apogee minerals ltd. technical report on the preliminary assessment

Transcripción

APOGEE MINERALS LTD.
TECHNICAL REPORT ON THE
PRELIMINARY ASSESSMENT OF THE
PULACAYO PROJECT,
PULACAYO TOWNSHIP, POTOSÍ DISTRICT,
QUIJARRO PROVINCE, BOLIVIA
June 25th, 2010
R. Pressacco, M.Sc.(A), P.Geo.
G. Harris, CEng, MIMMM
M. Godard, P.Eng.
C. Jacobs CEng, MIMMM
SUITE 900 - 390 BAY STREET, TORONTO ONTARIO, CANADA M5H 2Y2
Telephone (1) (416) 362-5135 Fax (1) (416) 362 5763
TABLE OF CONTENTS
Page
1.0 SUMMARY .................................................................................................................... 1 2.0 INTRODUCTION AND TERMS OF REFERENCE............................................... 13 3.0 RELIANCE ON OTHER EXPERTS ......................................................................... 15 4.0 PROPERTY DESCRIPTION AND LOCATION .................................................... 16 4.1 LOCATION ............................................................................................................... 16 4.2 PROPERTY STATUS ............................................................................................... 17 4.2.1 Overview of Bolivian Mining Law .................................................................... 17 4.2.2 Project Ownership .............................................................................................. 19 5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,
INFRASTRUCTURE AND PHYSIOGRAPHY ....................................................... 22 5.1 ACCESS..................................................................................................................... 22 5.2 CLIMATE AND PHYSIOGRAPHY ........................................................................ 22 5.3 LOCAL RESOURCES AND INFRASTRUCTURE ................................................ 23 6.0 HISTORY ..................................................................................................................... 25 7.0 GEOLOGICAL SETTING ......................................................................................... 28 7.1 REGIONAL GEOLOGY ........................................................................................... 28 7.2 DISTRICT GEOLOGY ............................................................................................. 30 7.3 LOCAL GEOLOGY .................................................................................................. 32 7.3.1 Structural Geology ............................................................................................. 34 7.3.2 Hydrothermal Alteration .................................................................................... 35 8.0 DEPOSIT TYPES ........................................................................................................ 36 9.0 MINERALIZATION ................................................................................................... 39 10.0 EXPLORATION .......................................................................................................... 43 10.1 TOPOGRAPHIC SURVEY ....................................................................................... 43 10.2 GEOLOGICAL MAPPING AND SAMPLING ........................................................ 44 10.3 GEOPHYSICAL SURVEY ....................................................................................... 45 11.0 DRILLING ................................................................................................................... 48 11.1 ASC BOLIVIA LDC (2002-2005) ............................................................................ 48 11.2 APOGEE (JAN 2006 – MAY 2008).......................................................................... 48 11.3 APOGEE (JUN 2008 – SEP 2009) ............................................................................ 49 12.0 SAMPLING METHOD AND APPROACH .............................................................. 51 13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY .................................. 53 ii
14.0 DATA VERIFICATION ............................................................................................. 56 15.0 ADJACENT PROPERTIES ....................................................................................... 61 15.1 SAN CRISTOBAL .................................................................................................... 61 15.2 SAN VINCENTE ....................................................................................................... 61 16.0 MINERAL PROCESSING AND METALLURGICAL TESTING ........................ 63 16.1 METALLURGICAL TESTWORK ........................................................................... 63 16.1.1 RDi Preliminary Metallurgical Results, March 2003 ........................................ 63 16.1.2 UTO Metallurgical Testwork, August 2009 ...................................................... 65 16.2 MINERAL PROCESSING ........................................................................................ 73 17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ................... 74 17.1 INTRODUCTION ..................................................................................................... 74 17.2 DESCRIPTION OF THE DATABASE .................................................................... 74 17.3 TOPOGRAPHIC SURFACE ..................................................................................... 75 17.4 HISTORICAL MINE WORKINGS .......................................................................... 75 17.5 METAL PRICE SELECTION ................................................................................... 77 17.6 DOMAIN MODELING ............................................................................................. 79 17.7 TREND ANALYSIS.................................................................................................. 83 17.8 GRADE CAPPING .................................................................................................... 85 17.9 COMPOSITING METHODS .................................................................................... 88 17.10 BULK DENSITY ....................................................................................................... 90 17.11 VARIOGRAPHY....................................................................................................... 90 17.12 BLOCK MODEL CONSTRUCTION ....................................................................... 91 17.13 BLOCK MODEL VALIDATION ............................................................................. 94 17.14 MINERAL RESOURCE CLASSIFICATION CRITERIA ....................................... 94 17.15 RESPONSIBILITY FOR THE ESTIMATE ............................................................. 95 17.16 MINERAL RESOURCE ESTIMATE ....................................................................... 95 18.0 OTHER RELEVANT DATA AND INFORMATION ............................................. 98 18.1 MINING ..................................................................................................................... 98 18.1.1 Mining Method and Design ............................................................................... 98 18.1.2 Mine Development and Production Schedule .................................................... 99 18.1.3 Mining Equipment ........................................................................................... 103 18.2 PROCESSING ......................................................................................................... 105 18.2.1 Pulacayo Process Plant Option ........................................................................ 105 18.2.2 Toll Milling Option .......................................................................................... 111 18.3 INFRASTRUCTURE .............................................................................................. 114 18.3.1 Power Supply ................................................................................................... 114 18.3.2 Water Supply.................................................................................................... 115 18.3.3 Ancillary Buildings .......................................................................................... 115 18.3.4 Roads ................................................................................................................ 115 18.4 ENVIRONMENTAL AND SOCIAL ASPECTS .................................................... 116 18.4.1 Environmental Conditions ............................................................................... 116 18.4.2 Social Conditions ............................................................................................. 117 iii
18.4.3 Impact Assessment, Mitigation, and Management .......................................... 118 18.4.4 Permitting Process............................................................................................ 119 18.4.5 International Financing .................................................................................... 120 18.4.6 Consultation ..................................................................................................... 120 18.4.7 Environmental and Social Capital and Operating Costs .................................. 121 18.5 PROJECT ECONOMICS ........................................................................................ 121 18.5.1 Macro-economic Assumptions ........................................................................ 122 18.5.2 Production Schedules ....................................................................................... 122 18.5.3 Revenue ............................................................................................................ 123 18.5.4 Capital Costs .................................................................................................... 124 18.5.5 Operating Costs ................................................................................................ 128 18.5.6 Project Schedule ............................................................................................... 131 18.5.7 Cash Flow Forecast .......................................................................................... 131 18.5.8 Sensitivity Studies ............................................................................................ 134 19.0 INTERPRETATION AND CONCLUSIONS ......................................................... 140 20.0 RECOMMENDATIONS ........................................................................................... 142 21.0 REFERENCES ........................................................................................................... 145 22.0 SIGNATURES ............................................................................................................ 148 23.0 CERTIFICATES ........................................................................................................ 149 24.0 APPENDICES ............................................................................................................ 154 iv
LIST OF TABLES
Page
Table 1.1 Summary of Mineral Resources, Pulacayo Deposit...........................................2 Table 1.2 LOM Production Forecast at a Cut off Value of 200 g/t Ag Eq. .......................3 Table 1.3 Concentrate Grades and Recovery at Forecast Average Head Grade ................3 Table 1.4 Summary of Base Case Capital Expenditure .....................................................5 Table 1.5 Summary of Base Case Operating Costs ...........................................................5 Table 1.6 Project Base Case - LOM Cash Flow Summary ................................................6 Table 1.7 Toll-Milling Option - LOM Cash Flow Summary.............................................9 Table 6.1 List of Significant Intersections (ASC, 2002) ..................................................27 Table 10.1 Summary Table of Rock Chip Sampling Completed by Apogee ....................45 Table 14.1 Comparison of Micon Check-Assay Results from Drill-hole PUD045 ...........59 Table 16.1 Head Assays of the Pulacayo Metallurgical Composites.................................63 Table 16.2 High/High Grade Locked-Cycle Flotation Tests .............................................65 Table 16.3 Locked-Cycle Flotation Tests Assay Results...................................................67 Table 16.4 Locked Cycle Test Results-No Desliming Prior to Flotation ..........................68 Table 16.5 Metallurgical Balance, Deslimed Prior to Float, Medium Grade Test 4 .........69 Table 16.6 Locked Cycle Test Results - Desliming Prior to Flotation ..............................71 Table 16.7 Concentrate Grades and Recovery at Forecast Average Head Grade ..............73 Table 17.1 Summary of the Pulacayo Drill Hole Database as at October 14, 2009 ..........75 Table 17.2 Summary of the Input Values and NSR Factors, Pulacayo Project .................80 Table 17.3 Summary Statistics for Raw Samples Contained within the Mineralized
