
North hosts deeper, stratabound, mineralisation where the feeder structures have
intersected relatively reactive, internally fractured sills, resulting in a
sharper mineralised boundary.
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Exploration on the Twangiza Property
has been divided into historical exploration and the recent work carried out
from October 2005 to December 2014. The recent exploration is subdivided into
five phases: October 2005 December 2006, January 2007 November 2008, January
2009 December 2010, January 2011 December 2012 and January 2013 December
2014.
9.1 |
Historical Exploration |
There have been three major field
exploration programs on the Twangiza Property prior to 2005.
The first was between 1957 and 1966 by
MGL and consisted of the driving of approximately 12,100 metres of adits and
8,200 metres of trenches at the Twangiza deposit. A total of 17,400 channel
samples were collected at two metre intervals from both the trenches and adits.
Secondly, from 1974 to 1976, Charter
Consolidated Limited undertook an evaluation program of the Twangiza area in
order to verify the results obtained by MGL and to look for possible extensions
to the mineralisation. Soil sampling was conducted over a 4.6 square kilometre
area to the north of the Twangiza deposit. Anomalous soil samples were tested by
11 pits, 6 trenches and 5 adits. Work also included the re-sampling of three MGL
adits (Levels 2100, 2130, and 2220).
The third historical program was
undertaken by Banro between August 15, 1997 and April 15, 1998. The program was
managed by CME and consisted of:
|
|
Topographical surveying, LANDSAT acquisition
and interpretation, Helicopter-supported airborne magnetic survey. |
|
|
|
|
|
Detailed geological mapping and rock sampling, grab
samples and channel samples from 16 adits were taken and after sample
preparation at Banros on site sample preparation laboratory were sent to
Acme Analytical Laboratory in Vancouver, Canada for analysis; |
|
|
|
|
|
Petrographic studies and density testing was
performed on behalf of Banro by CME |
9.2 |
Recent Exploration (October 2005 December
2006) |
Banro resumed its exploration programme
at Twangiza after the Congolese government had established control and authority
in the area in October 2005.
9.2.1 |
Soil geochemical programme |
A soil geochemical programme designed
to test the immediate northern, eastern, western and southern extensions of the
known Twangiza mineralisation was completed in 2006. The 7 km soil geochemical
grid had its base line orientated along the hinge of the anticline at 350º. Soil
sampling was undertaken at 40 m intervals on lines spaced at 80 m. The baseline
origin for the soil geochemical grid was pegged at UTM coordinate 9682698.2N /
693500.5E, which corresponds to a local grid coordinate of 10000N/20000E.
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9.2.2 |
Trenching programme |
A trenching programme was initiated to
test the gold-in-soil geochemical anomalies and the continuity of mineralisation
on the northern extension of the Twangiza Main deposit, as well as the southern
and northern extensions of the Lukungurhi workings. A total of 785 channel
samples were collected from 1.159 m of trenching.
9.2.3 |
Prospect scale mapping |
A detailed mapping project was carried
out in the Twangiza artisanal workings in order to gain a better understanding
of the geology and mineralisation controls and to verify and compliment the
diamond drilling data. The study reviewed all aspects of the geology including
lithology, structure and alteration.
9.3 |
Recent Exploration (January 2007 November
2008) |
|
|
9.3.1 |
Geophysical exploration |
Airborne magnetic and radiometric
surveys were completed over the entire Twangiza property, utilising a flight
line spacing of 100 m with tie lines at 1000 m intervals; several targets were
identified for follow-up work.
LIDAR to create accurate topographic
maps of the region. In addition, colour aerial digital photography has been
rectified to create accurate orthophotos.
9.3.3 |
Additional regional work |
During late 2007 Banro began
exploration of the Twangiza property outside the main trend. The following
targets were investigated:
Mufwa
Located 13 km northwest of the Twangiza
Main deposit, Mufwa is the focus of intense artisanal activity, where miners are
exploiting a series of quartz veins within mudstone.
Kaziba
Located 11 km east of the Twangiza Main
deposit, the Kaziba target was discovered towards the end of 2008.
Sulphide-associated, disseminated mineralisation similar in style to that at
Twangiza Main, occurs within mudstones and siltstones on the western limb of a
northerly plunging anticline. Exploration during 2008 consisted of mapping and
sampling of artisanal workings, and soil sampling of a 2 x 1 km area around the
workings on an 80 x 40 m grid. A total of 290 rock samples and 573 soil samples
were collected. The data indicates the presence of gold mineralisation over a
strike of 250 m, potentially with a thickness of up to 30m.
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Tshondo
Preliminary work at Tshondo, an old
colonial discovery located 9 km west of Twangiza Main, indicates that gold is
associated both with quartz veins and within the surrounding hydrothermally
silicified mudstone and siltstone. The mineralisation is associated with the
axis of a northerly plunging syncline. Artisanal mining is focusing on the
relatively high-grade quartz mineralisation over a strike of 500 m. Soil
geochemical sampling in tandem with rock chip sampling, trenching and adit
mapping and sampling was undertaken in late 2010. Analytical results are awaited
to warrant further work comprising diamond drilling.
Radiometric anomaly
This prospect is approximately 5 km
west of Twangiza Main, and was targeted due to the presence of a coincident
uranium-thorium radiometric anomaly similar to that at Twangiza, located on a
well-defined anticline axis. Work during 2008 comprised a programme of stream
sediment sampling (117 samples) and regional mapping. The stream results define
three anomalous areas with values of up to 1,840 ppb Au.
Southern anomaly
This target is located 10 km south of
Twangiza Main. It is situated on the axis of a tightly folded syncline, and is
associated with a coincident uranium-thorium radiometric anomaly. Historical
records indicate that alluvial gold was exploited in colonial times, and
alluvial artisanal mining is still carried out locally. The 2008 programme
comprised stream sediment sampling (185 samples) and regional mapping. Anomalous
values of up to 470 ppb Au will be followed up initially by soil sampling.
9.4 |
Recent Exploration (January 2009 December
2010) |
During this period, exploration focused
on mineralized structures flanking the Twangiza Main orebody and on geochemical
sampling to the west, north and east of the original 7 x 2 km grid. Regional
work continued at Kaziba and the Radiometric Anomaly, and also at Ntula, a new
discovery in the north of the Twangiza concession.
9.4.1 |
Twangiza West Zone |
|
|
|
This zone lies immediately to the west of the Twangiza
Main deposit. Work during 2009 included surface mapping and rock chip
sampling, and extensive auger drilling. |
|
|
9.4.2 |
Twangiza East Zone |
|
|
|
Mineralisation to the east of the Twangiza Main ore body
was tested by extensive auger drilling (221 holes, 866 m). |
|
|
9.4.3 |
Valley Fill |
|
|
|
An extensive pitting program was completed aimed at
estimating the gold resource of the Mwana River Valley Fill material, a
total of 44 pit were completed along 17 traverses. This was followed by 36
auger drill holes totalling of 119.15 m. |
|
|
9.4.4 |
Geochemical surveys |
|
|
|
Soil sampling, rock chip sampling and mapping coverage
was expanded. |
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Exploration continued at Kaziba,
Radiometric Anomaly and Ntula.
9.5 |
Recent Exploration (January 2011 December
2012) |
During this period, exploration focused
on infill drilling at Twangiza East and West flanks to provide a better
understanding of these deposits. Regional work focused on soil sampling, rock
chip sampling and mapping at Ntula, Luntukulu and Kaziba.
9.6 |
Recent Exploration (January 2013 December
2014) |
During the period 2013 2014, the
company reduced its exploration activities in the Twangiza concession to
grassroots activities focused on the Ntula Mufwa regional corridor and
prospects around Luntukulu area. In 2014, there was further reduced exploration
focussed on the Mufwa and Kadubo Prospects.
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Due to the extreme topography at
Twangiza, resource definition drilling was supported by an A-Star 350 B2
helicopter for moving drilling rigs, materials and personnel from site to site.
The majority of drilling has been diamond drilling focussing on Twangiza Main
and North deposits as summarised in Table 10-1 and shown in Figure 10-1.
|
Table 10-1: |
Drilling Summary
|
|
Diamond Drilling |
RC Drilling |
|
number |
length (m) |
number |
length (m) |
Twangiza North |
164 |
31,373 |
4 |
438
|
Twangiza Main |
169
|
40,427 |
15 |
2,936 |
Total |
333
|
71,800 |
19 |
3,374 |
The Mineral Resource model in this
report uses only the resource drilling; it has not been updated with the grade
control drilling undertaken since the mining started.
10.1 |
Drilling (September 1997 March
1998) |
A total of 20 diamond drilling holes
covering, 9,122 metres, 8,577 samples, HQ and NQ core size was completed between
September 4, 1997 and March 9, 1998. The drilling covered an 800 m strike length
of mineralisation within the hinge of the Twangiza Anticline, with holes drilled
at different orientations.
Drilling was performed by Rosond
International Limited of South Africa utilizing two Longyear 38 drill rigs with
a maximum depth potential of 600 m.
10.2 |
Drilling (February 2006 May
2008) |
Four drill rigs were deployed at the
Twangiza property, with two additional rigs mobilized in 2008.
A total of 17,037.34 metres of diamond
drilling involving 71 holes were completed between February and December, 2006.
In January 2007, a major drilling campaign commenced with the aim of converting
the remaining Inferred Mineral Resources at Twangiza Main into the Indicated and
Measured categories and to identify additional Inferred Mineral Resources
particularly at Twangiza North.
A total of 61 resource holes were
drilled at 40 m centres to infill the holes drilled in 1997/98 with the
objective of upgrading the Inferred Mineral Resources to the higher confidence
Measured and Indicated Resources. In addition, 39 exploration holes were drilled
to test the Twangiza North geochemical anomaly.
The remainder of the programme between
May 2007 and May 2008 consisted of a total of 24,231.3 metres of PQ, HQ and NQ
diamond drilling involving 116 holes. The focus of the drilling was to infill
the drilling grids and potential resources within the Twangiza North area of the
deposit. 98 holes were drilled at 40 m centres to infill the holes drilled in
2006 and early 2007 in the Twangiza North deposit and 18 holes were drilled in
Twangiza Main.
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The diamond drilling was performed by
Geosearch International Limited of South Africa utilizing four portable CS1000
and two Longyear 38 drill rigs with a maximum depth capability of 600 metres.
All drill hole collars were surveyed
with RTK GPS equipment. Drill hole collar azimuths and inclinations were
established at surface by using hand held compasses. Down-hole surveying of
drill holes utilized a Reflex Single Shot or Flexit instrument, which measures
both azimuth and inclination of the hole. All drill core was orientated.
Orientation was carried out by the "Spear" method or the Ezy Mark system. The
majority of the drill holes were drilled to the east on an azimuth of between 70
- 80°, and at inclinations of between 50 55°. A portion of the programme has
been drilled in the opposite direction on an azimuth of 260° to improve the
definition of the 3D wireframes.
10.3 |
Drilling (May 2008 November
2008) |
102 diamond drill holes totalling
21,952.26 metres of PQ, HQ and NQ were completed between May 2008 and November
2008. Drilling tested the mineralised interpretation at depth with the aim of
increasing confidence in the previous estimates, particularly in Twangiza Main.
Resource holes were drilled on 40 m centres to infill the holes drilled during
previous campaigns with the objective of upgrading the Inferred Mineral
Resources to the higher confidence Measured and Indicated Mineral Resources and
improving the estimation at depth. The same drilling procedures were used in
terms of rig set-up and rig movement as in the previous drilling campaign.
The majority of the holes in the
programme were drilled to the east on an azimuth of between 70 - 80° at dips of
between 50 55°. A portion of the programme has been drilled in the opposite
direction on an azimuth of 260° to improve the definition of the 3D
wireframes.
10.4 |
Drilling (December 2008 December
2012) |
138 diamond drill holes, 61 RC holes
and 1,515 Auger holes totalling 12,565.71 metres, 8,101 metres and 4,424.21
metres respectively were completed between December 2008 and December 2012.
Diamond drilling was mainly used for
infill drilling at Twangiza East (7 holes totalling 747.91metres) and Twangiza
West (28 holes totalling 2,638.45 metres) zones and also to further test the
Ntula (29 holes totalling 3,861.46 metres) and Lukugurhi (4 holes totalling
442.15 metres) mineralisation. Some 90 diamond drilling holes totalling 3,613.94
metres were also undertaken for geotechnical and electrical study of the
Tailings Dam and the plant site areas.
A total of 3 Reverse Circulation holes
(375 metres) had been completed in this zone by end of year 2010. RC and Auger
drilling undertaken mainly focused on exploration even through some RC holes
were drill for ground water monitoring.
The majority of the diamond and RC
drill holes in the programme were drilled to an azimuth of between 70 - 80° at
dips of between 50 55°. A portion of the programme has been drilled in the
opposite direction on an azimuth of 260° to improve the definition of the 3D
wireframes. All auger drill holes were drilled at a vertical angle as well as
diamond and RC holes drilled for geotechnical study and ground water monitoring.
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10.5 |
Drilling (January 2013 December
2014) |
Due to the significant scaling down in
exploration activities, there was no resource definition drilling undertaken in
this period. 20 auger holes totalling 43.6 m were drilled at Ntula in the
regional exploration area for target generation.
2 resource holes (183.9 m) were drilled
in the Twangiza main pit to test the extension of the main pit mineralisation in
2014. Also 9 geotechnical holes totalling 316 m were drilled for geotechnical
studies within the Twangiza main pit in 2014.
A summary of significant mineralised
intersections can be found in Appendix II of the 2009 Feasibility Study.
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11 |
SAMPLE PREPARATION, ANALYSES AND
SECURITY |
|
|
11.1 |
SAMPLING METHOD AND APPROACH |
|
|
11.1.1 |
Soil
Geochemistry |
Approximately 4 to 6 kg of soil was
taken from each sample site, below the upper soil horizon containing vegetative
matter. The average sample depth was 35 cm. Soil samples were collected and
placed in a plastic sample bag. A wet strength sample tag with a unique sample
number was placed in the bag.
Continuous channel samples were taken
from the floor of all trenches that intersected saprolite or weathered bedrock.
Each channel was approximately 10 centimetres wide and 5 centimetres deep.
Sample intervals were determined by geological features: in homogeneous rock,
the maximum sample interval was 1 metre, and the minimum sample interval
employed was 0.3 m. Veins, altered zones, or distinct geological units were
sampled so that the contacts were a standard 2 cm within the sample boundaries.
Sample weights varied between 3 and 6 kilograms.
11.1.3 |
Stream Sediment Samples |
The stream sediment sampling technique
employed by Twangiza Mining involved collecting the fine sediment fraction from
the top of the stream bed. The geologist took at least six 250 g sub-samples
from within 10 m upstream and downstream of the starting point. The sub-samples
were all placed in the same plastic sample bag, taking care to remove pebbles
and organic debris; any excess water was decanted. The bag was then sealed with
a cable tie and placed inside a second plastic bag. The sample ticket was placed
inside the second bag, and the second bag sealed.
11.1.4 |
Drill Core Samples |
Twangiza Mining has internal guidelines
and documentation which relate to the sampling of drill hole core samples. SRK
(UK) has reviewed the procedures and finds the methodology used to be acceptable
for the collection of representative samples. The entire length of each drill
hole was sampled, sample intervals were determined by geological features. In
homogeneous rock, the maximum sample interval was 1 m. The minimum sample
interval was 0.3 m. 1 m of split HQ core provides approximately 4.3 kg of
material for analysis. Samples were cut at site and shipped to Banros Bukavu
sample preparation laboratory for processing. The sample pulps were then sent to
the SGS laboratory in Mwanza, Tanzania to be analysed for gold by 50 g fire
assay.
Holes were drilled using window
sampling auger tubes, powered by an Atlas Copco percussion hammer. The diameter
of the upper sample was 64 mm, progressively reducing with depth to 49 mm, 39 mm
and 29 mm. The maximum hole depth was 7 m. Samples were logged and photographed
whilst in the window sampling tube, and then removed from the tube and bagged
for gold analysis. The logging and sampling principles employed were the same as
those used for diamond drill holes.
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11.2 |
Sample Preparation and Analysis |
|
|
11.2.1 |
Sample Preparation |
During the resource definition drilling
phase of work, a DRC subsidiary of Banro ran its own sample preparation facility
in Bukavu, DRC using its own full-time employees. ALS Chemex, Johannesburg built
the sample preparation facility and trained Banro staff. The facility was
commissioned in September 2005. Analysis of samples was undertaken by the SGS
laboratory in Mwanza, Tanzania while Genalysis in Western Australia served as
the umpire laboratory. Both SGS and Genalysis are internationally (NATA)
accredited and utilize conventional sample preparation, sample analysis and
associated quality control protocols.
The preparation of soil samples is
independently carried out to avoid possible contamination from the higher-grade
trench, adit or core samples.
Samples were delivered to the Banro
sample preparation facility in large white bags that hold between 20 and 30 kg
of samples. Each shipment between the field and Bukavu had a covering dispatch
form that was filled out in triplicate. Two copies were sent to Bukavu with the
samples and one remained at the project site. If there was any discrepancy
between the sample numbers and/or the number of samples recorded on the sample
sheets and those samples physically received at the Bukavu sample preparation
laboratory, the problem was immediately dealt with via HF radio communications
and a reconciliation report was sent by mail to the Senior Project
Geologist.
The in-house sample preparation
facility is a containerised laboratory specially designed by ALS Chemex but
managed by a subsidiary of Banro with periodic laboratory audits carried out by
external consultants.
The in-house sample preparation
facility comprises an electric oven, two jaw crushers, three disc pulverisers
and air compressor system all assembled in one 20 footer and one 40 footer
steel container.
All samples received from the field
were sorted and oven dried in labelled steel pans to optimise the resident
drying time. Using the jaw crushers, all adit, trench and drill core samples
were crushed to 80% passing a 2 mm screen. The crusher was thoroughly cleaned
between any two samples. After every 10th sample, the crusher was
flushed with barren granite, and the pulveriser was cleaned with similar
material between each sample. The cleaning process was enhanced by the use of
compressed air after each sample.
The crushed sample was split using a
riffle splitter to produce 800-1,500 g of material, which was pulverised using
B2000 Low Chrome Bowls for 90 to 300 seconds depending on the hardness of the
sample. The samples were pulverised to 90% passing a 75 micron screen.
The sample preparation laboratory has
organised areas/shelves designed for the storage of coarse and pulp rejects such
that the samples can be retrieved in a reasonable amount of time.
Pulp samples, of approximately 150 g,
are placed in brown packet envelopes, which in turn are placed in a rectangular
cardboard box that holds approximately 20 pulp samples. These boxes are shipped
in batches to the assay laboratory.
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Prior to 2005, limited adit re-sampling
by CME using ACME in Vancouver confirmed the repeatability of the original MGL
adit sample assay results determined by their laboratory in Kamituga.
Three laboratories have been used by
Banro for sample assaying since the commencement of exploration in Twangiza in
October 2005. The initial soil geochemical samples, adit re-samples and trench
samples were analysed by ALS Chemex, Johannesburg at which time SGS-Mwanza
served as the umpire laboratory.
For resource drilling Banro used
SGS-Mwanza as the principal analytical laboratory and Genalysis-Perth as the
umpire laboratory.
All gold analyses have been carried out
using conventional 50 g charge fire assay with atomic adsorption spectrography
(AAS) finish. The three laboratories involved have carried out internal checks,
which in the case of SGS-Mwanza are detailed in the section on quality control.
11.4 |
Quality Control Procedures |
In order to monitor the integrity of
the sample preparation and analytical data screen tests of crushed (5%) and
pulverized (10%) samples were routinely carried out to monitor the particle size
and percentage passing the 2 mm and 75 micron screens respectively.
To provide a measure of accuracy,
precision and confidence, a range of international reference materials,
duplicates and blanks were routinely but randomly inserted into each batch of
samples at a frequency of 12%. Blank samples were inserted during the routine
crushing and pulverising processes. Blanks were inserted into sample batches at
a frequency of 1 in 50 and a crush duplicate split was submitted 1 in 50.
Standard reference materials were inserted at a frequency of 4 in 50.
International reference materials were predominantly sourced from Rocklabs
Limited, New Zealand and occasionally from Geostats Pty Limited, Australia.
11.5 |
Assessment of Quality Control
Data |
Quality control procedures have been
implemented in all stages of the sample preparation and analytical process. The
quality control database is currently stored in a series of Century System
database which were provided to SRK (UK) for analysis. The quality control work
included the insertion of international reference samples, inter-laboratory
checks, sample preparation laboratory duplicates, blanks, and the analytical
laboratorys internal checks. These are all described in detail in the following
sections.
11.5.1 |
Certified Reference
Material |
Four certified reference material
samples (standards) were inserted in each batch of 50 samples. The standards
were sourced from Rocklabs Limited, New Zealand and a few came from Geostats,
Australia. The standard samples are in pulp form and are supplied in plastic
containers of 2.5 kg each, of both oxide and sulphide material with variable
grade ranges covering the expected grade range for the Twangiza deposit. A total
of seven different grade ranges of standards have been randomly inserted into
batches of samples submitted to the analytical laboratory during the latest
drilling programme; these cover the entire grade range and include both oxide
and sulphide material.
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The standards are randomly inserted
using the same quantity and sachets as the laboratory pulps, making them
difficult to be detected by the analytical laboratory. Statistical assessment of
the results is completed routinely using Twangiza Minings internal guidelines
and not using the parameters provided as part of the Rocklabs Quality Control
package. The mean assay values returned by the laboratory for all certified
reference material in relation to their respective values are considered to be
within acceptable limits. Results reporting out of assigned limits trigger a
re-assay of that batch of samples.
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|
Table 11-1: |
Statistics of Results of Standard Certified
Sample Submissions |
STD REFID |
REF VALUE
(g/tAu) |
NUMBER OF SUBMISSSIONS |
MIN (g/tAu) |
MAX
(g/tAu) |
MEAN (g/tAu) |
STD DEV |
CoV |
SP37 |
18.14 |
26 |
16.9 |
20.40 |
17.75 |
0.76 |
0.04 |
SN38 |
7.92 |
33 |
0.01 |
9.38 |
7.92 |
2.21 |
0.28 |
SL51 |
5.909 |
44 |
4.24 |
6.67 |
5.85 |
0.35 |
0.06 |
SL34 |
5.86 |
400 |
0.01 |
15.00 |
5.86 |
0.73 |
0.12 |
SK52 |
4.107 |
42 |
0.41 |
4.54 |
4.01 |
0.61 |
0.15 |
SK43 |
4.00 |
59 |
0.09 |
4.44 |
3.90 |
0.64 |
0.16 |
SK33 |
4.00 |
479 |
0.02 |
4.45 |
3.97 |
0.25 |
0.06 |
SI54 |
1.761 |
38 |
1.63 |
2.00 |
1.80 |
0.09 |
0.05 |
SI42 |
1.761 |
54 |
1.03 |
1.99 |
1.78 |
0.12 |
0.07 |
SH41 |
1.344 |
6 |
1.16 |
1.35 |
1.28 |
0.06 |
0.05 |
SG40 |
0.976 |
15 |
0.83 |
1.11 |
0.98 |
0.06 |
0.06 |
SE44 |
0.606 |
42 |
0.55 |
0.65 |
0.61 |
0.02 |
0.03 |
OXP61 |
14.92 |
35 |
13.9 |
15.80 |
15.00 |
0.39 |
0.03 |
OXN62 |
7.706 |
42 |
6.76 |
8.33 |
7.77 |
0.31 |
0.04 |
OXL63 |
5.865 |
8 |
5.62 |
6.27 |
5.94 |
0.21 |
0.04 |
OXL40 |
1.82 |
33 |
1.05 |
1.99 |
1.85 |
0.15 |
0.08 |
OXK69 |
3.583 |
31 |
3.24 |
3.73 |
3.58 |
0.10 |
0.03 |
OXJ68 |
2.342 |
55 |
2.22 |
2.55 |
2.35 |
0.07 |
0.03 |
OXJ64 |
2.366 |
36 |
2.18 |
2.72 |
2.35 |
0.09 |
0.04 |
OXI54 |
1.868 |
8 |
1.78 |
1.87 |
1.83 |
0.03 |
0.02 |
OXI 67 |
1.817 |
63 |
1.52 |
2.05 |
1.83 |
0.07 |
0.04 |
OXH55 |
1.282 |
13 |
1.05 |
1.35 |
1.28 |
0.09 |
0.07 |
OXH52 |
1.31 |
72 |
0.01 |
1.58 |
1.27 |
0.24 |
0.19 |
OXG60 |
1.07 |
8 |
0.99 |
1.16 |
1.07 |
0.06 |
0.06 |
OXG46 |
1.037 |
5 |
0.97 |
1.09 |
1.03 |
0.05 |
0.05 |
OXF65 |
0.805 |
39 |
0.74 |
0.94 |
0.83 |
0.03 |
0.04 |
TOTAL |
|
1686 |
|
|
|
|
|
A review of the charts shows the
laboratory historically produced accurate assays, however in the later stage of
the drilling programme the results were more variable and there was a low bias
but grades remain within acceptable limits. The limits shown on the charts are
based on the certified standard deviation with the upper and lower limits set at
three times the standard deviation value. A summary of the standard used in
submissions from the latest round of drilling are shown in Figure 11-1 and Table
11-1.
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11.5.2 |
Inter-laboratory check
assays |
Twangiza Mining completed
inter-laboratory check assays with the same assay techniques on a routine basis.
Sample pulp duplicates were sent at the end of every quarter in batches of
between 80 and 100 samples for analyses at both SGS-Mwanza and Genalysis-Perth.
The pulp samples cover a representative grade range.
Summary statistics for the two datasets
have been determined and the Mean Relative Differences (MRD) of the SGS and
Genalysis result have been calculated. Above 1.0 g/t Au there is no significant
bias with the mean relative difference (MRD) reporting between -2.6 % and 4.3 %.
11.5.3 |
Duplicate coarse split |
Before the 2009 Feasibility Study, a
total of 1,377 split duplicates were included in routine sample batches. These
are statistically reviewed in Table 11-2 below. After the 2009 Feasibility Study
a total of 158 split duplicate were submitted, these are summarised in Table
11-3.
The data is plotted in Figure 11-2.
There is generally close agreement although some pairs show considerable
variance indicating coarser gold in a minor proportion of the samples or some
potential for sample swaps during submission and analysis.
|
Table 11-2: |
Coarse Split Sample Pairs pre FS
|
|
ALL SAMPLES |
>0.5g/t Au |
ORIGINAL g/t Au
|
DUPLICATE g/t Au
|
ORIGINAL g/t Au
|
DUPLICATE g/t Au
|
NUMBER OF PAIRS |
1377 |
1377 |
274 |
274 |
MINIMUM |
0.004 |
0.005 |
0.50 |
0.01 |
MAXIMUM |
47.70 |
44.20 |
25.3 |
22.4 |
MEAN |
0.57 |
0.55 |
2.07 |
1.97 |
STANDARD DEVIATION |
2.39 |
2.43 |
2.80 |
2.53 |
|
Table 11-3: |
Coarse Split Sample Pairs Post FS
|
|
ALL SAMPLES |
>0.5g/t Au |
ORIGINAL g/t Au
|
DUPLICATE g/t Au
|
ORIGINAL g/t Au
|
DUPLICATE g/t Au
|
NUMBER OF PAIRS |
158 |
158 |
16 |
16 |
MINIMUM |
0.005 |
0.005 |
0.52 |
0.51 |
MAXIMUM |
2.99 |
3.15 |
2.99 |
3.15 |
MEAN |
0.18 |
0.17 |
1.31 |
1.20 |
STANDARD DEVIATION |
0.45 |
0.41 |
0.89 |
0.79 |
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The blank material is made from granite
and was purchased from ALS Chemex - Mwanza laboratory. There were 1587 blank
sample submissions which returned a mean value of 0.007 g/t; these are shown in
Figure 11-3. The more recent results suggest there was a minor contamination
issue; the grouped nature of slightly high results may indicate periods in which
the routine cleaning the equipment between samples was not undertaken
thoroughly. Overall this is not considered to be an issue for the data quality.
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12 |
DATA VERIFICATION |
|
|
12.1 |
Database and Data Quality |
Table 12-1 below shows the data type
and amount available for geological modelling and Mineral Resource Estimation.
|
Table 12-1: |
Data Type and Amount
of Data Employed In Current
Geological Modelling
|
DATA TYPE |
NUMBER |
TOTAL LENGTH |
NO OF SAMPLES |
HISTORICAL TRENCHES (MGL) |
71 |
8,323 |
4,088 |
HISTORICAL ADIT (MGL) |
202 |
13,207 |
13,716 |
1997/98 DIAMOND HOLES (CME) |
20 |
9,122 |
8,557 |
RE-SAMPLED ADITS (CME) |
16 |
1,689 |
1,613 |
FOLLOW-UP |
6 |
966.4 |
624 |
FOLLOW-UP (PHASE 1) DD HOLES |
33 |
6,421 |
6,881 |
FOLLOW-UP (PHASE 2) DD HOLES |
38 |
10,616 |
11,194 |
FOLLOW-UP (PHASE 3) DD HOLES |
29 |
6,835.78 |
7,602 |
FOLLOW-UP (PHASE 3 cont) DD HOLES |
116 |
22,114 |
24,534 |
FOLLOW-UP (PHASE 4 cont) DD HOLES |
102 |
21,952.26 |
20,004 |
FOLLOW-UP (PHASE 5) |
138 |
12,557 |
8,676 |
FOLLOW-UP (PHASE 5 Cont.) |
11 |
499.90 |
186 |
In the 1960s approximately 13,207
metres of adits were developed and sampled by MGL. A programme of data
verification was completed by CME by re-sampling 1,613 of the original 13,716
adit channel samples recorded by MGL. In total, 16 adits (including crosscuts)
were rehabilitated, mapped and channel sampled at one metre intervals. A total
of 1,613 channel samples were collected from both walls of the adits. A
graphical illustration of gold results for the MGL and CME sampling is provided
in Table 12-1.
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Composites of significant intersections
using a 1 g/t Au cut-off based on CMEs re-sampling are presented in Table 12-2
below.
|
Table 12-2: |
Adit sampling - statistics of composites
|
COMPANY |
NO OF
COMPOSITES |
MIN
(g/t Au) |
MAX
(g/t Au) |
AR.
MEAN (g/t Au) |
STD.
DEV (g/t Au) |
WEIG-MEAN (g/t Au) |
MGL |
40 |
0.25 |
6.37 |
3.05 |
1.57 |
3.76 |
CME |
40 |
1.15 |
10.24 |
3.59 |
1.99 |
3.45 |
MGL and CME results over the same
interval compare reasonably well; overall CME results are slightly higher than
the MGL results.
The primary sample data files were
validated in several ways:
|
|
visually checking summary statistics of all
data fields to ensure they do not exceed reasonable limits and
unrecognized rock codes; |
|
|
|
|
|
passing the data through a general validation process
that identifies, amongst others duplicate samples, reversed and
over-length intervals as well as overlapping sample limits; and |
|
|
|
|
|
plotting the samples in plan and in section and
comparing the sample locations and assay values. |
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A number of transcription errors were
originally found in the database due to sample numbers entered in the gold
column. The error was reported to Twangiza and the files corrected accordingly.
SRK (UK) has reviewed logging
procedures on site and the electronic files provided by Twangiza at the time of
the resource estimate. SRK (UK) noted the lower quality in the returned assays
during the third quarter of 2008 but nevertheless considers the data to be valid
and therefore suitable for the Mineral Resource estimate.
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13 |
MINERAL PROCESSING AND METALLURGICAL
TESTING |
|
|
13.1 |
Background |
Scoping study metallurgical tests were
conducted on oxide, transition and fresh, non-refractory, ore types from the
Twangiza Main and North ore bodies. The testwork evaluated a number of process
options to recover gold including flotation, gravity and cyanidation processes.
Comminution tests were also performed. The scoping study results were used to
develop the initial process flow sheet for the Twangiza ore body. The scoping
study testwork results are contained in the following reports:
|
|
September, 2007, SGS, Laboratory Testwork:
Scoping Study Banro Twangiza Project (Democratic Republic of Congo),
Report # MET 07/U82 TWA; |
|
|
|
|
|
October 5, 2006, SGS, Procedural Diagnostic
Leach Appraisal on 11 Gold Bearing Ore Samples, Report No. MET 06/S16;
|
|
|
|
|
|
April 20, 2007, SGS, Bulk Rougher Flotation
Testwork on Twenty Three Gold Ore Samples, Report No. Flotation 07-412;
|
|
|
|
|
|
July 3, 2006, SGS, Flotation Testwork on a
Gold Bearing Ore Blends from Banro- Twangiza Mine, Report No. Flotation
07/132; |
|
|
|
|
|
June 28, 2007, SGS, Mineralogical Characterization and
Gold Deportment Study on Five Gold Ore Samples from the Twangiza Gold
Deposit, DRC, Mineralogical Report No. MIN 0507/066; |
|
|
|
|
|
October 3, 2007, SGS, Bio-oxidation Test
Program on Twangiza Concentrate, Report No. BIOMET 07/06; |
|
|
|
|
|
October, 2007, Geobiotics, Phase 1, Report for
GEOCOAT® Amenability Test; and |
|
|
|
|
|
April 22, 2008, SGS, Cyanide Destruction Testwork on
Twangiza North and Main Composite Middlings and Tailings, Report No.
BIOMET 08/09. |
As part of the pre-feasibility study
SENET initiated and supervised additional metallurgical testwork to further
investigate and improve gold recoveries and reagent consumptions. These results
were incorporated in to the pre-feasibility study. A list of the reports
prepared by a number of laboratories is provided below:
|
|
April 22, 2008, SGS, Cyanide Destruction
Testwork on Twangiza North and Main Composite middlings and tailings,
Report # BIOMET 08/09; |
|
|
|
|
|
May 26, 2008, SGS, A General Mineralogical and
Gold Deportment Study on two Gold bearing ore Samples from the Twangiza
Project, Report # MIN 0208/038a; |
|
|
|
|
|
June 2008, Knelson Concentrators Africa (Pty) Ltd,
Extended Gravity Recoverable Gold (EGRG) Testwork on Two Ore Samples from
the Twangiza Project submitted by SENET (Pty) Ltd Report No. SENET/03/08;
|
|
|
|
|
|
July 2008, Knelson Concentrators Africa (Pty) Ltd,
Modelling the Proposed Gravity Gold Recovery Circuit at the Twangiza
Project Report No. SENET/03.2/08; and |
|
|
|
|
|
July 2008, Paterson and Cooke, Twangiza
Project: Results of Bench-Top Thickening and Rheology Testwork, Report
No. SEN-TWA-8144R01 Rev 0. |
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Further metallurgical tests were
conducted to raise the level of confidence to feasibility level and results of
these tests are contained in the reports listed below:
|
|
July 4, 2008, SGS, Metallurgical Test
Programme for Gold Bearing Ore Sources (Twangiza Non-Refractory FP
Samples, Report No. Met/08/131 Rev2; |
|
|
|
|
|
August 07, 2008, SGS, Cyanide Destruction Testwork on
Twangiza Main Fresh and Transitional Composite Middlings and Tailings,
Report No. BIOMET 08/15; |
|
|
|
|
|
August 08, 2008, SGS, Cyanide Destruction Testwork on
Twangiza North Fresh and Transitional Composite Middlings and Tailings,
Report No. BIOMET 08/18; |
|
|
|
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Extended Gravity Recoverable Gold (EGRG) Testwork on Two Main Ore Body
Fresh and Transition Ore Samples from the Twangiza Project submitted by
SENET (Pty) Ltd Report No. SENET/05/08; |
|
|
|
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Modelling the Proposed Gravity Recovery Circuit at the Twangiza Project
Main Ore Body Hard Ores Report No. SENET/05.1/08; |
|
|
|
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Extended Gravity Recoverable Gold (EGRG) Testwork on Two North Ore Body
Fresh and Transition Ore Samples from the Twangiza Project submitted by
SENET (Pty) Ltd Report No. SENET/06/08; |
|
|
|
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Modelling the Proposed Gravity Recovery Circuit at the Twangiza Project
North Ore Body Hard Ores Report No. SENET/06.1/08; |
|
|
|
|
|
November 20, 2008, SGS, Metallurgical Test Programme:
SENET Project Twangiza Oxide, Report No. Met07/BB27; |
|
|
|
|
|
December 10, 2008, Mintek, Twangiza Comminution Test
Work Report, Report No. MPC-584; |
|
|
|
|
|
January, 2009, Orway Mineral Consultants (OMC),
Twangiza Gold Project Comminution Circuit Design, Report No. 8282; and
|
|
|
|
|
|
February 10, 2009, SGS, Metallurgical Test
Programme: SENET Project Twangiza Non-Refractory Variability Samples,
Report No. Met07/BB27. |
The summarised findings of the testwork
are given in Table 13-1.
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|
Table 13-1: |
Ore Characteristics
|
ITEM
|
UNIT |
OXIDE
|
TRANSITIONAL |
|
MAIN |
NORTH |
MAIN |
NORTH |
ORE CHARACTERISTICS |
|
|
|
|
|
ORE HEAD GRADE (AU) |
g/t |
2.42 |
2.1 |
1.52 |
2.36 |
SPECIFC GRAVITY |
t/m3 |
2.79 |
2.88 |
2.79 |
2.99 |
UNCONFINED COMP STRENGTH |
MPa |
14.88 |
8 |
22.63 |
10.56 |
BOND CRUSHER WORK INDEX |
kWh/t |
3.1 |
2.7 |
9.41 |
6.54 |
BOND CRUSHER WORK INDEX |
kWh/t |
4.3 |
3.6 |
12.1 |
8.1 |
ROD MILL WORK INDEX |
kWh/t |
6.8 |
5.3 |
8.06 |
4.05 |
BALL MILL WORK INDEX |
kWh/t |
3.2 |
4.23 |
8.61 |
6.88 |
ABRASION INDEX |
Ai |
0.0078 |
0.034 |
0.196 |
0.1932 |
LIFE FACTOR |
|
9.16 |
10.87 |
2.19 |
2.19 |
JK TECH PARAMETERS |
|
|
|
|
|
A |
|
85.9 |
85.9 |
55.1 |
65.6 |
b |
|
1.25 |
1.25 |
2.49 |
5.22 |
Ta |
|
1.14 |
1.14 |
0.96 |
3.38 |
GRAVITY |
|
|
|
|
|
GRAVITY RECOVERY |
% of h/g |
16.5 |
36.4 |
11.8 |
41.6 |
GRAVITY CONCENTRATION RATIO |
|
1,100 |
2,500 |
1,100 |
1,100 |
CONC.MASS AS % OF FEED TO CONC |
% |
0.05 |
0.05 |
0.05 |
0.05 |
INTENSIVE LEACH DISSOLUTION |
|
98 |
98 |
98 |
9 |
CIL |
|
|
|
|
|
LAB LEACH DISSOLUTION |
% |
91.1 |
89.5 |
80.0 |
91.7 |
EFFICIENCY FACTOR |
% |
98.0 |
98.0 |
98.0 |
98.0 |
CIL DISSOLUTION |
% |
89.3 |
87.7 |
78.4 |
89.9 |
LEACH SOLIDS FEED % m/m |
% |
35 |
35 |
42 |
42 |
CYANIDE CONSUMPTION |
kg/t |
0.06 |
0.63 |
0.98 |
0.65 |
LIME CONSUMPTION AS 100% |
kg/t |
4.31 |
4.70 |
2.26 |
1.27 |
TOTAL LEACHING RESIDENCE TIME |
hrs |
15 |
15 |
24 |
24 |
CARBON TO GOLD LOADING RATIO |
|
2,500 |
2,500 |
1,750 |
1,750 |
GOLD PRODUCTION |
|
|
|
|
|
OVERALL RECOVERY |
% |
90.2 |
91.2 |
79.5 |
93.2 |
DETOXIFICATION |
|
|
|
|
|
REQUIRED RESIDENCE TIME |
hr |
2 |
2 |
2 |
2 |
RESIDUAL CYANIDE IN CIL TAILS |
ppm |
90 |
90 |
90 |
90 |
LIME CONSUMPTION AS 100% |
kg/t CaO |
0.06 |
0.06 |
0.06 |
0.06 |
SMBS USAGE RATE / RESIDUAL CYANIDE |
g/g NaCN |
4.44 |
4.18 |
4.57 |
4.62 |
CUSO4USAGE RATE / RESIDUAL CYANIDE |
g/g NaCN |
0.48 |
0.48 |
0.11 |
0.15
|
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All samples were selected and supplied
by Twangiza according to SENETs requests.
13.3 |
Review of Scoping Study Metallurgical
Testwork |
This section summarises the results of
the metallurgical tests performed by SGS in 2007. Sample sourcing and make-up
test methods and detailed results are contained in the reports from SGS
referenced above and are part of the scoping study technical report produced by
SENET dated 13th September, 2007 and entitled Preliminary Assessment
NI 43-101 Technical Report, Twangiza Gold Project, South Kivu Province,
Democratic Republic of Congo.
The main findings from the SGS scoping
study metallurgical tests were as follows:
|
|
Bond Ball Mill Work indices (BBWi) obtained ranged from
about 3.2 13.2 kWh/t metric. It was noted that the oxides were in the in
the very soft category and transition and fresh were in the medium band;
|
|
|
|
|
|
The abrasion results ranged from 0.0096 to 0.66 which
showed that the oxides were in the low abrasive category whilst the
transition and fresh was in the medium to high abrasive classes;
|
|
|
|
|
|
Diagnostic leach tests indicated that gold in oxides was
amenable to direct cyanidation while transition and fresh samples
displayed that some of the gold could easily be recovered with direct
cyanidation technique and some could hardly be recovered using this
method; |
|
|
|
|
|
Mineralogical investigations on transition and fresh
samples revealed that a fair proportion of the not so easy to recover
gold was associated with arsenopyrite and pyrite and that only 30-60% of
the gold would be amenable to direct cyanidation, a phenomenon that was
proved in the later stages of the scoping study testing; |
|
|
|
|
|
The oxides (main and north) responded well to gravity and
leach recovery testwork with overall recoveries of 90% being attained;
|
|
|
|
|
|
Bottle-roll tests conducted on transition and fresh ore
samples with excess cyanide indicated that there were distinct lithologies
within each ore type and each lithology responded differently to direct
cyanidation. Typically the gold dissolution achieved was:
|
|
o |
Transition FP 81% |
|
|
|
|
o |
Transition CMS 38% |
|
|
|
|
o |
Fresh FP 91.8% |
|
|
|
|
o |
Fresh CMS 54% |
These results led to the investigation
of an alternative processing methods utilising flotation followed by separate
treatment by cyanidation of the flotation concentrates and tailings of all
lithologies. In general the following results were achieved:
|
|
Generally >95% of the gold was recovered
into the concentrates in all instances; |
|
|
|
|
|
The mass yield to flotation concentrates were generally
high; ranging between 30 and 50%. In instances where low mass yields were
obtained (10-15%), low gold recoveries were also noted;
|
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|
For CMS, it was noted that there was no significant
improvement in gold dissolution from concentrates and flotation tails
compared to dissolutions obtained on the as Received samples; and
|
|
|
|
|
|
For FP, it was noted that gold dissolution remained the
same from the leaching of the concentrates and flotation tailings compared
to dissolutions obtained on the as Received samples.
|
Based on poor gold recoveries, seen for
both flotation concentrates and tailings from CMS material, bio-oxidation was
investigated as an alternative treatment route. Bio-oxidation resulted in a
sulphide oxidation of 95% after 29 days and final gold recovery of 88% from the
oxidised residue. This represented a 30% improvement in gold recovery over
direct cyanidation. While the tests demonstrated that the ore was amenable to
bio-oxidation, the high mass pulls (between 30 and 50%) coupled with the long
residence time meant that heap leach of the concentrates using a process called
GEOCOAT® was the only viable option. However, due to real estate limitations at
the Twangiza site this option was shelved.
13.4 |
Review of Pre-Feasibility Study Metallurgical
Testwork |
Following a review of the scoping study
metallurgical test results, a test programme for the pre-feasibility study was
implemented and was divided into two phases, namely:
Phase 1: Optimization tests aimed at
establishing optimum parameters for gold recovery and characterization of the
comminution properties;
Phase 2: Variability testwork aimed at
testing the optimum conditions / parameters on variability samples.
The following is a summary of the
pre-feasibility study testwork findings and detailed results are contained in
the reports from SGS, Knelson Concentrators Africa and Paterson and Cooke
referenced above and are part of the prefeasibility study technical report
produced by SENET dated August 13, 2008 and entitled Pre-feasibility Study
NI43-101 Technical Report, Twangiza Gold Project, South Kivu Province,
Democratic Republic of Congo.
The mineralogical testwork was
conducted on the main and north oxide composite samples. No additional
mineralogy was conducted on transition and fresh ore types as the work conducted
during scoping study phase was deemed to be adequate. The main findings were:
|
|
The Twangiza north oxide ore is characterised by coarse
(mostly >75µm ECD) liberated gold (>99%) and gravity gold recovery
could be an option for this type of ore; and |
|
|
|
|
|
Twangiza main oxide ore is characterised by finer-grained
gold (all <75µm ECD) with a substantial proportion locked in gangue
minerals (~20%) which could potentially exhibit recovery problems. Both
Twangiza main and north oxide ore contained considerable amounts of clay
minerals that could be detrimental to different parts of the process such
as mill selection, thickening and equipment sizing such as pumps,
agitators and inter- stage screens. |
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13.4.2 |
Comminution results of composite
samples |
The following are main findings from
the comminution tests performed on the composite samples:
|
|
The Bond Rod Mill Work Indices (BRWi) for main
oxide and north oxide was classified as soft; |
|
|
|
|
|
The abrasion indices (Ai) for the oxides were all
<0.05 which classified them as very low abrasive ores. The transition
and fresh abrasion indices were classified as moderately abrasive. This
indicates that medium grinding media consumptions and liner wear will be
expected; |
|
|
|
|
|
The Bond Crushability Work Indices was only
performed on the oxide and showed the ore to be very soft; and |
|
|
|
|
|
The UCS results for the oxides indicated that
the ore is very soft (<50MPa) and should be treated using a mineral
sizer. |
13.4.3 |
Comminution results of variability
samples |
Variability comminution testwork
included BBWi determinations only and the results are summarized below.
The BBWI results showed that the main
and north oxides exhibited very soft to soft behaviour (1.8 -7.4kWh/t);
The main transition FP non refractory
samples were in the medium hardness category (9.1 -14.5 kW/h) and north
transition FP non refractory in the medium to hard class (10.7 -18.2 kWh/t);
The Bond ball work indices of main
fresh FP non refractory indicated medium hardness (10.8 -13.4 kWh/t) whilst the
north fresh FP non refractory showed a range of medium to hard (14.3 -16.2
kWh/t).
The results obtained from the
variability testwork showed that the ore is variable in nature and this should
be taken into account when designing the comminution circuit.
13.4.4 |
Gold recovery testwork |
Optimisation gold recovery testwork was
performed on the main and north oxide composite samples during the scoping
study. The optimum conditions obtained during these tests were used to evaluate
the variability of the ore.
No optimisation gold recovery tests
were performed on the non-refractory transition and fresh samples during the
scoping study. It was therefore necessary to perform these tests during the
pre-feasibility study.
13.4.5 |
Gravity and intensive cyanidation testwork (oxides
only): |
Twangiza main oxides displayed less
gravity recoverable gold than Twangiza north oxides which is in line with
mineralogical investigations. However in both cases, the recoveries warranted
the installation of a centrifugal concentrator to recover the free gold ahead of
CIL.
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Twangiza main oxides obtained a GRG
(gravity recoverable gold) recovery of 28.7% which equated to a predicted plant
gravity recovery of 20.9% .
Twangiza north oxides achieved a higher
GRG recovery of 49.7% which equated to a predicted plant gravity recovery of
43.4% .
These gravity recoveries, obtained for
the oxides, were in line with the recoveries obtained during the scoping
testwork phase.
Intensive cyanidation testwork,
performed on the gravity concentrate gave gold dissolution >95% in less than
24 hours irrespective of the concentrate grades. This is an indication that
concentrates will be easily leached during plant operations. An intensive
cyanidation reactor followed by electrowinning was therefore considered as part
of the gravity gold recovery system.
13.4.6 |
Variability leach testwork (oxides
only) |
The calculated head grade of main oxide
varied from 1.99 g/t to 5.04 g/t with minimum and maximum recoveries of 83.66%
and 88.51% . This showed that this ore body, when leached, obtained similar
recoveries irrespective of head grade.
Cyanidation of north oxide variability
samples with head grades ranged from 0.74 g/t to 1.97 g/t resulted in gold
dissolutions between 70.12% and 89.77% . The varying range in the dissolution
demonstrated that this ore is highly variable which could be an indication of
presence of some refractory material in the oxides.
The sodium cyanide consumption for the
oxides was generally less than 1 kg/t which was much higher than that obtained
during the scoping study which could be an indication of the variable nature of
cyanide consumers associated with the oxide ore. Lime consumption was generally
very high however still comparable to what was achieved during the scoping
study, for all samples that were tested.
13.4.7 |
Cyanide Detoxification (Oxides
Only) |
Bench scale cyanide destruction testing
was performed on composite samples of oxide leach tailings generated from
metallurgical testwork.
The relative effectiveness of ferrous
sulphate, sodium metabisulphite, sodium metabisulphite plus copper, hydrogen
peroxide and alkaline chlorination at removing or destroying soluble cyanide
forms in the tailings were compared and results showed that the alkaline
chlorination and ferrous sulphate gave the best cyanide destruction results.
However this method was not selected due to environmental concerns. The use of
sodium metabisulphite and copper was chosen as the effective method, with 82%
cyanide destruction.
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13.4.8 |
Transition and fresh (main and north) optimisation
recovery testwork |
Bulk gravity tests were performed on
main and north (transition and fresh) non-refractory composite samples to
produce middlings and tailings for the leach optimisation testwork and gravity
concentrate for the intensive cyanidation testwork. The main findings were:
Gravity - The composite samples showed
gravity recoveries varying from 4.5 to 21%, with mass yields ranging from 0.33
to 0.5% . These test results are inconclusive and additional EGRG (Extended
Gravity Recoverable Gold) tests were scheduled to be performed during the
feasibility testwork phase;
Intensive Cyanidation - Intensive
cyanidation tests were performed on the primary gravity concentrates and
relatively high gravity concentrate dissolutions (>90%) were obtained for the
transition main and north samples. Approximately 85% gold dissolution was
obtained for the fresh main and north gravity concentrates which could be an
indication of the presence of some refractory material in the concentrates.
Overall, the concentrates responded well to intensive cyanidation;
Preg-robbing - The preg-robbing test
results showed that the transition and fresh ores displayed preg-robbing
characteristics which is in line with the mineralogical findings conducted
during scoping study phase. All further tests were therefore conducted using the
CIL treatment route;
Effect of Grind - It was noted for all
ore types that there was an increase in dissolution with finer grinding and an
optimum grind of 80% -75 µm was chosen which is in line with the selected grind
for oxides. This grind was used for all subsequent testwork;
Effect of cyanide addition - Transition
and fresh (main and north) recoveries increased with increase in cyanide
addition from 1.0 to 3.0 kg/t. However the increase in cyanide consumptions and
hence costs, erodes the benefits associated with extra recovery. It was
recommended that cyanide addition rates be maintained at 2 kg/t; and
Effect of Time and Oxygen - a small
increase in dissolution was noted with an increase in leach time above 24 hours.
However, 24 hours was selected as the leach time in order to reduce the number
and size of the leach tanks.
13.4.9 |
Settling and viscosity |
Settling testwork was conducted by
Paterson and Cooke on oxides, transition and fresh ore samples obtained from
Twangiza main and north ore bodies. The detailed report is entitled July, 2008,
Paterson and Cooke, Twangiza Project: Results of Bench-Top Thickening and
Rheology Testwork, Report No. SEN-TWA-8144R01 Rev 0. The main findings were:
|
|
The main and north oxide mill feed slurries exhibited
similar rheological characteristics and was classified as pastes at mass
solids concentrations greater than 52% w/w. Slightly higher yield stress
and viscosity values were achieved for the north oxide and it was found
that due to the high viscosities, the addition of viscosity modifiers to
the milling circuit might be required; and |
|
|
|
|
|
Whilst a thickener has not been included in the Twangiza
design, flocculant addition and thickener sizing information is included
in the Paterson and Cooke report should the results be required.
|
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13.5 |
Review of Feasibility Study Metallurgical
Testwork |
Following the review of the scoping
study and the pre-feasibility study metallurgical test results, SENET proposed a
test program for the feasibility study with the following aims and objectives:
|
|
To characterise the ore body further with respect to its
comminution characteristics by performing SAG mill tests to support the
flow sheet developed during the pre-feasibility study or to develop a new
flow sheet for the feasibility study; and |
|
|
|
|
|
To determine the variability of the ore by using the
optimum parameters achieved during the scoping and pre-feasibility
testwork phases for gold recovery. |
The results for the EGRG tests and
cyanide detoxification tests, performed on main and north transition and fresh
composite samples during the pre-feasibility testing phase, were only available
after completion of the pre-feasibility study and are included as a summary in
this section.
13.5.1 |
Gravity recovery main and north transition and fresh
samples |
Knelson Africa was commissioned to
perform GRG tests and simulation of the results using KCMOD*Pro model to predict
circuit recovery and the result and their findings is contained in four reports
noted below:
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Extended Gravity Recoverable Gold (EGRG) Testwork on Two Main Ore Body
Fresh and Transition Ore Samples from the Twangiza Project submitted by
SENET (Pty) Ltd Report No. SENET/05/08; |
|
|
|
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Modelling the Proposed Gravity Recovery Circuit at the Twangiza Project
Main Ore Body Hard Ores Report No. SENET/05.1/08; |
|
|
|
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Extended Gravity Recoverable Gold (EGRG) Testwork on Two North Ore Body
Fresh and Transition Ore Samples from the Twangiza Project submitted by
SENET (Pty) Ltd Report No. SENET/06/08; |
|
|
|
|
|
August, 2008, Knelson Concentrators Africa (Pty) Ltd,
Modelling the Proposed Gravity Recovery Circuit at the Twangiza Project
North Ore Body Hard Ores Report No. SENET/06.1/08.
|
The results are summarized in the Table
13-2.
|
Table 13-2 |
Predicted Plant GRG Recoveries
|
ORE TYPE |
GRG VALUE (%)
|
PREDICTED
PLANT GRG RECOVERY (%) |
MAIN TRANSITION |
16.7 |
11.8 |
MAIN FRESH |
11.6 |
6.6 |
NORTH TRANSITION |
53.8 |
41.6 |
NORTH FRESH |
36.1 |
26.5 |
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Free gold recovery from the Twangiza
main, transition and fresh ores, is problematical for gravity recovery as the
GRG values are low, the F80 is coarse and the free gold is
finely disseminated. These factors result in a difficult and low-value
opportunity for gravity recovery of gold from these ores as can be seen by the
results in Table 13-2.
In contrast, the Twangiza north
transition and fresh ore samples, exhibited a more ideal situation for gravity
recovery as the GRG values were high. Consequently, a gravity recovery stage was
included for the Twangiza process plant.
13.5.2 |
Cyanide destruction |
Bench scale cyanide destruction testing
was performed on samples of leach tailings generated from metallurgical
testwork. The samples were as follows:
|
|
Main Transition ore. |
|
|
|
|
|
Main Fresh ore. |
|
|
|
|
|
North Transition ore. |
|
|
|
|
|
North Fresh ore. |
SGS performed the cyanide destruction
tests and the results and findings are contained in two reports listed below:
|
|
August 07, 2008, SGS, Cyanide Destruction Testwork on
Twangiza Main Fresh and Transitional Composite Middlings and Tailings,
Report No. BIOMET 08/15; and |
|
|
|
|
|
August 08, 2008, SGS, Cyanide Destruction Testwork on
Twangiza North Fresh and Transitional Composite Middlings and Tailings,
Report No. BIOMET 08/18. |
Two methods were used on the transition
and fresh samples; sodium metabisulphite plus copper, and Caros Acid for
removing or destroying soluble cyanide forms in the tailings. A summary of WAD
Cyanide removal and reagent consumptions are shown in Table 13-3.
The WAD (Weak Acid Dissociable) cyanide
removal efficiency for all samples was very low and this was due to the high WAD
CN in the feed to the detoxification. The WAD cyanide removed ranged from 218 to
594 ppm which is higher than what will be expected in the plant. The cyanide
destruction process will be designed to be run on a continuous basis but was
performed batch wise at the lab and is less efficient and tend to require more
reagents for efficient cyanide removal. The reagent consumptions are therefore
exaggerated and should be lower.
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Table 13-3: |
Summary of Cyanide Destruction Tests
|
TEST METHOD |
UNIT |
MAIN
|
NORTH
|
TRANS. |
FRESH
|
TRANS.
|
FRESH
|
SMBS and CuSO4 |
|
|
|
|
|
START WAD CN (FEED TO TAILINGS) |
ppm |
744 |
793.6 |
345 |
322 |
END WAD CN (AFTER DETOX) |
ppm |
150.4 |
288 |
72 |
104 |
WAD CN REMOVED |
ppm |
594 |
506 |
273 |
218 |
WAD CN REMOVAL EFFICIENCY |
% |
79.78 |
63.71 |
79.13 |
67.7 |
g (SMBS) : g (WAD CN) |
g:g WAD |
4.57 |
5.73 |
4.62 |
5.41 |
g (CuSO4) : g (WAD CN) |
g:g WAD |
0.11 |
0.15 |
0.19 |
0.25 |
CAROS ACID |
|
|
|
|
|
START WAD CN (FEED TO TAILINGS) |
ppm |
744 |
793.6 |
345 |
322 |
END WAD CN (AFTER DETOX) |
ppm |
240 |
241.6 |
104 |
152 |
WAD CN REMOVED |
ppm |
504 |
552 |
241 |
170 |
WAD CN REMOVAL EFFICIENCY |
% |
67.74 |
69.56 |
69.86 |
52.8 |
g (CAROS ACID) : g (WAD CN) |
g:g WAD |
57.9 |
6.2 |
56.2 |
74.3 |
g (CuSO4) : g (WAD CN) |
g:g WAD |
0.1 |
- |
- |
- |
13.5.3 |
Comminution appraisal |
Comminution testwork was performed in
two phases. The first phase involved performing the tests on individual
composites of the transition and fresh ores types, main and north ore bodies.
These results were analysed and the data was provided for the second phase to
Orway Mineral Consultants (OMC), an internationally recognised specialist in
grinding circuit design, for interpretation and mill sizing by comparison
against their existing database.
Mintek carried out full bench scale
comminution tests that included the standard bond work index (BBWi) tests,
Advanced Media Competency Tests (AMCT) and JKTech drop weight testing. A summary
of the results are shown in the Table 13-4 and Table 13-5. The detailed results
can be found in the report by Mintek, December 10, 2008, Mintek, Twangiza
Comminution Test Work Report, Report No. MPC-584.
|
Table 13-4: |
Summary of Bond Work Index Tests
|
SAMPLE |
BBWi (kWh/t) |
BRWi (kWh/t) |
BRWi : BBWi |
BBWi CLASSIF. |
MAIN FRESH |
10.53 |
9.37 |
0.89 |
MEDIUM |
MAIN TRANSITION |
8.61 |
8.06 |
0.94 |
SOFT |
NORTH FRESH |
13.39 |
14.1 |
1.05 |
MEDIUM |
NORTH TRANSITION |
6.88 |
4.05 |
0.59 |
SOFT |
|
Table 13-5: |
Summary of AMCT And JKTech Drop Weight
Tests |
SAMPLE |
CBi |
UCS |
UCS
CLASSIF IC. |
JKMRC
PARAMETERS |
A |
b |
Axb |
Ta |
MAIN FRESH |
9.8+1.8 |
161.29 |
MEDIUM |
88.6 |
0.68 |
60.2 |
0.69 |
MAIN TRANSITION |
9.41+0.99 |
22.63 |
SOFT |
55.1 |
2.49 |
137.2 |
0.96 |
NORTH FRESH |
9.08+0.78 |
177.15 |
MEDIUM |
79.4 |
0.55 |
43.7 |
0.15 |
NORTH TRANSITION |
6.54+0.7 |
10.56 |
SOFT |
65.6 |
5.22 |
342.4 |
3.38 |
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The four ores were characterised into
two groups; main fresh and north fresh ores were considered medium hardness and
main transition and north transition ores were soft. The results showed that the
ores would not be amenable to fully autogenous milling. The BBWi:BRWi was less
than 1.1, indicating that the ore has a tendency to break down into a size that
can easily be handled by a secondary ball mill.
The bond crushability index results
showed that all ore types have low crushability indices and are characterised as
very soft (values <10kWh/t).
The low UCS values that were recorded
for main transition and north transition ores indicated that they are soft and
will fracture easily. Main fresh and north fresh ores were more competent and
were classified as hard ores.
The main parameters from the JKTech
testing, A*b, was consistent with the rest of the tests conducted and indicated
the fresh main and north ores were more competent than the transition ores.
Comminution testwork results were
forwarded to OMC for ore interpretation and modelling and was used for the
design of the comminution circuit for Twangiza.
A summary of OMCs interpretation,
findings and recommendations are as follows and detailed information can be
found in the report by OMC referenced January, 2009, Orway Mineral Consultants
(OMC), Twangiza Gold Project Comminution Circuit Design, Report No. 8282:
The ore interpretation showed that the
ore had a low competency and showed lower fracture energies when compared to the
database. This meant that Twangiza ore will not provide good lump media for AG
and SAG grinding;
A combination of the JK parameters and
OMC power modelling demonstrated that the North Fresh ore deposit exhibited the
highest SAG and ball mill specific grinding energies and was used as the
limiting factor for the design of the circuit;
OMC suggested that a SABC (SAG-Ball
mill-Crushing circuit) and SS SAG (Single stage SAG mill) comminution circuit
designs should be considered. The power at the pinion required for both options
is as follows:
|
|
SABC (4716 + 5070 kW) 9 786 kW |
|
|
|
|
|
SS SAG 9 900 kW |
Higher throughputs will be expected
when treating the oxide and transition material as they are softer. The maximum
throughput will be limited by the total volume flow in the circuit as excessive
water will be required to overcome the viscosity issue for the oxide ore;
The throughputs in the SABC circuit
selection will be from 505 700tph and 505 625t/h in the SS SAG circuit;
The SABC mill sizes selected were as
follows:
|
|
SAG mill 8.5mØ x 4.45mEGL (Grate discharge,
variable speed 6.2MW motor) |
|
|
|
|
|
Ball mill 6.1mØ x 9.07mEGL (Over-flow
discharge, fixed speed 6.2MW motor) |
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The mill size in the SS SAG circuit was
9.75mØ x 6.38mEGL (Grate discharge, fitted with a 12MW variable speed drive).
The liner and ball consumptions were as
follows as shown in Table 13-6.
|
Table 13-6: |
Liner and Ball Consumptions
|
ITEM |
UNIT |
SABC |
SS SAG |
SAG MILL |
|
|
|
BALL CONSUMPTION |
kg/t milled |
0.389 |
0.739 |
LINER CONSUMPTION |
kg/t milled |
0.077 |
0.146 |
BALL MILL |
|
|
|
BALL CONSUMPTION |
kg/t milled |
0.538 |
|
LINER CONSUMPTION |
kg/t milled |
0.071 |
|
Recommendations raised by OMC are shown
below:
|
|
The grinding circuit mill sizes should be reviewed once
the preferred circuit configuration has been selected and expected feed
blend has been defined from the mining schedule; |
|
|
|
|
|
The milling turndown scenarios need to be
investigated when softer oxide and transition ores are treated; and |
|
|
|
|
|
Viscosity modifiers should be considered when
treating oxide ore to ensure that higher milling densities can be used.
|
13.5.4 |
Recovery variability
testing |
The aim of the variability tests was to
establish the degree of variability within the ore zones identified with respect
to their metallurgical response using the optimum conditions determined during
the scoping and pre-feasibility testwork phases as listed below.
Oxides Grind of 80% -75µm, 24 hour
leach time, pH of 10.5 and cyanide addition of 1kg/t;
Transition and Fresh - Grind of 80%
-75µm, 24 hour leach time, pH of 10.5 and cyanide addition of 2kg/t.
A summary of the variability results is
shown below and the detailed reports containing the results is contained in the
following reports:
|
|
November 20, 2008, SGS, Metallurgical Test
Programme: Senet Project Twangiza Oxide, Report No. Met07/BB27; and
|
|
|
|
|
|
February 10, 2009, SGS, Metallurgical Test
Programme: Senet Project Twangiza Non- Refractory Variability Samples,
Report No. Met07/BB27. |
The results obtained during the
variability tests showed that the Twangiza ore is variable in terms of gold
recovery.
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Oxide variability testwork
Gravity and leach variability testwork
was performed during the prefeasibility and results showed that the samples were
highly variable obtaining recoveries less than 90% which was originally obtained
during the optimisation testwork phase. The lime consumptions were also very
high. This led to SENET requesting more samples (main and north) for repeat
variability tests to increase the confidence level of the results obtained. A
summary of the results is shown below in Table 13-7 and Table 13-8.
|
Table 13-7: |
Main Oxide Variability Results
|
SAMPLE ID
|
GRG
(g/t) |
GRAV
MASS PULL (%) |
GRAV
RECOV (%) |
CIL
FEED GRADE (g/t) |
NaCN
CONS. (kg/t) |
CaO
CONS. (kg/t) |
CIL
DISSOL. (%) |
OVER.
GOLD RECOV. (%) |
#68 |
2.15 |
0.07 |
2.8 |
2.14 |
0.56 |
4.48 |
76.5 |
77.1 |
#69 |
5.49 |
0.04 |
1.2 |
5.59 |
0.65 |
5.14 |
86.6 |
86.8 |
#70 |
5.17 |
0.07 |
1.2 |
5.29 |
0.59 |
4.91 |
83.9 |
84.1 |
#71 |
3.46 |
0.06 |
3.4 |
3.54 |
0.48 |
4.98 |
97.1 |
97.2 |
#72 |
2.70 |
0.05 |
19.4 |
2.20 |
0.65 |
6.01 |
87.6 |
90.0 |
#73 |
3.55 |
0.07 |
1.8 |
3.62 |
0.56 |
5.58 |
90.4 |
90.5 |
#74 |
3.08 |
0.05 |
4.1 |
2.99 |
0.54 |
2.91 |
94.8 |
95.0 |
#75 |
2.43 |
0.06 |
1.3 |
2.48 |
0.68 |
4.04 |
93.9 |
94.0 |
#76 |
4.23 |
0.04 |
2.6 |
4.27 |
0.71 |
3.74 |
92.6 |
92.8 |
#77 |
6.70 |
0.07 |
3.1 |
6.46 |
0.85 |
5.99 |
73.5 |
74.3 |
|
Table 13-8: |
North Oxide Variability Tests
|
SAMPLE ID |
GRG (g/t) |
GRAV MASS
PULL (%) |
GRAV RECOV
(%) |
CIL FEED
GRADE (g/t) |
NaCN CONS.
(kg/t) |
CaO CONS.
(kg/t) |
CIL
DISSOL. (%) |
OVER. GOLD
RECOV. (%) |
#78 |
0.54 |
0.05 |
7.0 |
0.50 |
0.56 |
3.11 |
93.9 |
94.4 |
#79 |
4.56 |
0.08 |
28.3 |
4.26 |
0.85 |
5.37 |
88.8 |
92.0 |
#80 |
3.65 |
0.08 |
18.7 |
3.06 |
0.48 |
5.10 |
84.7 |
87.6 |
#81 |
3.29 |
0.09 |
19.3 |
2.78 |
0.83 |
5.88 |
84.9 |
87.8 |
#82 |
2.26 |
0.04 |
42.7 |
1.35 |
0.54 |
4.80 |
89.7 |
94.1 |
#83 |
1.56 |
0.05 |
27.9 |
1.01 |
0.48 |
3.28 |
57.5 |
69.3 |
#84 |
2.44 |
0.05 |
12.2 |
2.24 |
0.45 |
5.80 |
78.2 |
80.8 |
#85 |
11.66 |
0.06 |
25.8 |
8.76 |
0.83 |
6.24 |
91.5 |
93.7 |
#86 |
2.61 |
0.09 |
2.4 |
2.76 |
0.65 |
6.07 |
82.4 |
82.8 |
#87 |
11.13 |
0.05 |
13.9 |
9.67 |
0.62 |
5.44 |
85.7 |
87.7 |
The cyanide consumption decreased from
the prefeasibility testwork and the lime consumption varied from 2.91 to 6.01
kg/t which was high but comparable to what was previously obtained. On average
the recoveries were similar to the scoping and optimisation results. The varying
range in the dissolutions obtained demonstrated that this ore is highly variable
which could be an indication of the presence of some refractory material in the
oxides. This applied to both the main and north ore bodies.
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Transition and fresh variability
testwork
The testwork was performed on the
initial samples (labelled sack) that were sent together with additional
samples requested to confirm results, recoveries and reagent consumptions.
Recovery variability testwork was performed on fourteen main transition and
fresh ore samples. Only one fresh sample was tested due to sample availability
and eight north transition and five north fresh samples underwent leach
variability testwork
The results showed that both the main
and north ore bodies are highly variable with regards to reagent consumptions
and gold dissolution. This could be due to the presence of refractory transition
and fresh material in the samples that were submitted as was noticed with the
oxides. The results are summarised in Table 13-9 and Table 13-10Table 13-9.
|
Table 13-9: |
Main Transitional and Fresh Variability
Results |
SAMPLE ID |
CIL
FEED GRADE (g/t) |
NaCN
CONS. (kg/t) |
CaO
CONS. (kg/t) |
CIL
DISSOL. (%) |
MAIN TRANSITION SACK 8 |
6.80 |
1.56 |
2.69 |
89.9 |
MAIN TRANSITION SACK 9 |
2.06 |
0.98 |
7.46 |
73.5 |
MAIN TRANSITION SACK 10 |
4.32 |
1.42 |
7.42 |
82.2 |
MAIN TRANSITION SACK 11 |
4.82 |
1.56 |
4.22 |
64.7 |
MAIN TRANSITION SACK 12 |
2.90 |
1.56 |
2.86 |
63.0 |
MAIN TRANSITION SACK 23 |
2.47 |
1.25 |
3.19 |
55.7 |
MAIN TRANSITION SACK 24 |
1.54 |
1.27 |
2.86 |
77.0 |
MAIN TRANSITION SACK 25 |
2.69 |
1.56 |
2.52 |
75.0 |
MAIN TRANSITION SACK 26 |
3.14 |
0.64 |
2.42 |
85.0 |
MAIN TRANSITION SACK 27 |
1.90 |
0.72 |
1.24 |
95.3 |
MAIN TRANSITION 56 |
0.91 |
0.84 |
1.55 |
83.7 |
MAIN TRANSITION 57 |
3.08 |
0.98 |
1.68 |
76.6 |
MAIN TRANSITION 58 |
3.37 |
1.13 |
2.78 |
68.0 |
MAIN FRESH 59 |
0.92 |
0.69 |
0.95 |
69.4 |
|
Table 13-10: |
North Transitional and Fresh Variability
Results |
SAMPLE ID |
CIL
FEED GRADE (g/t) |
NaCN
CONS. (kg/t) |
CaO
CONS. (kg/t) |
CIL
DISSOL. (%) |
NORTH TRANSITIONAL SACK 38 |
1.51 |
0.62 |
1.39 |
92.1 |
NORTH TRANSITIONAL SACK 39 |
3.62 |
0.71 |
1.09 |
79.5 |
NORTH TRANSITIONAL SACK 40 |
10.35 |
0.27 |
1.19 |
96.6 |
NORTH TRANSITIONAL SACK 41 |
3.56 |
0.71 |
0.79 |
92.1 |
NORTH TRANSITIONAL SACK 42 |
0.85 |
0.71 |
0.72 |
98.0 |
NORTH TRANSITION 60 |
0.73 |
0.65 |
3.11 |
89.1 |
NORTH TRANSITION 61 |
1.29 |
0.56 |
1.91 |
91.0 |
NORTH TRANSITION 62 |
2.72 |
0.68 |
3.15 |
81.8 |
NORTH FRESH 63 |
1.12 |
0.61 |
1.38 |
78.9 |
NORTH FRESH 64 |
1.50 |
0.67 |
1.35 |
78.4 |
NORTH FRESH 65 |
2.81 |
0.78 |
1.85 |
86.7 |
NORTH FRESH 66 |
1.36 |
0.58 |
1.35 |
79.4 |
NORTH FRESH 67 |
2.74 |
0.72 |
1.51 |
92.2 |
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13.5.5 |
Transition CMS and fresh CMS refractory ore
testwork |
Head analysis transition CMS and
fresh CMS refractory ores
Head analyses were completed on the
Twangiza Transition CMS and the Fresh CMS composite samples. They were fire
assayed for Au and Ag. The average gold content of the Transition CMS was
2.23g/t and 2.65 g/t for the fresh CMS. A portion of the head split was assayed
semi-quantitatively using ICP for an additional 21 elements. The two samples
were also analysed for total carbon, carbon speciation, total sulphur, sulphur
speciation, arsenic and mercury. The multi-elemental analysis indicated that
both the transition CMS and fresh CMS samples were high in arsenic (0.46% and
0.93%, respectively). The specific gravity (SG) of the two ore types were 2.82
and 2.99 for the transition and fresh ores respectively. The sulphur speciation
assays showed that the fresh CMS ore body is a massive sulphide and the
transition CMS is oxidised and still contains a significant amount of sulphide
mineralisation. The carbon analysis illustrated that there are large amounts of
preg-robbing material that have to be removed upfront during flotation prior to
the leach. A pre-float step for graphitic carbon removal was therefore included.
The results are summarised in Table 13-11 and Table 13-12.
|
Table 13-11: |
Head Assay Carbon and Sulphur Speciation
|
ELEMENT
|
UNIT
|
FRESH
CMS |
TRANSITION CMS |
S (t) |
% |
10.20 |
5.61 |
S (2-) |
% |
9.57 |
4.53 |
S (0) |
% |
<0.5 |
<0.5 |
C (t) |
% |
1.82 |
1.37 |
CO3 |
% |
2.11 |
0.76 |
C (org) |
% |
1.40 |
1.22 |
C (graph) |
% |
0.09 |
0.08 |
The flotation testwork was divided into
the following stages:
|
|
Scouting tests; |
|
|
|
|
|
Grind optimisation; |
|
|
|
|
|
Reagent optimisation; |
|
|
|
|
|
Rougher rate tests; |
|
|
|
|
|
Flotation test at optimum conditions; |
|
|
|
|
|
Bulk concentrates generation for leach
testwork. |
Scouting testwork
Scouting testwork was performed using a
generic reagent suite used during the early stage of the testwork, to determine
the mass pulls and recoveries that can be obtained to start the optimisation
testwork. The results from the scouting testwork were not in line with what was
obtained in the Pre-Feasibility Study as higher recoveries and mass pulls were
expected. The Fresh CMS gold recovery was above 90 % at a mass pull of 30 %. A
gold recovery of 75 % was achieved for Transition CMS at a mass pull of 15%.
Optimisation testwork followed to maximise recovery.
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|
Table 13-12: |
Full Elemental Analysis
|
ELEMENT |
UNIT
|
FRESH
CMS |
TRANSITION CMS |
Au |
g/t |
2.65 |
2.23 |
Ag |
g/t |
<1 |
<1 |
Hg |
ppm |
<3 |
<3 |
As |
% |
0.93 |
0.46 |
Al |
% |
6.20 |
5.80 |
Ba |
ppm |
140 |
210 |
Be |
ppm |
<1 |
<1 |
Ca |
% |
0.76 |
0.16 |
Cd |
ppm |
<10 |
<10 |
Co |
ppm |
32 |
32 |
Cu |
ppm |
<10 |
<10 |
Fe |
% |
10.9 |
7.9 |
K |
% |
0.24 |
0.32 |
Li |
ppm |
<10 |
<10 |
Mg |
% |
0.42 |
0.14 |
Mn |
% |
220 |
210 |
Mo |
ppm |
<10 |
<10 |
Na |
% |
3.9 |
3.5 |
Ni |
ppm |
52 |
54 |
P |
ppm |
310 |
330 |
Pb |
ppm |
33 |
<30 |
Sn |
ppm |
<20 |
<20 |
Sr |
ppm |
15 |
13 |
V |
ppm |
51 |
60 |
Y |
ppm |
13 |
17 |
Zn |
ppm |
16 |
41 |
Zr |
ppm |
100 |
150 |
Cr |
% |
<0.02 |
<0.02 |
Si |
% |
23 |
25 |
Ti |
% |
0.34 |
0.39 |
Two bulk composites, transition CMS and
fresh CMS, were made up and each crushed to -1.7mm, using a large scale
laboratory jaw crusher, and blended. These samples were used for the flotation
and leach recovery testwork shown below.
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Grind optimisation
The effect of flotation feed grind size
on metallurgical performance for each of the composites is presented in Figure
13-1. The summarised results are shown in Table 13-13 and Table 13-14. The
optimum grind chosen was 70% -75 µm. Grinding finer increased the mass pull but
recovery decreased for the fresh CMS. A slight improvement in gold recovery for
the transition CMS material was noted with a finer grinding from 70% to 80% -75
µm.
|
Table 13-13: |
Grind Optimisation Results Summary -
Fresh |
GRIND -75µm |
MASS PULL
(%) |
Au (g/t) |
S (%)
|
S2-
(%)
|
Au REC.