Domain Model .................................................................................................86 Table 17.4 Summary Statistics for 1.0 m Composite Samples Contained within the
Mineralized Domain Model .............................................................................89 Table 17.5 Summary of Variographic Parameters for 1.0 m Composite Samples,
Pulacayo Project ..............................................................................................91 Table 17.6 Summary of Block Model Parameters, Pulacayo Project ................................92 Table 17.7 Block Model Validation Results, Pulacayo Project .........................................94 Table 17.8 Summary of Mineral Resources, Pulacayo Deposit.........................................96 Table 17.9 Comparison of Capped vs Uncapped Grades, Pulacayo Deposit ....................96 Table 18.1 Mineral Resources above Silver Equivalent Cut off Values of 125 to 275
g/t Ag Eq. .......................................................................................................102 v
Table 18.2 Base Case LOM Production at a Cut off Value of 200 g/t Ag Eq. ................103 Table 18.3 Base Case LOM Production Schedule ...........................................................103 Table 18.4 Mobile Mine Equipment List .........................................................................104 Table 18.5 Pulacayo Process Design Criteria ..................................................................105 Table 18.6 Pulacayo Bond Work Index (kWh/st) ............................................................106 Table 18.7 Base Case LOM Processing Schedule ...........................................................122 Table 18.8 NSR Parameters .............................................................................................123 Table 18.9 Summary of Base Case Capital Expenditure .................................................124 Table 18.10 Mining Capital Costs .....................................................................................125 Table 18.11 Process Capital Expenditure ..........................................................................126 Table 18.12 Toll Milling-Pulacayo Site Capital Expenditure ............................................127 Table 18.13 General and Administrative Capital Expenditure ..........................................127 Table 18.14 Tailings Storage Facilities-Capital Expenses (from EPCM Report)..............128 Table 18.15 Average Unit Operating Costs for Mining ($/t mined) ..................................129 Table 18.16 Process Plant Labour Costs (from EPCM Report).........................................129 Table 18.17 Process Plant Cash Operating Costs ..............................................................130 Table 18.18 Processing Costs - Toll Milling at Don Diego Mill .......................................130 Table 18.19 General and Administrative Costs .................................................................130 Table 18.20 Environmental and Social Operating Costs ...................................................131 Table 18.21 Project Base Case - LOM Cash Flow Summary ............................................132 Table 18.22 Project Base Case Production and Cash Flow Projection ..............................133 Table 18.23 Toll-Milling Option - LOM Cash Flow Summary.........................................136 Table 18.24 Toll Milling Option – Production and Cash Flow Projection (275 g/t Ag
Eq cut off) ......................................................................................................137 Table 19.1 Summary of Mineral Resources, Pulacayo Deposit.......................................140 vi
LIST OF FIGURES
Figure 1.1 Page
Base Case Cash Flow Summary ........................................................................7 Figure 1.2 Base Case Sensitivity Chart (NPV After tax) ....................................................7 Figure 1.3 NPV versus Cut-Off Grade for On-Site Milling ................................................8 Figure 1.4 Toll-Milling Option – Annual Cash Flow Summary .........................................9 Figure 4.1 Location Map, Pulacayo Project ......................................................................16 Figure 4.2 Outline of Mineral Concessions, Pulacayo Project ..........................................20 Figure 4.3 Drill-hole Locations Relative to the Property Outline, Pulacayo Project. .......21 Figure 6.1 Huanchaca Mining Company of Bolivia, circa 1890 .......................................25 Figure 6.2 Schematic Longitudinal Projection of the Silver Grades, Veta Tajo ...............26 Figure 7.1 Regional Geology of Bolivia ...........................................................................28 Figure 7.2 General Geology of the Pulacayo Area, Potosí District, Bolivia .....................29 Figure 7.3 Local Geology of the Pulacayo-Paca Area, Potosí District, Bolivia................31 Figure 7.4 Detail Geology of the Pulacayo Area, Potosí District, Bolivia ........................33 Figure 8.1 Epithermal Mineral Deposit Model .................................................................36 Figure 8.2 Alteration Mineral Distribution in a Low Sulphidation System ......................37 Figure 9.1 Drusy Vein Containing Sphalerite, Galena and Pyrite ....................................39 Figure 9.2 Example of a Massive Sulphide-Filled Vein, Pulacayo Deposit .....................40 Figure 9.3 Example of Veinlet and Disseminated Mineralization ....................................40 Figure 9.4 Example of a Quartz-Galena-Sphalerite-Filled Vein, Pulacayo Deposit .........41 Figure 9.5 Longitudinal View of the Stratigraphic Sequence, Pulacayo Deposit .............42 Figure 10.1 Topographic Survey Crew, Pulacayo Project ..................................................44 Figure 10.2 Induced Polarization Survey Coverage Area, Pulacayo Project ......................45 Figure 10.3 Induced Polarization Chargeability Results, Pulacayo Project ........................46 Figure 12.1 Bulk Density Determinations, Pulacayo Project, Bolivia ................................52 Figure 13.1 Sample Preparation Flowsheet, Pulacayo Project, Bolivia ..............................53 Figure 13.2 Particle Size Analyses of Exploration Samples, Pulacayo Project ..................54 Figure 14.1 General View of the Diamond Drilling Operation, Pulacayo Project ..............56 Figure 14.2 Comparison of Silver Check-Assay Results, Pulacayo Project .......................59 Figure 14.3 Comparison of Zinc Check-Assay Results, Pulacayo Project .........................60 Figure 14.4 Comparison of Lead Check-Assay Results, Pulacayo Project .........................60 vii
Figure 15.1 Location of the San Cristobal property ............................................................61 Figure 15.2 Schematic diagram showing San Vicente property .........................................62 Figure 16.1 Sample Preparation at Pulacayo Core Shack ...................................................65 Figure 16.2 Bench Flotation Cells at Universidad Técnica de Oruro .................................66 Figure 16.3 Open Circuit Float Test Parameters - Desliming Prior to Float -Test 4...........70 Figure 16.4 Silver Grade vs % Recovery without Desliming prior to Flotation .................72 Figure 16.5 Silver Grade vs % Recovery when 75% Silver Recovered from
Deslimed Clays ................................................................................................72 Figure 17.1 Vertical Longitudinal Projection of the Mined Out Areas as at 1945,
Pulacayo Project ..............................................................................................77 Figure 17.2 Selected Views of Digital Models of Historical Workings, Pulacayo
Project ..............................................................................................................78 Figure 17.3 Plan and Longitudinal Views of the Nominal $40/t NSR Solid ......................81 Figure 17.4 Cross Section 740300E Showing the Outline of the Nominal $40/t NSR
Domain Model .................................................................................................82 Figure 17.5 Contoured Silver Values for the Nominal $40/t NSR Domain Model,
Pulacayo Project ..............................................................................................83 Figure 17.6 Contoured Zinc Values for the Nominal $40/t NSR Domain Model,
Pulacayo Project ..............................................................................................84 Figure 17.7 Contoured Lead Values for the $40/t NSR Domain Model, Pulacayo
Project ..............................................................................................................84 Figure 17.8 Contoured NSR Values for the Nominal $40/t NSR Domain Model,
Pulacayo Project ..............................................................................................85 Figure 17.9 Silver Frequency Histogram for Samples within the Mineralized
Domain, Pulacayo Project................................................................................86 Figure 17.10 Zinc Frequency Histogram for Samples within the Mineralized Domain,
Pulacayo Project ..............................................................................................87 Figure 17.11 Lead Frequency Histogram for Samples within the Mineralized Domain,
Pulacayo Project ..............................................................................................87 Figure 17.12 Copper Frequency Histogram for Samples within the Mineralized
Domain, Pulacayo Project................................................................................88 Figure 17.13 Sample Length Histogram for Samples within the Mineralized Domain,
Pulacayo Project ..............................................................................................89 Figure 17.14 Specific Gravity Histogram for Samples Within the Mineralized
Domain, Pulacayo Project................................................................................90 viii
Figure 17.15 Longitudinal and Isometric Views of the Mineral Resources, Pulacayo
Project ..............................................................................................................97 Figure 18.1 Plan View of the Existing Development Working and the Planned Mine
Infrastructure ..................................................................................................100 Figure 18.2 Isometric View Looking North and Showing the Resource Model the
Planned Development and the Mined Out Areas...........................................101 Figure 18.3 Grade-tonnage Curve for Mineral Resource vs Silver Equivalent Cut-off
Grade ..............................................................................................................102 Figure 18.4 Town of Pulacayo as Viewed From the Mill Site ..........................................106 Figure 18.5 Pulacayo Flowsheet (from EPCM report) ......................................................108 Figure 18.6 Diagramatic location of Tailings Storage Facility (NTS) ..............................109 Figure 18.7 Starter Dam Design (from EPCM report) ......................................................110 Figure 18.8 Conceptual Plan View of the Tailings Storage Facility .................................111 Figure 18.9 Don Diego Mill, Crushing and Fine Ore Bin .................................................112 Figure 18.10 Don Diego Plant Flowsheet ...........................................................................113 Figure 18.11 NSR Value of Payable Metals .......................................................................124 Figure 18.12 Base Case Cash Flow Summary ....................................................................132 Figure 18.13 Base Case Sensitivity Chart (NPV After tax) ................................................134 Figure 18.14 NPV versus Cut-Off Grade for On-Site Milling ............................................135 Figure 18.15 Toll-Milling Option - Cash Flow Summary ..................................................136 Figure 18.16 Toll-Milling Option - Sensitivity ...................................................................138 Figure 18.17 On-Site Milling versus Toll-Milling Option ..................................................139 ix
LIST OF ABBREVIATIONS
Item
Apogee Minerals Limited
Apogee Minerals Bolivia S.A.
ASC Bolivia LDC
Apex Silver Mines Corporation
Corporación Minera de Bolivia
Golden Minerals Company
Micon International Limited
EPCM Consoltores S.R.L.
Resource Development Inc.
Universidad Técnica de Oruro
Abbreviation
AML
Apogee
ASC
ASMC
COMIBOL
GMC
Micon
EPCM
RDi
UTO
United States Dollar
Bolivian Bolivianos
Canadian Institute of Mining, Metallurgy and Petroleum
Canadian National Instrument 43-101
Atomic Absorption Spectroscopy
centimetre(s)
Degrees, Celsius
gram(s), kilograms, milligrams
grams per metric tonne
hectare(s)
hour
Internal Rate of Return
litre(s)
Life of mine
metre(s), centimetre, millimetre, kilometre
Net Present Value (discounted at rate %/y)
Net Smelter Return
Not available/applicable
Ordinary Kriging
Ounce (troy)
Parts per million, part per billion
Parts per
Percent(age)
Pound (avoirdupois)
Programmable Logic Controler
Quality Assurance/Quality Control
Reduced Level
Second (time)
Silver Equivalent grade
Specific Gravity
Sub-level open stoping
Système International d’Unités
Tailings storage facility
Tonne (metric), thousands, millions
tonne per day, tonne per year
Universal Transverse Mercator
Year(s)
$ or US$
BOB
CIM
NI 43-101
AAS
cm
o
,C
g, kg, mg
g/t
ha
h
IRR
L
LOM
m, cm, mm, km
NPVrate
NSR
n.a.
OK
oz
ppm, ppb
ppm
%
lb
PLC
QA/QC
RL
s
Ag Eq
SG
SLOS
SI
TSF
t, t 000, Mt
t/d, t/y
UTM
y
x
1.0
SUMMARY
Introduction
At the request of Mr. Joaquin Merino-Marquez, Exploration Manager of the wholly-owned
subsidiary of Apogee Minerals Ltd. (AML), Apogee Minerals Bolivia S.A. (Apogee), Micon
International Limited (Micon) has been engaged to perform a preliminary assessment of the
Pulacayo project and prepare a Technical Report in compliance with the requirements set out
in Canadian National Instrument (NI) 43-101.
Previously, Micon prepared technical reports for AML dated March, 2007 and December,
2008 and describing its resource estimates on the Paca and Pulacayo properties, respectively.
Micon understands that: Apogee, under an agreement dated March 8, 2006, acquired the right
to earn a sixty percent (60%) interest in both the Paca and Pulacayo properties from ASC
Bolivia LDC (ASC), a subsidiary of Apex Silver Mines Corporation (ASMC) which had
previously conducted exploration on the properties; effective March 24, 2009, Golden
Minerals Company (GMC) became the successor to the assets of ASMC (renamed Golden
Service Company).
On January 26, 2010, AML announced that it had entered into a non-binding term sheet (the
"Term Sheet") with GMC to acquire the Pulacayo Deposit. Upon completion of the proposed
transaction, AML will be able to acquire a 100% interest in the property. Pursuant to the
Term Sheet, the Company would acquire all of the issued and outstanding shares of a
Cayman based company that is a wholly-owned subsidiary of GMC, which indirectly holds a
100% interest in the Pulacayo deposit. In consideration, AML would issue 5,000,000
common shares in AML upon closing of the transaction and an additional 3,000,000 common
shares in AML plus a cash fee in the amount of $500,000 eighteen (18) months following
closing of the transaction. Completion of the acquisition is subject to negotiation and
execution of a definitive agreement, necessary board approvals and receipt of all required
regulatory and securities approvals, including the approval of the TSX Venture Exchange,
along with other customary closing conditions.
Geology & Resources
The Pulacayo epithermal deposit is hosted by sedimentary and igneous rocks of Silurian and
Neogene age. The sedimentary rocks are composed of diamictites, sandstone and shale. The
Neogene-aged rocks are mostly of volcanic-sedimentary origin and are composed of
conglomerate, sandstones, reddish conglomerates, reddish-brown clay, whitish rhyolite tuff,
andesite lava flows, dacitic-rhyolite domes and andesite porphyry.
The principal mineralized structure at Pulacayo is known as “Veta Tajo”, which was
historically the main silver producer in the Pulacayo mine. The Veta Tajo is part of a larger
structural system that is oriented approximately east-west and dips 75° to 90° south. The
1
width of this vein varies from less than 1 metre to several metres. The structure is filled with
quartz, barite, pyrite, sphalerite, galena and silver sulpho-salts.
Apogee has carried out detail geological mapping and sampling at surface and in the old
underground workings, followed up by a topographic survey, geophysical survey, and
diamond drilling. Between January, 2006 and September, 2009, four phases of drilling were
carried out.
A simple, upright, whole-block model with the long axis of the blocks measuring 10 m
(strike) x 10 m (height) x 2 m (width) and oriented along an azimuth 100° was constructed
using the Gemcom-Surpac version 6.1.1 mine planning software package.
Micon then carried out a geostatistical analysis of the deposit using the results of this drilling.
Taking account of the topographic mapping, plans and sections pertaining to the extent of
previous mine workings, trend analysis, metal prices and potential metallurgical recoveries,
Micon then prepared an estimate of the mineral resource.
Mineral resources reported from within the mineralized domain are given in Table 1.1. The
effective date of this estimate is October 14, 2009.
Table 1.1
Summary of Mineral Resources, Pulacayo Deposit
Classification
Indicated
Inferred
Tonnes
4,892,000
6,026,000
Ag (g/t)
79.96
98.26
Pb (%)
0.79
0.78
Zn (%)
1.64
1.68
(1) Tonnages have been rounded to the nearest 1,000 tonnes. Average grades may not sum due to rounding.
(2) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of
mineral resources may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical,
marketing, or other relevant issues.
(3) The quantity and grade of reported inferred resources in this estimation are conceptual in nature and there has been
insufficient exploration to define these inferred resources as an indicated or measured mineral resource. And it is
uncertain if further exploration will result in upgrading them to an indicated or measured mineral resource category.
Mining
Sub-level open stoping (SLOS) with backfill is the mining method which Micon considers
most suitable for underground mining at Pulacayo. The average value of the resource justifies
the use of backfill as opposed to leaving pillars in-situ. SLOS mining with backfill also gives
a reduced risk of surface subsidence. SLOS is a more productive method, even in relatively
narrow stopes, when compared to cut and fill mining.
The mine is accessible through the San Leon Adit (4,130 m RL), which has a nominal arched
profile of 2.2 m (high) by 2.0 m (wide). It is Micon’s opinion that the most efficient method
of access and ore haulage is through the adit.
Two new inclined ramps and two decline ramps are planned. They will access the ore above
and below the 4130 m level, respectively. The inclined ramps will be developed from the
2
enlarged San Leon adit, starting from the FW to the south of the ore body. The decline ramps
will be developed from the enlarged San Leon adit.
Ventilation air will be exhausted through multiple vent raises to surface. Main fans will be
located on the surface end of each raise. Intake air will be drawn in through the north and
south ends of the San Leon adit. This will ensure that both the primary and emergency means
of egress are situated in intake air.
For mine planning purposes, a silver-equivalent cut-off grade of 200 g/t Ag Eq. was selected.
Silver equivalent grades were calculated using metal prices $14.66/oz for silver, $0.98/lb for
lead and $1.05/lb for zinc. The mineable portion of the mineral resources considered in this
preliminary assessment are as given in Table 1.2, and have been modified to account for
estimated mining losses and mining dilution.
Table 1.2
LOM Production Forecast at a Cut off Value of 200 g/t Ag Eq.
Class
Indicated
Inferred
Resource
t 000
1,793
2,456
Ag
g/t
143.4
162.2
Pb
%
1.0
1.0
Zn
%
2.1
1.9
Ag
Metal
kg
257,000
398,300
Pb
Metal
t’000
18.83
25.30
Zn
Metal
t 000
36.94
47.40
Mineral resources which are not mineral reserves do not have demonstrated economic viability.
Processing
The metallurgical testwork shows that a lead and zinc concentrate can be produced using a
conventional flotation flowsheet. The base case for this preliminary assessment considers a
milling capacity averaging 1,800 t/d over 360 d/y. Concentrate grades and recovery at the
forecast headgrade are shown in Table 1.3
Metallurgical testwork at the UTO laboratory was completed in February, 2010, testing low
medium and high grade composite samples. The medium grade assayed 181g/t silver, 0.69%
lead, and 2.45% zinc, approximating the average mill head grade forecast of 154.2 g/t silver.
Table 1.3
Concentrate Grades and Recovery at Forecast Average Head Grade
Product
Mill Feed
Lead concentrate
Zinc concentrate
Tailings
Mass
Yield
dmt/d
1800
29
59
1713
Grade
Percent Recovery (%)
Ag g/t
%Pb
%Zn
Ag
Pb
Zn
154.2
6220
873
28.5
1.0
51.0
0.85
0.22
2.0
3.72
53.0
0.19
100.0
63.9
18.6
17.6
100.0
77.6
2.7
19.7
100.0
3.0
87.7
9.3
There is a significant amount of clay generating rock in the material that negatively affects
concentrate grades and recoveries, and that could negatively affect reclaim water clarity and
3
deposition density in the Tailings Storage Facility. Further metallurgical tests for the clay
fraction are recommended for later stages of project development, to guide refinements to the
process flowsheet and equipment selection.
There are four deleterious elements in the Pulacayo material that will affect the lead
concentrate net smelter returns and so should be recalculated in future models. In the
medium grade lead concentrate, the concentration of these elements were arsenic (3,940 g/t),
copper (2.80%), antimony (3.44%), and zinc (4.19%).
The toll-milling option considered as an alternative scenario in the study was based on
processing of the Pulacayo material off site at the Don Diego lead/zinc mill located 40 km
east of Potosi. Although toll milling scenarios using higher cut-off grades were found to be
economic, the results were found to be inferior to the base case (with on-site milling) at all
cut-offs tested in the study up to 275 g/t Ag equivalent. The results suggest that toll milling
could be regarded as a potential fall-back position should on-site milling not be possible for
any reason.
Infrastructure
For the base case, a new Tailing Storage Facility (TSF) will be required at Pulacayo. A
capital cost estimate for this has been prepared on the basis of a conceptual layout. No design
has yet been prepared for the TSF.
Existing power supply is inadequate and will require upgrading. The option selected is to tie
into the San Cristobal-Punutuma 220 kV transmission line, the closest point to which is 10
km from Pulacayo.
The base case assumes that the current water supply pipeline serving Pulacayo town will
remain operational and can be used to supply the mine. Thus, the estimated cost of replacing
this line (US$ 1.41 million) has not been included in the base case.
Environment and Social
The project area is affected by historical mine workings that are causing acidic drainage and
metal contamination to enter the surrounding air and water. Project development needs to
take these into consideration. New development must consider whether operations can be
isolated from historical contaminants, be integrated with historical works to clean up some of
the contamination, or historical contaminants can be remediated prior to new development.
Regardless, definition and documentation of historical contamination is important so that the
company can manage its risks.
Potential social effects on Pulacayo include increased income from direct employment,
increased demand for local services and suppliers, and an influx of workers. Potential
adverse social effects and pressures on housing and infrastructure will need to be effectively
managed. A community development plan should be developed in concert with the
4
community to help ensure economic benefits are realized in the community. Social and
health effects should be considered for any small-scale miners who are still active in the
Pulacayo area.
Economics
The economics of the Pulacayo project have been assessed under two scenarios:

In the first scenario, which forms the base case for this report, a processing facility is
built on-site to treat the material mined from underground, and concentrates of zinc
and lead are produced for shipment to port. Silver credits are obtained for both
products.

Alternatively, no processing facility is constructed and, instead, the mine ships ROM
material to the Don Diego process plant, where it is toll treated. This scenario is
treated as a sensitivity case, the cash flow from which is then compared to the base
case.
In each case, production rates and all other assumptions are kept the same to allow the
relative value of each scenario to be determined.
Table 1.4 shows a summary of the capital required for the base case.
Table 1.4
Summary of Base Case Capital Expenditure
Capital Cost Summary
Initial
US$ (M)
18.48
27.77
3.00
2.53
2.00
15.63
69.41
Mining
Processing
Tailings
Infrastructure & indirect
Environmental & Social
Contingency
Total
Base case cash operating costs are given in Table 1.5.
Table 1.5
Summary of Base Case Operating Costs
Operating Cost Summary
Mining
Processing
General & Administrative
Total
5
US$/t
22.60
12.77
2.33
37.70
Sustaining
US$ (M)
15.06
0.15
4.35
2.22
5.87
27.65
Table 1.6 presents the project base case LOM cash flow summary, and Figure 1.1 provides a
summary of the main components of the annual cash flow for the base case.
Table 1.6
Project Base Case - LOM Cash Flow Summary
LOM
($ 000)
178,537
203,519
382,056
15,282
366,774
$/t
treated
42.02
47.90
89.92
3.60
86.32
$/oz Ag
11.81
13.46
25.27
1.01
24.26
NPV8
(2010)
115,503
131,665
247,168
9,887
237,281
Operating Costs
Mining costs
Processing costs
General & Administrative costs
Total cash operating cost
96,027
54,257
9,902
160,186
22.60
12.77
2.33
37.70
6.35
3.59
0.65
10.59
62,500
35,115
6,396
104,011
Net Operating Margin
206,587
48.62
13.66
133,270
Capital Expenditure
97,060
22.84
6.42
83,268
Pre-tax Cash Flow
109,527
25.78
7.24
50,002
Taxation
29,023
6.83
1.92
17,013
Net Cash Flow After Tax
80,504
18.95
5.32
32,988
NSR Silver only
NSR Co-products
NSR value
less Royalty
This preliminary assessment is preliminary in nature; it includes inferred mineral resources
that are considered too speculative geologically to have the economic considerations applied
to them that would enable them to be categorized as mineral reserves, and there is no
certainty that the preliminary assessment will be realized.
On a pre-tax basis, at a discount rate of 8 %/y, the base case cash flow evaluates to a net
present value (NPV8) of $50.0 million, and has in internal rate of return (IRR) of 24.0%.
After tax, the NPV and IRR are estimated to be $33.0 million and 19.6%, respectively.
Payback on the undiscounted cash flow after tax occurs in Year 3 and, over the life-of-mine
(LOM) period, the net cash flows before and after tax are $109.5 and $80.5 million,
respectively.
On an annual basis, the estimate of maximum funding required before positive cash flow is
$89.5 million.
6
Figure 1.1
Base Case Cash Flow Summary
80
Net Cash Flow
60
Royalty
40
Taxation
Working Capital
0
Capital
(20)
Opcosts
Yr9
Yr8
Yr7
Yr6
Yr5
Yr4
Cum C/Flow
Yr3
(80)
Yr2
Cum DCF
Yr1
(60)
Yr‐1
Net Revenue
Yr‐2
(40)
Yr‐3
USD million
20
The sensitivity of the base case cash flow after tax to changes in product pricing, operating
costs and capital expenditure is shown in Figure 1.2.
Figure 1.2
Base Case Sensitivity Chart (NPV After tax)
100
80
NPV (8%) USD million
60
40
20
0
(20)
70 75 80 85 Product Price (21.4 (12.1 (2.9)
6.1
90 95 100 105 110 115 120 125 130 15.1 24.1 33.0 41.9 50.8 59.7 68.5 77.4 86.2
Opcosts
56.9 52.9 49.0 45.0 41.0 37.0 33.0 29.0 25.0 21.0 17.0 13.0
9.0
Capex
57.4 53.3 49.2 45.2 41.1 37.1 33.0 28.9 24.9 20.8 16.7 12.7
8.6
7
Micon evaluated the base case (on site milling) using a series of cut-off grades to determine
the optimum grade/tonnage combination for the project. The results (Figure 1.3) show that
project NPV and IRR are maximized when applying a cut-off grade of 200 g/t Ag Eq. Micon
therefore selected this value of the cut-off grade for its base case economic assessment of the
project in this study.
35.0
35.0%
30.0
30.0%
25.0
25.0%
20.0
20.0%
15.0
15.0%
10.0
10.0%
IRR (%)
NPV ($ million) | Minng Cost ($/t)
Figure 1.3
NPV versus Cut-Off Grade for On-Site Milling
5.0%
5.0
0.0%
.0
125
150
175
200
225
Cutoff grade (g/t Ag Eq)
NPV
Mining $/t
250
275
IRR(%)
The study also considered an alternative to the on-site milling of material mined at Pulacayo.
For this purpose, it was assumed that crushed material was taken by road to Uyuni and
thence by rail to the Don Diego mill for toll treatment. Savings in the process plant and
tailings dam construction costs result in a reduction of approximately $33.1 million in capital
invested before positive cash flow, with $56.4 million required for toll-milling compared to
$89.5 million in the base case. Nevertheless, Table 1.7 and Figure 1.4 show that, although
payback on the undiscounted cash flow occurs in Year 4, the LOM net cash flow after tax of
$43.1 million is $37.4 million less than is forecast in the base case ($80.5 million).
Moreover, the toll-milling option does not appear to maximize project value since, at a cutoff
of 200 g/t Ag Eq, its NPV8 of $14.5 million is $15.5 million less than the base case ($33.0
million). Even at a cut-off of 275 g/t Ag Eq, the after-tax NPV8 of $16.4 million for the toll
milling option is $1.2 million less than for the base case ($17.6 million) at that cut-off.
Nevertheless, because of the reduction of capital, at cut-off grades above 250 g/t Ag Eq, toll
milling appears to offer an improved internal rate of return, with IRR of 19.4% and 20.7%
after tax at 250 g/t and 275 g/t respectively, compared to rates of 18.6% and 17.6%
respectively in the base case.
8
Table 1.7
Toll-Milling Option - LOM Cash Flow Summary
Using 200 g/t Ag Eq Cutoff
LOM
($ 000)
178,537
203,519
382,056
15,282
366,774
$/t
treated
42.02
47.90
89.92
3.60
86.32
$/oz
Ag
11.81
13.46
25.27
1.01
24.26
Operating Costs
Mining costs
Processing costs
96,027
145,486
9,902
251,415
22.60
34.24
2.33
59.17
6.35
9.62
0.65
16.63
62,500
94,121
6,396
163,017
115,359
27.15
7.63
74,264
Capital Expenditure
56,374
13.27
3.73
50,396
Pre-tax Cash Flow
58,985
13.88
3.90
23,868
Taxation
15,878
3.74
1.05
9,339
43,107
10.15
2.85
14,529
NSR Silver only
NSR Co-products
NSR value
less Royalty
NPV8 (2010)
($ 000)
115,503
131,665
247,168
9,887
237,281
Figure 1.4
Toll-Milling Option – Annual Cash Flow Summary
80
Net Cash Flow
60
Royalty
40
Taxation
Working Capital
0
Capital
(20)
Opcosts
(40)
Net Revenue
(60)
Cum DCF
9
Yr7
Yr6
Yr5
Yr4
Yr3
Yr2
Yr1
Yr‐1
Yr‐2
(80)
Yr‐3
USD million
20
Cum C/Flow
Interpretation and Conclusions
The project base case comprises the development of an underground mine connecting to
existing workings though a new adit portal, extraction using a sub-level open-stoping method
with backfill, feeding 1,800 t/d to a new milling and flotation plant on site, for the production
and sale of lead and zinc concentrates containing economically important silver values, and
storage of flotation tailings in a new, purpose-built facility adjacent to the new plant.
The preliminary assessment of this base case shows it to be economic, with an IRR of 24%
and NPV8 of $50.0 million before tax. Payback is in Year 3, leaving one further year of full
production.
An alternative scenario, with toll-milling of the underground mine production at the Don
Diego mill, is also shown to be potentially economic, albeit at higher cut-off grades. This
option has a reduced capital requirement, resulting in an improved IRR before tax of 27.5%,
although the NPV8 is lower ($16.4 million after tax).
Recommendations
A complete list of recommendations is given is Section 20. Micon recommends, inter alia:

Analysis of duplicate samples as part of the Quality Control program should be
carried out at a laboratory that is a separate corporate entity from the laboratory that
conducted the primary analyses.

Detailed modeling of the narrow higher grade vein structures, should a local estimate
of the amount of material amenable to underground mining be needed. In support of
this local estimate, additional information in the form of in-fill drilling should be
obtained.

The position, shape and content of the mined out stopes, and the position and
geometry of the existing development should be determined by appropriate methods
to an appropriate degree of accuracy as project development advances.

In consideration of the range of specific gravities observed in the sample data, Micon
recommends that should the project proceed to a more advanced state, additional
density measurements should be taken from samples chipped from the walls of the
existing mine workings to assist in filling in the gaps in the spacing of the
information. The density of each block in the model should be estimated so as to
provide a more accurate local estimate of tonnage. Care will need to be taken in
order to obtain an accurate specific gravity measurement for samples that are porous.

A program of geotechnical characterization of the wall rocks should be carried out in
support of mine design.
10

A detailed geotechnical study should be carried out, that will provide the basis of
more detailed mine planning.
With respect to metallurgy:

Complete pressure-filter moisture tests on the lead and zinc concentrates to confirm
concentrate moistures will be less than 8 wt%. This is required for transport by ship.
If this is limit is not attainable, a disk-filter, gas fired dryer or an atmospheric drying
pad may be required.

Review and modify the flowsheet and equipment selection to maximize silver
recovery from the clay fraction.

Bench tests should be completed to determine how the clay fraction responds in the
TSF, both for reclaim water clarity and deposition density for TSF volume
calculations.

For the toll milling option, recalculate the ore transport charges from the Pulacayo
mine to the Don Diego mill, after the road improvements to Highway 701 are
completed, which should be around the first quarter of 2011 (see description in
Section 18.3.4). The direct trucking of ore on this route would significantly reduce
the transport charges, since Highway 701 is the most direct route between Pulacayo
and Don Diego.
Environmental and Social Considerations:

Waste rock should not be used for construction.

The waste rock and tailings disposal design and water management plans need to
consider the acid generating and metal leaching properties of the waste rock and
tailings.

It is recommended that the impact assessment further document the extent of
historical contamination.

It is recommended that further project design take historical works into consideration
and remediate historical contaminants where possible.

Community consultation should continue and a Community Development Plan be
developed in concert with the community.
With respect to Project Development:

Project exploration and development should proceed together. The base case would
be significantly strengthened by additional mineral resources to extend the LOM
further beyond the payback period.
11