(%) |
S REC. (%)
|
S2
- REC. (%) |
60% |
27.95 |
10.14 |
28.00 |
24.54 |
89.94 |
88.94 |
97.90 |
70% |
32.58 |
7.53 |
29.80 |
22.14 |
93.34 |
90.97 |
89.17 |
80% |
36.93 |
6.45 |
32.40 |
19.00 |
91.29 |
91.34 |
86.88 |
90% |
36.91 |
6.32 |
30.23 |
13.87 |
91.13 |
86.04 |
78.44 |
|
Table 13-14: |
Grind Optimisation Results Summary -
Transition |
GRIND -75µm |
MASS PULL (%) |
Au (g/t) |
S (%) |
S2-
(%) |
Au REC. (%) |
S REC. (%) |
S2 -
REC. (%) |
60% |
14.15 |
14.46 |
30.20 |
24.90 |
76.8 |
74.55 |
80.41 |
70% |
15.00 |
14.33 |
30.39 |
27.90 |
77.35 |
80.85 |
91.62 |
80% |
16.40 |
12.29 |
35.75 |
24.02 |
79.02 |
85.18 |
87.88 |
90% |
19.93 |
9.49 |
22.38 |
20.95 |
78.68 |
84.01 |
86.27 |
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Reagent optimisation
Reagent optimisation testwork was
performed on fresh CMS by varying the collector and secondary collector addition
and keeping the copper sulphate (activator) the same at 100g/t at natural pH. It
was found that a secondary collector had to be used to maximise the recovery of
all sulphides. The results for the fresh CMS reagent optimisation tests by
varying the collector addition are shown in Table 13-15. An increase in the
collector addition increased the mass pull and improved the gold recovery
slightly.
|
Table 13-15: |
Reagent Optimisation at Natural pH
|
REAGENTS |
MASS PULL
(%) |
Au
RECOVERY (%) |
S2- RECOVERY (%)
|
100g/t PAX and 50g/t Aero407 |
34.82 |
88.86 |
89.45 |
200g/t PAX and 100g/t Aero407 |
39.44 |
90.99 |
91.64 |
300g/t PAX and 150g/t Aero407 |
45.87 |
93.08 |
95.86 |
400g/t PAX and 200g/t Aero407 |
46.70 |
93.85 |
95.59 |
Tests were performed at acidic pH,
Table 13-16, to determine the effect it will have on the mass pull and recovery
when the collector addition was reduced. It was found that the reduction in the
collector addition adversely affected gold recovery. More acidic conditions were
detrimental to both gold and sulphide recovery.
Further reagent optimisation scouting
testwork was performed to determine the impact that different collectors would
have on mass pull and recovery. The best results showed that PAX will have to be
used with a secondary collector, Aero 3302, Table 13-17.
|
Table 13-16: |
Reagent Optimisation at Acidic pH
|
REAGENTS |
MASS PULL
(%) |
Au
RECOVERY (%) |
S2- RECOVERY (%)
|
200g/t PAX and 100g/t Aero407 |
33.12 |
92.24 |
92.58 |
100g/t PAX and 50g/t Aero407 |
36.59 |
90.51 |
91.98 |
50g/t PAX and 25g/t Aero407 |
32.37 |
86.30 |
89.84 |
|
Table 13-17: |
Reagent Optimization by Varying Collector
Type |
REAGENTS |
MASS PULL
(%) |
Au
RECOVERY (%) |
S2-
RECOVERY (%) |
200g/t PAX and 100g/t Aero 3418A |
36.16 |
91.33 |
91.79 |
200g/t PAX and 100g/t Senkol 295 |
37.49 |
93.6 |
92.83 |
200g/t PAX and 100g/t Aero 3302 |
38.39 |
95.00 |
92.38 |
200g/t Aero 3418A only |
34.94 |
92.94 |
90.63 |
200g/t PAX and 100g/t Aero 407 |
41.59 |
92.75 |
91.83 |
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Rougher rate tests
A rougher rate test was performed using
the optimum reagent suite on the fresh and transition CMS and it was found that
most of the gold was recovered in the first 26 minutes. A total gold recovery of
93% was obtained for fresh ore and was in line with the recoveries obtained
previously. A poor recovery of 82% was achieved for transition ore and will be
reviewed at a later stage by performing mineralogical investigations of the feed
and tailings. It was noted however that the flotation test on the transition CMS
was a success with respect to the sulphide recovery as a recovery of 93% was
achieved. A summary of the results is presented in Table 13-18 and Table
13-19.
|
Table
13-19: |
Table 13-18: Rougher Rate Tests -
Fresh
|
ITEM |
TIME
(min) |
MASS PULL
(%) |
Au
RECOVERY (%) |
S2-
RECOVERY (%) |
SULPHIDE CONCENTRATE 1 |
26 |
35.6 |
88.34 |
85.77 |
SULPHIDE CONCENTRATE 2 |
6 |
2.95 |
3.56 |
6.38 |
SULPHIDE CONCENTRATE 3 |
8 |
1.51 |
0.61 |
1.53 |
SULPHIDE CONCENTRATE 4 |
10 |
0.97 |
0.37 |
0.69 |
CUMULATIVE/TOTAL |
50 |
41.03 |
92.88 |
94.37
|
|
Table 13-19: |
Table 13-18: Rougher Rate Tests -
Transitional
|
ITEM |
TIME (min)
|
MASS PULL
(%) |
Au
RECOVERY (%) |
S2-
RECOVERY (%) |
SULPHIDE CONCENTRATE 1 |
26 |
17.58 |
79.49 |
89.24 |
SULPHIDE CONCENTRATE 2 |
6 |
1.75 |
1.37 |
1.92 |
SULPHIDE CONCENTRATE 3 |
8 |
1.98 |
0.87 |
1.02 |
SULPHIDE CONCENTRATE 4 |
10 |
1.65 |
0.66 |
0.64 |
CUMULATIVE/TOTAL |
50 |
22.96 |
82.39 |
92.82
|
Bulk concentrate generation
A bulk concentrate was generated using
the optimum conditions obtained from all the flotation tests performed to date
for use during the leach optimisation testwork. A summary of the results is
shown in Table 13-20.
|
Table 13-20: |
Bulk Concentration
Results
|
MASS PULL (%) |
Au RECOVERY
(%) |
S2-
RECOVERY (%) |
56 |
95 |
98
|
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13.6 |
Review of Comminution Circuit Parameters to Process
1.7 Mtpa |
SENET requested OMC to undertake a
comminution equipment evaluation for the Twangiza Project.
The circuit configuration for
processing of 1.3 Mtpa Oxide ore using a tertiary crushed feed followed by two
stage ball milling had been reviewed previously. The use of a scrubber to remove
the clay from the primary crushed ore prior to the secondary crushing had been
included.
The following equipment was selected:
|
|
Scrubber - 3.0 m Ø x 8.0 m EGL, 300 kW, |
|
|
|
|
|
Secondary cone crusher Kawasaki KM2513G, 185kW,
|
|
|
|
|
|
Tertiary cone crusher Kawasaki KM1213G, 185kW,
|
|
|
|
|
|
Ball mill 1 - 3.81 m x 5.03 m EGL, 1,300 kW, |
|
|
|
|
|
Ball mill 2 - 3.23 m x 3.04 m EGL, 550 kW.
|
13.6.2 |
Summary of investigation by Orway Mineral
Consultants |
The maximum power achievable with the
selected scrubber is 1.09 kWh/t at a 275 t/h feed rate. This is expected to be
sufficient for most ores. The design is based on the North oxide ore
characteristics (BBWi = 4.3 kWh/t). The North Transition ore is 60% more
competent (BBWi = 6.88 kWh/t) and throughput will be affected by this material.
Modelling showed that the circuit is very sensitive to ore characteristics and
that this should be managed actively during operation. Feed particle size
distributions have been sourced from the OMC database for modelling.
Following the recommendations by OMC
after expansion, a number of key modifications were implemented to counter the
effect of size distribution and that of clay balls.
The grind in the ball mills is
sensitive to the feed size fraction; the coarser the feed, the coarser the grind
which results in a poor recovery. In-pit crushing and screening operations have
been added to the circuit to assist the plant crushing system to obtain -10mm
material to feed the ball mills.
The clay balls proportion depends on
the proportion of clay in the ore body. The deeper mining is done, the less clay
balls are formed and the easier the operations. The Plant crushing system alone
could not cope with the amount of clay balls. With additional capacity provided
by the in-pit crushing and screening, over 50%of clay balls by-pass completely
the Plant crushing system and report directly into the ball mills where they are
dealt with by using a varying proportion of steel balls and dilution rates.
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Active management of the closed side
setting of the crushers and the closing screen panels may be required. Accurate
measurement (including provision for calibration) of the -2mm feed to the mill
sump will be required for process control purposes. The mills are capable of
treating the expansion throughput at the selected design ore characteristics and
still maintain typical design margins. Full details of the review can be found
in the report no. 8623.30 Rev 1 titled Twangiza Gold Project 1.7 Mtpa
Throughput Option Oxide Only by OMC.
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14 |
MINERAL RESOURCE
ESTIMATE |
A Mineral Resource estimate was
originally produced by SRK (UK) for the 2009 Feasibility Study which was
undertaken by Martin Pittuck and Benjamin Parsons of SRK (UK), both of whom
signed as Qualified Persons. A modified version of the same resource model was
provided by Twangiza Mining and reviewed by SRK (UK) for the 2011 Phase 1
Economic Assessment compiled by SENET; Martin Pittuck signed off the resource as
Qualified Person.
The relatively minor modifications to
the resource model since 2009 have been prepared in house by Twangiza Minings
team under the supervision of Daniel Bansah (Head of Projects and Operations)
who is also a Qualified Person. The interim modifications have involved:
|
|
addition of minor deposits (Twangiza West and East);
|
|
|
|
|
|
conversion of Valley Fill deposit from Inferred to
Indicated; |
|
|
|
|
|
creation of diluted grade and recoverable grade
variables; |
|
|
|
|
|
changes to topographic survey to account for mining
depletion; and |
|
|
|
|
|
revisions to the density model in light of production
reconciliation results. |
All modifications have been reviewed
and approved by SRK (UK).
The adit, drill hole and trench data
were originally plotted on level plans and vertical sections; lithological and
mineralisation interpretations were then outlined on vertical sections and
flitch plans by Twangiza Mining supplied to SRK (UK) for creation of 3D
geological and mineralised wireframes.
The 3D model continuity was assessed by
SRK (UK) in Leapfrog mining software (Leapfrog). The mineralised domains were
largely based on a geological cut-off of 0.3 g/t Au but were also formed so as
to ensure continuity between sections and flitch plans. This process was
repeatedly verified until a robust 3D model of the mineralised zones had been
created.
The statistics and geostatistics for
the main deposits were completed in Isatis. Datamine was used for block
modelling and grade estimation. The resource data, survey data and block models
are housed in the Datamine software (Datamine).
14.3 |
Density determinations |
14.3.1 |
Feasibility Study Data collection and
analysis |
The initial bulk density testwork was
performed on behalf of Twangiza Mining by CME in 1998 on a variety of rock
types; a total of 165 samples were assessed by CME. Twangiza Mining subsequently
undertook more density determinations from drill core; a total of 2,031 relative
density determinations were undertaken by Twangiza Mining.
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Twangiza Minings density
determinations were undertaken at the following intervals in all drill holes:
|
|
Every 2 meters outside mineralized zones. |
|
|
|
|
|
Every 1 meter within mineralized zones
|
The following method was employed:
|
|
The geologist selects samples and marks each sample
position with an aluminium tag. The borehole number and depth are also
written on the selected piece of core with a marker pen. In order to avoid
bias when taking samples, the first piece of solid core weighing over 200
g after the meter mark is selected. |
|
|
|
|
|
The depth of each sample and rock type is recorded by the
geologist. |
|
|
|
|
|
The weight of the core sample is recorded (Weight 1)
|
|
|
|
|
|
The sample is checked for porosity by placing the sample
in water and recording the increase in weight over a 3-minute period
(Weight 2). If the sample absorbs more than a gram of water it is treated
as porous. |
|
|
|
|
|
The sample is dried in an oven at 105°C for 30 minutes,
and then re-weighed (Weight 3). This enables the moisture content of the
air dried core to be calculated: |
|
Moisture Content (%) = |
(Weight 1 Weight 3) x 100
|
|
|
Weight 1 |
|
|
If the sample is porous it is coated with varnish using a
brush, ensuring all cavities and irregularities are coated; the dry sample
is weighed whilst suspended in water (Weight 4) allowing the dry density
to be determined: |
|
Dry Density (t/m3) = |
Weight 3 |
|
|
(Weight 3 Weight 4) |
The density determinations were
statistically reviewed according to lithological type and position within the
weathering profile; a summary of block model dry density values used in the
Feasibility Study is given in Table 14-1.
|
Table 14-1: |
Feasibility Study Dry
Solid Densities
|
|
DRY DENSITY
(g/cm3) |
MATERIAL TYPE |
PORPHYRY |
SEDIMENT |
WASTE |
UPPER OXIDE LOWER OXIDE
TRANSIT. (SOFT) TRANSIT. (HARD) FRESH |
1.8 2.15 2.35
2.75 2.85 |
2.1 2.05 2.4
2.65 2.7 |
2 2.15 2.35
2.75 2.85 |
|
14.3.2 |
Pit Density Determinations |
During the course of the reconciliation
study undertaken as part of this technical report, a number of check density
samples were taken from the Twangiza open pit. In summary these comprised 5
samples taken from small pits and 1 bulk sample.
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Pit samples were taken by digging a
hole approximately 15cm by 15cm by 10cm deep to provide a sample volume of 2,400
cm3 on average. The volume of each sample was determined by assessing
the weight of sand required to fill the each sample pit and then calculating the
volume of sand in the pit by dividing the weight of the sand by the known bulk
density of the sand (1.25 g/cm3).
A 100g subsample of each pit sample was
weighted whilst wet and then reweighed after drying overnight in an oven at
105°C; this allowed moisture content to be determined.
Each pit sample was weighed as soon as
it had been extracted; this provided a wet weight which was then corrected by
the moisture content to determine its dry weight. The dry density of each sample
was calculated by dividing the dry weight by the volume of the pit. A summary of
the pit samples is given in Table 14-2.
|
Table 14-2: |
Pit Density
Determinations
|
|
WHITISH GREY MUDSTONE |
REDDISH BROWN MUDSTONE |
No.1 |
No.2 |
No.3 |
No.5 |
No.6 |
Wet Density (g/cm3) Moisture Content (%) Dry Density
(g/cm3) |
2.15 10%
1.92 |
1.88 18%
1.54 |
2.13 11%
1.90 |
1.90 25%
1.43 |
1.95 20%
1.56 |
14.3.3 |
Bulk Sample Density
Determinations |
During the course of the reconciliation
study undertaken as part of this technical report, the Twangiza Mining team used
a bulk sample method to determine density of material routinely mined from the
open pit. A block of ground, approximately 5m by 5m by 2.5m deep was excavated
and hauled directly to the crusher. The excavation was surveyed and determined
to have a volume of 67 m3. The mill weightometer reading was taken
before the trucks tipped this material into the crusher and after the material
had travelled up the belt over the weightometer.
A sub sample of the material was taken
weighed whilst wet and then weighted again after drying to determine the
moisture content.
The weightometer recorded 134 tonnes of
wet material (giving a wet density of 2.0 g/cm3) and this was then
corrected for moisture content (found to be 23.8%) to give a dry density of 1.53
g/cm3.
14.3.4 |
Revised Density Model |
Density in the ore is affected by
weathering intensity, hence modelling upper oxide, lower oxide, soft,
transition, hard transition and fresh weathering layers in the Mineral Resource
model. The recent density work suggests that very close to surface the density
of the ore is lower than previously estimated for the upper oxide.
At Twangiza Main, it is proposed by
Twangiza Mining and SRK that ore density has also been affected by material
collapsing into underground mining voids which were probably accessed from the
CME adits. The large artisanal open pit (Mbwega Pit) was originally thought by
Twangiza Mining and SRK to be the only artisanal mining site and the sole source
of the Valley Fill deposit. Mbwega Pit featured in the topographic survey used
in the Feasibility Study and the material mined from there did not add to any
resource models. However, collapsed underground workings are now thought to
exist, causing zones of broken material with low density but of sufficient bulk
so as not to be recognized as mine workings in the pit.
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The Feasibility Study density sampling
protocol required solid core samples to be taken which is common practice but it
results in low density material being under-represented in the core samples used
to estimate density, this is normally accounted for by rounding down density
estimates in oxide. However, in light of operational information, the bias in
the original estimate is now thought to be more considerable and so has been
specifically addressed in this technical report.
To reflect the most intense weathering
near surface that was under-represented in core density sampling, the block
model within 15m of the original topographic surface was reduced from 2.05
t/m3 to 1.80 t/m3 at Twangiza Main and to 1.89
t/m3 at Twangiza North. This affects approximately quarter to half of
the material in the oxide resource.
To reflect the collapsed material that
was under-represented in core density sampling, a 20% discount was applied to
oxide ore blocks in the part of the Twangiza Main Pit closest to the CME adits,
which is where underground mining is assumed to have taken place. This affects
approximately 15% of the Mineral Reserve.
The value of 20% is based on the
reconciliation study in which a historical tonnage shortfall was recorded at the
plant. Many possible explanations were considered, not all were studied through
to a satisfactory conclusion but the decision was taken to assume the balance of
adjustments should be attributed ore density.
The resultant adjusted ore densities
are given in Table 14-3.
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|
Table 14-3: |
Summary of
2015 Revised Dry Density Values
|
MATERIAL TYPE |
Resource
Category |
MAIN
DENSITY (g/cm3) |
NORTH
DENSITY (g/cm3) |
ALL
HARDNESS (MPa) |
PORPHYRY |
SEDIMENT |
WASTE |
PORPHYRY |
SEDIMENT |
WASTE |
PORPHYRY |
SEDIMENT |
WASTE |
UPPER 15m OXIDE |
Measured |
1.45 |
1.45 |
1.45 |
|
|
|
|
|
|
|
Indicated & Inferred
|
1.80 |
1.80 |
1.80 |
1.89 |
1.89 |
1.89 |
15 |
15 |
15 |
UPPER OXIDE |
Measured |
1.45 |
1.69 |
1.61 |
1.80 |
2.10 |
2.00 |
15 |
15 |
15 |
Indicated & Inferred
|
1.80 |
2.10 |
2.00 |
1.80 |
2.10 |
2.00 |
15 |
15 |
15 |
LOWER OXIDE |
Measured |
1.73 |
1.65 |
1.73 |
2.15 |
2.05 |
2.15 |
35 |
35 |
35 |
Indicated & Inferred
|
2.15 |
2.05 |
2.15 |
2.15 |
2.05 |
2.15 |
35 |
35 |
35 |
TRANSIT. (SOFT) |
Measured |
1.89 |
1.93 |
1.89 |
2.35 |
2.40 |
2.35 |
60 |
60 |
60 |
Indicated & Inferred
|
2.35 |
2.40 |
2.35 |
2.35 |
2.40 |
2.35 |
60 |
60 |
60 |
TRANSIT. (HARD) |
Measured |
2.21 |
2.13 |
2.21 |
2.75 |
2.65 |
2.75 |
175 |
175 |
175 |
Indicated & Inferred
|
2.75 |
2.65 |
2.75 |
2.75 |
2.65 |
2.75 |
175 |
175 |
175 |
SULPHIDE |
Measured |
2.29 |
2.17 |
2.29 |
2.85 |
2.70 |
2.85 |
210 |
210 |
210 |
Indicated & Inferred
|
2.85 |
2.70 |
2.85 |
2.85 |
2.70 |
2.85 |
210 |
210 |
210
|
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14.4 |
Descriptive statistics of assay
data |
Summary statistics have been calculated
for the different sampling phases and have been compiled in Table 14-4 below.
The results of the analysis were compared for similarities and differences to
determine whether the assay values could be combined for the modelling and
estimation processes.
|
Table 14-4: |
Summary of Raw
Statistics per Sampling Phase
|
DESCRIPTION |
NO |
MIN
(g/t Au) |
MAX
(g/t Au) |
MEAN
(g/t Au) |
VAR. |
CoV |
RECENT ( 2011
2014) DD HOLES |
2,658 |
0.005 |
18.8 |
0.19 |
0.61 |
4.14 |
RECENT ( 2009 2010) DD HOLES |
40,708 |
0.01 |
63.4 |
0.81 |
4.26 |
2.56 |
RECENT (PHASE FS INFILL) DD HOLES |
21,148 |
0.01 |
58.4 |
0.78 |
4.93 |
2.86 |
RECENT (PHASE 3 PFS) DD HOLES |
24,534 |
0.01 |
121.00 |
0.36 |
3.75 |
5.44 |
FOLLOW-UP DRILL HOLE SAMPLES |
25,677 |
0.01 |
280.00 |
0.70 |
7.74 |
3.95 |
1997/98 DRILLING SAMPLES |
8,884 |
0.01 |
310.28 |
0.95 |
16.79 |
4.31 |
ALL TRENCH |
4,660 |
0.01 |
93.80 |
2.61 |
15.15 |
1.49 |
1997/98 ADIT SAMPLES |
1,610 |
0.01 |
52.44 |
3.06 |
18.95 |
1.42 |
HISTORICAL ADIT SAMPLES |
13,328 |
0.01 |
157.70 |
3.56 |
25.61 |
1.42 |
The mean values of the data presented
in the table above indicate that the trench and adit samples have significantly
higher mean grades, which is due to the fact that they are preferentially
located in the oxide material. The drill hole data which tends to have been
lower mean grades have been drilled into all four material types (upper oxide,
lower oxide, transition and sulphide) but preferentially into the transition and
sulphide material and include both ore and waste sampling. More recent drilling
has a lower average grade owing to targeting the narrower and deeper
intersections in Twangiza North thus incurring a greater proportion of waste
samples.
14.5 |
Geological Modelling |
14.5.1 |
Geological wireframes |
In the Feasibility Study, metal
recoveries from different rock units were highly variable, especially within the
transitional zone; there were good recoveries from the porphyry material and
lower recoveries in the sediment material. With this in mind, care was taken to
model the lithologies in the resource block model.
The geological aspects considered
during interpretation were lithological and structural with respect to
anti-formal axis and some minor offset faulting. SRK (UK) reviewed the
geological data and concluded that the following geological factors should be
included in the 3D block model:
|
|
Separation of the rock-codes into porphyry and
sediment rock types; |
|
|
|
|
|
Construction and extension of fault planes
which impact on the mineralisation cutting the deposit; |
|
|
|
|
|
Creation of oxidation surfaces to include upper
and lower oxide, soft and hard transitional material and top of fresh
rock; |
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Evidence of a synclinal structure to
the east of the pit wall between X 693,525 693,700 and Y 682,800 683,000,
which due to its proximity to the potential final pit limits, will influence the
geotechnical issues.
The average distance between section
lines is approximately 40 50 m within the main drilling increasing to
approximately 150 m at the edge of the deposits. To ensure waste blocks were
assigned the correct oxidation code for mine planning, interpretations were
extended to the east and west beyond the current drilling information.
The anticline hinge zone plunges to the
south and the interpretation progresses northwards through the deposit the
mineralised part of the hinge zone daylights above the current topography. At
this point the interpretation has been changed from a single unit to separate
east and west mineralised limbs of the anticline. In general, the anticline
becomes more isoclinal with depth.
The location of North East South West
trending faults and East West trending faults in the northern portion of the
Twangiza Main deposit plays an important role in the structure of the anticline.
Interpretation of the faults was supplied by Twangiza Mining and extended to
intersect the main geological model by SRK (UK). SRK (UK) also imported bedding
measurements into the database in an attempt to improve the interpretation of
the dips within the porphyry units.
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14.5.2 |
Mineralisation wireframe |
The mineralised domain perimeters were
defined in the light of the available geological knowledge and using a 0.3 g/t
Au iso-surface. Plan interpretations created by Twangiza Mining were also used
to colour code the different mineralised lodes. Depth extrapolation beyond the
limits of drilling was carried out to ensure consistency in shape and
orientation, this is reflected in the classification of resource.
To ensure continuity in the mineralised
interpretation Leapfrog has been utilised to identify consistent mineralised
lenses. Semi-variograms and search ellipses have been generated which are
rotated into the general dip and strike of the deposit. Different anisotropies
were tested until the most continuous shape was created. The leapfrog wireframe
created agreed with and improved the historic interpretations of the mineralised
zone and therefore was used as a guideline during the digitisation of the
mineralised strings on each vertical drilling section. This technique provides
increased confidence in the potential ore / waste contacts between drill holes
and in areas of relatively low sampling density.
The subdivision of deposit into zones
is based on observed consistency of mineralisation. Two different styles of
mineralisation are present with the Twangiza deposit and can be summarised using
the following characteristics:
The Twangiza Main deposit is up to 300
m wide in places, width increases towards surface within the transitional and
oxide zones. At depth the mineralisation zones tend to be narrower and contain
zones of low grade material (i.e. less than 0.5 g/t Au).Twangiza East, West and
main extension mineralisation style is same as that of the Main.
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The East and Western flanking
structures are considered as part of the eastern and western limbs of the main
Twangiza anticlinal structure as they bear similar characteristics.
The Twangiza North deposit comprises
thinner zones (10 30 m wide) which are mainly confined to the porphyry units
and which have sharp contacts between ore and waste material.
The mineralisation contact is more
distinct at Twangiza North than the more diffuse distribution at Twangiza Main,
therefore a hard boundary at 0.4 g/t Au has been used at Twangiza North. In the
upper portions of the Twangiza Main zone the contacts are more difficult to
define and the mineralisation appears to be wider. A statistical analysis of the
alteration pattern associated with the mineralisation and general visual
appraisal suggests a natural cut-off at 0.3 to 0.5 g/t Au.
The Twangiza North style of
mineralisation is a series of thinner zones. The main characteristics of the
Twangiza North style ore bodies are the strong relationship between the
mineralised material and the lithology. The close relationship again stresses
the requirement for an accurate geological model as changes in the geology could
cause changes in the mineralisation model. The Twangiza North style
mineralisation begins to the north of a faulted zone (northing 682800 N), beyond
which mineralisation is limited to the porphyry. The figure above shows the
current interpretation of the geometry and situation of each of the mineralised
domains.
The mineralised domains were assigned a
numeric code. Attempts were made to follow the strike trends of the major zones
and to limit the number of mineralised zones (lodes). The main mineralisation
zone has been defined as domain 2. The domains were stored in the model under
the field KZONE as summarised in Table 14-5 below.
|
Table 14-5: |
Summary of Kriging Zones
Used in the Latest Block Model
|
KZONE |
DESCRIPTION
|
KZONE 1 |
UPPER OXIDE (CAP) |
KZONE 2 |
MAIN MINERALISATION ZONE |
KZONE 3 - 6 |
SECONDARY MINERALISED ZONES
|
14.5.3 |
Geological block model |
A 10 x 10 x 5 m prototype parent block
was created with sub-blocking allowed along the boundaries to a minimum of 2.5 m
along strike and 1 m across strike and in the vertical direction. Further
sub-blocking was used at surface.
|
Table 14-6: |
Details of Block Model
Dimensions For Geological Model
|
BLOCK EDGE |
ORIGIN |
BLOCK SIZE |
NO OF BLOCKS
|
MIN X |
692,000 |
10 |
250 |
MIN Y |
681,500 |
10 |
450 |
MIN Z |
1,500 |
5 |
220 |
Using the wireframes described above, a
series of codes were developed to describe each of the major geological
properties of the rock types; these were used for the pit optimisation exercise.
Table 14-7 summarises additional fields created within the geological model and
the codes used.
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Table 14-7: |
Summary of Fields Used
For Flagging Different Geological Properties
|
FIELD NAME |
CODE |
DESCRIPTION
|
LITH1 |
100 |
PORPHYRY |
|
101 |
SEDIMENT |
OXTRAN |
1 |
UPPER OXIDE |
|
2 |
LOWER OXIDE |
|
3 |
TRANSITIONAL |
|
4 |
FRESH ROCK |
SRKZONE |
1 |
TWANGIZA MAIN |
|
2 |
TWANGIZA NORTH |
HARD |
1 |
SOFT |
|
2 |
HARD |
HARDNESS |
VARIABLE (MPa) |
AS DESCRIBED IN Table 14-3
|
Further alphanumeric codes described in
Table 14-8 and Table 14-9 were stored in the new field ROCK. The model is
split into two components based on a northing of 682,800 N, marking the change
from Twangiza Main type mineralisation to Twangiza North mineralisation.
|
Table 14-8: |
Rock Codes Used in
Twangiza Main
|
|
MEASURED |
INDICATED |
INFERRED |
|
PORPHYRY |
SEDIMENT |
PORPHYRY |
SEDIMENT |
PORPHYRY |
SEDIMENT |
OXIDE 1 |
1001 |
1011 |
1002 |
1012 |
1003 |
1013 |
MMPS |
MMSO1 |
MMSO1 |
MDSO1 |
MFPO1 |
MFSO1 |
OXIDE |
1021 |
1031 |
1022 |
1032 |
1023 |
1033 |
MMPO |
MMSO |
MDPO |
MDSO |
MFPO |
MFSO |
TRANSITION |
1041 |
1051 |
1042 |
1052 |
1043 |
1053 |
MMPT |
MMST |
MDPT |
MDST |
MFPT |
MFST |
SULPHIDE |
1061 |
1071 |
1062 |
1072 |
1063 |
1073 |
MMPS |
MMSS |
MDPS |
MDSS |
MFPS |
MFSS |
|
Table 14-9: |
Rock Codes Used in
Twangiza North
|
|
MEASURED |
INDICATED |
INFERRED |
|
PORPHYRY |
SEDIMENT |
PORPHYRY |
SEDIMENT |
PORPHYRY |
SEDIMENT |
OXIDE 1 |
2001 |
2011 |
2002 |
2012 |
2003 |
2013 |
NMPO1 |
NMSO1 |
NDPO1 |
NDSO1 |
NFPO1 |
NFSO1 |
OXIDE |
2021 |
2031 |
2022 |
2032 |
2023 |
2033 |
NMPO |
NMSO |
NDPO |
NDSO |
NFPO |
NFSO |
TRANSITION |
2041 |
2051 |
2042 |
2052 |
2043 |
2053 |
NMPT |
NMST |
NDPT |
NDST |
NFPT |
NFST |
SULPHIDE |
2061 |
2071 |
2062 |
2072 |
2063 |
2073 |
NMPS |
NMSS |
NDPS |
NDSS |
NFPS |
NFSS
|
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14.6 |
Topography, oxide/transition
sub-models |
A LIDAR survey was used to create a
digital terrain model within Datamine to represent the original topography which
included the artisanal open pit in the oxide zone of the Twangiza Main
deposit.
This has been updated by the end
December 2014 mine survey to account for all depletion to date.
The oxide and transitional models were
created by linking the cross-sectional interpretations to form single surfaces
within Datamine. These surfaces have been used during the zoning process of
samples and in creating the final block model.
14.7 |
Statistical analysis of the mineralised
data |
14.7.1 |
Selection of composite lengths for
statistics |
Sample lengths are varied as shown in
Figure 14-5. To ensure no bias exists in the compilation of the statistics and
geostatistics a standard composite 2.0 m length was used.
14.7.2 |
Summary statistics and
histograms |
Each of the individual zones was
assessed independently as illustrated in Table 14-10 from the summary
statistics. On reviewing the summary statistics SRK (UK) took the decision to
combine the different deposit areas and to only subdivide zones based on the
oxidation state.
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Table 14-10: |
Summary Statistics of
2m Composites
|
FIELD |
N |
MIN |
MAX |
MEAN |
VAR |
CoV |
DESCRIPTION |
FIELD |
Au |
28,487 |
0.01 |
157.82 |
2.51 |
14.28 |
1.50 |
ALL 2m SAMPLES |
Au |
Au |
10,028 |
0.01 |
93.12 |
3.11 |
15.99 |
1.29 |
UPPER OXIDE |
Au |
Au |
4,772 |
0.01 |
69.00 |
3.13 |
15.62 |
1.26 |
LOWER OXIDE |
Au |
Au |
4,544 |
0.01 |
157.82 |
2.78 |
23.87 |
1.76 |
TRANSITIONAL |
Au |
Au |
9,143 |
0.01 |
60.54 |
1.41 |
5.08 |
1.60 |
SULPHIDE |
Au |
Au |
10,034 |
0.01 |
93.12 |
3.11 |
15,99 |
1.28 |
KZONE 1 |
Au |
Au |
15,369 |
0.01 |
157.82 |
2.39 |
14.36 |
1.59 |
KZONE 2 |
Au |
Au |
787 |
0.01 |
9.66 |
1.07 |
1.76 |
1.24 |
KZONE 3 |
Au |
Au |
2,175 |
0.01 |
60.54 |
1.23 |
6.71 |
2.10 |
KZONE 4 |
Au |
Au |
11 |
0.39 |
1.80 |
0.89 |
0.23 |
0.54 |
KZONE 5 |
Au |
Au |
111 |
0.03 |
9.31 |
1.60 |
2.69 |
1.03 |
KZONE 6 |
Au |
Au |
405 |
0.01 |
10.21 |
0.87 |
1.57 |
1.45 |
TW OX |
Au |
Au |
363 |
0.01 |
6.99 |
0.47 |
0.69 |
1.75 |
TW TR |
Au |
Au |
430 |
0.01 |
15.16 |
0.99 |
1.83 |
1.37 |
TE OX |
Au |
Au |
22 |
0.15 |
3.17 |
0.86 |
0.51 |
0.82 |
TE TR |
Au |
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The statistical distributions for each
of the individual zones display similar properties and have distributions which
tend towards log normal. The histograms and log histograms per domain are shown
in Figure 14-6 below. Descriptive statistics were calculated and statistical
graphs produced in both real and log space as a measure of assessing:
|
|
The population characteristics of the mineralised grade
distribution; |
|
|
|
|
|
Confirmation of the statistical domains, and possible
combining of zones for geostatistics; |
|
|
|
|
|
The need, if any, to apply a top-cut during grade
interpolation. |
14.7.3 |
High grade capping |
Plots of the composite assay grades
against the cumulative mean and cumulative CoV were produced for each of the
sample types within the different oxidation domains, an example is given in
Figure 14-7. The plots were used to distinguish the grade at which the
cumulative totals CoV becomes extreme. Using this methodology top-cuts were
defined for each domain as summarised in Table 14-11.
Furthermore, log-probability plots have
been checked to ensure the top-cuts applied are applicable and the spatial
occurrence of the extreme cut gold values was visually verified to determine if
they formed discrete zones.
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Table 14-11: |
High-Grade Capping
|
DOMAIN |
GRADE CAPPING (g/t)
|
UPPER OXIDE |
20 |
LOWER OXIDE |
25 |
TRANSITION |
20 |
SULPHIDE |
10 |
14.8 |
Geostatistical analysis |
Geostatistical analysis was carried out
on the selected composite samples and for the various material types (upper
oxide, lower oxide, transition and sulphide). Initially variographic analysis
was completed to establish any directional anisotropy. Based on the results of
the semi-variograms the search ellipse and the kriging parameters were
optimised.
Semi-variograms were estimated for each
of the four oxidation domains described earlier. Initially the data was
transformed using the Discrete Gaussian transformation algorithm. Semi-variogram
models were produced in this transformation for the experimental variograms and
these models were then back transformed into the untransformed data space for
use in the kriging routine.
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Variograms were produced in the three
principal directions along strike, down dip and across the strike. In addition
to these, an omni-directional down-hole variogram was produced, and used to
determine the nugget variance which was applied to the directional variograms.
Figure 14-8 gives an example of the down hole and directional Gaussian
variograms produced for the Lower Oxide domain.
The best variograms were produced from
the along strike direction. Those produced for the down dip directions were
affected by the limited data extent in this direction. The back-transformed
Gaussian variograms for the Lower Oxide domain are also included below. The
final semi-variograms per zone can be found in Appendix V of the 2009
Feasibility Study.
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The back-transformed Gaussian variogram
parameters determined for each of the domains are included in Table 14-12. All
zones are characterised by a nugget variance in the order of 40 45% of the
sill value, while the fresh rock semi-variograms display a nugget variance in
the order of 56%, which indicates a relatively high degree of variability.
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Table 14-12: |
Back-Transformed
Gaussian Variogram Parameters
|
VARIOGRAM PARAM. |
UPPER OXIDE |
LOWER
OXIDE |
TRANSITION |
SULPHIDE |
Co |
5.15 |
5.96 |
4.39 |
1.52 |
C1 |
3.26 |
2.45 |
2.27 |
0.47 |
a1 ALONG STRIKE (m) |
12 |
25 |
25 |
20 |
a1 DOWN DIP (m) |
13 |
12 |
10 |
15 |
a1 ACROSS STRIKE (m) |
20 |
30 |
10 |
20 |
C2 |
2.56 |
3.07 |
2.39 |
0.73 |
a2 ALONG STRIKE (m) |
55 |
60 |
30 |
80 |
a2 DOWN DIP (m) |
40 |
25 |
80 |
60 |
a2 ACROSS STRIKE (m) |
50 |
30 |
50 |
32 |
C3 |
1.29 |
1.81 |
1.89 |
- |
a3 ALONG STRIKE (m) |
160 |
120 |
160 |
- |
a3 DOWN DIP (m) |
150 |
100 |
80 |
- |
a3 ACROSS STRIKE (m) |
60 |
30 |
60 |
- |
NUGGET EFFECT (%) |
42.10 |
60.94 |
40.13 |
55.88
|
Grade estimation was performed using
Ordinary Kriging routines within the Datamine software package. A quantitative
Kriging Neighbourhood Analysis (QKNA) exercise was completed in order to
optimise the kriging parameters; this was completed within the Isatis software
package. Each of the four oxidation domains was optimised individually with the
search ellipses rotated into the plane of the ore body domain to take account of
the anisotropy identified during the semi-variogram analysis.
The Twangiza North domain search was
oriented with the major (x) search axis striking towards 340° and dipping 75° to
the west. A second rotation with the major (x) search axis striking towards 340°
and dipping 45° to the west, has also been tested as there was evidence within
the Twangiza Main deposit of high-grades running in this orientation. The
general anisotropy defined in the Geostatistical studies of the main zones were
used to define the orientated search ellipses for these zones as they are
consistent with the interpreted geology of the deposit. Dynamic anisotropy was
employed in the estimation process for Twangiza East and West estimation.
In general the results displayed good
slopes of regression (i.e. greater than 0.8) within the well informed areas for
all scenarios. A summary of the final parameters selected are shown in Table
14-13 below.
A minimum of 4 and a maximum of 18
composites were used to estimate a block in the first pass for all zones. This
represents an increase from a minimum of 2 and a maximum of 12 composites used
in the previous model. The block discretization for the interpolations was 4 x 4
x 4.
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Table 14-13: |
Search Radius For Pass
1
|
PASS 1 PARAMETERS |
UO DOMAIN |
LO DOMAIN |
TR DOMAIN |
FR DOMAIN |
RADIUSX (m) |
130 |
130 |
130 |
80 |
RADIUS Y (m) |
100 |
100 |
80 |
60 |
RADIUS Z (m) |
20 |
20 |
20 |
20 |
MIN. NO. OF SAMPLES |
4 |
4 |
4 |
4 |
NO. OF SECTORS |
4 |
4 |
4 |
4 |
MAX NO. OF SAMPLES |
18 |
18 |
18 |
18 |
DESCRITIZATION |
4x4x4 |
4x4x4 |
4x4x4 |
4x4x4
|
The first pass employed a search
ellipse which equates to approximately the semi-variogram range. The longest
dimension of the ellipse was approximately equal to one of the range of the
variogram. Approximately 90 % of the blocks within the current economic pit were
estimated using the first search range
Axis multipliers were set at 2 and 3
for the second and third search volumes. The number of samples used was
increased in the second search volume to produce more averaged grades.
The compiled block model was checked
within the Datamine software package for missing or duplicated estimates to
ensure there is no double accounting of ore tonnage.
14.9 |
Mineral Resource
Classification |
The definitions provided in the
following section are taken from the Canadian Institute of Mining, Metallurgy
and Petroleum (CIM) Standards on Mineral Resources and Reserves.
A Mineral Resource is a concentration
or occurrence of diamonds, natural solid inorganic material, or natural solid
fossilized organic material including base and precious metals, coal, and
industrial minerals in or on the Earths crust in such form and quantity and of
such a grade or quality that it has reasonable prospects for economic
extraction. The location, quantity, grade, geological characteristics and
continuity of a Mineral Resource are known, estimated or interpreted from
specific geological evidence and knowledge.
The term reasonable prospect for
economic extraction implies a judgement (albeit preliminary) by the Qualified
Person in respect of the technical and economic factors likely to influence the
prospect of economic extraction, including the approximate mining parameters. In
basic terms the code highlights that the Mineral Resource is not simply an
inventory of all mineralisation drilled or sampled, regardless of cut-off grade,
likely mining dimensions, location and continuity. The Mineral Resource is
therefore the portion of the Mineralised Block Model which under assumed and
justifiable technical and economic conditions in whole or in part may become
economically extractable.
14.9.1 |
Geological complexity |
As highlighted above, the oxidation and
lithological model at Twangiza is important in terms of metal recoveries of the
different units. The porphyry units provide reasonably large continuous zones
which have been modelled with a reasonable level of confidence within well
drilled areas. The geological knowledge, detailed interpretation and good data
density in well drilled areas and areas with adits, have allowed the resource to
be classified with high confidence in some places.
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14.9.2 |
Quality of data used in the
estimation |
Twangiza Mining used best methods of
sampling and sample preparation. Twangiza Mining has conducted a systematic
process of sample preparation within clean facilities, which is well documented.
The systems include regular insertion of blanks, standards and duplicates in all
sample submission to the laboratory in Mwanza. In general, the results from the
QA/QC programme indicate acceptable levels of errors are achieved at the
laboratory.
14.9.3 |
Results of the geostatistical
analysis |
Based on the variograms, the QKNA was
undertaken and the results showed a good slope of regression within the well
informed and reasonably informed blocks. The slope of regression values,
however, reduced in areas of limited sampling where the sample support is
reduced.
14.9.4 |
Classification Method |
The classification was carried out
using a combination of drill hole spacing, slopes of regression, geological and
wireframe confidence. Classification was applied to the model by digitizing
areas on 80m spaced vertical sections based on
|
|
Measured Mineral Resource consists of kriged model blocks
which have been interpolated by data within 20m and are limited to the
areas surrounding the adit sampling within the Twangiza Main proportion of
the deposit, with extensions of 10m20m below the deepest adit; |
|
|
|
|
|
Indicated Mineral Resources are those kriged blocks which
have been interpolated by adit and drill hole data using an average drill
hole spacing of 40 x 40m within the search area. A minimum number of
points used to estimate a block grade has also been reviewed as have the
search volumes and the slope of regression; and |
|
|
|
|
|
Inferred Mineral Resources are model blocks lying outside
the first search area estimated by OK and those blocks which are deemed to
be poorly informed and therefore have a low associated slope of
regression. The down-dip extensions of the Inferred Mineral Resources have
been limited to 50m100m from the Indicated boundary and to remove
interpretations based on a single drill hole spacing of 150m.
|
Areas which were modelled as part of
the geological model but which fell outside of the Inferred boundary limits were
flagged in the model as areas for potential additions to the Mineral Resources.
These areas require additional drilling before being classified as a Mineral
Resource.
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The Mwana River Valley Fill deposit
comprises tailings that resulted from historical artisanal mining and washing of
oxide ore in the Mbwega pit at Twangiza Main. The deposit runs to the west and
northwest of Twangiza Main as shown in Figure 14-10.
To build on the Inferred Mineral
Resource in the 2009 Feasibility Study, further sampling has been undertaken in
the last quarter of 2009 and the end of 2010 to obtain a better understanding of
the grade distribution of the gold, average gold grade and the thickness of the
deposit.
50 auger holes were drilled to a
maximum of 7m length, totalling 161m generating 113 samples. 14 pits up to 4.2m
deep totalling 35.1 m were dug further downstream generating 50 samples. Sample
spacing is approximately 100m along strike by 40m across sections and covering a
total area of some 260,000m2.
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Based on the sampling, sectional
interpretations were digitized along the auger and pitting lines and the
perimeters were linked together to generate a 3D wireframe in Datamine Studio
software. The total volume of the model was 1,876,157m3 and the model had an
average thickness of 7m. A 3D block model was created from the wireframe as
summarised in Table 14-14.
|
Table 14-14: |
Valley Fill Block Model
Parameters
|
Model Parameters |
Block Dimension
|
Min |
Max |
Range |
X |
50 |
691580 |
692980 |
1400 |
Y |
50 |
681950 |
683950 |
2000 |
Z |
2 |
1850 |
2000 |
150 |
Samples were composited to 2m and
composited sample grades were used to interpolate grade into the blocks using
ordinary Kriging to generate a 3D resource model. A density factor of 1.65t/m
3, determined by weighing dried samples from known volumes, was
applied for tonnage estimation.
By the end of 2013, 0.33 Mt of Valley
Fill material grading 3.13 g/t Au for 0.03 Moz of gold had been mined from the
upstream part of the deposit which was removed to allow construction of the
TMF1A dam wall and avoid sterilisation. At the end of this mining activity the
mined out area of approximately 42,810m2 was surveyed and the outlines were used
to deplete the Valley Fill Mineral Resource model. During the mining process
some areas were mined to a depth of up to 10m.
Between February and March 2014, some
17 additional pits were sampled in the Mwana river bed. A total of 46 samples
were collected and assayed and the results were used to update the Valley Fill
Mineral Resource model. This additional sampling along with the good agreement
of the mined grades enabled the Valley Fill Mineral Resource to be upgraded to
the Indicated category and to be included in the Mineral Reserve.
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The remaining Valley Fill Mineral
Resource reported at 0.40g/t insitu Au cut-off is 2.18 Mt at an average grade of
1.93g/t for 0.13 Moz gold. Whilst the grades further downstream are relatively
low, being further from the point of deposition, all the block grades in the
resource model are above cut-off. The Valley Fill resource contributes 2% of the
Measured and Indicated Mineral Resource at Twangiza.
14.11 |
Mineral Resource Statement |
An updated Mineral Resource estimate as
at end December of 2014, has been prepared as reported in Banros press release
dated June 08, 2015. As a basis for the update, the Mineral Resource model was
modified to reflect changes to material densities (mainly upper oxide zone),
artisanal mining, topography and mining depletion as determined by an historical
reconciliation review and survey update as at end December 2014. The estimate
had a 0.40 g/t gold cut-off grade applied and was constrained to a USD1,600/oz
gold price optimum pit shell, which was considered an appropriate price for this
purpose, being 33% above the prevailing forecast gold pricing.
The updated resource reflects the
remaining resource at the end of year 2014 after all depletions due to mining.
Table 14-15 below details the Oxide and Non-Oxide components of the 2014
Twangiza Mineral Resource estimate, split by confidence category, reported at a
cut-off grade of 0.40 g/t gold.
|
Table
14-15: |
Mineral Resource Estimate as at December 31, 2014
|
OXIDE MINERAL RESOURCE CATEGORY |
Tonnes |
Grade |
Gold |
|
(Mt) |
(g/t Au ) |
(Moz) |
MEASURED |
3.72 |
2.3 |
0.28 |
INDICATED |
8.76 |
1.88 |
0.53 |
MEASURED AND INDICATED |
12.48 |
2.02 |
0.81 |
INFERRED |
1.34 |
1.32 |
0.06 |
|
Tonnes |
Grade |
Gold |
NON-OXIDE MINERAL RESOURCE CATEGORY |
(Mt) |
(g/t Au ) |
(Moz) |
MEASURED |
3.8 |
2.23 |
0.27 |
INDICATED |
93 |
1.4 |
4.18 |
MEASURED AND INDICATED |
96.8 |
1.43 |
4.45 |
INFERRED |
11.65 |
1.12 |
0.42 |
NB: Any apparent errors are due to
rounding and are therefore not considered material to the estimate
The details of the Twangiza Mineral
Resource by pit, material type (oxidation state) and confidence category are
outlined in Table 14-16.
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Table 14-16: |
Mineral Resource Estimate by Pit, Material Type and Confidence Category as at December 31, 2014
|
PROSPECT |
MATERIAL |
MEASURED
MINERAL RESOURCES |
INDICATED MINERAL RESOURCES |
INFERRED
MINERAL RESOURCES |
|
Average |
Gold |
|
Average |
|
|
Average |
Gold |
Tonnes (Mt) |
Grade
(g/t) |
Content |
Tonnes (Mt) |
Grade |
Gold
Content |
Tonnes |
Grade |
Content |
|
|
(Moz) |
|
(g/t) |
(Moz) |
(Mt) |
(g/t) |
(Moz)
|
Twangiza Main |
Oxide |
3.14 |
2.54 |
0.26 |
2.15 |
1.43 |
0.10 |
0.39 |
0.93 |
0.01 |
Transition |
3.58 |
2.25 |
0.26 |
9.55 |
1.63 |
0.50 |
0.76 |
0.96 |
0.02 |
Fresh |
0.22 |
1.80 |
0.01 |
75.96 |
1.30 |
3.18 |
8.76 |
1.15 |
0.32 |
Twangiza Main Sub Total |
6.94 |
2.37 |
0.53 |
87.66 |
1.34 |
3.78 |
9.91 |
1.13 |
0.36 |
Twangiza North |
Oxide |
|
|
|
1.75 |
2.63 |
0.15 |
0.00 |
1.24 |
0.00 |
Transition |
|
|
|
2.12 |
2.21 |
0.15 |
0.04 |
1.53 |
0.00 |
Fresh |
|
|
|
1.57 |
2.43 |
0.12 |
0.01 |
3.04 |
0.00 |
Twangiza North Sub Total |
|
|
|
5.44 |
2.41 |
0.42 |
0.05 |
1.86 |
0.00 |
Twangiza Central |
Oxide |
0.01 |
1.16 |
0.00 |
2.11 |
1.86 |
0.13 |
0.41 |
2.08 |
0.03 |
Transition |
|
|
|
1.19 |
1.80 |
0.07 |
0.42 |
2.01 |
0.03 |
Fresh |
|
|
|
1.57 |
2.43 |
0.12 |
0.01 |
3.04 |
0.00 |
Twangiza Central Sub Total |
0.01 |
1.16 |
0.00 |
4.87 |
2.03 |
0.32 |
0.85 |
2.06 |
0.06 |
Twangiza East |
Oxide |
|
|
|
|
|
|
0.07 |
0.75 |
0.00 |
Transition |
|
|
|
|
|
|
|
|
|
Fresh |
|
|
|
|
|
|
|
|
|
Twangiza East Sub Total |
|
|
|
|
|
|
0.07 |
0.75 |
0.00 |
Twangiza West |
Oxide |
0.00 |
0.00 |
0.00 |
0.58 |
1.11 |
0.02 |
0.46 |
0.95 |
0.01 |
Transition |
0.00 |
0.00 |
0.00 |
0.36 |
1.26 |
0.02 |
0.56 |
0.81 |
0.01 |
Fresh |
0.00 |
0.00 |
0.00 |
0.67 |
1.33 |
0.02 |
1.08 |
0.77 |
0.03 |
Twangiza West Sub Total |
0.00 |
0.00 |
0.00 |
1.61 |
1.11 |
0.06 |
2.10 |
0.82 |
0.06 |
Valley Fill |
|
|
|
|
2.18 |
1.93 |
0.14 |
|
|
|
Stockpile |
|
0.57 |
1.07 |
0.02 |
|
|
|
|
|
|
TOTAL |
7.52 |
2.27 |
0.55 |
101.76 |
1.44 |
4.71 |
12.99 |
1.14 |
0.48
|
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15 |
MINERAL RESERVE
ESTIMATE |
The current pit optimisation and design
applied for estimation of the updated Mineral Reserve relies on the geotechnical
analysis completed as part of the PFS.
A full description of the PFS
geotechnical investigation and slope stability analysis is given in Section 16.1
of the SRK (UK) report entitled Pre-Feasibility Study NI43-101 Technical
Report, Twangiza Gold Project, South Kivu Province, Democratic Republic of
Congo, dated August 13, 2008.
Based on the slope parameters derived
by SRK (UK) for the PFS, the following overall slope angles were used for the
pit optimization taking into consideration ramp sections:
|
|
Overall slope angle in soil/saprolite material
of 27.5 to 30°; |
|
|
|
|
|
Overall slope angle in oxide material of 34 to
38°; |
|
|
|
|
|
Overall slope angle in transitional material of
41 to 42°; and, |
|
|
|
|
|
Overall slope angle in fresh material of 53°.
|
These parameters assume dry mining
conditions.
15.2 |
Open pit optimization |
The 2015 open pit optimisation assumes
a 1.7Mtpa throughput rate and is based on the resource block model generated by
Twangiza Mining as presented in this report.
A conventional open pit shovel and
truck method will be used for the mining of sufficient ore to supply 1.7Mtpa of
ore throughput. The mining functions of the operation will be owner mining as
per the current practise.
The Whittle process requires various
input data including the resource block model, unit costs and other physical
parameters such as the slope angles at which the pit can be mined. Unit costs
specific to the Twangiza operation were determined by Twangiza Mining from a
zero based budgeting exercise and an analysis of historical costs.
The open-pit optimization study was
performed using the Whittle/Gemcom Four-X Analyser (Whittle 4X) software package
to provide guidance to the potential economic final pit geometries. Whittle 4X
compares the estimated value of the individual mining blocks at the pit boundary
versus the cost for waste stripping. It establishes the pit walls where the ore
revenue and waste stripping cost balance for maximum net revenue.
The optimum pit was considered by
Twangiza Mining to be sufficiently similar to the pit designs generated in early
2014 so as not to require a re-design. The ore/waste tonnages in the design pits
have been calculated and scheduled to determine the ore production and the waste
stripping requirements.
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The following sections describe the
methodology and derivation of the Whittle input parameters and assumptions.
Cost input parameters were based on the
Twangiza mine historic operating cost as consolidated in the 2015 production
year Budget. For the purpose of the Whittle optimization, capital costs,
depreciation, amortisation and other interest / finance charges have been
excluded.
Mining costs
For the pit optimization analysis, a
base unit mining cost of USD 3.28 /t mined for ore and USD 3.51/t mined for
waste has been assumed. A USD 0.30 /t mined adjustment for weathered material
has been applied to take into account the reduced drilling and blasting required
in the upper formations. The mining cost has also been adjusted for the
additional pumping cost of USD0.0005/tonne mined for blocks below the valley
floor.
Mining unit costs are outlined in Table
15-1.
|
Table 15-1: |
Mining Cost Summary
|
|
|
|
|
Ore Unit Mining
Cost |
Waste Unit Mining Cost |
Category |
(USD/t) |
(USD/t) |
|
|
|
Load
& Haul (Free Dig) |
2.61 |
2.84 |
Drill
& Blast |
0.30 |
0.30 |
Grade Control |
0.36 |
0.36 |
Total Mining |
3.28 |
3.51
|
Processing cost
Process costs have been based on the
2015 Budget which is derived from a zero based approach with reference to the
2014 actual costs and adjusted to account for planned cost saving measures, the
increased proportion of transition and fresh ore over the LOM and the increase
to a 1.7Mtpa throughput rate.
The total process operating cost
encompasses metallurgy and process plant maintenance. The average direct process
cost was estimated at USD 18.59 /t processed.
General Administration cost
The General and Administration
(G&A) cost was also based on 2014 historical estimates as provided under the
2015 Budget adjusted to reflect planned cost savings and increase in production.
The average G&A cost was estimated at USD 12.29/t processed.
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Revenue and Selling costs
A gold price of USD1,200 per ounce was
used for the purpose of reserves estimation and life of mine planning. It was
judged to be consistent with prevailing industry estimates. A royalty calculated
at the rate of 1% of gross revenue is payable to the Government of DRC.
Selling costs cover all direct and
indirect operating costs incurred in carrying out the gold sales and off-site
activities. They include Refinery & Shipment, Government Royalties, Head
Office Management in the DRC and in Toronto. The average selling cost is USD
63.19 /ounce.
A summary of the Government royalties,
refining and selling costs is provided in Table 15-2.
|
Table 15-2: |
Government Royalties,
Refining and Selling Costs Summary
|
|
Costs (USD/ounce
produced) |
Refinery and Shipment |
15.20 |
Government Royalties |
12.50 |
Head office cost (Toronto) |
12.93 |
Head office cost (DRC) |
21.55 |
Banro Foundation |
1.00 |
Total Royalties, Refining & Selling |
63.19 |
A breakdown of the costs and parameters
used in the Whittle optimization runs are shown in the Table 15-4.
For the pit optimisation, the mining
grade dilution factor was set at 5% (at zero grade) and mining recovery at 100%.
A recoverable gold cut-off Grade,
(COG) has been calculated by dividing the overall marginal operating cost of
one tonne of ore by the recovered value of the gold contained therein. This is
applied to the recoverable gold grade variable housed in the resource block
model. This approach simplified the reporting process by incorporating variable
recovery by rock type within the recoverable grade value of the resource model.
For the various deposits the COG was determined as shown in Table 15-3
|
Table 15-3: |
Recoverable Gold Cut-Off
Grade by Deposit
|
|
Unit |
Total
Mine |
Main |
Central |
North |
East |
West |
Cut-Off grade |
g/t |
0.84 |
0.84 |
0.84 |
0.86 |
0.84 |
0.85
|
A range of optimizations was prepared
with the results for a gold price of $1,200 per ounce and a mining cost of
USD3.51/t as shown in Figure 15-1, Figure 15-3, Figure 15-5 and Figure 15-7 for
each of the deposits.
The updated Whittle shells compare well
with the 2014 pit designs so these pit designs have been used for reporting of
the end December 2014 Mineral Reserves. Figure 15-2, Figure 15-4, Figure 15-6
and Figure 15-8 below show the block model and the practical pit design.
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Table 15-4: |
Whittle Parameters for
Open Pit Optimization
|
|
Unit |
Total
Mine |
Main |
Central
|
North
|
West
|
Plant Capacity |
Processed tpa |
1,700,000 |
|
|
|
|
Mining Costs |
|
|
|
|
|
|
Ore |
USD/tonne mined |
2.92 |
2.92 |
2.92 |
3.22 |
3.20 |
Waste |
USD/tonne mined |
3.15 |
3.15 |
3.15 |
3.15 |
3.15 |
Grade control
COSTM |
USD/tonne mined USD/tonne
mined |
0.36 3.51 |
0.36 3.51 |
0.36 3.51 |
0.36 3.51 |
0.36 3.51 |
Rehabilitation Costs |
|
|
|
|
|
|
Waste dumps
|
USD/waste tonnes mined |
0.10 |
0.10 |
0.10 |
0.10 |
0.10 |
Post Mining Costs |
|
|
|
|
|
|
Process Plant Costs - Carbon In Leach |
USD/tonne treated |
7.73 |
7.73 |
7.73 |
7.73 |
7.73 |
Assay |
USD/tonne treated |
0.71
|
0.71
|
0.71
|
0.71
|
0.71
|
Power Engineering (Maintenance) Costs |
USD/tonne
treated USD/tonne treated |
7.35
3.21 |
7.35
3.21 |
7.35
3.21 |
7.35
3.21 |
7.35
3.21 |
Additional ore cost |
USD/tonne treated |
(0.41) |
(0.41) |
(0.41) |
0.12 |
0.09 |
Processing Costs |
USD/tonne treated |
18.59 |
18.59 |
18.59 |
19.11 |
19.08 |
Infrastructure, Overheads and Sundries (G&A) |
USD/tonne treated |
10.49 |
10.49 |
10.49 |
10.49 |
10.49 |
Sustaining Capital (Tailings Dam Lifts, Pad Expansions) |
USD/tonne
treated |
1.80 |
1.80 |
1.80 |
1.80 |
1.80 |
G&A costs |
USD/tonne treated
|
12.29 |
12.29 |
12.29 |
12.29 |
12.29 |
|
|
|
|
|
|
|
COSTP |
USD/tonne treated
|
30.87 |
30.87 |
30.87 |
31.40 |
31.37 |
Selling Costs |
|
|
|
|
|
|
Refinery and Shipment |
USD/Ounce Produced |
15.20 |
15.20 |
15.20 |
15.20 |
15.20 |
Government Royalty |
USD/Ounce Produced |
12.50 |
12.50 |
12.50 |
12.50 |
12.50 |
H/O Management Fee (Toronto) |
USD/Ounce Produced |
12.93
|
12.93
|
12.93
|
12.93
|
12.93
|
H/O Management Fee (Banro
Congo Mining) |
USD/Ounce
Produced |
21.55 |
21.55 |
21.55 |
21.55 |
21.55 |
Banro Foundation |
USD/Ounce Produced |
1.00 |
1.00 |
1.00 |
1.00 |
1.00 |
Selling cost / oz |
USD/Ounce
Produced |
63.19 |
63.19 |
63.19 |
63.19 |
63.19 |
Rehandle - Owner Mining |
USD/tonne treated
|
0.39 |
0.39 |
0.39 |
0.39 |
0.39 |
|
|
|
|
|
|
|
Gold Price (USD/oz) |
1200 |
|
|
|
|
|
Cut-Off grade = (costp - rehab dumps)/((Gold
Price/31.103475)*(1-Selling Cost %)) |
Cut-Off grade |
g/t |
0.84 |
0.84 |
0.84 |
0.86 |
0.85 |
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Table 15-5: |
Comparison of Inventory Between
EOY2014 Optimum Pit and 2014
Practical Pit Design |
Item |
Description |
Unit |
Twangiza
Main |
Twangiza
Central |
Twangiza
North |
Twangiza West |
WHITTLE INPUTS FOR
OPTIMISATION |
Gold price |
USD/ounce |
1200 |
1200 |
1200 |
1200 |
Mining costs |
USD/tonne mined |
3.51 |
3.51 |
3.51 |
3.51 |
Processing costs |
USD/tonne processed |
18.59 |
18.59 |
19.11 |
19.08 |
General and administration
costs |
USD/tonne processed |
12.29 |
12.29 |
12.29 |
12.29 |
Royalties and selling costs
|
USD/ounce |
63.19 |
63.19 |
63.19 |
63.19 |
|
|
|
|
|
|
|
EOY2014 OPTIMUM WHITTLE
PIT INVENTORY |
Ore mined to process |
000 tonnes |
17,701 |
3,974 |
4,588 |
472 |
Insitu grade |
g/t |
2.29 |
2.23 |
2.84 |
2.04 |
Diluted grade |
g/t |
2.18 |
2.12 |
2.70 |
1.93 |
Recoverable grade |
g/t |
1.59 |
1.78 |
2.24 |
1.57 |
Contained gold |
000 ounces |
1,238 |
271 |
398 |
29 |
Recoverable gold |
000 ounces |
906 |
227 |
330 |
24 |
Waste mined |
000 tonnes |
19,687 |
4,959 |
16,645 |
401 |
Total material |
000 tonnes |
37,388 |
8,933 |
21,233 |
873 |
Strip ratio |
t/t |
1.11 |
1.25 |
3.63 |
0.85 |
Undiscounted Cashflow |
Million USD |
352 |
104 |
157 |
9 |
|
|
|
|
|
|
|
2014 PRACTICAL PIT
DESIGN INVENTORY |
Ore mined to process |
000 tonnes |
13,872 |
2,826 |
2,971 |
304 |
Insitu grade |
g/t |
2.42 |
2.16 |
2.99 |
2.05 |
Diluted grade |
g/t |
2.30 |
2.05 |
2.84 |
1.95 |
Recoverable grade |
g/t |
1.61 |
1.67 |
2.34 |
1.55 |
Contained gold |
000 ounces |
1,026 |
187 |
271 |
19 |
Recoverable gold |
000 ounces |
717 |
152 |
224 |
15 |
Waste mined |
000 tonnes |
28,180 |
9,418 |
30,809 |
1,225 |
Total material |
000 tonnes |
42,053 |
12,244 |
33,780 |
1,529 |
Strip ratio |
t/t |
2.03 |
3.33 |
10.37 |
4.02 |
Undiscounted Cashflow |
Million USD |
240 |
42 |
43 |
2 |
Selection of optimized pit shell
The full optimization generates nested
pit shells using an incremental revenue factor ranging from 0.5 to 1.5 with a
step size of 0.0125 (1.25% of revenue) to generate a maximum of 81 pit shells.
The optimization algorithm uses a cut-off grade ore selection method. A revenue
factor of 1.0 equates to the gold price of USD 1,200/oz adopted in the
optimization. As part of the optimization process the block model is flagged
with a number (1-81) representing the range of nested pit shells. The flagged
model is then exported from Whittle and imported into Surpac for graphical
inspection and further evaluation.
Conventionally, the pit number with
revenue factor equal to 1 (one) is referred to as the optimal pit. Selection of
the optimum pit will also depend on consideration of net cashflow (discounted
and undiscounted), average mining cost, marginal mining cost and total ounces.
The selection of the optimized final pit shell was based on the maximum
undiscounted net cash flow.
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15.3 |
Practical pit design |
Practical pit designs were prepared
using optimized pit shells as templates for all deposits except for the Valley
Fill which is already accessible and will be free dig mined to remove sand fill
from the river valley.
Surpac software was used to prepare the
practical pit, and to incorporate the haul roads, ramps and berms together with
appropriate allowance for inter-ramp slope angles. The final economic shell
chosen was used as a guide to select pushbacks within the ultimate pit in order
to schedule the mining of the pit in a continuously profitable sequence. The
pushbacks and ultimate pit were then processed using a minimum mining width
algorithm in Whittle in order to apply appropriate practical mining constraints
to generate a pit shell for use in pit engineering design work.
Detailed and practical pit designs were
produced from the chosen pushbacks to confirm the practicability of mining the
deposits using the optimal pit shells as a guide. Care was taken to keep
designed strip ratios within the optimal shell strip ratios and a series of
options for optimal positioning of ramps was reviewed in order to avoid losses
and excessive waste introduction to the final engineered pit design. In-pit and
ex-pit haul roads were designed based on these criteria with a continuous
gradient of -/+10% at a width of 20 m to provide sufficient room for two way
traffic flow. Engineering pit design parameters were as follows;
|
|
Nominal Bench height 10m |
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|
Berm width 6m |
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Ramp gradient 10% |
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Ramp width (with safety berms) 20m |
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Bench face angle (Upper Oxides) 32° |
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|
Bench face angle (Lower Oxides) 38° |
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|
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Bench face angle (Transition) 60° |
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|
|
Bench face angle (Fresh) 70° |
A total of five pits have been designed
as follows:
|
|
Main Intermediate 1 is an interim pit within the
Twangiza ultimate pit exploiting the oxide zone using Whittle pit number
12 as a template. This pit has been designed to minimize the initial
scheduled stripping ratio. A plan view is shown in Figure 15-10 |
|
|
|
|
|
Main Intermediate 2 is a second interim pit also within
the main Twangiza ultimate shell targeted to extract all the Oxide
material to blend with the Fresh and Transition material contained within
the optimized shell which forms part of this reserve. This is referred to
as cut 2 using Whittle pit number 23 as a template as shown in Figure
15-11. |
|
|
|
|
|
Main Final the design is based on Whittle pit number 36
as a template. This pit forms part of the phase two programme for the
Twangiza operations as discussed in the 2009 Feasibility Study. |
|
|
|
|
|
Twangiza North this pit exploits the North ore body
using Whittle pit 29 as a template, as shown in Figure 15-12. |
|
|
|
|
|
Twangiza Central this pit exploits the southern portion
of the North ore body as shown in Figure 15-13. It uses Whittle pit 38 as
a template. |
|
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|
|
|
Twangiza West this pit lies to the North-West of the
Twangiza Main Pit and is shown in Figure 15-14. |
|
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|
|
Twangiza Valley Fill this pit lies to the North-West of
the Twangiza Main Pit and is shown in Figure 15-15
|
Principal haul roads have been designed
to connect the working areas to the primary crusher and the waste dumps. The
final layout of pit designs is shown in Figure 15-16.
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15.4 |
Mine production schedule |
The scheduling process consisted of
developing a mine plan that is economically optimum using the inventory included
in the practical pits. The schedule methodology adopted a simple bench by bench
approach with selective mining of ore from waste.
The material mined is classified into
three categories namely: High Grade (HG), Low Grade (LG) and Waste (Wst). This
is accomplished by applying cut-off grades to the material in the pit.
The schedule by material type is given
in Table 15-7 and the overall mine production schedule is given in Table 15-8.
An ore reserve COG of 0.84g/t, based on
the recoverable gold grade, is used to differentiate waste from Run-Of-Mine
(RoM) ore. Ore is further categorised into HG and LG by applying an HG COG of
1.0g/t recoverable. Both HG and LG comprise RoM material that can be blended to
meet the total plant throughput rate of 1.7 Mtpa.
Grade control mark-up in the pit is
undertaken using coloured flagging tapes defining the various categories of
material on each mining flitch or bench. The current practice is for the HG, LG
and Wst material to be flagged with Red, Yellow and Blue tapes respectively.
All HG material is sent to either the
RoM stockpile or direct tipped into the ROM bin whereas all LG material is sent
to the LG stockpile.
The material movement is focused on
maintaining a smooth total mined profile whilst achieving the required 1.7 Mtpa
throughput target.
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During the first five years a blend of
the oxide and transition material is delivered to the plant to target a process
recovery above 82%. Any excess of transition and fresh material is stockpiled to
be processed when all the oxide RoM is depleted.
The current LG stockpile is
predominantly oxide material which has been accumulated since the commencement
of operations. The stockpile closing balance at End-of-Year 2014 is provided in
Table 15-6.
|
Table 15-6: |
Stockpile Closing
Balance as at December 31, 2014
|
Stockpiles |
Tonnes |
Gold Grade (g/t) |
RoM
Stockpiles |
69,905 |
2.44 |
Low Grade Stockpiles |
501,050 |
0.88 |
Total Stockpile Balance |
570,955 |
1.07 |
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|
Table 15-7: |
Annual Mine Production Schedule by Material Type
|
Material Type |
Units |
2015 |
2016 |
2017 |
2018 |
2019 |
2020 |
2021 |
2022 |
2023 |
2024 |
2025 |
2026 |
Total |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Oxide Quantities |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Waste Volume |
kbcm |
115 |
1,658 |
4,775 |
2,682 |
3,758 |
1,905 |
1,602 |
2,473 |
269 |
29 |
|
|
19,268 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Waste Tonnage |
kt |
188 |
3,313 |
9,537 |
5,497 |
7,683 |
3,983 |
2,500 |
4,185 |
452 |
56 |
|
|
37,393 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Volume |
kbcm |
4 |
52 |
11 |
9 |
20 |
7 |
5 |
52 |
9 |
0 |
|
|
171 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Tonnage |
kt |
7 |
103 |
23 |
19 |
42 |
14 |
9 |
97 |
17 |
1 |
|
|
332 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG In Situ Grade |
g/t |
1.15 |
1.26 |
1.11 |
1.19 |
1.21 |
1.1 |
1.13 |
1.11 |
1.13 |
1.06 |
|
|
1.17 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Diluted Grade |
g/t |
1.09 |
1.19 |
1.05 |
1.13 |
1.15 |
1.04 |
1.07 |
1.05 |
1.07 |
1.01 |
|
|
1.11 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Ore Recoverable Grade |
g/t |
0.96 |
1.05 |
0.93 |
0.93 |
1.02 |
0.93 |
0.94 |
0.92 |
0.94 |
0.89 |
|
|
0.98 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Mined Ounces |
koz |
0 |
4 |
1 |
1 |
2 |
0 |
0 |
3 |
1 |
0 |
|
|
12 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Recoverable Ounces |
koz |
0 |
3 |
1 |
1 |
1 |
0 |
0 |
3 |
1 |
0 |
|
|
10 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Volume |
kbcm |
639 |
583 |
653 |
283 |
258 |
168 |
86 |
577 |
446 |
32 |
|
|
3,724 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Tonnage |
kt |
1,017 |
1,048 |
1,324 |
591 |
484 |
350 |
159 |
985 |
765 |
61 |
|
|
6,784 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG In Situ Grade |
g/t |
2.9 |
2.42 |
2.7 |
2.59 |
2.36 |
3.64 |
1.82 |
2.26 |
2.19 |
1.65 |
|
|
2.55 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Diluted Grade |
g/t |
2.76 |
2.3 |
2.56 |
2.46 |
2.24 |
3.46 |
1.73 |
2.14 |
2.08 |
1.57 |
|
|
2.42 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Recoverable Grade |
g/t |
2.35 |
1.98 |
2.24 |
2.16 |
1.92 |
3.03 |
1.47 |
1.86 |
1.74 |
1.31 |
|
|
2.09 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Mined Ounces |
koz |
90 |
77 |
109 |
47 |
35 |
39 |
9 |
68 |
51 |
3 |
|
|
528 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Recoverable Ounces |
koz |
77 |
67 |
95 |
41 |
30 |
34 |
8 |
59 |
43 |
3 |
|
|
456 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Oxide Volume |
kbcm |
759 |
2,292 |
5,439 |
2,975 |
4,036 |
2,080 |
1,694 |
3,102 |
725 |
62 |
|
|
23,164 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total
Oxide Tonnage |
kt
|
1,212
|
4,464
|
10,884 |
6,107
|
8,209
|
4,347
|
2,668
|
5,267
|
1,234
|
118
|
|
|
44,509
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Transition Quantities |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Waste Volume |
kbcm |
383 |
299 |
346 |
1,107 |
757 |
1,362 |
232 |
951 |
456 |
354 |
44 |
|
6,291 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Waste Tonnage |
kt |
843 |
705 |
841 |
2,758 |
1,798 |
3,533 |
599 |
2,149 |
1,060 |
906 |
115 |
|
15,308 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG
Volume |
kbcm
|
6 |
11 |
2 |
2 |
6 |
6 |
4 |
23 |
10 |
2 |
|
|
72
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Tonnage |
kt |
14 |
24 |
4 |
5 |
13 |
16 |
8 |
54 |
24 |
6 |
|
|
167 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG In Situ Grade |
g/t |
1.89 |
1.72 |
1.19 |
1.61 |
1.58 |
1.32 |
1.31 |
1.26 |
1.4 |
1.45 |
|
|
1.44 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Diluted Grade |
g/t |
1.8 |
1.63 |
1.13 |
1.53 |
1.5 |
1.25 |
1.24 |
1.19 |
1.33 |
1.37 |
|
|
1.37 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Ore Recoverable Grade |
g/t |
0.95 |
0.93 |
0.92 |
0.98 |
1.05 |
0.93 |
0.98 |
0.93 |
0.95 |
0.96 |
|
|
0.95 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG
Mined Ounces |
koz
|
1 |
1 |
0 |
0 |
1 |
1 |
0 |
2 |
1 |
0 |
|
|
7,357
|
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43-101 Twangiza Main Report |
Material Type |
Units |
2015 |
2016 |
2017 |
2018 |
2019 |
2020 |
2021 |
2022 |
2023 |
2024 |
2025 |
2026 |
Total |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG
Recoverable Ounces |
koz
|
0 |
1 |
0 |
0 |
0 |
0 |
0 |
2 |
1 |
0 |
|
|
5,097
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Volume |
kbcm |
390 |
318 |
112 |
448 |
292 |
341 |
79 |
245 |
366 |
369 |
195 |
|
3,156 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Tonnage |
kt |
762 |
696 |
271 |
1,144 |
684 |
891 |
208 |
550 |
862 |
953 |
512 |
|
7,534 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG In Situ Grade |
g/t |
2.86 |
3.26 |
2.39 |
2.68 |
2.46 |
2.6 |
2.56 |
2.54 |
2.36 |
2.45 |
2.7 |
|
2.63 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Diluted Grade |
g/t |
2.72 |
3.1 |
2.27 |
2.54 |
2.33 |
2.47 |
2.44 |
2.41 |
2.24 |
2.33 |
2.56 |
|
2.5 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Recoverable Grade |
g/t |
1.78 |
1.98 |
1.76 |
1.99 |
1.59 |
1.72 |
1.95 |
1.63 |
1.38 |
1.4 |
1.5 |
|
1.69 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Mined Ounces |
koz |
67 |
69 |
20 |
93 |
51 |
71 |
16 |
43 |
62 |
71 |
42 |
|
606 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Recoverable Ounces |
koz |
44 |
44 |
15 |
73 |
35 |
49 |
13 |
29 |
38 |
43 |
25 |
|
408 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Transition Volume |
kbcm |
780 |
628 |
459 |
1,557 |
1,054 |
1,709 |
315 |
1,219 |
832 |
725 |
239 |
|
9,518 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total
Transition Tonnage |
kt
|
1,619
|
1,425
|
1,116
|
3,907
|
2,495
|
4,439
|
815
|
2,753
|
1,946
|
1,865
|
628
|
|
23,009
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Fresh Quantities |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Waste Volume |
kbcm |
44 |
190 |
0 |
159 |
435 |
848 |
1,278 |
333 |
358 |
833 |
1,429 |
331 |
6,239 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Waste Tonnage |
kt |
114 |
515 |
0 |
446 |
1,174 |
2,337 |
3,516 |
879 |
944 |
2,227 |
3,874 |
907 |
16,931 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Volume |
kbcm |
7 |
1 |
0 |
2 |
10 |
8 |
34 |
2 |
12 |
24 |
50 |
17 |
166 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Tonnage |
kt |
19 |
3 |
0 |
6 |
27 |
20 |
93 |
5 |
32 |
64 |
139 |
45 |
453 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG In Situ Grade |
g/t |
1.27 |
1.89 |
0 |
1.98 |
1.76 |
1.44 |
1.52 |
1.65 |
1.63 |
1.57 |
1.42 |
1.62 |
1.53 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Diluted Grade |
g/t |
1.2 |
1.79 |
0 |
1.88 |
1.67 |
1.37 |
1.44 |
1.56 |
1.55 |
1.49 |
1.35 |
1.54 |
1.45 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Ore Recoverable Grade |
g/t |
0.88 |
1.02 |
0 |
0.92 |
0.97 |
0.93 |
0.94 |
0.96 |
0.94 |
0.94 |
0.94 |
0.94 |
0.94 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG Mined Ounces |
koz |
1 |
0 |
0 |
0 |
1 |
1 |
4 |
0 |
2 |
3 |
6 |
2 |
21 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
LG
Recoverable Ounces |
koz
|
1 |
0 |
0 |
0 |
1 |
1 |
3 |
0 |
1 |
2 |
4 |
1 |
14
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Volume |
kbcm |
10 |
35 |
0 |
83 |
128 |
141 |
462 |
36 |
27 |
252 |
432 |
108 |
1,715 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Tonnage |
kt |
21 |
93 |
0 |
234 |
338 |
386 |
1,288 |
96 |
74 |
685 |
1,188 |
301 |
4,703 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG In Situ Grade |
g/t |
3.35 |
2.32 |
0 |
2.96 |
2.71 |
2.2 |
2.21 |
2.19 |
2.3 |
2.41 |
2.16 |
1.97 |
2.29 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Diluted Grade |
g/t |
3.18 |
2.21 |
0 |
2.81 |
2.57 |
2.09 |
2.1 |
2.08 |
2.19 |
2.29 |
2.05 |
1.87 |
2.18 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Recoverable Grade |
g/t |
1.42 |
1.38 |
0 |
2.17 |
1.6 |
1.39 |
1.4 |
1.34 |
1.33 |
1.38 |
1.33 |
1.3 |
1.42 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Mined Ounces |
koz |
2 |
7 |
0 |
21 |
28 |
26 |
87 |
6 |
5 |
50 |
78 |
18 |
329 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Recoverable Ounces |
koz |
1 |
4 |
0 |
16 |
17 |
17 |
58 |
4 |
3 |
30 |
51 |
13 |
215 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Fresh Volume |
kbcm |
61 |
227 |
0 |
244 |
573 |
996 |
1,774 |
371 |
397 |
1,109 |
1,912 |
455 |
8,120 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Fresh
Tonnage |
kt
|
153
|
611
|
0 |
686
|
1,539
|
2,743
|
4,897
|
980
|
1,049
|
2,976
|
5,200
|
1,253
|
22,088
|
U6391 Twangiza
Reserve Tech_Report Final.docx |
July
2015 |
Page 115 of 187 |
SRK Consulting |
NI
43-101 Twangiza Main Report |
Material Type |
Units |
2015 |
2016 |
2017 |
2018 |
2019 |
2020 |
2021 |
2022 |
2023 |
2024 |
2025 |
2026 |
Total |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Mwana Valley Fill Deposit |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG
Volume |
kbcm
|
|
197
|
|
484
|
447
|
|
|
|
|
|
|
|
1,128
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Tonnage |
kt |
|
326 |
|
799 |
737 |
|
|
|
|
|
|
|
1,862 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG In Situ Grade |
g/t |
|
2.21 |
|
2.21 |
2.21 |
|
|
|
|
|
|
|
2.21 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Diluted Grade |
g/t |
|
2.1 |
|
2.1 |
2.1 |
|
|
|
|
|
|
|
2.1 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Recoverable Grade |
g/t |
|
1.85 |
|
1.85 |
1.85 |
|
|
|
|
|
|
|
1.85 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG Mined Ounces |
oz |
|
21,964 |
|
53,855 |
49,688 |
|
|
|
|
|
|
|
125,507 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
HG
Recoverable Ounces |
oz
|
|
19,328
|
|
47,393
|
43,726
|
|
|
|
|
|
|
|
110,447 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Ore proportion by degree of Oxidation (Weathering)
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Oxides |
% |
55 |
58 |
83 |
29 |
33 |
22 |
10 |
61 |
44 |
4 |
|
|
36 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Transition |
% |
43 |
37 |
17 |
60 |
44 |
54 |
12 |
34 |
50 |
54 |
28 |
|
39 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Fresh |
% |
2 |
5 |
0 |
11 |
23 |
24 |
78 |
6 |
6 |
42 |
72 |
100 |
26 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total
|
% |
100
|
100
|
100
|
100
|
100
|
100
|
100
|
100
|
100
|
100
|
100
|
100
|
100
|
U6391 Twangiza
Reserve Tech_Report Final.docx |
July
2015 |
Page 116 of 187 |
SRK Consulting |
NI
43-101 Twangiza Main Report |
|
Table 15-8: |
Mine Production Schedule Summary
|
Material Type |
Units |
2015 |
2016 |
2017 |
2018 |
2019 |
2020 |
2021 |
2022 |
2023 |
2024 |
2025 |
2026 |
2027 |
2028 |
Total |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Mining Schedule |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
ROM Mined |
kt |
1,800 |
2,164 |
1,595 |
2,768 |
2,243 |
1,626 |
1,655 |
1,631 |
1,700 |
1,700 |
1,700 |
301 |
|
|
20,883 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
In Situ Grade |
g/t |
2.89 |
2.65 |
2.65 |
2.55 |
2.39 |
2.73 |
2.22 |
2.35 |
2.28 |
2.4 |
2.32 |
1.97 |
|
|
2.49 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Diluted Grade |
g/t |
2.74 |
2.52 |
2.51 |
2.42 |
2.27 |
2.59 |
2.11 |
2.23 |
2.16 |
2.28 |
2.2 |
1.87 |
|
|
2.37 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Recoverable Grade |
g/t |
2.1 |
1.94 |
2.16 |
2 |
1.74 |
1.92 |
1.48 |
1.75 |
1.54 |
1.39 |
1.38 |
1.3 |
|
|
1.77 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Low Grade Mined |
kt |
39 |
130 |
27 |
30 |
82 |
51 |
110 |
156 |
72 |
71 |
139 |
45 |
|
|
953 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
In Situ Grade |
g/t |
1.46 |
1.36 |
1.12 |
1.42 |
1.45 |
1.31 |
1.47 |
1.18 |
1.44 |
1.55 |
1.42 |
1.62 |
|
|
1.39 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Diluted Grade |
g/t |
1.39 |
1.29 |
1.06 |
1.35 |
1.37 |
1.24 |
1.4 |
1.12 |
1.37 |
1.47 |
1.35 |
1.54 |
|
|
1.32 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Recoverable Grade |
g/t |
0.92 |
1.03 |
0.93 |
0.94 |
1.01 |
0.93 |
0.94 |
0.93 |
0.94 |
0.94 |
0.94 |
0.94 |
|
|
0.95 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Ore mined |
kt |
1,839 |
2,293 |
1,622 |
2,798 |
2,326 |
1,677 |
1,765 |
1,787 |
1,772 |
1,771 |
1,839 |
346 |
|
|
21,835 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
In Situ Grade |
g/t |
2.86 |
2.58 |
2.62 |
2.53 |
2.36 |
2.69 |
2.17 |
2.25 |
2.24 |
2.37 |
2.25 |
1.92 |
|
|
2.44 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Diluted Grade |
g/t |
2.72 |
2.45 |
2.49 |
2.41 |
2.24 |
2.55 |
2.06 |
2.13 |
2.13 |
2.25 |
2.14 |
1.83 |
|
|
2.32 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Recoverable Grade |
g/t |
2.07 |
1.88 |
2.14 |
1.99 |
1.72 |
1.89 |
1.44 |
1.68 |
1.52 |
1.37 |
1.35 |
1.25 |
|
|
1.74 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Opex Waste Mined |
kt |
1,144 |
4,533 |
10,378 |
8,701 |
10,656 |
9,852 |
6,615 |
7,213 |
2,457 |
3,189 |
3,989 |
907 |
|
|
69,633 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Waste Mined |
kt |
1,144 |
4,533 |
10,378 |
8,701 |
10,656 |
9,852 |
6,615 |
7,213 |
2,457 |
3,189 |
3,989 |
907 |
|
|
69,633 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Tonnes Mined |
kt |
2,984 |
6,826 |
12,000 |
11,499 |
12,981 |
11,529 |
8,380 |
9,000 |
4,229 |
4,960 |
5,828 |
1,253 |
|
|
91,468 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Strip Ratio |
w:o
|
0.62
|
1.98
|
6.4
|
3.11
|
4.58
|
5.87
|
3.75
|
4.04
|
1.39
|
1.8
|
2.17
|
2.63
|
|
|
3.19
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Process Schedule |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Total Tonnes Processed |
kt |
1,600 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
1,700 |
403 |
22,406 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Head Grade |
g/t |
2.68 |
2.39 |
2.52 |
2.32 |
2.13 |
2.71 |
2.48 |
2.21 |
2.16 |
1.84 |
2.15 |
2.17 |
2.22 |
1.45 |
2.29 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Contained Gold |
kg |
4,291 |
4,067 |
4,279 |
3,938 |
3,618 |
4,610 |
4,212 |
3,750 |
3,680 |
3,121 |
3,649 |
3,695 |
3,767 |
583 |
51,270 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Contained Gold |
koz |
138 |
131 |
138 |
127 |
116 |
148 |
135 |
121 |
118 |
100 |
117 |
119 |
121 |
19 |
1,648 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
CIL Plant Recovery |
% |
84 |
81 |
85 |
86 |
83 |
75 |
75 |
78 |
71 |
66 |
63 |
66 |
66 |
65 |
75 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Gold Recovered |
koz
|
116
|
106
|
117
|
109
|
96 |
112
|
102
|
94 |
84 |
66 |
74 |
79 |
79 |
12 |
1,246
|
U6391 Twangiza
Reserve Tech_Report Final.docx |
July
2015 |
Page 117 of 187 |
SRK Consulting |
NI
43-101 Twangiza Main Report |
As part of the Mineral Resource and
Reserve review process, a detailed analysis of the historical mine
reconciliation from the commencement of mining in October 2011 through to
December 2014 was completed by Twangiza Mining with assistance from SRK (UK).