The toll-milling scenario remains attractive while resource tonnage is limited – this
can therefore be viewed as a fall-back scenario should exploration meet with only
limited success in locating additional resources.
12
2.0
INTRODUCTION AND TERMS OF REFERENCE
At the request of Mr. Joaquin Merino-Marquez, Exploration Manager of the wholly-owned
subsidiary of Apogee Minerals Ltd. (AML), Apogee Minerals Bolivia S.A. (Apogee), Micon
International Limited (Micon) has been engaged to perform a preliminary assessment of the
Pulacayo project and prepare a Technical Report in compliance with the requirements set out
in Canadian National Instrument (NI) 43-101.
Previously, Micon prepared technical reports for AML dated March, 2007 and December,
2008 and describing its resource estimates on the Paca and Pulacayo properties, respectively.
Micon understands that: Apogee, under an agreement dated March 8, 2006, acquired the right
to earn a sixty percent (60%) interest in both the Paca and Pulacayo properties from ASC
Bolivia LDC (ASC), a subsidiary of Apex Silver Mines Corporation (ASMC) which had
previously conducted exploration on the properties; effective March 24, 2009, Golden
Minerals Company (GMC) became the successor to the assets of ASMC (renamed Golden
Service Company).
On January 26, 2010, AML announced that it had entered into a non-binding term sheet (the
"Term Sheet") with GMC to acquire the Pulacayo Deposit. Upon completion of the proposed
transaction, AML will be able to acquire a 100% interest in the property. Pursuant to the
Term Sheet, the Company would acquire all of the issued and outstanding shares of a
Cayman based company that is a wholly-owned subsidiary of GMC, which indirectly holds a
100% interest in the Pulacayo deposit. In consideration, AML would issue 5,000,000
common shares in AML upon closing of the transaction and an additional 3,000,000 common
shares in AML plus a cash fee in the amount of $500,000 eighteen (18) months following
closing of the transaction. Completion of the acquisition is subject to negotiation and
execution of a definitive agreement, necessary board approvals and receipt of all required
regulatory and securities approvals, including the approval of the TSX Venture Exchange,
along with other customary closing conditions.
In the present study, Micon has performed a preliminary assessment of the potential for
underground mining of the higher grade portions of the Pulacayo resource. It is reasoned that
limiting the scope of the study to this aspect will allow a baseline evaluation to be
established, against which any additional value to be gained from a larger-scale open-pit
mining operation can be measured at a later date. Accordingly, this study does not consider
any potential open-pit mining.
The study does consider an important trade-off, between (i) the construction of a new
treatment plant and tailings storage facility near the mine and (ii) contracting out processing
of the resource to a toll milling plant. In either case, the product of this processing is assumed
to be concentrates which will be sold to a third party for further processing, and hence project
revenues are the net smelter return, after deduction of concentrate transport costs.
13
The study has been carried out by Micon personnel, utilizing technical information warranted
by the client and which it has reviewed and found to be reasonable and appropriate, within
the +/-30% level of accuracy expected of a scoping study. Mr. Reno Pressacco P.Geo.,
Micon’s senior geologist at that time, conducted a site visit to the Pulacayo project area
between March 26 and 29, 2008, while drilling was in progress. Mr. Pressacco was
responsible for preparation of the mineral resource estimate upon which this preliminary
assessment is based. Micon’s senior metallurgist, Mr. Michael Godard P.Eng., and senior
mining engineer, Mr. Geraint Harris, CEng., visited the project area between July 29 and 30,
2009, and between August 6 and 7, 2009, respectively. They were able to hold discussions
with Apogee personnel on site and make an independent assessment of the project area and
associated infrastructure before preparing, respectively, the processing and mining sections
of this preliminary assessment.
Unless otherwise indicated, all currency amounts are stated in United States dollars ($ or
US$) or Bolivian Bolivianos (BOB). For the 12 months ending 31 March, 2010, the average
rate of exchange was approximately BOB 7.17/US$. The project has employed the metric
system of measurement, consequently weight will be expressed in metric tonnes (tonnes),
frequencies in Hertz (Hz), distance in metres (m) or kilometres (km), area in hectares (ha)
and silver values in grams per metric tonne (g/t Ag). In some cases, equipment is sized in
imperial lengths (feet or inches), horsepower (hp) and kilowatt hours per short ton (kWh/st).
14
3.0
RELIANCE ON OTHER EXPERTS
Micon has reviewed and evaluated the data pertaining to the mineralization found on the
Pulacayo project located in Pulacayo Township, Potosí District, Bolivia that was provided to
it by Apogee and its consultants, and has drawn its own conclusions therefrom. Micon has
not carried out any independent exploration work, drilled any holes or carried out any
sampling and assaying other than described in this report.
While exercising all reasonable diligence in checking, confirming and testing it, Micon has
relied upon the data presented by Apogee, and found in public domain documents in
conducting its technical review. Micon is pleased to acknowledge the helpful cooperation of
Apogee’s management including Mr. Joaquin Merino Marquez, all of whom made available
any and all data requested, and responded promptly, openly and helpfully to all questions,
queries and requests for material.
The status of the mineral concessions under which Apogee holds title to the surface and
mineral rights for these properties has not been investigated or confirmed by Micon, and
Micon offers no opinion as to the validity of the mineral title claimed by Apogee. The
description of the property, and ownership thereof, as set out in this report, is provided for
general information purposes only. As well, the substance of the various option agreements
has not been investigated or confirmed by Micon, and Micon offers no opinion as to the
validity of the terms set out therein. The essential terms of these agreements outlined in this
report are provided for general information purposes only.
Micon has relied on the estimate of environmental remediation (mine closure) costs provided
in a report prepared by EPCM Consoltores S.R.L. of Bolivia (EPCM) for Apogee, Micon
understands that EPCM has relevant local experience in the requirements of Bolivian
environmental legislation as it pertains to mine closure.
15
4.0
4.1
PROPERTY DESCRIPTION AND LOCATION
LOCATION
The Pulacayo prospect is located 18 km east of the city of Uyuni (Canton of Pulacayo,
Quijarro Province) in the Department of Potosí in southwestern Bolivia, 460 km southsoutheast of the capital city, La Paz, and 130 km southwest of Potosí, the department capital
(Figure 4.1).
Figure 4.1
Location Map, Pulacayo Project
Pulacayo is accessible by paved and good gravel highways from La Paz via Oruro (560 km),
and by good gravel road from Potosí (189 km). Unpaved sections are generally navigable the
whole year although they may present some level of difficulty during the rainy season.
The tourist town of Uyuni, on the edge of the large Salar de Uyuni (salt lake) provides
limited local services. It has railway connections with the cities of Oruro, Potosí, Villazon,
and to the borders with Argentina and Chile. Uyuni has a small gravel airstrip which permits
the operation of light aircraft. There are several small hotels, hostels, restaurants, schools,
16
medical and dental facilities and internet cafes. Apex’s San Cristóbal Mining Company has
constructed a gravel road from San Cristóbal, approximately 100 km southwest of Uyuni, to
the border with Chile.
4.2
PROPERTY STATUS
4.2.1
Overview of Bolivian Mining Law
The granting of mining concessions in Bolivia is governed by the Constitution (Constitución
Política del Estado), the Mining Code (Código de Minería) supplemented by certain Supreme
Decrees that rule taxation, environmental policies, administrative matters, and the like.
Rights to mineral resources, which are fundamentally the property of the Bolivian state, can
be granted for their exploitation but the Bolivian state is prohibited from transferring title to
them, according to Article 136 of the Constitution.
Bolivian companies, foreign companies or individuals, with the exception of minors,
government agents, armed forces members, policemen, or their relatives, may own mining
concessions. Foreigners, pursuant to Article 25 of the Constitution and Article 17 of the
Mining Code, are not authorized to own mining concessions or real estate property within a
buffer zone of 50 km surrounding the Bolivian international borders, but they may enter into
joint venture agreements on the frontier regions.
In March, 1997, Bolivia enacted Law No. 1777 to revise its CODIGO DE MINERIA
(Mining Code), to promote private ownership of mineral properties and to enable
COMIBOL, the state-owned mining corporation, to lease or joint venture mineral properties
which are subject to state-owned mineral leases.
The Codigo de Mineria (1997) is available in an official Spanish-English side-by-side
version which facilitates understanding the Bolivian mining code. Key features are:

There is only one type of mining license, a “La Concesion Minera”, which is
comprised of 25 ha units, named “cuadricula minera”. A maximum of 2,500 units is
allowed for a mining concession. There is no limitation to the number of concessions
that can be held by a company or an individual.

Field staking is not required; concessions are applied for on 1:50,000 scale base
maps.

The owner of the concession has exclusive rights to all minerals within the
concession.

Annual rents, payable in January of each year, are BOB 9/ha in the first 5 years and
BOB 18/ha thereafter (approximately $1.00/ha and $2.00/ha, respectively).

If the title holder continues to make the “patentes payment” on time the term of the
mining concession is indefinite.
17

Mining concessions cannot be transferred, sold or mortgaged.

Provision is made for surface access, compensation and arbitration with private land
owners, if any. (NB: private ownership of surface lands outside of major cities is
limited).

Historical mining concessions, 1 ha “pertenencia minera”, applied for and granted
according to the system governed by the old, pre-1967, Mining Code remain valid if
the owners have complied with the “Catastro Minero”, an obligatory registration of
the mining concessions that existed prior to the implementation of the new Mining
Code. This registration involves the legal audit of the titles and the verification of the
technical information of the mining concessions, to be included in a digital format on
the database of the Bolivian National Service of Geology and Technical of Mines
(SERGEOTECMIN).

Mining concessions, both “cuadrículas” and “pertenencias” must have their “Título
Ejecutorial” registered with the “Mining Registry” that is part of the
SERGEOTECMIN and before the Real State Registration Office.

Simultaneous with the introduction of the new mining code in 1997 were a number of
taxation reforms. Bolivian taxes are now fully deductible by foreign mining
companies under US corporate income tax regulations.
Taxes applicable are:

Mining Royalty (Regalía Minera) equivalent to 1-7% of the gross sales value of the
mineral. The tax is paid before the mineral is exported or sold in the local market (in
this case only 60% of the tax is paid).

Profits tax of 25% on net profits [Gross income – (expenses+costs)]; losses can be
carried forward indefinitely. An additional 12.5% is paid when metals/minerals reach
extraordinary market prices.

Mineral production is subject to a Value Added Tax of 13%.
The Ministry of Mining and Metallurgy is responsible for mining policy. Servicio Geologico
y Tecnico Minero de Bolivia (SERGEOTECMIN) – the Bolivian Geological Survey, a
branch of the Ministry, is responsible for management of the mineral titles system.
SERGEOTECMIN also provides geological and technical information and maintains a
USGS-donated geological library and publications distribution centre. Also, tenement maps
are available from SERGEOTECMIN, which has a GIS based, computerized map system.
Exploration and subsequent development activities require various degrees of environmental
permits, which various company representatives have advised are within normal international
standards. Permits for drill road construction, drilling and other ground disturbing activities
18
can be readily obtained in 2 – 4 months, or less, upon submission of a simple declaration of
intent and plan of activities.
4.2.2
Project Ownership
Details of ownership of the Pulacayo project properties are complicated by multi-layered
option and joint venture agreements. Apogee’s Option/JV agreement with Apex’s Bolivian
subsidiary, ASC Bolivia LDC (Agreement I), allows Apogee to earn interests in two earlier
agreements that ASC had established with the Pulacayo Mining Cooperative and COMIBOL
(Agreements II, III). A brief summary of this information, as provided by Apogee’s legal
counsel, is given as Appendix I to this report.
The project’s environmental requirements have been completed in compliance with the
Environment Law (Law Nº 1333) and the Environmental Regulation for the Mining
Activities. A certificate of exemption has been obtained for the exploration phase.
An audit of the Environmental Base Line (ALBA) was carried out between December, 2007
and July, 2008 by Mining Consulting & Engineering “MINCO S.R.L.” Its audit report
summarizes the work carried out during the Environmental Assessment, and includes:

A compilation of information on the local vegetation, animals, soil, water, air, etc.
More than 500 samples collected in the area of interest support the conclusions and
recommendations of the report;

An evaluation of the social impact of the project;