This work was prompted by identification of a significant negative tonnage
reconciliation issue, of the order of -42% when compared with the previous
resource model.
The analysis resulted in identification
of key adjustments to be applied both to the previous resource block model (as
detailed in Section 14.3.4) and additional modifying factors to be applied to
the adjusted resource block model to estimate Mineral Reserves which are given
in Table 15-9.
|
Table 15-9: |
Resource to Reserve
Modifying Factors
|
Factor |
Tonnage |
Grade |
Metal |
Adjusted resource to actual Grade Control model |
0.86 |
1.03 |
0.88 |
Survey pick-up error |
0.97 |
1.00 |
0.97 |
Topo effect 2007 and 2010 |
0.99 |
1.00 |
0.99 |
Mining losses and Dilution |
1.00 |
0.98 |
0.98 |
Combined Factors to convert Adjusted
Resource to Reserve |
0.83 |
1.00 |
0.83
|
Key sources of variance were identified
as:
|
|
Density and underground artisanal mining adjustments in
the near-surface oxide as detailed in Section 14.3.4. |
|
|
|
|
|
Grade control model variance: -12% tonnage, +3% grade,
-9% metal due to slight shape change and higher cut-off grade |
|
|
|
|
|
Minor losses due to survey markup, topographical
variance, mining losses and dilution and stockpile reconciliation
accounted for a further -7% of tonnage. This was offset by 4% addition of
material from valley fill. |
|
|
|
|
|
A relatively low level of mining dilution (2.5%) and ore
loss (98%) has been used based on wide average ore widths, good ore
continuity, low dig line length to areas ratio and a low strip ratio.
|
The potential for an underestimation of
tonnage arising from a mill weightometer error was examined but remains
unconfirmed. This may have occurred where short term throughput rates exceeded
the 200t/hour weightometer limit. This requires further assessment as it impacts
historical (and therefore forecast) plant recovery estimates.
A detailed series of recommendations
were developed to monitor the mine reconciliation and management system.
15.6 |
Mineral Reserve Estimate |
The Mineral Reserve Statement uses the
definitions and guidelines given in CIM Definition Standards on Mineral
Resources and Mineral Reserves and is reported in accordance with NI 43-101
requirements.
Following a technical review of
operating performance in 2014, Twangiza Mining has included transitional and fresh ore types in the
updated estimate based on the proven ability of the current plant to
economically process non-oxide materials contained within the reserve pit shell.
An update of the Mineral Reserves was prepared as at the end of year 2014. The
new Reserve incorporates
U6391 Twangiza
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|
Addition of non-oxide material; |
|
|
|
|
|
Mining depletion from 1 October, 2011 to 31
December, 2014; |
|
|
|
|
|
Revised density as described in Section 14.3.4;
|
|
|
|
|
|
Revised modifying factors as described in 15.5;
|
|
|
|
|
|
Additional resource that has been added from
Twangiza West; and |
|
|
|
|
|
Changes in economic assumptions.
|
The Mineral Reserve is reported using a
breakeven cut-off grade of 0.84g/t Au applied to the recoverable gold grade and
a gold price of USD1,200 per ounce. The Mineral Resource was modified using the
factors provided in Table 15-9; the Mineral Resource is inclusive of the Mineral
Reserve.
The table below shows the updated
Mineral Reserves estimated to be contained within the Twangiza practical pit
design and associated production schedule.
|
Table 15-10: |
Summary of Twangiza
Mineral Reserves as at December 31, 2014
|
Category |
Deposit |
Tonnes
(Mt) |
Grade
(g/t Au) |
Ounces
(Moz Au) |
Proven |
Twangiza Main + North + Central + West |
7.47 |
2.41 |
0.58
|
Probable |
Twangiza Main + North + Central + West |
14.91 |
2.22
|
1.06 |
Proven + Probable |
|
22.38 |
2.28 |
1.64 |
A breakdown of the Mineral Reserves by
deposit and stockpile is provided in Table 15-11.
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|
Table 15-11: |
Mineral Reserve by Pit
as at December 31, 2014
|
|
|
|
PROVEN |
|
|
PROBABLE |
|
Deposit/Pit |
Material |
|
|
|
|
Gold |
|
|
|
Tonnes |
Gold Grade |
Gold |
Tonnes |
Grade |
Gold |
|
|
(Mt) |
(g/t) |
(Moz) |
(Mt) |
(g/t) |
(Moz) |
North |
Oxide |
0.00 |
0.00 |
0.00 |
1.17 |
3.25 |
0.12
|
|
Transition |
0.00 |
0.00 |
0.00 |
1.38 |
2.57 |
0.11
|
|
Fresh |
0.00 |
0.00 |
0.00 |
0.43 |
2.60 |
0.04
|
|
Subtotal |
0.00
|
0.00
|
0.00
|
2.97
|
2.84
|
0.27 |
Main |
Oxide |
0.01 |
1.24 |
0.00 |
1.97 |
1.97 |
0.12
|
Extension |
Transition |
0.00 |
0.00 |
0.00 |
0.75 |
2.29 |
0.05
|
(Central) |
Fresh |
0.00 |
0.00 |
0.00 |
0.10 |
2.10 |
0.01
|
|
Subtotal |
0.01
|
1.24
|
0.00
|
2.82
|
2.06
|
0.19 |
Main |
Oxide Stockpile |
0.57 |
1.07 |
0.02 |
0.00 |
0.00 |
0.00
|
|
Oxide |
4.56 |
2.36 |
0.35 |
1.07 |
1.51 |
0.05
|
|
Transition |
2.25 |
2.83 |
0.20 |
3.73 |
2.30 |
0.28
|
|
Fresh |
0.08 |
2.77 |
0.01 |
4.02 |
2.00 |
0.26
|
|
Subtotal |
7.46
|
2.41
|
0.58
|
8.82
|
2.07
|
0.59 |
West |
Oxide |
0.00 |
0.00 |
0.00 |
0.18 |
1.94 |
0.01
|
|
Transition |
0.00 |
0.00 |
0.00 |
0.00 |
1.21 |
0.00
|
|
Fresh |
0.00 |
0.00 |
0.00 |
0.12 |
1.97 |
0.01
|
|
Subtotal |
0.00
|
0.00
|
0.00
|
0.30
|
1.95
|
0.02 |
Totals |
Oxide Stockpile |
0.57 |
1.07 |
0.02 |
0.00 |
0.00 |
0.00 |
|
Oxide |
4.57 |
2.36 |
0.35 |
4.39 |
2.20 |
0.31 |
|
Transition |
2.25 |
2.83 |
0.20 |
5.85 |
2.36 |
0.44 |
|
Fresh |
0.08 |
2.77 |
0.01 |
4.68 |
2.05 |
0.31 |
|
Subtotal |
7.47 |
2.41 |
0.58 |
14.91 |
2.22 |
1.06 |
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16.1 |
Mining Method and Site
Layout |
Mining operations are based on
conventional open cast techniques. Excavation of the ore and waste rock on 2.5m
high mining benches will be performed by hydraulic excavators in backhoe
configuration loading out to 40 tonne nominal capacity articulated dump trucks
(Bell B40D or Cat equiv.). Mining is free dig in the oxide zone with any harder
transition and fresh materials drilled and blasted on 5.0 m benches.
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RoM will be direct fed to the mill with
any excess RoM stockpiled on the RoM pad and low grade ore stockpiles located
within 500 m of the pit. The stockpiles will be rehandled using a
front-end-loader (Cat 966) on a just-in-time basis to meet the plant throughput
requirements during the active mining period. Any remaining stockpile will be
treated after mining comes to an end for the Phase 1. Waste material mined is
hauled to the TMF 1 wall or a waste dump within a 500 m range from the pit exit.
A site layout plan is provided below in Figure 16-1and shows the layout of pits,
haul roads, access roads, stockpiles, waste dumps, process plant, tailings dam
facility (TMF) and site infrastructure.
Site preparation requires construction
of access roads, vegetation clearing, tree grubbing and topsoil stripping. The
main haul road is advanced to allow access between pits and then areas are
prepared for stockpiles, waste dumps and field workshops as required for each
separate mining area. The haul roads have a 10% gradient and 20 m width
considering the type of equipment used in the mining operations.
Vegetation clearing, tree grubbing and
top soil stripping are done in conformity to the direction of mining activities
and in consonance to the established DRC and/or Internationally accepted mining
and environmental regulations. The disturbance of vegetation and top soil
stripping are also done progressively according to the mining schedule of
activities. Top soil stripping is done to a depth of 200 mm and stockpiled at
designated dump locations close to major disturbed areas for future
rehabilitation works.
In general the Phase 1 of Twangiza has
not required drilling and blasting as mining has been concentrated in the
friable and oxidized material. On advancing mining activities into the more
competent transition material, minor drilling and blasting will be required
where the surface oxidation no longer permits free digging and loading. Drill
and blast activities have already commenced at Twangiza although the highest
proportion of mill feed ore is still derived from the oxide and upper transition
material, reducing the drill and blast activities significantly. In 2015,
transition and fresh material requiring drilling and blasting will comprise
approximately 59% of total material, reducing thereafter (with increased
production from shallower oxide pits and Valley Fill material) to average about
30% over the five year period.
Drill and blast design parameters and
input costs have been based on actual operating estimates for the Namoya Mining
SA operation in the DRC. Based on this analysis a drill & blast unit cost of
USD 0.30 /t has been estimated. A Sandvik Pantera DP 1500i rig will be used to
drill the blast holes. Operations technical expertise and training will be
sourced from in-house and the Namoya operation.
Loading and hauling activities will be
conducted with two hydraulic excavators in backhoe configuration equipped with
3.7 m3 nominal capacity buckets. Haulage will be carried out with
articulated dump trucks with a rated payload of 40 tonnes.
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Excavating and loading of the 5 m
mining benches will be undertaken in two flitches, of 2.5 m depth. Mining
geologists will closely supervise and monitor the selective mining of ore to
minimize dilution and ore loss. They will also supervise and control dispatch of
haulage trucks to appropriate destinations by material type - either to RoM pad,
low grade stockpile or waste dump in accordance with grade control ore mark-ups
based on cutoff grades provided by the Mine Planning Team. The services of
other technical support sections including the Survey, Geotechnical and
Dewatering teams will also be utilized for such routine mining activities as
survey of pit floors, and setting outs, mapping, pit wall monitoring and for
water control.
Apart from the main mining fleet,
ancillary equipment will be used to support the mining operations. Ancillary
equipment encompasses mine units which are not directly responsible for drilling
and blasting, loading and haulage of high grade ore, low grade ore and waste
materials, but are used to support the major production units and provide safe
and clean working areas. Such activities include dozing, batter trimming, road
maintenance, floor level clean-up and levelling, safety berm and bund
installation and dewatering activities. Pit supervisors are engaged to ensure
that the mining personal and fleets are effectively managed and utilized for
optimal production.
Some of the waste material from the
Main and Main Extended (Central) pit is scheduled to be used for the
construction of the TMF Phase 1 wall while the remainder, including the waste
mined from other mine pits will be dumped at designated waste dumps within the
catchment of the TMF for eventual re-handling and construction of the tailings
dam walls, if wet weather conditions do not permit direct tipping onto the
pad.
The waste dump development will be
based on a geotechnical design which conforms to the mountainous terrain.
Parameters used in the design of the waste dumps are described below:
|
|
Dump face angle - 35° |
|
|
|
|
|
Batter angle (rehabilitation) - 25° |
|
|
|
|
|
Berm width - 21 m |
|
|
|
|
|
Lift Height 15 m |
|
|
|
|
|
Average Dump Height 60 m |
|
|
|
|
|
Average Standoff Distance from Pit 350 m
|
|
|
|
|
|
Mining Work Schedule |
Mining activities will be scheduled
according to a continuous 12 hour shift roster 7 days per week, 2 shifts daily
for 365 days of operation. Ore will be preferentially mined on day shift due to
higher visibility and to allow for greater supervision. In order to achieve
this, a two crew system will be adopted for all direct operations personnel,
thus, a shift on day-shift, another on night-shift.
The operating time per shift will be
the actual time during the shift that the equipment is productively working and
this is equal to the total mechanically available time less all scheduled and
unscheduled delays.
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The effect of weather on mining
operations has been factored into the determination of effective working time
and equipment productivity. This means effective working hours have been reduced
to reflect rainfall and other weather delays. Estimation of rainfall delays was
based on 5-year rainfall average values recorded for the Twangiza Area.
Key elements of the mining work
schedule are presented in Table 16-1.
|
Table 16-1: |
Mining Work Schedule
|
|
Scheduled |
|
|
Activity |
Time |
|
Units |
Shift Change |
90 |
|
Minutes
per day |
Chop time/Break |
90 |
|
Minutes
per day |
Working Period |
21 |
|
Hours
per day |
Shift Duration |
10.5 |
|
Hours
per Shift |
Number of Shifts |
730 |
|
Shifts
per year |
Working Time |
7665 |
|
Hours
per year |
Weather Delays |
669 |
|
Hours
per year |
Effective Working Period |
6996
|
|
Hours per year |
Operations personnel include all paid
staff working with Twangiza Mining. It includes labour requirements for the
various departments comprising Process Metallurgy, Finance & Administration,
Process Engineering, Mining Operations, Mineral Resources Management, Community
Relations, Human Resources, Health, Safety & Environment, General
Management, Civil Engineering, and Mobile Fleet Maintenance. Additional services
required on short term or temporal basis may be sourced from contractors or
local hire companies.
Both introductory and extensive
training will be required for the national labour due to a lack of local
operational and technical skills.
A number of experienced expatriate
staff, comprising first level supervisory staff to senior management, will be
engaged for both operations and training. Other experts will also be engaged to
train the local workforce in technical, operational and maintenance skills on an
as-required basis.
Table 16-2 details the annual labour
required per department over life of the mine.
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Table 16-2: |
Labour Schedule
|
Department |
2015 |
2016 |
2017 |
2018 |
2019 |
2020 |
2021 |
2022 |
2023 |
2024 |
2025 |
2026 |
2027 |
2028 |
PROCESS
METALLURGY |
122 |
122 |
122 |
122 |
122 |
122 |
122 |
122 |
122 |
122 |
122 |
122 |
122 |
29 |
FINANCE &
ADMINISTRATION |
69 |
69 |
69 |
69 |
69 |
69 |
69 |
69 |
69 |
69 |
69 |
69 |
69 |
16
|
PROCESS ENGINEERING
|
99 |
99 |
99 |
99 |
99 |
99 |
99 |
99 |
99 |
99 |
99 |
99 |
99 |
23
|
MINING OPERATIONS
|
87 |
114 |
186 |
186 |
177 |
174 |
159 |
150 |
102 |
102 |
102 |
48 |
|
|
MINERAL RESOURCES
MANAGEMENT |
77 |
77 |
77 |
77 |
77 |
77 |
77 |
77 |
77 |
77 |
77 |
77 |
|
|
COMMUNITY RELATIONS
|
30 |
30 |
30 |
30 |
30 |
30 |
30 |
30 |
30 |
30 |
30 |
30 |
30 |
7
|
HUMAN RESOURCES |
28 |
28 |
28 |
28 |
28 |
28 |
28 |
28 |
28 |
28 |
28 |
28 |
28 |
7
|
HEALTH, SAFETY &
ENVIRONMENT |
14 |
14 |
14 |
14 |
14 |
14 |
14 |
14 |
14 |
14 |
14 |
14 |
14 |
3
|
GENERAL MANAGEMENT
|
5 |
5 |
5 |
5 |
5 |
5 |
5 |
5 |
5 |
5 |
5 |
5 |
5 |
1
|
CIVIL ENGINEERING
|
117 |
117 |
117 |
117 |
117 |
117 |
117 |
117 |
117 |
117 |
117 |
117 |
117 |
28
|
MOBILE FLEET MAINTENANCE |
49
|
49
|
49
|
49
|
49
|
49
|
49
|
49
|
49
|
49
|
49
|
49
|
|
|
TOTAL LABOUR |
697
|
724
|
796
|
796
|
787
|
784
|
769
|
760
|
712
|
712
|
712
|
658
|
484
|
115 |
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The mine equipment has been selected
based upon the annual mine production schedule and equipment productivity
estimates. The size and type of mining equipment is consistent with the project
scale with peak annual material movements averaging 12 Mt from Production Year
2017 to Year 2020. An asset table and pricing is provided in Table 16-3 while
the fleet requirements and a replacement schedule are provided in Table 16-4 and
Table 16-5 respectively.
|
Table 16-3: |
Mining Equipment Asset
Table
|
Asset Table CY2015 |
No. of
Units |
Price per Unit
(USD) |
Major
Mining Fleet |
|
|
Hydraulic Excavator
Hitachi ZX670 |
2 |
643,894 |
Truck Cat ADT745C
|
4 |
504,347 |
Drill Rig Sandvik
Pantera 1500 DR |
1 |
629,640 |
Track Dozer CatD8R
|
1 |
629,640 |
Grader CAT14M |
2 |
486,216 |
Minor Mining
Fleet |
|
|
Hydraulic Excavator
Kato 1430 |
2 |
298,768 |
FEL CAT966H |
2 |
346,680 |
Service Truck Bell
B20D |
1 |
271,562 |
Diesel Truck Bell
B20D |
1 |
271,562 |
Water Truck Bell
B20D |
1 |
271,562 |
Lighting Plants |
10 |
18,000
|
Light Vehicle |
10 |
53,500
|
Water Truck Actros MERC |
1 |
197,649 |
The replacement equipment schedule
provided in Table 16-5 takes into account the useful life of each item of plant
as scheduled using Xeras software. No replacement after 2023 reflects both a
reduction in material movements and therefore fleet numbers combined with
parking up of equipment as the remaining useful life of various items of plant
is expended.
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Table 16-4: |
Mining Equipment Schedule Matched to Production Requirements
|
Description |
|
2015 |
2016 |
2017 |
2018 |
2019 |
2020 |
2021 |
2022 |
2023 |
2024 |
2025 |
2026 |
2027 |
Hydraulic Excavator
Hitachi ZX670 |
|
2 |
2 |
5 |
5 |
5 |
4 |
3 |
3 |
2 |
2 |
2 |
|
|
Truck Cat ADT745C
|
|
4 |
9 |
17 |
17 |
17 |
17 |
13 |
13 |
8 |
8 |
8 |
2 |
|
Track Dozer Cat D8R
|
|
2 |
4 |
7 |
7 |
7 |
7 |
7 |
5 |
3 |
3 |
3 |
1 |
|
Water Truck Bell
B20D |
|
1 |
1 |
3 |
3 |
2 |
2 |
2 |
2 |
1 |
1 |
1 |
|
|
Service Truck Bell
B20D |
|
1 |
1 |
2 |
2 |
2 |
2 |
2 |
2 |
1 |
1 |
1 |
|
|
Diesel Truck Bell
B20D |
|
1 |
1 |
2 |
2 |
2 |
2 |
2 |
2 |
1 |
1 |
1 |
|
|
Lighting Plants |
|
10 |
10 |
14 |
14 |
14 |
13 |
10 |
10 |
7 |
7 |
7 |
1 |
|
Light Vehicle |
|
10 |
10 |
13 |
13 |
13 |
12 |
10 |
10 |
6 |
6 |
6 |
1 |
|
Hydraulic Excavator
Kato 1430 |
2 |
2 |
2 |
2 |
2 |
2 |
2 |
|
2 |
1 |
1 |
1 |
|
|
Water Truck Actros
Mercedes |
1 |
|
1 |
3 |
3 |
2 |
2 |
2 |
2 |
1 |
1 |
1 |
|
|
FEL Cat 966H |
2 |
2 |
2 |
2 |
2 |
2 |
2 |
2 |
2 |
2 |
2 |
2 |
|
2 |
Sanvik Panterra 1500
DR |
1 |
1 |
2 |
2 |
2 |
2 |
2 |
2 |
1 |
1 |
1 |
|
|
|
Grader Cat 14M |
2 |
|
4 |
7 |
7 |
6 |
6 |
6 |
5 |
3 |
3 |
3 |
1 |
|
Table 16-5: |
Replacement Mining Equipment Schedule
|
Replacement Units |
2015 |
2016 |
2017 |
2018 |
2019 |
2020 |
2021 |
2022 |
2023 |
Hydraulic
Excavator Hitachi ZX670 |
|
1 |
|
1 |
|
|
|
|
|
Water Truck
Bell B20D |
|
1 |
|
|
|
|
|
2 |
|
Service Truck
Bell B20D |
|
|
|
|
1 |
|
|
|
|
Diesel Truck
Bell B20D |
|
|
|
|
1 |
|
|
|
|
Hydraulic
Excavator Kato 1430 |
|
|
|
|
1 |
|
1 |
|
|
Water Truck
Actros Mercedes |
|
|
|
|
|
|
|
|
1 |
FEL Cat 966H |
|
|
|
|
|
|
|
1 |
1
|
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The process plant at Twangiza was
originally designed to process oxides ores at a nominal throughput of 1.3 Mtpa.
The flowsheet incorporated primary crushing of oxide ore using a mineral sizer,
wet scrubbing to remove fine clays, conventional secondary and tertiary crushing
to nominally -10 mm, two stages of grinding to 80% passing 75 µm, gravity
concentration to recover free gold and intensive cyanidation of the gravity
concentrates, cyanide leaching and carbon adsorption of gold, acid washing of
loaded carbon, elution and electrowinning of gold and smelting. Cyanide tailings
are detoxified using sodium metabisulphite and copper sulphate and the
detoxified tailings gravitate to the tailings dam.
Evaluations of the circuit indicated
that the plant could be modified to process an increased throughput of 1.7 Mtpa
and following treatment of a blend of oxide and transition ore, in a ratio of
3:1, a number of changes were identified and implemented. During this period the
grind was slightly coarser and the gold recovery was acceptable at approximately
82%.
17.2 |
Process Plant Design |
Overview of modifications completed to
increase throughput from 1.3 Mtpa to 1.7 Mtpa
The initial evaluation of the
processing capacity of the existing plant indicated that an annual tonnage of
1.3 Mtpa was possible. Subsequent in-depth investigations by a specialist
company established that the processing plant could, with modifications, be
modified to increase the annual throughput to 1.7 Mtpa. In addition the
modifications took account of the likely increase in the proportion of harder
transition ore in the plant feed.
The aim of the process design component
of the 2011 economic assessment was to complete a detailed investigation into
modifications targeted for upgrading the plant to 1.7 Mtpa and to establish a
capital cost associated with these modifications.
Priority was given to the minimizing of
production downtime during equipment installation and the construction
philosophy.
The installation of new plant equipment
was planned with the majority of the installation work taking place during the
ramp-up phase of the processing plant towards achieving nameplate capacity of
1.3 Mtpa. The steel structures and pipe work in the areas requiring more
extensive modifications would be pre-erected where practical and installed
during shutdowns specifically planned for these events, or during periods of
planned maintenance.
In order to achieve the increased
throughput target each section of the process plant was reviewed and upgrades
and modifications identified. The modifications identified were:
|
|
Install a secondary feed system for the addition of ore
crushed in the open pit or reclaimed from stockpile. |
|
|
|
|
|
Install a cover over the ROM stockpile to reduce the
impact of rain on the materials handling characteristics of high clay
content ores. |
|
|
|
|
|
Install a Rock Breaker at the primary crusher dump
hopper. |
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|
|
Install additional Power Pack on existing Apron
Feeder Drive to increase the speed of the unit. |
|
|
|
|
|
Install a larger capacity Mineral Sizer - ALP
650 Series with 2410mm wide throat. |
|
|
|
|
|
Increase the speed of various conveyors in the
crushing circuit. |
|
|
|
|
|
Install a more robust Secondary Crusher. |
|
|
|
|
|
Install a more robust Tertiary Crusher. |
|
|
|
|
|
Modify the system for transferring the scrubber
fines between the scrubber and the mil by upgrading the transfer pumps and
rerouting the piping to eliminate the numerous bends which restrict the
flow of pulp. |
|
|
|
|
|
Upgrade the Mill Cyclone Feed pumps,
Hydrocyclone Cluster and all associated piping, including the cyclone
overflow to CIL. |
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Add/ Upgrade Process Water Pumps. |
|
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Install an additional Linear Trash screen.
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|
|
|
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Install 4 new CIL Leach Tanks complete with
agitators, mechanical screens and carbon transfer pumps to increase the
leach residence time. |
|
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|
|
|
Change static and air powered equipment in
existing CIL (old CIP) tanks to electrically powered units - mechanical
screens and carbon transfer pumps. |
|
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Install larger capacity Air Blowers
/Compressors. |
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Install an additional Tower Crane in the CIL
circuit to improve maintenance downtime. |
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|
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|
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Install an additional Elution Circuit complete
with Regeneration Kiln to increase the amount of carbon that can be
treated and to improve equipment availability. |
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Upgrade Raw Water Supply Pump System and
Piping. |
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Upgrade Gland Service Water System. |
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Upgrade Tailings Return Water System (Pumps
& Pipe Line). |
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Install a new electrical MCC for the new items
of equipment. |
17.2.1 |
Front End / Crushing and
Screening |
Utilizing the existing equipment as it
now stands, throughput tonnages of 300 wet tonnes per hour were considered
achievable
The existing primary MMD Mineral Sizer
was replaced with a larger, more powerful ALP 625 Series, double drive unit,
allowing for an increase in throughput up to 300 wet tonnes per hour.
Operating experience with the new ALP
Mineral Sizer quickly highlighted shortcomings in the tooth design and tooth
replacement strategy, particularly when processing mudstone ore during the wet
season. The tooth change out regime on the new ALP Mineral Sizer was radically
changed from the design even tooth pattern to an uneven one increasing the
nip angle between the teeth significantly thereby improving throughput and
reducing down time and maintenance costs.
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The existing apron feeder drive and
primary conveyor feed conveyor [21-CVR-01] drive units were found to be adequate
and did not need upgrading as had been planned to achieve the throughput of 1.7
Mtpa. The speed of the apron feeder feeding the ALP Mineral Sizer was capped at
72% to eliminate overloading of the primary feed conveyor. The conveyor belt
width was increased from 750 mm to 800 mm and the installation of additional
training idlers on the carry and return portions of the belt greatly increased
belt operating time and reduced spillage.
The scrubber fines pumps were upgraded
with new 75 kW motors and upgraded VSDs (variable speed drives).
A single 315mm OD pipeline with smooth
long radius bends was installed to replace the original twin 250mm lines with
sharp 90 degree elbows that were prone to sanding out.
The original Zenith HPC220, 220 kW
Secondary Cone Crusher and the Zenith HPF220, 220 kW Tertiary cone crusher, were
removed and replaced with, more robust Secondary and Tertiary, FLSmidth Raptor
XL300 Crushers. Installation was completed in June 2013.
17.2.2 |
Secondary Plant Feed Point |
During the implementation of the
expansion project it became evident that overall plant availability and
utilization could be enhanced by feeding crushed and screened ore into the plant
via a secondary feed point. Originally a small feed hopper was installed over
23-CVR-01 in the crushing area to facilitate the loading of alluvial ore mined
from below the TMF into the plant. This allowed ore to be fed directly to the
grinding circuit via the mill feed bin, bypassing the front end altogether. This
concept was developed further during 2014 through the installation of an
additional feed conveyor located adjacent to 23-CVR-08, the mill feed bin
conveyor. The use of available mobile crushing and screening equipment to
provide -25 mm ore to the grinding circuit has now become permanent operating
practice providing supplementary feed on an hourly basis as required to keep the
grinding circuit operating effectively while the front end of the plant is
shutdown.
17.2.3 |
Upgrade of feed
weightometers |
At the higher feed tonnages the
scrubber feed weightometer and the Mill no. 1 feed weightometer were out of
range which potentially introduced significant metallurgical accounting errors
in terms of measurement of the feed tonnage. Both weightometers have been
changed to larger units.
The implementation of the expansion to
1.7 Mtpa was incremental in nature and it was realized that an upgrade to the
existing grinding circuit was not required at the time but recognized that it
would be needed for any further expansion.
However a number of minor changes were
made to the installed equipment in the grinding circuit to enhance circuit
performance and to effectively operate at 1.7 Mtpa.
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The trommel panels on both No.1 and
No.2 ball mill were changed from 10 mm polyurethane panels to 10 mm wire screens
to increase the effective open area and screening efficiency. The cyclone feed
pipelines have been extensively replaced with durable rubber mining hose with
smooth long radius elbows for improved pumping efficiency and longer service
life. A fourth cyclone was added to the cyclopack increasing the number
available from three to four.
An additional 9m2 Linear
Trash Screen, required to handle the increase in flow rate with the planned
tonnage increase was ordered and installed in parallel with the current
operating Linear Trash Screen.
The existing Linear Trash Screen will
be relocated to feed the new #2 CIL Tank to be constructed, and the CIL will be
operated as two parallel trains.
The construction of four new,
412m3 CIL Tanks (same size as existing # 1 and 2 CIL Tanks) was
completed in September 2013. The new tanks have been constructed at the head of
the existing CIL train, in order to increase the residence time at the original
tonnage and enhance the residence time with the future increase in tonnage.
Four new Kemix agitators are installed
in the new CIL tanks. Ten complete Kemix MPS700(P) mechanically swept interstage
screens have also been installed throughout the CIL circuit. All the screens
will have 800 µm baskets.
The existing static, air swept, wedge
wire screens in CIL tanks #8 thru 13 were prone to choking and have been
replaced with mechanically swept Kemix MPS700(P) screens, the same as installed
in the first eight CIL tanks. All screens will be fitted with 800 µm screens
which will allow the tanks to be operated at higher pulp densities than before,
thereby allowing better mixing of the carbon in pulp, improved adsorption
efficiencies, reduced dissolved gold loses, savings in reagents, increased
residence time and improved performance of the CIL in general.
The airlifts for carbon transfer in CIL
tanks #8 through 13 have been replaced with vertical spindle pumps, thereby
eliminating the excessive use of blower air, which is better utilized for
agitation and leaching in the CIL. A total of ten FLSmidth Krebs 4x4 1.8L
vertical carbon transfer pumps have been installed throughout the CIL
circuit.
Two used (110 operating hours) high
capacity Atlas Copco ZE4 200kW 3.5bar blowers have been installed and tied into
the existing CIL air sparging system to enable oxygen levels in the CIL to be
increased, thereby improving leach kinetics and overall recoveries.
A second tower crane (used) has been
installed for construction and servicing of the new CIL tanks. This crane will
be positioned so as to enable it to also be utilised to service the milling and
cyclone area as well.
17.2.6 |
Acid Wash, Elution and
Regeneration |
The existing acid wash and elution
system will remain unchanged.
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The original vertical carbon
regeneration kiln was removed to make room for the new Kemix 300 kg/hour diesel
fired horizontal regeneration kiln. A new (second) elution system, complete with
horizontal regeneration kiln was installed and commissioned in August 2014 and
now operates in parallel with the existing carbon treatment system. This allows
for more efficient and more complete stripping of gold from the carbon, ensuring
all stripped carbon is regenerated before it is returned to the CIL circuit,
thereby improving recoveries and maximizing gold production.
17.2.7 |
Electro-winning and
Smelting |
The new elution system included two
complete electro-winning cells, gold sludge recovery system, calcine oven,
smelting furnace, gold scale, doré handling tools, a bullion safe, work bench
and associated fume hoods and extraction fans.
The existing tailings carbon safety
(catch) screens capacity will be assessed as tonnage throughput increases, to
establish if there is a need for additional equipment to accommodate the
increase of the capacity of the system.
The present Detox and tailings delivery
system will be evaluated at high tonnage throughput.
17.2.9 |
Water Supply Systems |
The capacity of the raw water and
tailings return water systems has proven adequate for the expansion to 1.7
Mtpa.
17.2.10 |
Summary and Conclusions |
Overall the expansion from 1.3 Mtpa to
1.7 Mtpa can be considered successful in its overall scope to increase
throughput as designed. However, while the expansion project is essentially
complete, a number of processing issues have become apparent and will need to be
addressed as part of the on-going optimization program.
The operation of a total of 17 CIL
tanks in series has proven very difficult from the carbon management
perspective. Transferring carbon sequentially upstream through all tanks has
proven difficult to manage and control effectively. Numerous operating
strategies for slurry movement through the tanks and carbon transfer regimens
have been attempted. At present the slurry flows sequentially through all 17
tanks in series. Attempts to operate two parallel slurry streams proved
impractical when one single tank was down for maintenance. As of writing this
report the first four tanks in the leach train, the new tanks installed in the
expansion, are operated as leach tanks. Loaded carbon is transferred up stream
only to the old No.1 CIL tank. Thus the circuit now operates as a CIP plant.
Further changes to this arrangement may be adopted in the future to fully
optimize gold leaching, carbon adsorption and carbon management and effectively
process the different ore types present in the ore body.
Residence time in the CIL circuit has
been the subject of much debate and deliberation since start-up and ramp-up to
1.3 Mtpa and through the expansion to 1.7 Mtpa. The determination of the optimum
residence time required for the oxide, transition and fresh ore is necessary
before any further tankage additions are contemplated.
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17.3 |
General Process Plant
Description |
The following describes the flow sheet
of the process plant as of December 2014 treating 1.7 Mtpa of a blend of soft
oxide and harder transitional ore. The plant now consists of the following
sub-sections:
|
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Mobile Crushing and Screening |
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|
ROM Pad Storage Area and Primary Feed Point
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In Plant Crushing and Scrubbing |
|
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Milling and Classification |
|
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Gravity and Intensive Cyanidation |
|
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Trash Removal |
|
|
|
|
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Leach and Adsorption Carbon-in-Leach
(CIL)/Carbon-in-Pulp (CIP) |
|
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Carbon Safety and Detoxification |
|
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Tailings Dam Storage and Return |
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Acid Wash |
|
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Elution |
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Electrowinning |
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Regeneration |
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Gold Room |
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Reagents and Consumables |
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Air Services |
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|
|
Water Services |
17.3.2 |
Mobile Crushing and
Screening |
In order to control the size of ROM fed
to the Plant, mobile crushers and screens organized under the outside section or
ore re-handling section in the pit are used for the realization of pre-crushing
and pre-screening of ROM ore which is then selectively routed either to the ROM
pad (-100mm to +25mm) or the secondary feed point (-25 mm) for plant feed.
17.3.3 |
ROM Pad Storage Area Primary Feed
Point |
The ROM pad as the primary feed point
is covered with a roofing structure since the end of Q2 2014. The completion of
the ROM pad roof allows dry storage of up to 25,000 tonnes of a mixture of both
ROM ore and pre-crushed and pre-screened ore products from the rehandling
section in the pit. This solution was implemented to protect stockpiled ore from
the rain and thereby control the moisture content of the ore. Historically, ore
processed through the plant exhibits significant changes in overall moisture
content from dry to wet season. Wet sticky clay ore is extremely difficult to
process in the crushing section of the process plant. Controlling the moisture
content of ore, especially ores with high fines or clay content, minimizes the
effect of changes in the viscosity of the ore. Drier high clay content ore has
better flow and processing characteristics than ore that has been exposed to
rainfall.
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17.3.4 |
In-Plant Crushing and
Scrubbing |
Blended ore from either the ROM storage
pad or direct from the pit are fed into the in-plant crushing circuit.
The front end of the plant will treat ore at a rate of up to 300 wet
t/h to a product of size 100% passing 100 mm, through the ALP mineral sizer.
The blended ore is tipped into the ROM
bin using rear dump trucks or fed from the ROM pad stockpile by front end
loader. The ROM bin has a capacity of 50m³. A variable speed apron feeder
withdraws the ore from the bin at a controlled rate and discharges it into the
mineral sizer. A small amount of oversize ore (rejects) from the crusher are
collected in a sizer rejects stockpile. The mineral sizer reduces the ore to a
product size of 100% passing 100 mm. The crushed product is discharged onto the
crusher discharge conveyor which, in turn feeds into a rotary scrubber.
Process water is added to the scrubber
to produce discharge slurry with a 50% solids density by weight. The scrubber
discharge slurry overflows onto a vibrating double deck scrubber discharge
screen. The upper screen deck is set at 50mm x 10 mm slotted. The lower deck is
2 mm.
The secondary crusher feed conveyor
transfers the screen top deck oversize to the secondary crusher. The bottom deck
oversize is conveyed into the mill feed bin by means of a scrubber screen
discharge conveyor, transfer conveyor and crusher product conveyor. The screen
bottom deck undersize slurry gravitates into a 12m³ scrubber discharge sump,
from where it is pumped to the mill discharge sump.
The secondary crusher reduces the ore
size to a product of 100% passing 37 mm. The crusher discharge conveyor
transfers the crushed ore to the mill feed bin which has a capacity of 480
m3.
17.3.5 |
Secondary Feed Point |
The secondary feed point has been
developed to provide a means of feeding the process plant when the front end of
the plant, the mineral sizer, scrubber, scrubber discharge screen, fines pumps,
in plant crushing and conveyor belts, are shut down.
Operation of the plant and gold
production can be maintained, albeit at a lower overall tonnage, during such
times. On regular operating days the secondary feed point is used to supplement
the tonnage processed by the front end of the plant.
The -25mm screened fines from the
re-handling section or direct from the pit are stockpiled in the security
airlock. These fines are fed to the plant through a grasshopper conveyor which
passes through the security fence discharging directly onto the mill feed bin
feed conveyor.
17.3.6 |
Milling and Classification |
The milling circuit consists of one
primary ball mill (Mill 1) and one secondary ball mill (Mill 2), which are
capable of treating up to 250 t/h of blended ore.
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Crushed ore, from either or both
primary and secondary feed points is fed into the No.1 ball mill using a
variable speed reclaim conveyor. The dimensions of this ball mill are 3.638mØ
(inside liners) x 4.660m EGL and its installed power 1300kW. The ball mill
discharge slurry overflows onto No.1 mill trommel screen with a screen size of
10mm. The oversize from the screen is collected in No. 1 ball mill scats bunker.
The undersize slurry gravitates into a 17 m³ mill discharge hopper, where it
combines with the slurry from the No.2 ball mill. The combined mill discharge
slurry, with a solids density of 52% by weight, is pumped by one of the cyclone
feed pumps to the mill discharge cyclone cluster.
The cyclone cluster classifies the
slurry to produce an overflow at a target of 80% passing 75 µm, with a solids
density of 36% by weight. The cyclone overflow gravitates to a linear trash
screen.
The cyclone underflow, at 67% solids
density by weight, gravitates to the 2m³ cyclone underflow splitter box 1, where
there is an off take to the gravity circuit which is currently not used. The
cyclone underflow then reports to a 2 m³ cyclone underflow splitter box 2, from
where 33% of the stream is sent to the No. 2 ball mill via a velocity break box.
The dimensions of the second ball mill are 3.088 m Ø (inside liners) x 3.035 m
EGL and its installed power 550 kW. 67% of the stream from the splitter box 2
gravitates back to the no. 1 ball mill.
Milling and cyclone spillage is
contained in a bunded area that has a sloping floor to direct spillage to two
sumps. One sump is allocated to the mill feed spillage and the other the mill
discharge and cyclone spillage. Each sump has a vertical spindle spillage pump.
Both spillage pumps discharge into the mill discharge sump. There are two more
spillage pumps, one for the No.1 ball mill scats bunker and another for the No.2
mill bunker scats bunker area.
17.3.7 |
Gravity and Intensive
Cyanidation |
At the current plant throughput of 1.7
Mtpa it is not possible to operate the gravity circuit effectively while
maintaining the plant throughput. Effective operation of the Knelson
concentrator requires a specific and large volume of flush water. A high
throughput thus upsets the water balance in the ball mill discharge pump box
limiting overall throughput.
Although this section of the plant is
currently not used the following process description still applies.
A portion of the cyclone underflow can
be diverted from splitter box No.1 over the 1.12 mW x 2.40 mL gravity feed
screen to remove the +2mm material. Dilution water is added onto the screen to
dilute the feed slurry to a solids density of 45% by weight. The screen oversize
is recycled back to the no.1 ball mill. The screen undersize is fed to the
Knelson concentrator to recover the free gold. The tailings from the Knelson
concentrator is returned to the cyclone feed sump.
The concentrates from the Knelson
concentrator is periodically discharged into a concentrate cone for dewatering,
with excess water overflowing to the floor, from where it is pumped to the mill
discharge sump using the gravity spillage pump. Concentrates are stored in the
concentrate cone until tabling or the intensive leaching cycle is ready.
Presently gravity concentrates are treated by means of a Johnson Table and the
gold concentrate collected and smelted directly after calcining.
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An Intense Cyanide Leach reactor has
been installed and is yet to be properly commissioned once the pregnant solution
tanks for the system have been relocated to make space for the second complete
elution system.
The Intense Cyanide Leach Reactor (ILR)
- Gekko will process the gravity concentrates. At the beginning of each batch
leach cycle, the entire contents of the feed cone will be discharged into the
ILR drum. Excess water from the drum, during the loading cycle, will overflow
into the ILR drum sump and will be pumped by the recirculation/transfer pump to
the ILR solution storage tank.
The levels in the solution storage tank
will be adjusted through the addition of raw water, caustic and cyanide
solutions before the commencement of the leaching step. Leaching will be
effected by re-circulating the 2% cyanide solution through the rotating reactor
drum. The overflow will gravitate to the sump and will be pumped back to the
solution storage tank. Hydrogen peroxide will be added to the ILR sump to
provide oxygen for the leaching step.
At the end of the leach cycle, which
will range between 1416 hours, the drum will be stopped and the solution in the
drum allowed to drain into the ILR sump and pumped to the solution storage tank
where it will be clarified by adding flocculant. The clarified solution will be
pumped to the gravity pregnant tanks. Wash water will be added to the drum to
wash entrained solution from the solids and allowed to clarify in the solution
tank before being pumped to the leach tanks.
The leached and washed solids will be
emptied by running the reactor drum in reverse and pumped to the mill discharge
hopper.
The pregnant solution stored in the
gravity electrowinning tanks will be pumped to a dedicated electrowinning cell,
wherein gold will be deposited onto steel cathodes.
The fume hood fan is installed to
remove potential poisonous and explosive gases evolving during electrowinning.
Owing to the use of strong caustic and
sodium cyanide solution in the ILR, a safety shower is provided in this area. It
is activated by a foot pedal and equipped with an eye bath.
The leach feed slurry is passed over a
9 m2 linear trash screen to remove any tramp material before it is
fed to the leach circuit. The screen undersize is sampled, for metallurgical
accounting and control, by the leach feed sampler as it gravitates to the leach
and CIL circuit. There is a provision for a second 9 m2 trash linear
screen to feed the leach circuit.
Oversize from the trash screen is
collected in a trash basket.
17.3.9 |
Leach and Adsorption
Circuit |
The trash screen undersize gravitates
into a 4 m³ leach feed boil box. The boil box channels the slurry into the first
leach tank, and a provision is made to feed the second leach tank from the box
when leach tank 1 is off-line.
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The leach and adsorption circuit has a
total of seventeen tanks, with a total residence time of 11.5 hours. The circuit
is operating currently as a CIP plant. The CIP configuration is composed of 4
leach tanks and 13 adsorption tanks. The slurry and carbon in the tanks is
maintained in suspension by the action of the dual impeller tank agitators, each
with 22 kW of installed power.
All leach and adsorption tanks are
equipped with Kemix interstage wedge wire screens, which prevent any migration
of carbon from one tank to another during slurry inter-tank flow. The flow of
slurry from one tank to another is affected by the pumping action of the
internal impeller mechanism of the interstage screens through an interconnecting
launder system.
Each screen is periodically lifted from
the tank onto a wash frame for cleaning. A spare screen is provided to replace
other screens during cleaning or maintenance of any screen. A high pressure, low
volume wash pump is used to clean blocked screens. A tower crane is utilised to
lift screens and for general maintenance purposes.
The cyanide solution is added into the
feed boil box from a ring main. A TAC1000 automatic cyanide analyser is used to
ensure efficient addition of cyanide solution. A provision is also made to
manually add cyanide into the first three CIL tanks in the event that cyanide
concentration is too low.
Lime slurry is added into the leach
feed boil box for pH adjustment. Blower air, at 250 kPa, is introduced (sparged)
into the tanks for oxidation during gold dissolution.
Interstage carbon transfer pumps
transfer carbon from the last tank through to the first tank in the adsorption
circuit. Loaded carbon is pumped from the first adsorption tank using a loaded
carbon recovery pump to a 1.8mL x 0.9mW loaded carbon screen. With the expansion
programme an extra Elution circuit is installed for more flexibility in gold
stripping from loaded carbon. There is therefore provision to route the loaded
carbon either to the old elution circuit or to the new one.
Eluted and regenerated carbon is added
to the last tank or the second last tank if the last tank is off-line.
Spillage in the Leach and Adsorption
circuit area is contained in a bund and pumped by two dissolution area sump
pumps.
Three safety showers are provided and
each is activated by a foot pedal and equipped with an eye bath.
17.3.10 |
Carbon Safety and
Detoxification |
The tailings slurry from the CIL
gravitates to the first of 2 cyanide destruction tanks in series. A bypass
facility is provided to direct the tailings slurry to the second tank whenever
tank 1 is offline.
Sodium metabisulphite, copper sulphate
and blower air (at 250 kPa) is added to the circuit to provide sulphur dioxide,
copper catalyst and oxygen, respectively. The free cyanide and/or weakly bound
metal cyanide complexes present in the tailings slurry oxidizes to the less
toxic cyanate (OCN) according to the reaction:
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Lime slurry is added into the circuit
to neutralise the sulphuric acid that is generated during the reaction, and thus
maintain the pH within a range of 8 10.
The slurry exiting the second cyanide
destruction tank gravitates into a 2 m³ safety screen feed splitter, where it is
split and fed to two 4.8 mL x 1.5 mW carbon safety screens. Any carbon that
escapes from the adsorption section, as a result of damaged interstage screens,
is recovered as screen oversize and collected in a safety screen oversize
basket. The screen undersize slurry gravitates into a 20 m³ guard screen
undersize hopper. The final tails slurry gravitates from the hopper to the
tailings dam. A sample is taken by a cross-cut tails slurry sampler.
Spillage in the carbon safety and
detoxification area is contained in a bund and pumped to the feed splitter using
a spillage pump.
Two safety showers are provided in this
area. Each shower is activated by a foot pedal and equipped with an eye bath.
17.3.11 |
Tailings Dam Storage and
Return |
Process water from the tailings dam is
pumped by pontoon pumps to a 200 m³ tank. Return water pumps transfer water from
this tank to another 200 m³ tank, from where it is pumped to a 1,000 m³ tank
plant process water tank.
Loaded carbon from the first CIL tank
is pumped to a 1.8 mL x 0.9 mW loaded carbon screen. Spray water is added onto
the screen to wash the slurry off the loaded carbon. The slurry (screen
undersize) is returned to the first adsorption tank. The washed loaded carbon
gravitates into a 7.6 m³ acid wash hopper, and then to the acid wash column. The
column accommodates a loaded carbon batch of 3t.
During acid washing the acid wash pump
circulates dilute hydrochloric acid solution (~3% HCl) from a 10 m³ acid wash
tank through the acid wash column at a rate of 2 bed volumes per hour for a
period of 1 hour. The acid exiting the column is return to the acid wash tank
via internal strainers that prevents any carryover of carbon.
At the end of one hour of acid washing,
the acid wash pump is stopped. The content of the acid wash column is rinsed
with a volume of raw water equivalent to 2 bed volumes. The rinse effluent
exiting the column is directed to the carbon safety and detoxification circuit.
Rinsed carbon is hydraulically transferred to the elution column.
Periodically, typically after every 4
acid washes, the dilute acid in the acid wash tank is too contaminated for
further use. The solution is pumped to the carbon safety screen. Fresh acid wash
solution is prepared by filling the acid wash tank with raw water to a
pre-determined level and then pumping approximately 909 litres of 33% HCl in
order to make up a 3% HCl solution strength.
Spillage in the acid make-up area is
contained in a bund that has a spillage pump which pumps the spillage to the
carbon safety screen.
A safety shower is provided as a result
of the use of acid solution in this area. It is activated by a foot pedal and is
equipped with an eye bath.
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Two elution circuits (old and new
elution circuit) are installed to process loaded carbon at an acceptable rate to
recover the gold. The two circuits are independently connected to two
electrowinning circuits to deposit the gold, followed by two independent
regeneration kilns for carbon re-activation.
Gold elution from the loaded carbon is
carried out using the Zadra method. The elution liquor (12.5 - 2.9 % caustic
solution) is pumped from the eluate tank, via secondary and primary heat
exchangers into the elution column under elevated temperature and pressure. This
promotes a chemical reaction where gold adsorbed onto the carbon is removed by
washing with a number of bed volumes of hot water. The gold is thus stripped
from the carbon into the eluate solution.
The elution column is operated under a
pressure of typically 300 to 350 kPa for the old elution and 400 to 450 kPa for
the new elution. Heat required for the elution cycle is provided by diesel fired
elution heaters. The heater burners heat up thermal oil, which is used to
transfer heat to the eluant solution entering the primary heat exchanger en
route to the elution column. The pregnant electrolyte leaving the column is
cooled by the fresh eluant solution in the secondary heat exchanger. The primary
and reclaim heat exchangers are plate and frame type.
The pregnant electrolyte flows out of
the elution column to the electrowinning cells. In the cells gold is deposited
on the stainless steel wool cathodes. The electrolyte gravitates from the cells
back to the eluent solution tank from where it continues to be cycled through
the elution column until the elution is completed. The barren eluent is utilized
for a number of elutions before it is discarded to the Leach circuit and a fresh
batch of strip solution made up.
Once gold has been stripped from a
batch of loaded carbon, the elution cycle is complete. The eluted carbon is
hydraulically transferred from the elution column to the regeneration circuit or
to the last tank.
To reduce scale build up in the heat
exchangers a sulphamic acid solution is periodically circulated by the sulphamic
acid dosing pump through the heat exchangers. Spillage in the elution area is
contained in a bund that has a spillage pump which pumps spillage to leach.
A safety shower is provided in this
area. It is activated by a foot pedal and equipped with an eye bath.
The pregnant solution (eluate) from the
elution column is circulated through the electrowinning circuit until acceptably
low gold values are achieved on both the eluted carbon and the electrolyte. Gold
is electrowon from the eluate in cells using stainless steel wool cathodes. It
is deposited onto the cathodes as loosely adhering sludge. The circulation of
the projected 63.5 m³ of eluate through the elution column and electrowinning
cells continues for about 16 hours during which time the pregnant solution grade
decreases from a target of 120 to 5 ppm gold.
Loaded cathodes from electrowinning are
periodically remove from the cells to the cathode wash table in the gold room,
where the gold sludge is washed off by means of a high pressure washer. Gold
sludge from the cells is collected in a concentrate cone and drained into
buckets, from where it is transferred to the filter press in the gold room.
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A fume hood fan is installed to extract
potentially poisonous and explosive gases that evolve during electrowinning from
the cells and discharge outside the gold room.
Owing to the use of caustic in this
area, two safety showers are provided. Each safety shower is activated by a foot
pedal and equipped with an eye bath.
Spillage in the barren solution tank
area is contained in a bund and pumped to the Leaching circuit. Spillage
generated in the electrowinning cell area is also pumped to head of the CIL.
At the end of an elution cycle, eluted
carbon is hydraulically transferred from the elution column to the barren carbon
storage tank for regeneration. A provision is made for eluted carbon to bypass
the regeneration. The eluted carbon gravitates from the barren carbon storage
tank to the carbon regeneration kiln feed box which is fitted with a small
dewatering screen. Excess water is drained off as screen undersize, and the
oversize (carbon) is fed by spiral screw feeder to the 300 kg/hour. Horizontal
Diesel Fired Regeneration Kiln.
The kiln, operating at 700 °C, will
treat the entire carbon batch in a period of 10 hours.
The regenerated carbon is quenched with
water, before it can react with atmospheric oxygen, in a 7.2 m³ barren carbon
transfer tank, from where it will be hydraulically transferred to a vibrating
1.2 mL x 0.6 mW re-activated carbon fines screen. The screen undersize (carbon
fines) will gravitate into the guard screen undersize hopper. The screen
oversize (activated carbon) will gravitate into the last adsorption tank.
The filtered concentrate (gold sludge)
from the vacuum pan filter is placed in stainless steel calcining trays, up to 6
trays at a time and the trays are loaded into the calcining furnace operated at
800°C for drying. The calcining trays are then removed from the furnace and
placed on a cooling table and allowed to cool down.
When the dried product from the
calcining furnace has cooled down, it is weighed, mixed with fluxes (stored in
the flux storage box) in determined proportions. The gold sludge/flux mixture is
transferred to a smelting crucible in a diesel-fired smelting furnace operated
at 1,200 to 1400 C.
During smelting metal oxides form slag
and at the end of smelting the furnace crucible contents are poured into
cascading moulds mounted on a cascade trolley. The bullion collects in the first
mould with any excess collected in the second mould while slag flows and is
collected in a slag collection crucible. A steel slab is included to protect the
concrete floor.
Once cooled, the gold bars are cleaned,
sampled, stamped and stored in the safe prior to dispatch.
Two safety showers are installed in the
electrowinning and gold room area. They are activated by a foot pedal and
equipped with an eye bath.
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Grinding media, 75 and 50mm Ø balls are
used in No.1 ball mill and 50 and 25mm Ø balls are used in No. 2 ball mill. For
ease of transportation, grinding media is delivered in 200 litre drums. Grinding
media to mill 1 and ball mill 2 are charged using a bottom discharge skip and a
hoist.
Sodium cyanide is delivered to site in
1t bulk bags and packed into wooden crates, to limit the danger of spillages
during transportation. The bulk bags are transported using a fork lift from the
cyanide storage area to the cyanide mixing area, located in the reagent make-up
area.
Prior to the addition of cyanide
briquettes into the cyanide tank, the pH of the water in this tank is adjusted
to about 10 using caustic soda solution in order to prevent any formation of
hydrogen cyanide at low pH values.
When the pH has been adjusted, the
hoist lifts the cyanide bags to the bag breaker from where they are discharged
into a 21 m3 covered cyanide mix tank. The tank agitator ensures that
cyanide briquettes are completely dissolved during the make-up process to form a
25% cyanide solution by weight.
The cyanide solution is transferred
from the make-up tank to a 65 m3 cyanide solution storage tank using
the transfer pump.
One of the cyanide solution feed pumps
transfers cyanide solution into the cyanide ring main from which cyanide is
tapped off into the leach feed boil box or CIL tanks. A safety shower is
installed in the cyanide make-up area. It is activated by a foot pedal and
equipped with an eye bath.
Spillage in the cyanide make-up area is
contained in a bund and a spillage pump is used to pump cyanide into the leach
feed boil box.
Caustic is delivered to site in 25 kg
bags packed onto wooden pallets. The 25 kg bags on pallets are transported using
a fork lift from the caustic storage area to the caustic tank for make-up area.
The dissolution of caustic takes place in a 5m3 covered caustic tank.
Once delivered to the caustic area, the reagent hoist is used to lift the
caustic pallet to a platform on top of the caustic tank from where an operator
manually lift one 25 kg bag at a time onto the bag breaker. The caustic pearls
discharged into the caustic tank. The caustic tank agitator ensures that the
caustic pearls are completely dissolved during the make-up process to form a 20
% caustic solution by weight.
The caustic dosing pumps only runs for
the time required to deliver the various quantities of the reagent to various
distribution points, including the cyanide make-up, acid wash, and barren eluent
tank (elution).
A safety shower is installed in the
caustic make-up area. It is activated by a foot pedal and equipped with an eye
bath.
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Spillage in the caustic make-up area is
contained in a bund and a spillage pump is used to pump it into the leach feed
boil box.
Un-slaked lime (quick lime) is
delivered to site in 1t bags and kept in a lime store. The bags are hoisted and
transferred to the lime loading hopper by overhead crane, where the bags are cut
open. The hopper will be equipped with a dust filter that will help keep the
operation dust free.
Lime powder is transferred from the
hopper to a 45 m³ silo by means of a rotary feeder and a transfer blower. The
lime silo is also equipped with a dust collector. A rotary feeder and an
inclined screw feeder transfer lime from the silo, at a rate of 4 t/h, to the
lime slaker during the slaking period. An electric silo vibrator is fitted at
the bottom of the transfer hopper to enhance discharge of lime from the silo.
The lime slaker continuously convert
calcium oxide (quicklime) into calcium hydroxide or slaked lime in the form of
slurry of controlled consistency, and dilutes it to the required density of 15%
by weight.
Water and lime is added to the first
compartment of the slaker in measured proportions and vigorously agitated by
slaker mixers.
The shape of the vessel promotes
efficient slaking and the overflow into the adjacent compartment ensures the
necessary retention time, thus preventing the discharge of unslaked material.
The temperature is between 70° and 75°C.
Slaked lime overflows from the second
compartment of the lime slaker onto a vibrating grit classifier for the removal
of grit. The oversize grit is stockpiled and removed by a front end loader,
while the screen undersize gravitate into the slaked lime sump fitted with a
mixer to keep the lime particles in suspension. A lime transfer pump transfers
slaked lime to a 140 m³ lime storage and dosing tank, where it will be kept in
suspension by a mixer.
Lime slurry is pumped to the ring main
by the lime dosing pumps, one on-line and another on standby, to the milling,
CIL and cyanide detoxification circuits. Unused lime slurry in the ring main is
returned to the dosing tank.
The spillage in the lime make-up area
is contained in a bund and transferred by a spillage pump to the grit
classifier.
Owing to the presence of hot lime
slurry in the make-up and dosing areas, a safety shower is provided. It is
activated by a foot pedal and equipped with an eye bath.
17.4.5 |
Sodium Metabisulphite |
Presently the sodium metabisulphite
mixing and dosing system is out of order and the reagent lifted by tower crane
and is manually added to the detoxification tanks.
Under normal operating conditions
Sodium metabisulphite is delivered to site in 1 tonne bulk bags. The bulk bags
are transported using a fork lift from the storage area to the sodium
metabisulphite make-up area.
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The detoxification reagents hoist is
used to lift the bags to a bag breaker, from where they are discharged into a 19
m³ covered sodium metabisulphite mixing tank. The tank is equipped with a mixer
that ensures that the powder is dissolved completely during the make-up process
to form a 25 % sodium bisulphite solution by weight.
Sodium metabisulphite solution is
transferred from the make-up tank to a 21 m³ dosing tank using a transfer pump.
Duty and standby variable speed hose
dosing pumps are used to pump sodium metabisulphite solution to the carbon
safety and detoxification circuit at a controlled rate.
A mixing tank fan and a dosing tank fan
are used to extract any sulphur dioxide that is generated during the make-up of
sodium metabisulphite solution.
Presently the Copper Sulphate mixing
and dosing system is out of order and the reagent lifted by tower crane and is
manually added to the detoxification tanks.
Copper sulphate is delivered to site in
25 kg bags on 1 tonne pallets. The pallets are transported using a fork lift
from the storage area to the copper sulphate mixing area.
Once delivered to the makeup area, a
hoist is used to lift the pallet onto the platform where an operator loads the
bags onto a bag breaker, from where they are discharged into a 3 m3
covered copper sulphate mixing tank equipped with a mixer which ensures that the
crystals are dissolved completely during the make-up process to form a 15 %
copper sulphate solution by weight.
The 15 % copper sulphate solution is
transferred from a make-up tank to a 7 m³ dosing tank using the transfer pump.
Duty and standby variable speed hose
pumps transfer the copper sulphate solution to the carbon safety and
detoxification circuit at a controlled rate.
A safety shower is installed in the
sodium metabisulphite and copper sulphate make-up area. It is activated by a
foot pedal and equipped with an eye bath.
Spillage is contained in a bunded area
and spillage pumps transfer the spillage to the safety screen feed splitter in
the carbon safety and detoxification area.
Plant diesel is transferred from the
main diesel storage facility to a 7 m³ diesel storage tank from where it is
pumped by the diesel pump to a 2.2 m³ header tank. Diesel gravitates from the
header tank to the elution heater, regeneration kiln and smelting furnace.
Diesel for use in the laboratory is pumped from the main diesel storage facility
to a separate 2.2 m³ header tank.
Excess diesel from the elution heater
and regeneration kiln is return to the storage tank.
Activated carbon is delivered to site
in 500 kg bulk bags. The bulk bags are transported using a fork lift from the
storage area to the Adsorption circuit. When required, the carbon is added into
the last adsorption tank.
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Hydrochloric acid is delivered in 290
kg (210 litres) plastic drums at 33% strength by weight. The drums are
shrink-wrapped and palletized for safety reasons and easy storage. When
required, the palletized hydrochloric acid drums are transported using a
forklift from the storage area to the acid wash area. Acid is pumped from the
drum using a drum pump.
Compressed Air
Two air compressors, one working and
one on standby, deliver 180 Nm3/h of compressed air, at a pressure of
750 kPa, to the plant air receiver via one of the air filters. The compressed
air is distributed to the lime make-up area, workshop and various points in the
plant for general usage.
Instrument air is supplied by a
dedicated compressor. An instrument air drier and filters are provided in order
to ensure that instrument air is moisture-free and of good quality prior to
storage in an air receiver.
A provision is made to supply air from
the discharge of the plant air compressor to the instrument air filters and
dryer, if the instrument air compressor is off-line.
All three compressors are housed in a
compressor shed.
Blower Air
There are three aeration blowers, two
on-line and one common standby. The first blower provides oxidation air (at 250
kPa) to the first ten CIL tanks and the three cyanide destruction tanks. The
second blower provides air to the airlifts in the last six CIL tanks for carbon
transfer.
Leach aeration blowers, one on-line and
another on standby, deliver 1981 m3/h of blower air (at 350 kPa),
which is distributed to the leach and CIP and carbon safety and detoxification
circuits.
Water Abstraction and Storage
Dam
Raw water from a storage dam is drawn
by two centrifugal pumps (both on-line) and transferred to a raw water tank and
a process water tank in the plant. A third pump is provided as a standby pump.
Process Water
Return water from the tailings storage
facility and raw water top-up from the raw water storage tank constitutes
process water. This water is directed to a 1,000m³ process water tank.
Duty and standby low pressure high
volume process water pumps transfers process water to the scrubber, ball mills
and the gravity scalping screen.
High pressure spray water pumps, one
on-line and another on standby, distribute spray water to several screens
throughout the plant.
A high pressure hosing water pump is
used to supply hosing water to the plant.
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Raw Water
Raw water from a storage dam is routed
to a 1,000m³ raw water tank located close to the processing plant and alongside
the process water tank
Raw water pumps are used to distribute
water to the crushing and milling areas (for dust suppression), intensive
cyanidation, reagents make-up, potable water supply, carbon transfer, process
water top-up, gland water tank top-up, final tails sump flushing water and high
pressure wash pumps.
Gravity concentrator fluidising water
requirements are provided through the use of dedicated pumps.
Potable Water
A raw water break tank receives raw
water from the raw water pumps. The raw water undergoes purification in the
water treatment plant. Purified (potable) water is pumped to a 100 m³ potable
water storage tank. The backwash/rinse water from the water treatment plant is
directed to the storm water drain.
Two potable water distribution pumps,
one on-line and one on standby, supply water to the safety shower and potable
water headers via a hydrosphere, which is used to maintain pressure in the
potable water system.
Gland Water and Safety Shower Water
Distribution
Gland water pumps, one on-line and
another on standby, will transfer gland water from a 10 m³ tank to the glands of
all slurry pumps that require gland water.
Potable water from the hydrosphere is
distributed to various safety showers throughout the plant.
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18 |
PROJECT INFRASTRUCTURE |
Road access to the Twangiza Project had
been maintained along an existing 31 km section of access track linking the N2
national route from Bukavu, to the existing Twangiza exploration camp on the
Twangiza property.
This section of road which has
undergone extensive upgrade through widening of the road surface and adjustment
of the alignment to insure minimum turning radii and maximum gradients is
currently under constant maintenance to allow for the transport of all materials
and equipment required for the successful continuous operation of the mine. The
maintenance work of this portion includes the widening of some portions of the
road surface, gravelling, and repairs of bridges. Suitable layer works are being
done to ensure the strength of the road and to provide an all-weather surface.
Adequate drainage has been provided to ensure free flow of rain water. This
portion of road is made up of 3 bridges, namely bridges 4-6 which are currently
in good condition.
A further 5 km of road linking the
exploration camp and the process plant has been constructed to the same
specification as the access road.
To allow for the construction of the
tailings management facility, a road approximately 4.5 km long has also been
constructed linking the mine haul roads to the embankment location of this
facility. As this road is being used for the hauling of materials for embankment
construction by the mine ADT fleet, a road width of approximately 20 m has been
put in place.
18.1.2 |
Process Plant Buildings |
Buildings including the mine offices,
mine laundry and canteen, first-aid/consultation room, assay laboratory and
security gatehouses are in place within the process plant site. Some of these
buildings are of concrete and block work construction and local building
materials. Local Bukavu based building contractors were used to carry out these
works.
18.1.3 |
Process Plant Warehousing and
Workshop |
Warehouses and workshops are of steel
framed, steel-sheet cladding construction founding on concrete floor slabs and
bases.
The warehouses areas have been
calculated based on the consumption of the relevant process plant consumables,
as well as reasonable storage durations based on origin of supply and transport
risk. Warehouse offices are within the general warehouse building.
As per Item 18.1.6 a separate contained
area inside the general warehouse have been dedicated (for storage) of special
lubricants and greases, for the mining operation. In addition, a designated
caged area is in place for flammable products such as solvents and paints.
18.1.4 |
Process Plant Ancillary
Infrastructure |
Additional ancillary infrastructure
such as compressor house, a helipad and sewage and water treatment plants with
associated reticulation are in place for successful mining operations.
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Five camps are currently in operation
at Twangiza; these are in the following distribution:
Construction camp: |
130 rooms with 227 residents; |
|
|
Training Centre Accommodation: |
26 rooms with 28 residents; |
|
|
Dispatch Camp: |
12 rooms with 17 residents; |
|
|
Operators Camp: |
110 rooms with 110 residents; and |
|
|
Exploration Camp: |
81 rooms and 39 tents with 114 residents.
|
These housing units are in 3 varying
styles dependant on employment status, namely Senior Management, Supervisor and
Junior staff.
Three out of the five camps include
kitchen and dining facilities, a laundry, offices, an infirmary, a
security/gatehouse and recreational facilities.
Most of the buildings are made up of
imported prefabricated panel construction in order to minimise construction time
and ensure early availability of accommodation and office requirements for mine
construction. The 50mm thick panels are made of pre-painted chromadeck exterior
with polyurethane insulated infill. Some however, are made of locally
manufactured hydro-form bricks provided by local contractors to generate
employment for the local community.
Services including electric, water and
sewage reticulation as well as on-site water and sewage treatment plants have
been provided.
Various levels of security have been
established within the Process Plant area.
The entire mineral processing plant
from primary crushing through to the gold-smelting is completely enclosed to
create a high-security area. The 6m high fence comprises galvanised, anti-cut,
anti-climb panels including a barbed wire coil on the top. A no-go zone has been
created by the re-location of existing diamond mesh /barbed wire fencing to the
outside of the primary fence. Ultimately this fence will be replaced by a fence
of similar specification to the primary. The electrowinning and gold room area
are further enclosed within the high-security area. Storm water and Sewage
systems within this area exit the boundary fence via gold-trap structures.
A medium security area adjacent to the
process plant has been established containing the reagent storage and make-up
areas, the fuel farm and power plant. The external perimeter to this area has
been established utilising the diamond mesh/barbed wire fencing currently used
for construction purposes. Ultimately this fence will be replaced by a fence of
a higher specification. Within this medium security area a secondary fence is to
be established around the plant stores area.
The administration area comprising
offices, training centre and clinic is contained by a fence similar to that used
at the process plant and plant stores area, but of a lower specification.
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A plant wide CCTV system has been
established, together with a complimentary lighting system, in order to provide
for adequate perimeter monitoring. The security control room is equipped with
recording and storage equipment and remote monitoring of the CCTV system.
A single access point to the process
plant area is the primary access control system comprising vehicular boom gates,
manned 24 hours per day. Secondary access control is provided to the
administration area, the plant store and the high-security areas. Tertiary
access control has been established for the gold-room
The increase of throughput from 1.3
Mtpa to 1.7 Mtpa did not impact the security requirements of the process plant.
The power generating plant at the
Twangiza operation is rented from Aggreko, and Aggreko is also operating the
plant. A total of 11 containerised, 1250 KVA generator modules, are installed to
provide the power requirements for the increased throughput to 1.7 Mtpa. The
modules have been de-rated by 17% due to the altitude of the site.
The existing electrical infrastructure
is sufficient for the upgrade to 1.7Mtpa. This includes medium voltage
switchgear, transformers and cabling. The exception was the 400 volt motor
control centres which had to be extended. The size of the mobile switch rooms,
housing these motor control centres, was adequate to include all changes.
The power cost from Year 4 to the life
of the mine is expected to be reduced by 30% from USD 7.35/tonne processed to
USD 5.15/tonne processed, following the planned supply of hydroelectric power
from the Ulindi 1 river. The Hydro power plant is planned to be installed and
run by a third party to supply power to Twangiza Mining on a usage billing
agreement.
Studies by Knight Piesold show the
economic gain in switching from diesel-generated power to hydroelectric power.
The hydroelectric project will require an estimated capital of USD 40 million to
install a capacity of 6.5 MW plant.
External telephone communications are
currently provided by two cell phone operators and their services are
satisfactory.