An evaluation of the area contaminated during previous mining activities, including
tailings, abandoned facilities, acid waters, scrap, etc;

An evaluation of other environmental liabilities.
The location of the various concessions comprising the property holdings for the Pulacayo
project is presented in Figure 4.2. The location of the drill-holes and the Pulacayo deposit
relative to the concession boundaries are shown in Figure 4.3.
19
Figure 4.2
Outline of Mineral Concessions, Pulacayo Project
N
20
Figure 4.3
Drill-hole Locations Relative to the Property Outline, Pulacayo Project.
Paca Deposit
Drill-holes shown by black triangles.
21
5.0
5.1
ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE
AND PHYSIOGRAPHY
ACCESS
Bolivia, is the highest and most isolated country in South America, with diverse geographic
and climatic conditions ranging from snow capped peaks and high altitude plateaus to vast,
low-lying grasslands and rainforests. Climate and geography have influenced population
settlement, mineral discovery and subsequent development of transportation and
infrastructure.
Bolivia is landlocked, without direct seaport access. A reasonably well developed rail system
exists with connections south to Argentina, east to Brazil and west to Chile and the port of
Antofagasta. Rail service from Uyuni connects with Oruro, Atocha, Tupiza, and Villazon
(on the border with Argentina). Uyuni is also connected by railway to Chile through
Estación Abaroa. Disused rail lines exist between Uyuni-Potosí and Oruro-La Paz.
International air travel to Bolivia is via Miami (American Airlines), Mexico City, Brazil,
Chile (LAN), Argentina and Peru (Taca). Bolivian airlines AeroSur fly regular internal
flights between major cities, with three flights a week to Uyuni city. In 2008, ASC began
constructing a lighted airstrip at the San Cristóbal Mine, approximately one hour’s drive to
the west on a well maintained gravel road, with completion expected in 2012.
The principal highways are generally of a quite reasonable standard; heavy trucks and buses
dominate the road traffic outside of the major cities and, for the most part, road freight
service functions adequately even to small remote villages. However, secondary roads can be
best described as “tracks” and winding, single lane roads are often precariously carved out of
steep slopes.
The Pulacayo project is accessed from La Paz by means of a paved road, which runs to the
area of Huari, passing through Oruro. It can also be accessed by the road between Oruro
(gravel) and Potosí (paved) and from Potosí to Uyuni by a good quality gravel road. Paving
of the road from Potosí to Uyuni began in 2007 and was scheduled for completion in 2010.
5.2
CLIMATE AND PHYSIOGRAPHY
Two Andean mountain chains run through western Bolivia, with many peaks rising to higher
than 6,000 m. The western Cordillera Occidental Real forms the western boundary with Peru
and Chile; the Cordillera Oriental Real runs southeast from Lake Titicaca then turns south
across central Bolivia to join with the Cordillera Central along the southern border with
Argentina. Between these two mountain chains is the Altiplano, a high plain with an altitude
of 3,500 – 4,000 m. East of the Cordillera Central is a lower altitude region of rolling hills
and fertile basins with a Mediterranean climate. To the north, the Andes fall away into the
tropical lowlands of the Amazon Basin in Brazil.
22
Climate within Bolivia is altitude related. The rainy period lasts from November to March
(summer). Of the major cities, only Potosí receives regular snowfalls between February and
April at the end of the rainy season, although La Paz and Oruro do occasionally receive light
snow flurries. On the Altiplano and in higher altitude areas, sub-zero temperatures are
frequent at night throughout the year. Snow capped peaks are present year round at
elevations greater than 5,250 m.
The Pulacayo project area is located within the Fourth Geomorphological Province of Bolivia
(Eastern Andes), immediately to the southwest of the Cosuño Caldera. Topographical relief
is gentle to moderate, with elevations between 4,000 and 4,500 m amsl. The Paca and
Pulacayo Domes stand out as topographic highs.
Major valleys are shaped by permanent rivers fed by water draining the snow capped
volcanic peaks, such as the Cosuño Caldera, and thermal hot springs scattered about the same
volcanic complexes. The principal river is the Pucamayu (Río Rojo), which empties directly
into the Salar de Uyuni. A small tributary, the Irpa Mayu River, collects runoff from local
gorges like the Pacasmayo, Phusamayu, and others, and joins the Totora Mayu to form the
Capilla River which eventually joins the regional Pucamayu.
Pulacayo has a semi-arid climate, with annual rainfall around 100 mm and mean summer
temperatures of 12°C between October and March. Winters are cold with minimum
temperatures of -26°C and maximums of 18°C between June and July; mean temperature is
5.5°C. In general, the monthly annual mean temperatures range from a minimum of 1.5°C to
a maximum of 21.5°C (Source: SENAMHI interpolated from nearby stations:
www.senamhi.gov.bo and response to a request by Apogee).
5.3
LOCAL RESOURCES AND INFRASTRUCTURE
Bolivia is a natural resources rich country with a significant production history of silver and
tin and secondary production of gold, copper, antimony, bismuth, tungsten, sulphur and iron.
To the south and east, sizeable reserves of natural gas exist but their development and export
is a contentious national issue, exacerbated by the absence of a seaport.
The country has an abundance of hydroelectric power and transmission lines which parallel
the road system connecting most of the major settlements. Remote villages generally have
diesel generators which run only infrequently during evening hours. Power from the
hydroelectric plants of Landara, Punutuma, and Yura (reconditioned by a joint venture
between COMIBOL and the Valle Hermoso Electrical Company) passes a few kilometres
south of Pulacayo via the national network and a high tension line constructed by the San
Cristóbal Company.
Telephone service and internet access are available in most areas and cellular telephone
service is widespread, although coverage is patchy and international connectivity unreliable.
Local communication services in the area are better than average for rural Bolivia. There is
an ENTEL-based long-distance telephone service, a GSM signal for cell phones, and two
23
antennae for reception and transmission of signals from national television stations. Apogee
has installed a satellite receiver to provide internet access; this service is shared with the
Cooperative Social del Riesgo Compartido (Shared Risk Cooperative).
Exploration in Bolivia by international companies has been minimal in recent years;
Newmont, Coeur d’Alene, Pan American Silver, Glencore and Apex (now Golden Minerals
Company) are the most notable international companies represented in Bolivia in recent
years. Several junior mining companies are also reported to have been active more recently.
Due to perceived political instability and threatened changes to mining taxation there has
been a decline in foreign investment to the mining sector. Basic exploration services are
available within Bolivia: there are several small diamond core drilling contractors;
ALS/Chemex, an international geochemical laboratory group, operates a sample preparation
facility in Oruro; SGS operates an inspectorate service in La Paz; and other local assay
facilities also exist. The Bolivian National School of Engineering operates a technical college
in Oruro (Universidad Técnica de Oruro), including a Mineral Processing department with
laboratory testing facilities, which provides commercial services to the mining industry.
Competent junior-intermediate geologists, metallurgists, mining engineers and chemists are
available in the country.
Used mining equipment is plentiful, although most new equipment is imported from Chile or
Peru where abundant mining services and supplies are available.
At the peak of mining activity, more than 100 years ago, the population of Pulacayo
surpassed 10,000. Today, there remain about 600 people. Demographically, the population is
divided into a civil sector (people outside of mining activities) and a sector composed of
“cooperativistas” (people dedicated to mining activities: Cooperativa Minera Pulacayo Ltda.
(Pulacayo Mining Cooperative).
The village of Pulacayo has a state-run school, and medical services are provided by the state
Caja Nacional de Seguros (National Insurance Fund): a hospital and a clinic each function
independently. The encampment has numerous dwellings, some of which are the property of
COMIBOL while others belong to private individuals. Some COMIBOL properties have
been donated to the “cooperativistas”; some are partly paid for by people who do not reside
in the encampment, but keep the dwellings. As part of the Shared Risk Contract, COMIBOL
puts the use of this infrastructure at the disposal of the project (Figure 5.5).
Potable water for the encampment is supplied from a long-established dam (Yana Pollera)
located 28 km from Pulacayo, in the Cerro Cosuño.
24
6.0
HISTORY
The discovery of mineralization and the subsequent mining of rich silver deposits at
Pulacayo date back to the Spanish Colonial Period (c. 1545). Details of actual production
during this period are unknown. However, examination of the remnants of Colonial town
sites suggests there was a sizable workforce on the property.
In 1833, Mariano Ramírez rediscovered the Pulacayo deposit and, in 1857, Aniceto Arce
founded the Huanchaca Mining Company of Bolivia (Figure 6.1) with the support of French