Internet connectivity is provided by
two satellite systems, one at the plant site and the other at the exploration
camp. The plant, administration area and stores area are linked via a fibre
optic network. A microwave link between the two satellite systems at the
exploration camp and plant site provides a failover as a backup should one
system breakdown.
18.6 |
Sewage collection and
treatment |
Various ablution block/s have been
provided in the relevant areas within the process plant and associated
infrastructure, to which potable water from a water treatment plant is provided.
All sewage is piped to septic tank and French drain systems outside the
boundaries of the existing terraces. Gold-traps are installed on the lines where
sewage pipes cross the high security fencing of the process plant area.
The blend throughput rate of 1.7 Mtpa
is not expected to impact the sewage requirements of the process plant.
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18.7 |
Fuel and lubricant storage and
distribution |
Fuel is delivered to the mine from
Mombasa or Dar-es-Salaam by road tankers. The current fuel facility at Twangiza
consists of three 80,000 litre tanks adjacent to the bonded lay down area and a
further two 80,000 litre tanks south of the plant site in the proposed mine
workshop area.
The main fuel farm is operational with
two 1 million litre vertical tanks for storage purposes and two 80,000 litre day
tanks providing fuel to the plant and generators.
A complete fuel management system will
be installed at a later date.
18.8 |
Tailings Management
Facility |
As part of the economic assessment, an
evaluation of tailings storage capacity has been undertaken. A revised strategy
for tailings management has been adopted by Twangiza Mining, in that the
construction of the current Tailings Management Facility (TMF), namely TMF 1A,
will be halted at an elevation of 2060.9 m (AMSL) and a new facility, TMF1, will
be constructed in an adjacent valley.
The benefit will be an improved
tailings to fill ratio of 5.7:1 (LOM) compared to TMF 1A, which has a planned
final tailings to fill ratio of 1.135:1. The final volume of TMF1A will be 7.37
Mm3, or a storage capacity of some 8.18 Mt, which is based on an
average tailings dry density of 1.109 tonnes/m3, the density which
has been achieved to date.
Design values used by AMEC based on the
assumption that a maximum recorded 24 hour rainfall event of 120 mm at Bukavu is
equivalent to a 1 in 10 year event. Precipitation at the Twangiza site is
generally higher than at Bukavu so this value was increased to 155 mm. AMEC
produced rainfall estimates based on these assumptions.
However, SLR reviewed actual rainfall
data and revised the estimations upwards as part of their design report for
TMF1A. Both sets of estimates, for a 24 hour storm period, are summarised in
Table 18-1.
|
Table 18-1: |
Hour Design Storm
Event Estimates
|
|
Return Period |
Intensity (mm) |
|
1:10 |
155 |
|
1:100 |
180 |
AMEC |
1:200 |
185 |
|
1:1,000 |
210 |
|
1:10,000 |
240 |
|
1:50 |
235 |
SLR |
1:100 |
289 |
|
1:200 |
356 |
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18.8.3 |
Production rates and design
life |
Based on the current mine plan, the
required storage capacity for tailings is 23Mt oxide and non-oxides ore at a
rate of1.7 Mtpa over a 14 year mine life. Two TMFs are planned, the current
TMF1A and TMF1, which will be constructed in an adjacent valley.
TMF1A was originally designed to store
14.3Mtonnes (13.62 m3) at a nominal dry density of 1.05 tonnes per
cubic metre by Metago PTY of South Africa (now SLR Pty). The tailings dry
density has currently exceeded this target having achieved 1.109 tonnes per
cubic metre. However, due to unfavourable tailings to fill ratio of 1.135:1 and
the inability to store revised LOM reserves, a revision to the tailings
containment plan has been carried out by Twangiza Mining and a decision to
construct an additional tailings management facility, TMF1, has been taken. At
the same time it was decided to halt the construction of TMF1A in mid-2016 when
it will have reached a level of 2060.9 m and have a capacity of some 8.18 Mt
(7.37 Mm3) at 1.109 Mg/m3.
TMF1 will be constructed over the life
of the mine and will have a capacity of 22.80 Mt (20.56 Mm3) at a dam
height of 70 m. Notwithstanding the above, there is an option to raise the dam
height to 110 m at a later date should this be required to store potential
additional reserves. This would provide a tailings capacity of some 65 Mt for
TMF1.
18.8.4 |
Particle size distribution and specific
gravity |
The tailings particle size
distribution, adopted from the 2011, 1.3 Mtpa study, was summarised as follows:
|
|
100% passing 2mm (by mass). |
|
|
|
|
|
96% passing 0.85mm (by mass). |
|
|
|
|
|
69% passing 0.075mm (by mass). |
|
|
|
|
|
55% passing 0.020mm (by mass).
|
More recent gradings, carried out at
Twangiza in August 2014 are shown in Table 18-2.
|
Table 18-2: |
Particle Size
Distribution and Sieve Analysis
|
|
Sieve Size |
|
Passing% |
25µ |
38µ |
45µ |
53µ |
63µ |
75µ |
106µ |
150µ |
|
Finest |
73 |
75 |
78 |
82 |
84 |
88 |
93 |
97 |
Jun-14 |
Mean |
62 |
66 |
68 |
71 |
73 |
76 |
83 |
89 |
|
Coarsest |
43 |
48 |
51 |
54 |
58 |
60 |
75 |
84 |
|
Finest |
74 |
77 |
78 |
80 |
82 |
83 |
89 |
94 |
Aug-14 |
Mean |
64 |
67 |
69 |
72 |
74 |
77 |
84 |
89 |
|
Coarsest |
51 |
57 |
59 |
64 |
67 |
70 |
77 |
84 |
Furthermore, hydrometer testing carried
out in house recorded values of between 7% and 21% of material less than 2
µm.
Based on the above grading analyses and
previous laboratory test work, the material classifies as a CL material
according to the Unified Spoil Classification System i.e. a low plasticity clay
although it borders on being a ML (silt).
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18.8.5 |
Slurry characteristics |
The tailings slurry, adopted from the
1.3 Mtpa study, was taken as 37% solids by mass (provided by SENET) which at a
particle specific gravity of 2.84 t/m3 yields a slurry bulk density
of 1.32 t/m3.
Initial work carried out by Patterson
and Cooke found the specific gravity of samples to be as follows:
|
|
Main Pit: Main Oxide: 2.80 t/m3,
Main Transition: 3.04 t/m3, Main Fresh: 3.33 t/m3.
|
|
|
|
|
|
North Pit: North Oxide: 2.74 t/m3,
North Transition: 2.79 t/m3, North Fresh: 2.99 t/m3.
|
Additional work soil testing work
carried out for other purposes has indicated specific gravities in the range
2.82 to 2.85 t/m3. Based on these an average specific gravity of 2.83
t/m3 has been adopted for calculation purposes. Considering the 37%
solids suggested in the 1.3 Mtpa study, this yields an average slurry density of
1.315 t/m3.
However, tailings solids are currently
being discharged at between 30% and 33% by mass, which at a particle specific
gravity of 2.83 t/m3 yields a slurry bulk density in the range 1.24
to 1.27 t/m3. With more transition material being processed, average
slurry bulk density is likely to increase marginally to between 1.25 and 1.28
t/m3.
Based on existing data from TMF1A and
considering the high rates of rise, a minimum average dry density of 1.09
t/m3 is expected in TMF1, for a pool size of 30%. A lower starting
value of 0.75 t/m3, rising to 1.109 t/m3 over the first
two years of operation, has been used for the rate-of-rise assessment.
Table 18-3, based on 1.7 Mtpa, presents
the expected geotechnical parameters that the tailings will exhibit through its
life.
|
Table 18-3: |
Tailings Parameters
|
Source |
Unit |
Value |
Tailings Slurry |
Solids by mass |
% bulk density (t/m3 ) |
30-33 |
Particle Specific Gravity |
t/m3 |
2.83 |
Slurry Bulk Density |
t/m3 |
1.24-1.27 |
Dry density |
t/m3 |
0.486 |
Moisture Content |
% |
203-233 |
Slurry Void Ratio |
|
5.75-6.6 |
Deposited Tailings |
Average Bulk Density (Saturated) |
t/m3 |
1.485 |
Average Dry Density |
t/m3 |
0.75 |
Moisture Content |
% |
98.12 |
In situ void ratio |
|
2.77 |
Deposited Freshly Consolidated
Tailings |
Average Bulk Density (Saturated) |
t/m3 |
1.717 |
Average Dry Density |
t/m3 |
1.109 |
Moisture Content |
% |
54.96 |
In situ void ratio |
|
1.57 |
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18.8.6 |
Survey information |
Survey information used in the study
has been provided by Twangiza Minings in-house survey department as point data
and as an electronic contour map. Aerial photography of the area concerned has
also been provided. The survey is based on a combination of the aerial survey
completed in 2008 and a more recent ground survey.
18.8.7 |
Climate data summary |
Details on the climate are provided in
the report Surface Water Investigations for the Twangiza Gold Project by
Metago dated April 2010. In summary the area has relatively high rainfall of
about 1.35 m/year, evaporation of 1.2m/year, with the rainfall season lasting
from mid-October to April, while the 4.5 month dry spell is from May to
mid-September.
The design storm events calculated by
SLR are as follows:
|
|
1:50 year 24 hour storm event is 235mm. |
|
|
|
|
|
1:100 year 24 hour storm event is 289mm. |
|
|
|
|
|
1:200 year 24 hour storm event is 356mm.
|
18.8.8 |
Rate of rise of tailings |
The study requires the development of
stage capacity relationships for the tailings dam in order to define the
relationships between volume, area, height, rate of rise and production rate of
the facility at any point in time during the design life. The stage capacity
relationships are used to monitor the rate of rise of the tailings dams and to
determine the required height of the start embankment of the tailings
facility.
The stage capacity curves shown in
Figure 18-1 indicates that a starter wall (TMF1) with a crest elevation of 1977
m AMSL is required for a 22.5 month capacity or 3.14Mt. TMF1A, completed to
Stage 5, will contain 8.18Mt.
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18.8.9 |
TMF construction works |
A schedule of quantities has been
drafted for the construction works for both TMF1A and TMF1, as the stream
diversions required remains the same as for the 1.3 Mtpa study. Construction
rates have been based on rates calculated in-house. However, these rates are
generally in accordance with estimates provided by external contractors. The
assumption has been made that an average of 10 m of material must be removed
under the foundation of the wall which must be replaced with suitable compacted
fill. Additional assumptions include:
|
|
Crest width of 40m. Appendix XIV of the DRC Mining
Regulations requires a crest width of greater than H/5 + 3m. This
translates into a crest width of approximately 11.0 m for the 2 year
starter wall, increasing to 17.0m for the LOM. However, this would prove
problematic to construct without specialist plant, therefore a 30 m wide
pad plus sand drain and impermeable clay curtain has been added resulting
in a provisional width of 40 m. This has significant impact on the wall
volumes and a review of construction techniques is required as significant
savings are possible. |
|
|
|
|
|
Wall fill will predominantly be won from mine waste mixed
with borrow material from within a 1.5 km radius to ensure fill moisture
content is suitable for compaction. |
The TMF costing excludes:
|
|
Any lining of the TMF basin or any other ARD mitigation
measure. Water quality is monitored on a daily basis and TMF1A is not seen
to be impacting local ground water. The presence of a significant
artesian/sub-artesian head underlying the TMF basin provides a degree of
protection towards groundwater quality. |
|
|
|
|
|
Detoxification plant; To date, on TMF1A, detoxification
carried out ex-process has proved adequate to control TMF water quality.
|
|
|
|
|
|
The tailings delivery pipeline from the plant to the
basin of the TMF1. |
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|
The tailings delivery and return water pumping system
(floating buildings, pumps, motors, and all associated mechanicals and
return water pipeline. However, it is anticipated that the infrastructure
already in place for TMF1A will remain in operation and return water will
only require pumping from TMF 1 into TMF1A, minimising additional works.
|
|
|
|
|
|
A grout curtain being required under the base of the
wall. A grout curtain was not necessary for TMF1A and it is considered
unlikely to be required for TMF1. |
|
|
|
|
|
Access Roads. Almost all necessary access roads are
already in place. Maintenance is expected to be the primary cost.
|
The main conclusions in the TMF study
are summarised as follows:
|
|
The Construction of TMF1A will cease at a height of
2060.9 mASL resulting in a volumetric capacity of 7.37 Mm3 or 8.18Mt. This
will provide tailings storage until the end of October 2017. |
|
|
|
|
|
A starter wall for the new tailings management facility,
TMF1, with a crest elevation of 1977 m AMSL (40 m crest height) is
required to store 22.5 months of tailings, 3.14 Mt. |
|
|
|
|
|
A production rate of 1.7 Mtpa is anticipated, giving a
further 13.4 year LOM life for TMF1 following the filling of TMF1A
(predicted October 2017), brings the total tailings capacity to 15.4
years. |
|
|
|
|
|
TMF1 wall will be raised during the dry seasons of 2018
and 2019 to a height of 1987 mAMSL (7.9 Mt capacity) with a further rise
to 1995 mAMSL during the dry seasons of 2020 and 2021, which will result
in storage for 13.0 Mt. The final rise, carried out during the dry seasons
of 2023 and 2024 to an elevation of 2005 mAMSL providing a final volume
for tailings in TMF1 of 20.56 Mm3 or 22.80 Mt. |
|
|
|
|
|
The capital cost estimate for the TMF1 starter wall is
approximately USD 19.7M. |
|
|
|
|
|
The total capital cost required for the construction of
the remaining TMF1A and TMF1 is approximately USD 76.6M.
|
It must be stressed that a starter wall
for a start-up basin capacity/life of 22.5 months as opposed to the recommended
minimum of 2 years increases the risk of potentially not being able to deposit
tailings at the end of the 22.5 months if the wall raise is not in place at that
time. Typical factors that need to be considered in assessing this increased
risk include:
|
|
The vast majority of each wall raise can only be
undertaken the 5 driest months of the year, although during the earlier
part of the wet season some daytime working is generally possible.
Nevertheless, two dry seasons will be available for construction during
the 22.5 month period for the first wall raise. |
|
|
|
|
|
Sufficient plant and equipment must be provided based on
the application of an efficiency factor due to maintenance, break downs,
rain related delays, and where the actual borrow areas will be located
along with their haul road development etc. |
|
|
|
|
|
The footprint or size of each wall raise is limited even
though the volume placed is large, i.e. there is limited working area or
space to move plant (place, spread, compact, test etc.) and this needs to
be evaluated in detail. Generally, a 30 m wide pad width is necessary to
maximise efficiency unless specialised plant is used.
|
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19 |
MARKET STUDIES AND
CONTRACTS |
The gold price has shown a downward
trend since the Twangiza Mine started operations as can be seen in Figure 19-1,
which illustrates the gold prices performance over a fourteen year period.
SRK subscribes to a market consensus
forecast which is updated quarterly; the 5 year (long term) gold price forecast
has changed from 1,200 USD/oz to 1,150 USD/oz in the six months preceding the
date of this report.
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20 |
ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR
COMMUNITY IMPACT |
20.1 |
Environmental and Social Impact Assessment
status |
SRK (SA) conducted an Environmental and
Social Impact Assessment as part of the 2009 Feasibility Study. Neither SRK (SA)
nor SRK (UK) have done any further Environmental Investigations or additional
Social Impact Assessments since the report was filed.
The ESIA was written to meet potential
lenders Equator Principles standards, which cover DRC requirements. Since some
specific elements were to be formulated slightly differently in order to meet
DRC requirements for reporting, an update ESIA was compiled in French and
submitted to the DRC Ministry of Mines.
Twangiza Mining has in the interim
undertaken a relocation program as part of the mine construction and has also
considered an alternative site for the Tailings Storage Facility.
In November of 2009, Twangiza Mining
appointed Metago Environmental Engineers (Pty) Ltd (now part of the SLR Group),
an independent environmental consultancy based in South Africa, to compile an
interim ESIA report focussing specifically on the 1.3 Mt per annum Phase I of
the Twangiza Gold Project based on the information and specialist studies
available at that time,
Metago identified scope of work in
their March 2010 interim ESIA report which were completed by Twangiza Mining,
enabling SLR Consulting Ltd to prepare an updated ESIA report for Phase I of the
Twangiza Gold Project which put forward a conceptual environmental and social
management plan. The updated ESIA report drew significantly upon the content of
the following documents:
|
|
the interim environmental impact assessment (EIA) and
associated specialist studies and stakeholder involvement process
conducted by SRK Consulting SA (Pty) Ltd (SRK) for the full extent of the
5.0 million tonne per annum operation (SRK Consulting 2009); |
|
|
|
|
|
additional specialist investigations as amended in 2010
to suit the Phase I scope, and |
|
|
|
|
|
the social impact assessment and resettlement action plan
reports compiled by Hilde Van Vlaenderen for the Twangiza project (H. Van
Vlaenderen 2010a and 2010b, respectively). |
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Table 20-1: |
Summary of Key
Legislation and Relevant Compliance
|
LEGISLATION |
COMPLIANCE |
EXPLOITATION (MINING) PERMIT |
In terms of Title V of Decree no 038/2003, the |
|
company must apply for an Exploitation (mining) |
|
Permit. In order to apply for such a permit the |
Twangiza Mining holds exploitation permits |
company must be the title holder of a valid |
|
Exploration Permit(s) |
|
ENVIRONMENTAL IMPACT STUDY (EIS) AND
ENVIRONMENTAL MANAGEMENT PLAN (EMP) |
|
An Environmental and Social Impact |
|
Assessment (ESIA)* and Environmental and |
The environmental obligations are set out in Title |
Social Management Plan (ESMP*) have been |
XVIII of decree no 038/2003. With the exception of |
completed and an updated French version |
temporary quarrying, any mining operation requires |
submitted to DPEM (Ministry of Mines) in |
an approved Environmental Impact Study (EIS*) |
2013. |
and an Environmental Management Plan for the |
|
Project (EMPP*). |
The following management plans have been |
Schedule IX (Contents of EIS and EMPP) sets out |
prepared as part of the Environmental and |
the contents of the EIS and the EMPP and provides |
Social Management plan (ESMP equivalent |
detail regarding specific management measures |
to EMPP): |
and standards that are required. |
Resettlement Action Plan |
|
Stakeholder Engagement Plan |
* EIS = ESIA; EMPP = ESMP, ESIA and ESMP |
Community Development Plan |
being IFC/WBG/international terminology |
Conceptual Rehabilitation and Closure Plan |
|
|
20.2 |
Key risks to the
project |
20.2.1 |
Social conflict arising from
resettlement |
Risks associated with social conflict
due to involuntary resettlement have been managed to date through implementation
of a resettlement action plan (RAP) based on data from the social survey and
formal and effective consultation with the community. Based on the positive
results of the implemented program, Twangiza intends to continue carrying out
similar programs to address required resettlement based on the footprint of the
operations. These required activities will continue to be carried out through
programs that rely on ongoing community forums and constant involvement of local
leadership and state officials. In addition, employment opportunities and
sustainable development efforts by Twangiza are carried out in a manner to
target and prioritize affected communities where possible.
20.2.2 |
Twangiza North
Resettlement |
A survey of households carried out in
early 2015 has indicated that the area identified to be the affected area from
the development of the North Pit has approximately 763 households. This reflects
an influx to prior evaluations as a result of the operations of Twangiza and the
opportunity for economic benefit.
A significant proportion of the 1,276
artisanal miners who ceased mining on the Twangiza footprint in mid-2010, are
engaged in artisanal mining activities at the Twangiza North Pit, and around the
mine periphery on the alluvial deposits in the river systems. These original
miners have been joined by an influx of transient artisanal miners and current
estimates are that there are between 2,000 and 2,500 artisanal miners who will
require to be relocated from the North Pit by Quarter 2 of 2016.
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Twangiza Mining has identified some
minor deposits on its concessions which could be made available to artisanal
miners and has established with the national and provincial mining authorities
that it may be possible to take legal steps to make the deposits concerned
accessible to registered artisanal mining cooperatives.
Based on the mine profile, there is a
need to resettle approximately 763 households on a phased programme between
Quarter 4 of 2015 and Quarter 1 of 2017. A new Luhwindja Community Forum,
similar to previous resettlement activities, has been established to facilitate
the ongoing consultative engagements with the community. The community social
surveys have been updated for 142 households closest to Twangiza Main Pit, and
it is planned to update the social data for some 621 households in the areas
close to Twangiza North Pit by the end of Quarter 3 2015.
The households will not be resettled at
the Cinjira Resettlement Village, but at alternative land blocks identified
within Luhwindja. The options have been narrowed down to 11 sites. The costs
associated with acquiring and developing the land are not yet established.
20.2.3 |
New TMF Resettlement |
An additional resettlement project will
be required at Twangiza in order to construct a new tailings dam, because of the
need to replace the current TMF1A. Resettlement activities for the new TMF1 will
be carried out in a manner consistent with past and present resettlements
including a focus on community forums and working closely with local and
regional leadership. The new TMF1 site community has not yet been surveyed.
20.2.4 |
General Resettlement
Considerations |
The total resettlement requirement for
the project, including the new TMF, is estimated by Twangiza Mining to cover 830
households with appropriate allowance made in capital cost estimates. The cost
estimate may increase to overcome additional challenges which may arise when
executing the resettlement process with the various stakeholders involved,
including those who have not been a party to historical resettlement activities.
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20.3 |
Local political
instability |
There is a longstanding local political
conflict in the Twangiza area relating to the traditional chieftainship
succession, which has affected Twangiza activities from inception in the
exploration phase and has periodically continued to negatively affect community
stability.
There are potential liabilities
associated with the residual impacts of the extensive artisanal mining that has
taken place in the project area. Metal and metalloid contamination (arsenic,
chromium, lead, iron, manganese) is severe at many of the sampling points
downstream of artisanal mining areas, although no mercury has been detected.
Mercury may be adsorbed onto riverbed sediments, but these have not been tested.
20.5 |
Environmental monitoring |
Routine monitoring is being conducted
in the general Twangiza catchment area to assess pollution from both artisanal
and commercial mining activities, in addition to any other source of pollution
that may arise in the region. Surface and ground water quality are being
assessed on a monthly basis for the full spectrum of analysis including the
ICP-MS Scan and a few parameters like pH, conductivity, free CN are being
assessed on a daily basis at some key points.
See Figure 20-1 for the Twangiza Water
Monitoring Point.
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21 |
CAPITAL AND OPERATING
COST |
21.1 |
Operating Cost
Estimate |
Costs have been estimated based on a
zero-based (first principles) cost analysis following review of 2014
historical costs, with appropriate allowance for the future operating
environment, productivity and cost saving initiatives.
A number of changes are reflected in
the forecast costs, notably there are savings expected from:
|
|
lower oil price |
|
|
|
|
|
transition to hydro power source |
|
|
|
|
|
restructured staffing, |
|
|
|
|
|
renegotiated contracts, and |
|
|
|
|
|
more direct supply chain. |
The unit savings per processed tonne
are enhanced by the higher plant production rate, however the unit savings per
ounce produced is tempered by the lower gold recoveries in later years.
A summary of the unit operating costs
actually incurred in 2014 and forecast operating costs in the mine plan are
given in Table 21-1.
During 2014, there was a significant
price decline in the global oil market which subsequently reduced the cost of
diesel at Twangiza based on the nature of supply contracts. During 2014 the
average diesel price delivered to site was USD 1.52/litre which has been
renegotiated by Twangiza Mining to USD 0.84/litre for 2015. In addition, mining
diesel costs are expected to benefit from new mining fleet additions with
improved fuel efficiency.
From 2018 onwards, power is planned to
be supplied to Twangiza from a planned hydroelectric plant at Ulindi 1 River.
This is projected to reduce the total power cost from USD 7.35/ tonne processed
to USD 5.15/tonne processed; or 30% lower.
A preliminary engineering report and
Solicitation of Interest (SOI) document has been finalised and will be
distributed to potential financing and developing partners soon. The power
supply agreement will be a usage billing agreement between Twangiza Mining and
the successful power supply company; the initial capital requirement is planned
to be repaid over 10 years which forms part of the power cost of USD 5.15/tonne
processed.
21.1.3 |
Sundry and Expenses |
During 2014, Twangiza initiated a
number of in-house modification and improvement projects associated with ongoing
process optimization as well as the plant expansion which started in 2013. As a
result of these activities, additional resources were required by Twangiza
Mining which resulted in increased employee related costs, such as legal and
compliance, to be incurred during the period.
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During 2014, the Namoya Mining SA gold
project, another Banro subsidiary, progressed forward with gold production
levels which allowed for the renegotiation of bullion transportation agreements
to allow for certain costs to be shared due to the relative proximity of the two
operations.
In addition to the above, certain
non-recurring costs were incurred by Twangiza Mining in 2014 including fees
associated with building certifications at the Cinjira resettlement and
consumable storage costs that resulted from a previously employed procurement
methodology.
During the early stages of operations,
as is typical with new operations in regions without a history of industrial
mining, the Twangiza Mining management structure included a large number of
expatriates. During 2014, an organizational delayering process was executed
which ultimately led to a reduction of expatriate personnel along with the
delayering of the organogram to create a flatter structure which has been
operational in 2015. In addition to this delayering process, a review of all
contractor labour was undertaken to eliminate non-essential staff and costs that
existed from the capital projects and a reduction in contract labour to reflect
the standards of an operating mine.
21.1.5 |
Processing Materials and
Contracts |
Previously, most chemicals were
purchased from intermediate handlers who had mark-ups on product costs and fees.
New arrangements are now in place to purchase chemicals directly from the
manufacturers. This process was assisted by the increased scale of operations.
Furthermore, management decided to contract a transporter for in-country haulage
which offers cheaper transport costs and no management and personnel fees. In
addition to the review on the in-country haulage, all the external haulage which
was originally coming through Mombasa port was rerouted through Dar es Salaam
port and this reduced the haulage cost significantly.
Even though this process started in
late 2013, there were stocks of consumables which were bought at the higher
pricing. Those stocks were used during the 2014 production year and from 2015
onwards the benefits of the newly negotiated contracts are being seen.
21.1.6 |
Mining Materials and
Contracts |
Due to limited funds, the stock of
spares to maintain the mining fleet was limited which necessitated the rental of
certain equipment from contractors. In addition to the higher cost of contractor
equipment, the cost of spares to fix the owner operated fleet increased due to
working capital limitations.
Furthermore, Twangiza had a haulage
contract involving six Bell B40 ADT trucks, the terms of which have been
renegotiated with better pricing.
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Table 21-1: |
Summary of Unit
Operating Costs
|
|
ActualH22014
(Annualised) |
Forecast in Mine Plan |
Item |
|
|
USD |
|
USD/ |
USD |
|
USD/ |
|
million/ |
USD/t |
oz |
million/ |
USD/t |
oz |
|
annum |
processed |
poured |
annum |
processed |
poured |
Mining |
|
|
|
|
|
|
Payroll |
5.40 |
3.53 |
48 |
3.99 |
2.50 |
45 |
Materials |
3.71 |
2.43 |
33 |
2.60 |
1.62 |
29 |
Contractors&Consultants |
3.13 |
2.05 |
28 |
0.90 |
0.56 |
10 |
Diesel&Power |
3.42 |
2.23 |
30 |
1.89 |
1.18 |
21 |
SundryExpenses |
0.39 |
0.25 |
3 |
0.15 |
0.09 |
2 |
|
|
|
|
|
|
|
MiningTotal |
16.06 |
10.49 |
142 |
9.53 |
5.95 |
107 |
Processing |
|
|
|
|
|
|
Payroll |
6.15 |
4.02 |
54 |
5.08 |
3.17 |
57 |
Materials |
9.11 |
5.95 |
80 |
11.76 |
7.35 |
132 |
Contractors&Consultants |
1.96 |
1.28 |
17 |
1.28 |
0.80 |
14 |
Diesel&Power |
15.16 |
9.91 |
134 |
11.80 |
7.37 |
133 |
SundryExpenses |
3.61 |
2.36 |
32 |
2.17 |
1.35 |
24 |
|
|
|
|
|
|
|
ProcessingTotal |
36.00 |
23.52 |
318 |
32.09 |
20.04 |
360 |
G&A |
|
|
|
|
|
|
Payroll |
5.39 |
3.52 |
48 |
5.68 |
3.55 |
64 |
Materials |
2.07 |
1.35 |
18 |
0.73 |
0.46 |
8 |
Contractors&Consultants |
6.78 |
4.43 |
60 |
6.40 |
4.00 |
72 |
Diesel&Power |
0.74 |
0.49 |
7 |
0.69 |
0.43 |
8 |
SundryExpenses |
6.63 |
4.33 |
59 |
7.12 |
4.45 |
80 |
|
|
|
|
|
|
|
G&ATotal |
21.61 |
14.12 |
191 |
20.62 |
12.89 |
232 |
|
|
|
|
|
|
|
TotalOperatingCosts
|
73.67 |
48.12 |
651 |
62.24 |
38.88 |
699 |
21.1.7 |
Basis of Estimates |
Mining general
The mining operating costs were based
on an owner operated system as opposed a contractor operated system. Unit costs
were obtained from an in-house zero based budget prepared for the production
year 2015. The estimation of mining operating costs by period was based on
projected equipment performance and unit operating costs.
Plant operating labour
Derived from first principles using
assumptions and costs estimated in the in-house zero based budgets prepared for
the production year 2015.
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Plant maintenance labour
This was derived from first principles
using assumptions and costs estimated in the in-house zero based budgets
prepared for the production year 2015.
Plant consumable materials
This was derived through a combination
of projected reagent consumption and the in-house zero based budgets prepared
for the production year 2015.
Power
The average continuous monthly power
consumption for process plant was determined by taking into account the power
rating for each piece of major equipment and the projected running times as
outlined in the design criteria. The power costs were then determined by taking
into account the operating costs and the calculated monthly electrical
consumption in kWh.
Plant maintenance costs and supplies
This cost has been obtained from the
in-house zero based budgets prepared for the production year 2015.
General and Administration costs
(including assay costs)
This caters for administration labour;
which has been derived from first principles and a range of other costs
associated with administration such as camp costs, office supplies, telephones,
computers, safety supplies, clinic supplies, vehicles, insurance and head office
expenses as captured in the in-house zero based budget prepared for the
production year 2015.
Administration tax
An administration tax of 5% to cover
the cost of importation of plant, machinery and consumables has been included in
the projected capital and operating costs.
21.2 |
Capital Cost Estimate |
As part of the economic assessment,
capital costs were estimated for the upgrade of the Twangiza process plant to
facilitate processing of a blend of oxides and non-oxide material types to meet
the projected plant throughput rate of 1.7 Mtpa. Capital costs were also
estimated to purchase additional/replacement mining fleet to meet the material
movement requirements over the Life of Mine. Due to the increase in the ore to
be processed due to the addition of the non-oxides, the tailings capacity needs
to be increased to accommodate the increase in projected tailings that will be
generated over the life of mine. Additional capital provisions have been made to
cover the cost of constructing an alternative TMF, which is expected to be ready
by mid-2017. This approach is designed to significantly reduce the cost of the
alternative option to raise the wall of the existing tailings dam (TMF1A).
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|
Table 21-2: |
Twangiza Capital Cost
Summary
|
|
Cost |
|
Item |
(USD |
Comment on Estimate Accuracy |
|
million) |
|
Capitalised Expenditure |
|
|
|
|
(+/5%accuracy) Based on quotations received and
|
Mining Sustaining Capital |
26 |
used in2 015 zero based budget |
Processing Sustaining |
|
(+/5%accuracy) Based on quotations received and
|
Capital |
6 |
used in 2015 zero based budget |
|
|
(+/10%accuracy) Based on Design BOQs and costs
|
TMF Construction Capital |
46 |
inputs from 2015 zero based budget |
|
|
(+/10%accuracy) Based on Design BOQs and costs
|
Tailings Sustaining Capital |
30 |
inputs from 2015 zero based budget |
|
|
(+/5%accuracy) Based onq uotations and |
General & Administration - |
21 |
reviewed compensation rates and used in 2015
|
Sustaining Capital |
|
zero based budget |
|
|
|
Banro Foundation |
1 |
Statutory rate of US$1.00/ounce |
Total Capitalised Expenditure
|
130 |
|
21.3.1 |
Basis of Estimates |
The accuracy of the capital estimate is
considered by Twangiza Mining to be within ±10%. Most items are based on
historic data which have been generated since the beginning of operations, some
are based on quotations received from vendors. The TMF is based on Twangiza
Minings estimate which is factored from previous dedicated technical studies.
Mining Sustaining Capital
The additional initial and replacement
mining fleet requirements have been estimated based a multi-pit operation
approach sufficient to achieve peak total material movements requirements
sustain a suitable ore blend to feed the plant at a 1.7 Mtpa throughput target.
New mining fleet capital amounts to USD19 million, fleet replacement capital is
estimated at USD5 million with a further USD1.5 million allocated for grade
control, heavy duty workshop construction and tooling.
Process Sustaining Capital
Due to the increased hardness of the
plant feed, measures have been implemented to replace all liners and wear plates
to economically and safely handle the projected feed rate. A full time primary
crushing unit has been designed to be installed to crush and screen RoM to the
required -10mm before it is fed through a mineral sizer and auxiliary feed
point.
TMF Construction Capital
Total capital of USD 76 million will be
required to construct TMF facilities, split in two portions. The first portion
of USD 46 million will be required for TMF construction capital and will be used
to procure material, geotechnical studies, land compensation and relocation and
starter wall construction. The second portion totalling USD 30.5 million will be
expended in Years 2015 2025 to cover the cost of continuous rising of the
wall.
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General and Administration
Sustaining Capital
Provision has been made on the capital
estimates for North Pit artisanal miners and Kantambwe compensation and
relocation.
No escalation has been allowed for in
the estimate. Rates are based on historical cost from the Twangiza operations
and the zero-based budget for the 2015 production year.
In early 1997, Banro, SOMINKI and the
government of the DRC ratified a new 30 year mining convention that provided for
SOMINKI to transfer its gold assets to a newly created company, Société Aurifère
du Kivu et du Maniema, SARL (SAKIMA). In addition to this asset transfer, the
new mining convention included a ten year tax moratorium from the start of
commercial production, the ability to export all gold production, the ability to
operate in US currency, the elimination of import duties and title confirmation
for all of the concessions.
After the 10 year tax holiday, as
described above, the project is subject to corporate taxation at the rate of 30%
on profit.
Royalty payments, as agreed with the
Government are paid at the rate of 1% of gross revenue.
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The base case was developed using a
long-term gold price of USD1,200 per ounce and a 5% discount rate. The financial
model also reflects the favourable fiscal aspects of the mining convention
governing the Twangiza project, which includes 100% equity interest and a 10
year tax holiday from the start of production in September 2012. An
administrative tax of 5% for the importation of plant, machinery and consumables
has been included in the projected capital and operating costs. Calculated
sensitivities show the significant upside leverage to gold prices and the robust
nature of the projected economics to operating assumptions.
|
Table 22-1: |
Financial Analysis
Summary
|
ITEM |
UNIT |
AMOUNT |
LIFE OF MINE GOLD PRODUCTION |
koz |
1,246 |
PRODUCTION PERIOD |
years |
14 |
ANNUAL GOLD PRODUCTION FOR FIRST 5 YEARS |
koz |
109 |
TOTAL CAPITAL COSTS |
USD/oz |
104 |
ALL IN COSTS |
USD/oz |
888 |
POST-TAX NET PRESENT VALUE |
USD million |
285 |
NET CASHFLOW AFTER TAX AND CAPEX |
USD million |
395 |
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The assumptions used in the financial
analysis are given in the table below.
|
Table 22-2: |
Financial Model
Assumptions
|
Item |
Unit |
Value |
Revenue |
|
|
Plant Throughput |
000tpa |
1,700 |
Gold Price (Lower Limit) |
USD/oz |
1,000 |
Gold Price (Base Case) |
USD/oz |
1,200 |
Gold Price (Upper Limit) |
USD/oz |
1,600 |
Discount Rate |
% |
5% |
|
|
|
Fuel Price |
|
|
Diesel |
USD/litre |
0.84 |
|
|
|
Fiscal |
|
|
Tax Free Holiday |
years |
10 |
Tax Rate (Year 1 10) |
% |
|
Tax Year (Beyond Year 10) |
% |
|
Government Royalty |
% |
1.00 |
Depreciation |
% |
|
|
|
|
Conversion Factors |
|
|
Kilograms To Ounces |
kg/ troy ounce |
32.1505 |
Diesel Fuel Density |
t/m3 |
0.85 |
Exchange Rate |
ZAR : USD |
7.496 |
|
|
|
Other |
|
|
Refining Charges, Dore Transport and Insurance |
USD/oz |
15.20 |
Percent Of Capital Expenditure (Year 2015) |
% |
35% |
22.3 |
Sensitivity Analysis |
A sensitivity analysis was performed on
the after tax profits by varying the gold price between USD1,000 and USD1,600
per ounce. The results are summarized below.
|
Table 22-3: |
Sensitivity Analysis on
Gold Price
|
GOLD PRICE |
NET PRESENT VALUE (USD
million) |
(USD/oz) |
1600 |
1500 |
1400 |
1300 |
1200 |
1100 |
1000 |
NPV 0.0% |
847 |
734 |
621 |
508 |
395 |
282 |
169 |
NPV 5.0% |
612 |
530 |
449 |
367 |
285 |
203 |
122 |
NPV 8.0% |
516 |
447 |
378 |
310 |
241 |
172 |
103 |
NPV 9.5% |
476 |
413 |
350 |
286 |
223 |
159 |
96 |
NPV 10.0% |
464 |
402 |
341 |
279 |
217 |
156 |
94 |
NPV 12.5% |
411 |
356 |
302 |
247 |
193 |
139 |
84 |
NPV 15.0% |
366 |
318 |
270 |
222 |
173 |
125 |
77 |
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A sensitivity analysis of NPV (5%
discount rate) to total operating and capital costs +30% has also been
prepared as provided in Figure 22-2. Included in the sensitivity analysis are
comparative estimates for project NPV calculated using historical costs based on
full calendar year 2014 ($62M) and H2 2014 ($177M). It is evident that the
changes implemented during 2014 generally indicate an improving trend for unit
cost reduction and productivity improvement.
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22.3.1 |
Project life cash-flow |
Table 22-4: |
Cash Flow Summary
|
Revenues |
|
Year_2015 |
Year_2016 |
Year_2017 |
Year_2018 |
Year_2019 |
Year_2020 |
Year_2021 |
Year_2022 |
Year_2023 |
Year_2024 |
Year_2025 |
Year_2026 |
Year_2027 |
Year_2028 |
Total |
Gold Produced |
oz |
115,897 |
105,752 |
117,121 |
108,591 |
96,305 |
111,790 |
102,160 |
94,396 |
84,289 |
65,832 |
73,675 |
78,647 |
79,463 |
12,149 |
1,246,311 |
Gold Price |
USD/oz |
1,250 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,200 |
1,202 |
Revenue |
USD 000 |
144,871 |
126,903 |
140,545 |
130,309 |
115,566 |
134,148 |
122,592 |
113,275 |
101,147 |
78,999 |
88,409 |
94,376 |
95,355 |
14,578 |
1,501,368 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Operating Costs |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Mining cost |
USD 000 |
9,436 |
21,497 |
37,793 |
36,215 |
40,884 |
36,311 |
26,392 |
28,345 |
13,319 |
15,620 |
18,354 |
3,945 |
0 |
0 |
288,108 |
Grade Control cost |
USD 000 |
1,091 |
2,352 |
4,341 |
3,871 |
4,430 |
4,171 |
3,032 |
3,256 |
1,530 |
1,794 |
2,108 |
453 |
0 |
0 |
32,429 |
Process Plant Costs - |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Carbon In Leach |
USD 000 |
12,364 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
13,498 |
3,196 |
177,569 |
Assay |
USD 000 |
1,142 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
1,246 |
295 |
16,387 |
Power |
USD 000 |
11,780 |
12,499 |
12,499 |
8,749 |
8,749 |
8,749 |
8,749 |
8,749 |
8,749 |
8,749 |
8,749 |
8,749 |
8,749 |
2,072 |
126,366 |
Engineering (Maintenance) Costs |
USD 000 |
5,128 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
5,601 |
1,326 |
73,676 |
Rehab Provision (Pits & Dumps, TMF, Ponds, Demobilization)
|
USD 000 |
64 |
1,086 |
1,070 |
1,202 |
320 |
1,369 |
1,292 |
981 |
1,039 |
582 |
652 |
729 |
433 |
346 |
11,249 |
Infrastructure, Overheads and Sundries (G&A) |
USD 000 |
19,949 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
12,478 |
2,955 |
172,665 |
Refinery and Shipment cost |
USD 000 |
1,854 |
1,608 |
1,781 |
1,651 |
1,464 |
1,700 |
1,553 |
1,435 |
1,282 |
1,001 |
1,120 |
1,196 |
1,208 |
185 |
19,041 |
Royalty & Government Charges |
USD 000 |
1,449 |
1,322 |
1,464 |
1,357 |
1,204 |
1,397 |
1,277 |
1,180 |
1,054 |
823 |
921 |
983 |
993 |
152 |
15,579 |
H/O Management Fee (Toronto) |
USD 000 |
1,577 |
1,368 |
1,515 |
1,404 |
1,246 |
1,446 |
1,321 |
1,221 |
1,090 |
851 |
953 |
1,017 |
1,028 |
157 |
16,197 |
H/O Management Fee (Banro Congo Mining) |
USD 000 |
2,629 |
2,279 |
2,525 |
2,341 |
2,076 |
2,410 |
2,202 |
2,035 |
1,817 |
1,419 |
1,588 |
1,695 |
1,713 |
262 |
26,994 |
TOTAL OPERATING COST |
USD 000 |
68,462 |
76,833 |
95,810 |
89,612 |
93,194 |
90,375 |
78,640 |
80,024 |
62,701 |
63,662 |
67,268 |
51,590 |
46,947 |
10,946 |
976,260 |
Total Operating Cost per ounce |
USD/oz |
591 |
727 |
818 |
825 |
968 |
808 |
770 |
848 |
744 |
967 |
913 |
656 |
591 |
901 |
783 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
EBITDA |
USD 000 |
76,409 |
50,070 |
44,735 |
40,697 |
22,372 |
43,773 |
43,952 |
33,251 |
38,446 |
15,337 |
21,141 |
42,786 |
48,408 |
3,632 |
525,108 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Capital Expenditure |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Project Capital - TMF Construction |
USD 000 |
17,950 |
5,000 |
0 |
8,000 |
0 |
0 |
5,000 |
5,000 |
5,000 |
0 |
0 |
0 |
0 |
0 |
45,950 |
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Mining Capital (New & Replacement Fleet and Others)
|
USD 000 |
4,694 |
5,397 |
4,801 |
4,728 |
3,886 |
299 |
543 |
347 |
544 |
0 |
0 |
0 |
0 |
0 |
25,239 |
Process Capital (Plant Modification and Others) |
USD 000 |
941 |
2,000 |
3,500 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
6,441 |
Grade Control Capital |
USD 000 |
271 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
271 |
General & Admin Capital |
USD 000 |
10,562 |
10,000 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
0 |
20,562 |
Sustaining Capital (Tailings Wall Lifts) |
USD 000 |
2,884 |
3,064 |
3,064 |
3,064 |
3,064 |
3,064 |
3,064 |
3,064 |
3,064 |
3,064 |
0 |
0 |
0 |
0 |
30,467 |
Banro Foundation |
USD 000 |
116 |
106 |
117 |
109 |
96 |
112 |
102 |
94 |
84 |
66 |
74 |
79 |
79 |
12 |
1,246 |
TOTAL CAPITAL TAXES AND LEVY |
USD 000 |
37,418 |
25,567 |
11,483 |
15,900 |
7,046 |
3,475 |
8,709 |
8,505 |
8,693 |
3,130 |
74 |
79 |
79 |
12 |
130,177 |
Total Financing Cost per ounce |
USD/oz |
323 |
242 |
98 |
146 |
73 |
31 |
85 |
90 |
103 |
48 |
1 |
1 |
1 |
1 |
104 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
ALL-IN COST (Cash plus Capital) |
USD 000 |
105,880 |
102,400 |
107,292 |
105,512 |
100,240 |
93,849 |
87,350 |
88,529 |
71,393 |
66,792 |
67,342 |
51,669 |
47,026 |
10,958 |
1,106,437 |
All-In Cost per ounce |
USD/oz |
914 |
968 |
916 |
972 |
1,041 |
840 |
855 |
938 |
847 |
1,015 |
914 |
657 |
592 |
902 |
888 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Project Cashflow |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
Net Cashflow |
USD 000 |
38,991 |
24,502 |
33,253 |
24,797 |
15,326 |
40,298 |
35,242 |
24,746 |
29,753 |
12,207 |
21,068 |
42,707 |
48,329 |
3,620 |
394,931 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV Sensitivities |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV 0.0% |
$ m |
395 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV 5.0% |
$ m |
285 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV 8.0% |
$ m |
241 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV 9.5% |
$ m |
223 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV 10.0% |
$ m |
217 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV 12.5% |
$ m |
193 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
NPV 15.0% |
$ m |
173 |
|
|
|
|
|
|
|
|
|
|
|
|
|
|
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There are no adjacent properties
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24 |
OTHER RELEVANT DATA AND
INFORMATION |
No other information is considered
necessary.
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25 |
INTERPRETATION AND
CONCLUSIONS |
SRK (UK) considers the geology at
Twangiza Main and North to be well understood. The folded geometry of the
stratigraphy and sills has been modelled in detail allowing a confident
interpretation of the lithological distinction of ore types. The weathering that
has affected the deposit is also modelled in a series of continuous layers which
differentiated ore types of differing metallurgical behaviour and densities in
sufficient detail for the resource model.
The zone of mineralisation within the
fold hinge is reasonably continuous and well sampled by adits and resource
drilling. Data quality and quantity have been reviewed and are sufficient in SRK
(UK)s opinion to support the level of confidence in the Mineral Resource.
Density data from drilling samples has
been factored to bring the block model tonnage estimate in line with actual
tonnages recorded in historical production. The adjustment was supported by a
recent small programme of in pit density check samples. This programme should be
continued to consolidate new views on the ore density.
The model has been appropriately
depleted to the end of December 2014 pit survey.
Continue to monitor dry density and
moisture content of ore by sampling daily and analysing on a monthly basis.
25.3 |
Metallurgy and
Processing |
25.3.1 |
Historical Plant
Performance |
While the plant has been running for
approximately three years it has not been operated at the design conditions due
to issues relating to shortage of funds. Oxide ore handling has been
problematical especially in the wet season; a roof has only recently been
installed over the ROM crusher and re-handling area. A lack of spare parts
resulted in:
|
|
low availability of primary, secondary and
tertiary crushing |
|
|
|
|
|
ball mills where new liners have not been
available, and |
|
|
|
|
|
some CIL tanks were off-line due to
non-availability of spares for the agitators. |
The supply of reagents and other plant
consumables has been limited resulting in:
|
|
a lack of steel balls for the mills which has
impacted throughput, grinding circuit product |
|
|
|
|
|
size and an excessive amount of mill oversize; |
|
|
|
|
|
cyanide addition has been limited which has
resulted in poor leaching conditions; and |
|
|
|
|
|
the lack of supply of activated carbon has
reduced the adsorption efficiency in the CIL |
|
|
|
|
|
circuit and increased the soluble gold losses
in the tailings. |
Overall these issues have resulted in
lower than expected gold recovery, 82.5% gold recovery compared to 88 to 90% in
the feasibility study.
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25.3.2 |
Recent Plant Modifications |
The plant has been modified to process
1.7 Mtpa ore. In the latter half of 2014 the crushing circuit, grinding circuit
and the downstream cyanidation and adsorption circuit all operated at the
increased rate, so from a simple capacity perspective, as opposed to circuit
performance, they can handle the increased tonnages.
The installation of the covered RoM
stockpile has improved ore handleability. The modification to the layout of the
primary crusher has resulted in more consistent feeding of the crusher and has
improved access for maintenance. In addition a revised replacement strategy for
the sizer teeth has reduced equipment downtime and teeth replacement operating
costs.
In addition to these modifications, the
expected increase in the proportion of more competent non-oxide ore in the feed
blend in the next few years should also improve ore handleability issues. SRK
(UK) is of the opinion that these changes will increase the circuit utilisation
with resultant increases in plant throughput going forward.
SRK (UK) is of the opinion that once
the equipment is operating according to design and with sufficient reagents and
consumables the overall gold recovery from oxide ores will increase from the
historically achieved figures.
Pit optimisations have been conducted
based upon resource models revised for historical reconciliation to December
2014 and as such, are understood to reflect the current understanding of the
resource model and its exploitation performance on mining.
Optimisations have also been based
largely on costs generated from first principles, a zero-based approach,
following a review of historical costs. The cost structure is estimated to be
considerably lower than it has been historically and is dependent on the
achievement of key productivity and performance targets and operating strategies
(hydroelectric supply, process throughput and recovery of harder ores, drill and
blast performance, mining fleet operating strategies).
As a consequence of the lower forecast
costs, the cut-off grades are lower and optimisation and design inventories are
higher. This is mitigated to a degree by the resource grade profile, low
stripping ratios and relatively flat optimisation cashflow profiles.
25.5 |
Operating & Capital
Costs |
Costs have largely been generated from
first principles, a zero-based approach, following a review of historical
costs. There are a number of key reasons for this approach:
|
|
Diesel fuel price five year forecasts are
significantly lower than recent 2014 performance at USD1.52/l,
representing a reduction in a significant input cost; |
|
|
|
|
|
The operation proposes to source future power
from a hydroelectric supply at an estimated 30% lower price than currently
obtained from diesel sources; |
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The mining fleet basis has been upgraded to meet the
future increase in stripping ratio from less than 1.0 historically to an
average 3.2 in the future on a waste tonnage to ore tonnage basis, an
escalation to an average 12 Mtpa over the five year period. As such this
represents a higher productivity, economy of scale and potentially lower
cost structure going forward; |
|
|
|
|
|
Procurement and some contracts for supply of materials
and services have been rationalised to provide a potentially lower cost
structure going forward; |
|
|
|
|
|
Increased working capital will remove restrictions,
additional costs and operating inefficiencies; |
|
|
|
|
|
Historical reconciliation of both ore and waste movements
indicates that while ore reconciliation has been negative, strong positive
reconciliation of waste movement has occurred. While the strongly positive
waste movement reconciliation likely includes rehandle and borrow pit
movements, a downwards correction of truck factors for historical
movements of the order of 20-30% will yield a corresponding escalation in
historical unit waste mining costs; and, |
|
|
|
|
|
An increased proportion of transition and fresh material
types will incur additional mining costs (drill and blast, excavation
costs, secondary breakage) and process costs (throughput, wear and
recovery). |
An analysis of historical costs for
2014 indicates that the proposed zero-based cost structure approach represents a
significant cost reduction of up to 30% on certain unit cost estimates going
forward. This is a risk to the project if the proposed future operating
strategy, cost reductions, operating efficiencies and mining fleet operating
efficiencies are not achieved. This will require monitoring of unit costs for
mining, processing and G&A. It will also require further technical
investigation and management to ensure that:
|
|
Additional tailings capacity is achieved at projected
capital construction and operating costs. Approvals and contracts will
need to be negotiated and agreed; |
|
|
|
|
|
Hydroelectric power supply is sourced at projected costs
in accordance with the projected timeline. Approvals and contracts will
need to be established in a timely fashion; |
|
|
|
|
|
Process throughput rates and recoveries are achieved
and/or plant modifications implemented in order to meet targets; and, |
|
|
|
|
|
Drill and blast production rates and costs are achieved
based on proposed patterns. This will depend on operating performance. The
supply and cost of explosives and accessories will need to be managed in
order to ensure that the cost and timely supply is achieved on a regular
basis. |
During 2014, the nature and scale of
the Twangiza operation changed significantly with the completion and
commissioning of the process plant expansion project. Following commissioning in
mid-2014, operating levels increased to reach the current 1.7Mtpa design
capacity by the fourth quarter. Recognizing managements priority during H2 2014
was to maintain targeted LOM throughput, many cost savings in the LOM base case
included in this report had yet to be implemented and realized as at December
31, 2014; SRK (UK) requested a sensitivity on the financial model to understand
the consequence of using various historical operating cost profiles on a going
forward basis. While this cost profile is not assumed to be indicative of future
operations, it quantifies the implications if the forecasted cost profile is not
realized. This sensitivity indicates that the project would return a
considerably lower NPV over the LOM when considering solely the historical
costs. The achievements made by the Twangiza Mine while advancing from a 1.3Mtpa
throughput to the 1.7Mtpa LOM throughput lends support to the operations
ability to reduce future costs and improve productivities, but further delivery
of anticipated cost savings must be realized to achieve the forecast. This
highlights the importance of achieving the cost, productivity and performance
programmes and planning timelines to achieve the base case NPV in this report.
Operating costs and mine reconciliation will need to be monitored stringently to
ensure that unit cost performance targets are being achieved.
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Capital costs for key initiatives in
respect of the tailings dam and hydroelectric supply will need to be confirmed
by firm quotation and tendering.
There is no engineered construction
schedule for TMF1 at the moment, so there is a risk of interruption to the
production schedule if TMF1 construction takes longer than currently expected.
Whilst Twangiza Mining has a good track
record of relocating families and artisanal miners, significant challenges
remain before mining in Twangiza North and TMF1 construction can begin given
that there will need to be dealings with a different community group from
before.
The costs associated with the
relocation are significant, to some extent these costs can be based on similar
historical activity and previously established community relationships; but
there are also new challenges which may require greater effort. The project can
afford the estimated cost; however there is clearly a need to properly plan and
execute the relocation exercise to minimise any delay to the project or
additional costs.
A discount rate of 5% has been applied
for calculation of project NPV over the LOM. SRK (UK) regards this rate as
slightly low given the operating environment and country risk rating. Generally
the discount rate should incorporate allowance for the risk-free interest rate,
operational and country risk premiums. A discount rate of 8-10% is regarded as
more appropriate.
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Further work will eventually be
required to continue exploration at previously identified targets on the
Twangiza concessions, should the need arise to find more oxide ore at any
time.
The resource model has been modified to
reflect new ideas concerning ore density which have been mainly based on
historical tonnage reconciliation and also on a few in-pit check density
samples. SRK (UK) recommends continuing with in pit density sampling using the
same method, seeking to build a geo-referenced database of samples on each
flitch taken on a square grid spacing of approximately 20m.
The team on site needs to revise and
unify its understanding of wet density, moisture content and dry density as
applied to ore material. Further, the distinction between density and specific
gravity values used by the plant for slurry calculations also needs to be
clearly understood.
It will be important to continue
monitoring production grades, gold recoveries and densities on a monthly basis
going forward. Periodically, (at least monthly or quarterly), a reconciliation
exercise should be completed to assess whether further modifications need to be
made to the resource block-model.
26.2.3 |
Grade Control Modelling |
The creation of wireframes based on
grade control sampling and mapping is an area that could be improved in terms of
geological accuracy, 3D continuity and speed with which wireframes can be
created. The grade control block models should incorporate newly captured in-pit
dry density data. New and improved ways of interfacing with mine survey data
should be considered. Better block model version control should assist with
historical production analysis in the future.
26.3 |
Metallurgy and
Processing |
26.3.1 |
Measuring and Reporting
Production |
|
|
Instigate a more accurate measurement of
solids (scats) removed from the circuit returned to the stockpile. |
|
|
|
|
|
Determine the true density of the oxide and
transition ore types for use in the mass flow loop calculation. The
density of feed solids should be checked at least every 6 months as the
ore type change over time. |
|
|
|
|
|
The feed grades to the plant are variable
include the modelled ore hardness in the mine-plan to try and achieve a
more consistent feed blend to the plant. |
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|
The new feed weightometers will improve the accuracy of
the measurement of the feed tonnage to an acceptable level. Review of the
plant operating statistics indicated that the measurement of the scrubber
underflow slurry feed reporting to mill no.1 is potentially erroneous. The
slurry flow rate and the slurry density are measured and computed using a
standard mass flow loop to give a percentage solids and a solids tonnage.
This calculation uses the true solids density. It is apparent that the
density used by operations is too high and does not represent the true
density of oxide and/or transition ore. This can potentially introduce a
10% error in to the mass flow calculation. Further true solids density
measurements should be made to correct this situation. SRK (UK) recommends
that this is checked at least every six months to reflect the changing ore
blend with time. |
|
|
|
|
|
The grinding circuit can theoretically process 1.7 Mtpa
of mixed ore to the required grind size provided sufficient grinding media
is available. However, as the proportion of harder non-oxide ore increases
it is likely that the circuit product size will coarsen and this may
impact the leaching efficiency slightly. SRK (UK) recommends that the
effect of the changes in ore blend on grind size, including an assessment
of the effect on leaching performance, is closely monitored in the future. |
26.3.2 |
Technical Improvements |
|
|
Review the reasons for the low percent solids in the
leach feed. This reduces the residence time in the leach/CIP circuit and
will be affecting overall gold extraction in the circuit. If the clay
content is causing viscosity issues in the pulp necessitating lower
percent solids consider use of viscosity modifiers. |
|
|
|
|
|
Target lower gold in solution losses from CIP by
adjustment of carbon inventory and carbon movement. |
|
|
|
|
|
The crushing circuit has historically been operated to
produce a coarse product as feed to the mill. At times the top size has
been 30mm which is too coarse for a ball mill and has a detrimental effect
on mill performance. With adequate supply of crusher wear parts the
crushing circuit should be operated to produce the design product,
grinding circuit feed, of nominally -10 mm. |
|
|
|
|
|
The CIL circuit volumetric capacity has been increased by
the addition of tanks in order to maintain the CIL residence time
residence time at approximately 15 hours at the revised throughput of 1.7
Mtpa. The testwork performed indicates that 15 hours residence time is at
the lower end of what is required especially for some of the transition
and fresh ore types, therefore further increases in residence time should
be researched and implemented. |
|
|
|
|
|
Operating data indicates that the percentage of solids in
the leach feed is variable. A low percent solids in the grinding circuit
product from the hydrocyclone overflow or the leach feed increases
volumetric flow and reduces leaching residence time. This may be due to
the feed characteristics, especially the clay content of the ore, but may
also have been influenced by the less than optimal operation of the two
mills. If clay is a problem this should either be addressed as part of the
ore blending strategy to maintain a maximum level of clay in the feed, or
through the addition of viscosity modifiers to the circuit. The latter
would incur additional operating costs. |
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The gravity and intensive cyanidation
circuit has not been operated continuously to date. Testwork indicated that GRG
was present in oxide and transition ores and thus this circuit should be
recommissioned and brought in to operation as soon as possible. If the circuit
is operated without the gravity circuit there is a risk that coarser gold
particles may not be fully leached within the leaching circuit and there may be
losses to tailings. In addition it is noted that the adsorption circuit is
designed to process ore post gravity and thus operation of the gravity circuit
will reduce the gold load on the carbon adsorption system and prevent potential
losses to tailings.
26.3.3 |
Fund Fully to Maximise
Performance |
|
|
Avoid the spares issue with the crushing and grinding
circuit as it results in very inefficient operation of the circuit
(excessive sized solids in mill feed). |
|
|
|
|
|
Maintain the correct levels of steel balls, cyanide and
carbon in the circuit. |
|
|
|
|
|
The grinding mills have been operating with reduced steel
loads. Reduced steel load in the grinding mills affects mill power draw
and thus circuit performance. With adequate supply of steel grinding media
the correct design steel loadings in the two grinding mills should be
maintained which in turn will allow the grinding circuit product to
achieve the design 80% -75 microns. |
|
|
|
|
|
Twangiza Mining has operated the CIL circuit with
starvation amounts of cyanide. Once sufficient cyanide is available the
performance of the leach should be evaluated to determine the optimum
cyanide concentration and thereby optimum leaching performance. |
Carbon inventory in the CIP circuit
affects soluble gold losses to tailings. Once adequate supplies of activated
carbon are available Twangiza should maintain the design carbon concentrations
in the circuit. Once stable operation is achieved the soluble gold losses should
be assessed and optimised.
26.3.4 |
Additional Testwork |
|
|
Review historical metallurgical testwork and
identify and perform additional testwork on potential feed solids that
will be processed in the next few years. This should include: |
|
a. |
testing and studies to further evaluate the grinding
requirements of future ore types, work index and power draw variations,
grinding equipment size and throughput of the mills, grind size
achievable, etc. |
|
|
|
|
b. |
the leaching requirements of future ore types to
especially the residence times required for satisfactory leaching. This
may result in additional leaching tanks being required. |
|
|
|
|
c. |
Further testwork is required to identify an effective
treatment route for CMS material. |
|
|
Further comminution and leaching tests should be
performed on samples of fresh (sulphide) ore in the next two years in
order to assess the impact on circuit performance and identify
modifications required for successful treatment of ore blends containing
an increased proportion of harder fresh ore. |
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|
Further testwork should be performed on samples
of transition and fresh ores from all zones in order to be able to
optimise the leaching circuit for future ore blends. This may necessitate
the addition of new leaching tank capacity but this should be assessed
from both a metallurgical and an economic perspective. |
26.3.5 |
Ore Type Definition and
Management |
|
|
Metallurgical testwork performed in the study phase of
the project indicated that CMS ore is refractory and does not exhibit
acceptable leaching characteristics. This ore should not be added to the
blended feed and if mined should be stockpiled separately. |
|
|
|
|
|
The feed grade to the plant has been variable. Better
communication between geology- mining-plant should be instigated to
improve ore blending and feed consistency. |
SRK (UK) recommends further pit
optimisation, design and sensitivity analysis to better assess Mineral Reserve
and project cashflow sensitivity to operating cost escalation.
Review of current optimisation shells
compared to the December 2013 pit designs used for the December 2014 Mineral
Reserve shows that there are some differences that should be better accounted
for by redesign of the pits based on the revised resource models and
optimisation sensitivity analyses.
In particular, SRK (UK) recommends
creating or revising pit designs for:
|
|
Twangiza Main Intermediate Pit (Cut 2) to
provide for ramp access in the design; |
|
|
|
|
|
Valley Fill pit design; and, |
|
|
|
|
|
Waste dump designs over the LOM. |
The current schedule has different
proportions of hard and soft ore types and in particular first / transitional /
oxide ore type each year.
SRK (UK) is of the opinion that there
may be a risk associated with this variability and that the risk can be
mitigated to some extent by rescheduling to even out the ore types in the blend
each month.
No designs for waste dumps have been
presented to SRK (UK) for review, it is recommended that designs are completed
to assist with planning going forward.
26.5 |
Operating & Capital
Costs |
Continue monitoring operating costs on
a monthly basis and periodically reviewing these at least annually.
Some of the larger items in the capital
costs going forward, such as the TMF, require additional design and engineering
work to increase confidence in the cost estimate.
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TMF1 requires additional design work
and there should be a dedicated engineering and construction schedule to
increase confidence in successful delivery of expanded tailings storage capacity
in time for the mine plan; without this there is a risk that production will be
interrupted.
SRK (UK) recommends that the on-going
relocation process be given a high priority; otherwise parts of the mine plan
will be at risk or delayed.
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In addition to references in the table
below, a number of references to metallurgical testwork are given throughout
Section 13.
SRK (UK) has accumulated numerous
spreadsheet and monthly report documents from the mining operation; these are
not individually detailed here.
AUTHOR |
DATE |
TITLE |
SOURCE |
|
|
|
|
|
13th |
Preliminary Assessment NI 43-101 |
|
SENET |
September, |
Technical Report, Twangiza Gold |
|
|
2007 |
|
Banro |
|
|
Project, South Kivu Province, |
|
|
|
Democratic Republic of Congo |
|
|
|
UPDATED FEASIBILITY STUDY |
|
|
|
NI 43-101 TECHNICAL REPORT, |
|
SENET |
17TH July 2009 |
TWANGIZA GOLD PROJECT, |
www.sedar.com |
|
|
South Kivu Province, |
|
|
|
Democratic Republic of Congo |
|
|
|
ECONOMIC ASSESSMENT NI 43-101 |
|
|
|
TECHNICAL REPORT, |
|
|
24th March |
|
|
SENET |
|
TWANGIZA PHASE 1 GOLD PROJECT, |
www.sedar.com |
|
2011 |
|
|
|
|
South Kivu Province, Democratic |
|
|
|
Republic of the Congo |
|
|
|
Independent National Instrument 43-101 |
|
|
|
Technical |
|
Venmyn |
|
Report on the Namoya Gold Project, |
|
Deloitte |
12th May 2014 |
Maniema Province, Democratic Republic of |
www.sedar.com |
(Ref:D1417R) |
|
the Congo |
|
|
|
Prepared for Namoya Mining SARL (a |
|
|
|
subsidiary of Banro Corporation) |
|
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28 |
DATE AND SIGNATURE PAGE |
The independent Qualified Persons
(within the meaning of NI 43-101) for the purposes of this report are Martin
Pittuck and David Pattinson. Further review and authoring has been undertaken by
Allan Blair; these authors represent SRK Consulting (UK) Limited.
The SRK authors have undertaken an
extensive review of Twangiza Minings technical and economic data and have
reviewed Twangiza Minings contributions to this report.
Daniel Bansah, Head of Projects and
Operations of Banro Corporation has provided a QP certificate; he has supervised
the compilation of material provided by Twangiza Mining and has overall
responsibility for certain sections in this report.
Signed, this 29th day of
July, 2015.
Martin Pittuck, MSc, CEng, MIMMM
Director,
Corporate Consultant (Mining Geology)
SRK Consulting (UK) Limited
Allan Blair, B.App.Sc, MBA, MAusIMM
Principal Consultant
(Mining Engineering)
SRK Consulting (UK) Limited
David Pattinson, BSc, PhD, CEng, MIMMM,
Corporate
Consultant (Minerals Processing & Metallurgy)
SRK Consulting (UK)
Limited
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29 |
CERTIFICATES OF QUALIFIED
PERSONS |
CERTIFICATE OF QUALIFIED PERSON
Martin Pittuck
I, Martin Frank Pittuck, MSc., C.Eng, MIMMM do
hereby certify that: |
|
1. |
I am Director and Corporate Consultant (Mining Geology)
of SRK Consulting (UK) Limited with an office at 5th Floor, Churchill
House, Churchill Way, Cardiff CF10 2HH, United Kingdom. |
|
|
2. |
This certificate applies to the technical report with an
effective date of July 29, 2015 and titled NI 43-101 Technical Report,
Mineral Resource and Reserve Update, December 31 2014, Twangiza Gold Mine,
Democratic Republic of the Congo (the Technical Report). |
|
|
3. |
I am a graduate with a Master of Science in Mineral
Resources gained from Cardiff College, University of Wales in 1996 and I
have practised my profession continuously since that time. Since
graduating I have worked as a consultant at SRK on a wide range of mineral
projects, specializing in precious and rare metals. I have undertaken many
geological investigations, resource estimations, mine evaluation technical
studies and due diligence reports. I am a member of the Institute of
Materials, Minerals and Mining (Membership Number 49186) and I am a
Chartered Engineer. |
|
|
4. |
I have read the definition of Qualified Person set out
in National Instrument 43-101 (NI 43-101) and certify that by reason of my
education, affiliation with a professional association (as defined in NI
43-101) and past relevant work experience, I am a Qualified Person for
purposes of NI 43- 101. |
|
|
5. |
I visited the Twangiza property between 12th
and 15th March, 2015. |
|
|
6. |
I am responsible for Sections 2 to 16, 20 and 22 to 29 of
the Technical Report. |
|
|
7. |
I am independent of the issuer as described in section
1.5 of NI 43-101. |
|
|
8. |
I have not had prior involvement with the property that
is the subject of the Technical Report. |
|
|
9. |
I have read NI 43-101 and the sections of the Technical
Report I am responsible for have been prepared in compliance with
NI43-101. |
|
|
10. |
As at the effective date of the Technical Report, to the
best of my knowledge, information and belief, the sections of the
Technical Report I am responsible for contain all scientific and technical
information that is required to be disclosed to make the Technical Report
not misleading. |
Signed the 29th day of July, 2015.
Martin Frank Pittuck, MSc. C.Eng, MIMMM Director and Corporate
Consultant (Mining Geology)
U6391 Twangiza
Reserve Tech_Report Final.docx |
July
2015 |
Page 185 of 187
SRK Consulting |
NI 43-101 Twangiza Main Report |
CERTIFICATE OF QUALIFIED PERSON
David Pattinson
I, David Pattinson, PhD., BSc., C.Eng, MIMMM
do hereby certify that: |
|
1. |
I am a Corporate Consultant (Minerals Processing) of SRK
Consulting (UK) ) Limited with an office at 5th Floor, Churchill House,
Churchill Way, Cardiff CF10 2HH, United Kingdom. |
|
|
2. |
This certificate applies to the technical report with an
effective date of July 29, 2015 and titled NI 43-101 Technical Report,
Mineral Resource and Reserve Update, December 31 2014, Twangiza Gold Mine,
Democratic Republic of the Congo (the Technical Report). |
|
|
3. |
I am a graduate with a Doctor of Philosophy degree in
Minerals Engineering gained from Birmingham University, UK in 1982 and I
have practised my profession continuously since that time. Since
graduating I have worked for an international engineering company for 23
years and then as a consultant at SRK working on a wide range of mineral
projects including design and commissioning activities, technical studies
and numerous due diligence reports in gold and base metal plants. I am a
member of the Institute of Materials, Minerals and Mining (Membership
Number 46888) and I am a Chartered Engineer. |
|
|
4. |
I have read the definition of Qualified Person set out
in National Instrument 43-101 (NI 43-101) and certify that by reason of my
education, affiliation with a professional association (as defined in NI
43-101) and past relevant work experience, I am a Qualified Person for
purposes of NI 43- 101. |
|
|
5. |
I visited the Twangiza property between 12th
and 15th March, 2015. |
|
|
6. |
I am responsible for sections 17, 18 and 19.1 to 18.7 of
the Technical Report. |
|
|
7. |
I am independent of the issuer as described in section
1.5 of NI 43-101. |
|
|
8. |
I have not had prior involvement with the property that
is the subject of the Technical Report. |
|
|
9. |
I have read NI 43-101 and the sections of the Technical
Report I am responsible for have been prepared in compliance with
NI43-101. |
|
|
10. |
As at the effective date of the Technical Report, to the
best of my knowledge, information and belief, the sections of the
Technical Report I am responsible for contain all scientific and technical
information that is required to be disclosed to make the Technical Report
not misleading. |
Signed the 29th day of July, 2015.
David Pattinson, BSc, PhD, C.Eng, MIMMM
Corporate
Consultant (Minerals Processing)
U6391 Twangiza
Reserve Tech_Report Final.docx |
July
2015 |
Page 186 of 187
SRK Consulting |
NI
43-101 Twangiza Main Report |
CERTIFICATE OF QUALIFIED PERSON
Daniel Bansah
I, Daniel Kenneth Bansah, MSc.(MinEx),
MAusIMM(CP), do hereby certify that: |
|
1. |
I am the Head of Projects and Operations of Banro
Corporation, 1 First Canadian Place, Suite 7070, 100 King Street West,
Toronto, Ontario, M5X 1E3, Canada. |
|
|
2. |
This certificate applies to the technical report with an
effective date of July 29, 2015 and titled NI 43-101 Technical Report,
Mineral Resource and Reserve Update, December 31 2014, Twangiza Gold Mine,
Democratic Republic of the Congo (the Technical Report). |
|
|
3. |
I am a graduate of University of Science and Technology,
School of Mines, Ghana with a degree in Geological Engineering (1988). I
also have a Master of Science in Mineral Exploration with Distinction
gained from Leicester University, United Kingdom, and I have over 26 years
of continuous professional experience in the gold mining industry. I was
Banro Corporations Vice President of Exploration from 2007 to 2013. Prior
to joining Banro in 2004, I was the Group Mineral Resource Manager with
Ashanti Goldfields, with responsibilities for the coordination, auditing
and compilation of Ashanti's Mineral Resources and Ore Reserves in Africa.
I am a Member and a Chartered Professional of the Australasian Institute
of Mining and Metallurgy. |
|
|
4. |
I have read the definition of qualified person set out
in National Instrument 43-101 (NI 43-101) and certify that by reason of
my education, affiliation with a professional association (as defined in
NI 43-101) and past relevant work experience, I am a qualified person
for purposes of NI 43- 101. |
|
|
5. |
I have had prior involvement with the property that is
the subject of the Technical Report, in that I have worked on the Twangiza
property as an employee since September 2005 to the date of this
certificate. I have personally visited the property on many occasions with
the last visit on July 8, 2015. |
|
|
6. |
I am responsible for items 18.8 and 20 of the Technical
Report. |
|
|
7. |
I am not independent of the issuer as described in
section 1.5 of NI 43-101 by virtue of being Head of Projects and
Operations of Banro Corporation and a director of Twangiza Mining SA. |
|
|
8. |
I have read NI 43-101 and the sections of the Technical
Report I am responsible for have been prepared in compliance with NI
43-101. |
|
|
9. |
As at the effective date of the Technical Report, to the
best of my knowledge, information and belief, the sections of the
Technical Report I am responsible for contain all scientific and technical
information that is required to be disclosed to make the Technical Report
not misleading. |
Signed the 29th day of July, 2015.
Daniel Kenneth Bansah, MSc.(MinEx), MAusIMM(CP)
Head of
Projects and Operations
Banro Corporation
U6391 Twangiza
Reserve Tech_Report Final.docx |
July
2015 |
Page 187 of 187
CONSENT OF QUALIFIED PERSON
TO: |
Alberta Securities Commission |
|
British Columbia Securities Commission |
|
Manitoba Securities Commission |
|
Saskatchewan Financial Services Commission
|
|
Ontario Securities Commission |
|
New Brunswick Securities Commission |
|
Nova Scotia Securities Commission |
|
Prince Edward Island Securities Office |
|
Securities Commission of Newfoundland and
Labrador |
|
|
AND TO: |
Banro Corporation |
|
|
RE: |
Technical report of SRK Consulting (UK) Limited prepared
for Twangiza Mining SA (a subsidiary of Banro Corporation) dated July 29,
2015 and entitled "NI 43-101 Technical Report, Mineral Resource and
Reserve Update, December 31, 2014, Twangiza Gold Mine, Democratic Republic
of the Congo" (the "Technical Report") |
|
|
I, Daniel Bansah, hereby consent to the public filing of the
Technical Report by way of SEDAR and EDGAR by Banro Corporation. I also confirm
that I have read the press release of Banro Corporation dated July 29, 2015 and
such press release fairly and accurately represents the information in the
sections of the Technical Report for which I am responsible.
DATED the 29th day of July, 2015.
(signed)
"Daniel Bansah" |
|
Daniel Bansah |
|
|
SRK Consulting (UK) Limited
5th Floor Churchill House
17 Churchill Way
City and County of Cardiff
CF10 2HH, Wales
United Kingdom
E-mail: enquiries@srk.co.uk
URL: www.srk.co.uk
Tel: + 44 (0) 2920 348 150
Fax: + 44 (0) 2920 348 199 |
CONSENT OF QUALIFIED PERSON
TO: |
Alberta Securities Commission |
|
British Columbia Securities Commission |
|
Manitoba Securities Commission |
|
Saskatchewan Financial Services Commission
|
|
Ontario Securities Commission |
|
New Brunswick Securities Commission |
|
Nova Scotia Securities Commission |
|
Prince Edward Island Securities Office |
|
Securities Commission of Newfoundland and
Labrador |
|
|
AND TO: |
Banro Corporation |
|
|
RE: |
Technical report of SRK Consulting (UK) Limited prepared
for Twangiza Mining SA (a subsidiary of Banro Corporation) dated July 29,
2015 and entitled "NI 43- 101 Technical Report, Mineral Resource and
Reserve Update, December 31, 2014, Twangiza Gold Mine, Democratic Republic
of the Congo" (the " Technical Report") |
|
|
I, Martin Pittuck, hereby consent to the public filing of the
Technical Report by way of SEDAR and EDGAR by Banro Corporation. I also confirm
that I have read the press release of Banro Corporation dated July 29, 2015 and
such press release fairly and accurately represents the information in the
sections of the Technical Report for which I am responsible.
DATED the 29th day of July, 2015.
|
|
|
|
Martin Pittuck |
|
|
Registered Address: 21 Gold Tops, City and County
of Newport, NP20 4PG,
Wales, United Kingdom.
SRK
Consulting (UK) Limited Reg No 01575403 (England and Wales)
|
Group
Offices: Africa
Asia
Australia
Europe
North America
South America
|
|
SRK Consulting (UK) Limited
5th Floor Churchill House
17 Churchill Way
City and County of Cardiff
CF10 2HH, Wales
United Kingdom
E-mail: enquiries@srk.co.uk
URL: www.srk.co.uk
Tel: + 44 (0) 2920 348 150
Fax: + 44 (0) 2920 348 199 |
CONSENT OF QUALIFIED PERSON
TO: |
Alberta Securities Commission |
|
British Columbia Securities Commission |
|
Manitoba Securities Commission |
|
Saskatchewan Financial Services Commission
|
|
Ontario Securities Commission |
|
New Brunswick Securities Commission |
|
Nova Scotia Securities Commission |
|
Prince Edward Island Securities Office |
|
Securities Commission of Newfoundland and
Labrador |
|
|
AND TO: |
Banro Corporation |
|
|
RE: |
Technical report of SRK Consulting (UK) Limited prepared
for Twangiza Mining SA (a subsidiary of Banro Corporation) dated July 29,
2015 and entitled "NI 43- 101 Technical Report, Mineral Resource and
Reserve Update, December 31, 2014, Twangiza Gold Mine, Democratic Republic
of the Congo" (the " Technical Report") |
|
|
I, David Pattinson, hereby consent to the public filing of the
Technical Report by way of SEDAR and EDGAR by Banro Corporation. I also confirm
that I have read the press release of Banro Corporation dated July 29, 2015 and
such press release fairly and accurately represents the information in the
sections of the Technical Report for which I am responsible.
DATED the 29th day of July, 2015.
|
|
|
|
David Pattinson |
|
|
Registered Address: 21 Gold Tops, City and County
of Newport, NP20 4PG,
Wales, United Kingdom.
SRK
Consulting (UK) Limited Reg No 01575403 (England and Wales)
|
Group
Offices: Africa
Asia
Australia
Europe
North America
South America
|
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