investors. The operation ceased at the beginning of the 20th century when water problems in
lower levels of the mine halted mining activity.
Figure 6.1
Huanchaca Mining Company of Bolivia, circa 1890
In 1927, Mauricio Hochschild bought the property. The Veta Cuatro vein was intersected at
an elevation of approximately -266 m, and continued down-dip to the -776 m elevation
where it had a strike length of 750 m. During this time the 2.8 km long San Leon access
tunnel was developed to facilitate ore haulage, and the first recorded exploration work in the
area was undertaken.
In 1952 the Bolivian government nationalized the mines and administration of the Pulacayo
deposit passed into the hands of COMIBOL (the National mining enterprise), which
continued operating the mine until closing in 1959 due to “exhaustion of the reserves and
rising costs of exploitation”. COMIBOL also imposed cutbacks on exploration at this time.
25
Historical data from face and back sampling were used by COMIBOL to identify trends of
the high grade silver ore shoots for the Veta Tajo, and its contour map of the silver grades is
presented in longitudinal view in Figure 6.2.
Figure 6.2
Schematic Longitudinal Projection of the Silver Grades, Veta Tajo
(Source: COMIBOL Internal Report, 1958)
In 1962 the Cooperativa Minera Pulacayo (a local group) was founded and leased the mine
from COMIBOL. The Cooperative continues to conduct rudimentary, small scale mining to
the present day, exploiting narrow, very high grade silver mineralization in the upper levels
of the mine, above the San Leon adit level.
Exploration of the Pulacayo area recommenced toward the end of the 1980’s with various
mining and exploration companies targeting epithermal silver and gold mineralization in the
volcanic-intrusive system in the Pulacayo area. In 2001, ASC initiated an exploration
program in the district and signed agreements with the Pulacayo Mining Cooperative and
COMIBOL (see Appendix I). ASC completed regional and detailed geological mapping,
topographic surveying and sampling of the old historical workings. Subsequently ASC
completed 3 drill campaigns at Pulacayo, totalling 3,130 m of diamond drilling.
ASC concluded that silver-lead-zinc mineralization and hydrothermal alteration in the district
are controlled by a strong east-west fracturing system developed in the andesitic rocks
hosting the Tajo Vein.
Significant results from the drilling programs were reported by ASC in a press release dated
October 23, 2002 and are summarized in Table 6.1.
26
Table 6.1
List of Significant Intersections (ASC, 2002)
Hole N°
PUD004
PUD005
PUD006
PUD007
PUD007
PUD010
PUD011
PUD013
PUD013
PUD015
PUD015
PUD018
PUD019
PUD019
PUD020
PUD021
PUD022
PUD022
PUD022
PUD022
PUD022
PUD024
PUD024
PUD024
PUD024
PUD024
PUD024
PUD025
From
(m)
284.80
96.15
105.55
60.00
70.00
364.40
110.00
36.55
104.40
159.10
172.75
105.50
85.00
83.50
141.80
131.55
160.80
62.00
128.55
138.40
148.00
97.00
99.65
110.40
149.00
189.40
212.40
89.60
To
(m)
285.70
114.00
108.00
63.45
96.80
375.35
121.00
58.00
106.50
160.85
174.00
109.60
87.00
94.00
142.50
178.40
163.50
63.30
130.00
140.45
148.56
100.35
100.35
111.60
151.00
190.70
216.30
91.70
Intercept
(m)
True Width
(m)
0.90
17.85
2.45
3.45
26.80
11.35
11.00
21.45
2.10
1.75
1.25
4.10
2.00
10.50
0.70
46.85
2.25
1.30
1.45
2.05
0.56
3.35
0.70
1.20
2.00
1.30
3.90
2.10
0.64
8.00
2.00
2.00
20.00
6.50
5.50
10.50
1.50
1.20
1.00
4.00
2.00
10.00
0.50
25.00
2.00
1.00
1.20
1.50
0.50
2.50
0.50
1.00
1.00
1.00
3.00
2.00
Ag
(g/t)
712.8
521.2
2676.8
1178.3
517.2
181.9
191.6
54.2
653.9
101.7
162.8
132.0
143.0
85.2
144.0
37.0
250.4
103.0
131.0
229.0
354.0
295.0
1120.0
602.0
412.0
181.0
379.4
186.0
Pb
(%)
1.0
2.2
5.9
3.4
2.3
1.4
9.6
2.1
4.3
1.0
2.3
1.1
4.0
1.0
0.5
2.1
2.5
3.5
6.1
5.7
20.3
15.4
0.1
2.1
1.3
3.5
Zn
(%)
1.7
2.4
2.5
3.6
4.2
5.0
9.8
4.3
6.7
2.9
5.5
4.4
5.8
4.9
7.5
5.8
2.6
2.7
5.9
5.4
17.8
6.3
0.3
1.7
3.9
5.7
In March, 2006, Apogee entered into an Option/Joint Venture agreement with ASC (see
Appendix I) and commenced the exploration of the Pulacayo-Paca project shortly thereafter.
Details regarding the exploration work carried out by Apogee are presented in Chapters 10
and 11, below.
27
7.0
7.1
GEOLOGICAL SETTING
REGIONAL GEOLOGY
The geology of Bolivia is well described in various Bolivian government reports, including
Soruco (2000), and various international journals and publications. National and regional
scale geological maps are available from SERGEOTECMIN in La Paz. Some historical
exploration reports are also held in its library.
The principal geological provinces of Bolivia are shown in Figure 7.1, and the regional
geology of central and southwestern Bolivia and the Pulacayo/Paca area is shown in Figure
7.2. The following paragraphs are from the US Geological Survey in USGS Bulletin 1975
(edited).
Figure 7.1
Regional Geology of Bolivia
28
Figure 7.2
General Geology of the Pulacayo Area, Potosí District, Bolivia
“In southwestern Bolivia, the Andes Mountains consist of three contiguous
morphotectonic provinces, which are, from west to east, the Cordillera Occidental, the
Altiplano, and the Cordillera Oriental. The basement beneath the area, which is as thick
as 70 km, is believed to be similar to the rocks exposed immediately to the east, in the
Cordillera Oriental, where a polygenic Phanerozoic fold and thrust belt consists largely
of Paleozoic and Mesozoic marine shales and sandstones.
Deposited mostly on Precambrian basement, the rocks of the Cordillera Oriental were
deformed during at least three tectonic-orogenic cycles, the Caledonian (Ordovician),
the Hercynian (Devonian to Triassic), and the Andean (Cretaceous to Cenozoic). The
Altiplano is a series of high, intermontane basins that formed primarily during the
Andean cycle, apparently in response to folding and thrusting. Its formation involved
the eastward underthrusting of the Proterozoic and Paleozoic basement of the Cordillera
Occidental, concurrent with the westward overthrusting of the Paleozoic
miogeosynclinal rocks of the Cordillera Oriental. These thrusts resulted in continental
foreland basins that received as much as 15,000 m of sediment and interlayered
volcanic rocks during the Cenozoic.
Igneous activity accompanying early Andean deformation was primarily focused further
west, in Chile. During the main (Incaico) pulse of Andean deformation, beginning in the
Oligocene and continuing at least until the middle Miocene, a number of volcano-
29
plutonic complexes were emplaced at several localities on the Altiplano, particularly
along its eastern margin with the Cordillera Oriental, and to the south.
In Pleistocene time, most of the Altiplano was covered by large glacial lakes. The great
salars of Uyuni and Coipasa are Holocene remnants of these lakes. The Cordillera
Occidental consists of late Miocene to Recent volcanic rocks, both lava flows and ashflow tuffs, primarily of andesitic to dacitic composition, that have been erupted in
response to the subduction of the Nazca plate beneath the continent of South America.
This underthrusting continues, and many of the volcanoes that form the crest of the
Andes and mark the international border with Chile are presently active”.
Soruco (2000) describes the geology of the Cordillera Oriental in some detail (edited):
“The Bolivian Cordillera Oriental is a well defined geographic, geomorphological and
geological unit. It is an extension of the same chain in Peru and continues southwards
into Argentina. It is limited to the west by the Coniri and San Vicente faults, which
separate it from the Altiplano, and to the east by the Main Front Thrust as the limit with
the Subandean Ranges. This cordillera has the highest elevations in the Bolivian
territory, reaching altitudes close to 6,500 m above sea level, with the presence of
sectors of permanent snow and glacial development
Tectonically, the Cordillera Oriental can be divided into two sectors, separated by a
deep lineament formed by the Cordillera Real Fault Zone. This lineament possibly
pertains to a reactivated paleosuture. The sector west from this lineament pertains to the
Huarina Fold-Thrust Belt.
Geologically, the Cordillera Oriental holds the country’s most complete stratigraphic
sequence, with Proterozoic to Recent rock outcrops and marine to continental
sequences. The facies are also varied, mostly clastic, but with the development of
carbonate shelves in the Upper Carboniferous and Permian and volcanic and
volcanoclastics in different systems, but mostly in the Cenozoic.
During most of the Lower Paleozoic, it constituted an intracratonic basin, from shallow
to deep, with some compressive and extension phases separating the main tectonic
sedimentary cycles. It goes on later to make up foreland and backarc continental basins,
with important compressive phases with intense associated magmatism”.
7.2
DISTRICT GEOLOGY
The following description of the regional geological framework is based on work done by
geologists of companies which have explored the area and from GEOBOL (now
SERGEOTECMIN) publications. In particular, interpretations by ASC from the Hoja
Geológica Uyuni (Uyuni Geological Leaflet), published by GEOBOL on a scale of 1:250,000
and by Apogee’s geologists, who have worked in the region for many years (Figure 7.3).
30
Figure 7.3
Local Geology of the Pulacayo-Paca Area, Potosí District, Bolivia
The Pulacayo project is located on the western flank of a regional anticline in a geological
environment composed of sedimentary and igneous rocks of the Silurian, Tertiary and
Quaternary ages on the western flank of the Cordillera Oriental, very close to the CordilleraAltiplano boundary. The following major regional structures and geological features have
influenced the local geology and mineralization:

Uyuni-Khenayani Fault: This structure is located about 4 km west of Pulacayo. It is a
reverse fault which appears to control the position of volcanic complexes (Cuzco,
Cosuño, Pulacayo and San Cristóbal) and their respective mineralized areas
31
(Pulacayo, Cosuño, El Asiento, Carguaycollu and San Cristóbal). This structure
places Tertiary sediments in contact with the Paleozoic formations.

Cosuño Caldera: Located a few kilometres north-northeast of Paca (a mineralized
dome 10 km north of Pulacayo). This is a prominent, collapsed elliptical caldera
structure with associated subsidiary domes and extensive ignimbrite deposits which
partially cover the area of study.

Anticlinal Axis: The mineralized zones are almost all positioned on the west flank of a
north-south striking anticline, which is primarily comprised of Silurian sediments
overlain by Tertiary lacustrine formations. Within the anticline structure, a
sedimentary sequence of clay, sandstone, and conglomerates of reddish colour,
located between the Upper Oligocene (Chatiann) and the Lower Miocene
(Aquitanian) time periods, forms the base of the stratigraphic column.

Intrusive bodies: Prominent Lower Miocene dacitic-andesitic domes and stocks that
are associated with phases of resurgence of the calderas (Pulacayo, Tazna, Ubina and
Chorolque calderas) stand out and intrude the sedimentary units. A later volcanic
phase of the Miocene and Pliocene ages is represented in the anticlinal structure by
volcanic pyroclastic and outflows of lava of andesitic and rhyolitic composition. The
upper limit of the lithologic sequence is represented by well-developed ignimbrites,
as a product of the intense activity of the Cosuño Caldera.
The radiometric age dates for the intrusive centres of Animas, Chorolque, Tazna, and
Santa Ana, located from 40 km to 60 km to the southeast of the Pulacayo district are
between 13.8 and 16.8 Ma.
7.3
LOCAL GEOLOGY
SERGEOTECMIN has mapped and named the most relevant Tertiary volcanic-sedimentary
formations in the Pulacayo area. Apogee geologists have remapped the Pulacayo area at
1:1,000 scale and the detailed geology is presented in Figure 7.4.
The stratigraphic sequence that outcrops in the Pulacayo area is comprised of three Tertiary
sedimentary units: Potoco Formation – Ciclo Andino I, bottom-, (Pérez, 1963), the San
Vicente (Courty, 1907) and Quehua Formation – Ciclo Andino II, top-, (Geobol, 1963).
The economic mineralization in Pulacayo is hosted by the sediments of the Quehua
Formation and the Pulacayo andesites.
32
Figure 7.4
Detail Geology of the Pulacayo Area, Potosí District, Bolivia
The following paragraphs describe briefly the geology of the various formations.
Potoco Formation (Tpo) (Eocene, 50 Ma – Oligocene 30 Ma)
The unit was deposited in the backarc and foreland basin of the Eastern Cordillera. The
Potoco Formation forms the base of the Tertiary sequence. It consists of dark red-purple
colour interbedded conglomerates, clays and sandstone lenses up to 2 m thick. The total
thickness of this unit is unknown and it is present only in the Pulacayo area.
33
San Vicente Formation (TSV) (Oligocene, 30 Ma – 25 Ma)
This unit outcrops to the north of the Pulacayo (Cosuño) and forms the base of the sequence
identified in the northern area. It comprises a thick layer of clast-supported polymictic
conglomerate, containing dominantly subrounded clasts of quartzite of Palaeozoic origin that
are up to 25 cm in diameter. The conglomerate matrix is composed of medium-grained
sandstone and clay. Fragments of Cretaceous sandstones and/or of the Potoco Formation are
also seen. The unit as a whole is coloured dark purple to dark reddish and its thickness is
approximately 150 m. The San Vicente formation presents a diversity of continental
environments: alluvial fan prograding facies, braided river fluvial and lacustrine. All these
facies have marked volcanic influence.
A good exposure of the contact with intrusive rocks can be observed at the exit of the San
Leon tunnel at Pacamayo where the sediments are partially altered close to the contact with
the Pulacayo Dome.
Quehua Formation (TQH) (Geobol’s Quechua Fm) (Lower Miocene, 20 Ma – 15 Ma)
Unconformably overlying the San Vicente Formation, this formation is composed of an
intercalation of layers of clay and tuffaceous sandstone, reddish brown in colour to whitish
green-grey in the altered areas, containing isolated conglomerate lenses and coarse-grained
sandstone. Near the top of the formation there is a 20-m thick conglomerate layer.
The sedimentary sequence is intruded by different subvolcanic pulses, all of which constitute
the Pulacayo dome complex. J. Pinto in 1988 described the andesitic rocks of the Rotchild
and Megacristal units as pre-mineral and the dacitic-andesitic rocks of the Paisano unit as
post-mineralization in age. Hydrothermal breccias bodies were also mapped within the dome
complex.
7.3.1
Structural Geology
The mineralized systems in Pulacayo are hosted by the Tertiary sediments and volcanic rocks
of the Pulacayo dome complex. The complex, which is tens of kilometres in length,
constitutes a corridor of several domes having a close spatial relationship with a north-south
oriented regional fault.
Polymetallic mineralization occurs along east-west oriented fault systems, of which the best
known is the Tajo Vein System (TVS) emplaced in the southern side of the Pulacayo dome
complex., The Tajo vein bifurcates in the andesitic rocks to form separate veins, which
collectively form a dense network of veinlets along strike. The bifurcating, polymetallic
veins are commonly separated by altered andesitic rock that contains disseminated sulphide
mineralization.
The TVS is almost 2,700 m along strike at surface and continues to depth of at least 1,000 m,
the lowest level in the underground mine. In the upper levels the vein system is about 120 m
34
in width. The polymetallic veins exhibit a sigmoidal geometry along strike, believed to be the
result of sinistral movement along the north-south oriented regional fault
7.3.2
Hydrothermal Alteration
A local scale alteration system, which can be observed over an area of at least 3.0 km in
length by 2.0 km in width has been mapped at Pulacayo. The hydrothermal activity is
interpreted to have begun in the upper Tertiary, as can be implied from observation of the
Oligocene altered sediments. There is no absolute dating on hydrothermal minerals,
consequently the age of the hydrothermal alteration can only be inferred from cross-cutting
relationships.
Several assemblages of hydrothermal alteration have been recognized: propylitic, sericitic,
moderate-advanced argillic, and siliceous alterations. It is possible to observe and map the
different alteration assemblages in a traverse through the San León tunnel.
Different alteration assemblages and intensities appear in different lithologies. The premineral domes (Rotchild and Megacristal) typically contain an extensive moderate argillic
alteration that changes to an intense argillic alteration with closer proximity to the veins and
disseminated-stockwork zones. A halo of intense silicification measuring a few centimetres
in width is developed in the veins and veinlets walls. The moderate argillic alteration
disappears gradually into a propylitic alteration halo at the borders of the Rotchild and
Megacristal domes. The Paisano dome unit, interpreted as post mineralization in age, does
not contain any observable hydrothermal alteration.
The sedimentary sequence often contains a symmetric alteration halo related to the sulphide
mineralized veins. From the centre of the vein outwards, the alteration grades to a silicified
halo at the wall contacts and gradually into an argillic alteration with further distance from
the vein wall. There is a very distinct change in colour from light green to dark red when the
rock is fresh.
The Pulacayo deposit is a typical polymetallic deposit where galena, sphalerite, barite, sulfosalts, pyrite, quartz and minor chalcopyrite are the major minerals.
The sulphide mineralization occurs in veins, veinlets, disseminations in argillic altered rock
and stockworks. The veinlets, disseminations and stockwork predominate in the andesitic
rocks of the Rotchild and Megacristal domes. In the sediment the mineralization is restricted
to narrow veins. The main vein exploited to date is the Tajo vein, which is rarely wider than
3 m.
35
8.0
DEPOSIT TYPES
Epithermal mineral deposits are found in numerous locales world-wide. This class of mineral
deposit has been recognized since the seminal work of Lindgren (1922) but has only received
focused attention as exploration targets over the last 20 years. This work has revealed that
precious metal mineralization occurs in two basic and distinct end-member styles. Both
styles of mineralization are the result of heated, circulating water that is associated with the
intrusion of an igneous body to within 2-4 km of the topographic surface. These intrusions
are typically felsic to intermediate in composition, are of calc-alkaline affinity and are
currently believed to play an important role in the source of the precious metals and the
initial hydrothermal fluid composition. As these waters rise towards the surface they
undergo physical and chemical changes that control the style of alteration and the locations at
which the precious metals deposit. Many epithermal mineral deposits have been discovered
in the western parts of North and South America and have been found to have been formed
during three discrete geologic periods – Cretaceous (approximately 100 – 65 million years),
Eocene (55 – 40 million years) and Pliocene (5 – 2 million years). Figure 8.1 presents an
illustration of the relationship of these intrusions to the location of epithermal mineral
deposits.
Figure 8.1
Epithermal Mineral Deposit Model
The two end member mineralization styles differ in fundamental aspects. The High
Sulphidation (HS) deposits are formed by very acidic hydrothermal solutions and have
36
characteristic alteration assemblages that include quartz, alunite, and kaolinite. These
deposits are generally hosted by rock units that exhibit the effects of interaction with
extremely acidic solutions. In general terms this style of mineralization is found in close
association to the source of heat and is typically found in spatial relation to bi-modal volcanic
rocks (rhyolite and andesite) that reflect the presence of the underlying intrusions.
Low Sulphidation (LS) deposits are formed by the circulation of hydrothermal solutions that
are near-neutral in pH, resulting in very little acidic alteration with the host rock units. The
characteristic alteration assemblages include illite, sericite and adularia that are typically
hosted by either the veins themselves or in the vein wall rocks. The hydrothermal fluid can
travel either along discrete fractures where it may create vein deposits or it can travel through
a permeable lithology such as a poorly welded ignimbrite flow, where it may deposit its load
of precious metals in a disseminated deposit. In general terms this style of mineralization is
found at some distance from the source of heat. Figure 8.2 illustrates the spatial distribution
of the alteration and veining found in a hypothetical low-sulphidation hydrothermal system.
Figure 8.2
Alteration Mineral Distribution in a Low Sulphidation System
A great body of academic research has been completed on this deposit type in the past 20
years. A recent review of the salient geological features is provided in Sillitoe and
Hedenquist (2003) and a summary of exploration techniques and approaches for these types
of deposits is provided in Hendquist et al. (2000).
The mineralization at Pulacayo is a typical low sulphidation epithermal deposit containing
precious and base metals associated with volcanic rocks. The main geological characteristics
of Pulacayo are:
37

The sulphide mineralization is hosted by Tertiary volcanic rocks of intermediate
composition. These rocks form part of a dome complex, which outcrops at surface.
The mineralized body is composed of stockwork, narrow veins and veinlets, and
disseminations in the argillic-altered rock controlled by an east-west oriented normal
fault system. The width of the mineralization varies from 40 m to 120 m.

Sedimentary rocks intruded by the dome complex constitute the host rock for a
bonanza type, high grade vein (Veta Tajo), with high silver and base metals content.
The vein structure rarely is wider than 3 m and continues into the overlying
stockwork and disseminated zone in the volcanic rocks.

The sulphide mineralization extends along strike for 2,700 m and by almost 1,000 m
to depth, of which 450 m are hosted in the volcanic unit and 550 m are hosted in the
sedimentary unit.

The mineral assemblage is relatively simple: barite, quartz, pyrite, calcite as gangue
minerals; and galena, sphalerite, tetrahedrite, and other silver sulfo-salts as ore
minerals. There is also minor chalcopyrite and jamesonite. The internal texture of the
veins is generally banded and drusy with segments containing almost massive
sulphides. A vertical zonation appears to exist where base metals increase at depth
and silver content is higher at mid levels.
38
9.0
MINERALIZATION
The Pulacayo epithermal deposit is hosted by sedimentary and igneous rocks of Silurian and
Neogene age. The sedimentary rocks are composed of diamictites, sandstone and shale. The
Neogene-aged rocks are mostly of volcanic-sedimentary origin and are composed of
conglomerate, sandstones, reddish conglomerates, reddish-brown clay, whitish rhyolite tuff,
andesite lava flows, dacitic rhyolite domes and andesite porphyry.
The hydrothermal alteration assemblage is characterized by different mineral associations
that can be classified as propylitic, argillic, sericitic, silicification, and opaline alteration
styles. These alteration types form a semi-concentric zoning, which starts from the centre of
the dome and moves towards the outer edge. The spatial distribution of the alteration halos is
sinter silica, silica zone, sericitic zone, argillic zone and propylitic zone. However, these
alteration halos are influenced locally by the presence of mineralized structures.
The spatial distribution of the hydrothermal alteration is used as an indicator for the presence
of mineralized structures. As was mentioned above, there is a strong relationship between the
type of alteration and the presence of veins. The advanced argillic alteration is usually found
in the walls of the veins, grading into less intense argillic alteration and then into a propylitic
zone away from the vein walls (Figure 9.1). The thickness of the advance argillic alteration
envelope varies from few centimetres to several metres in width.
Figure 9.1
Drusy Vein Containing Sphalerite, Galena and Pyrite
Note the advanced argillic alteration halo in the vein walls.
At the Pulacayo deposit, as in many other hydrothermal deposits, the existence of a system of
normal faults is interpreted to have acted as the conduit (feeder) for the mineralizing fluids.
As the fluids circulate along the fractures, changes in temperature, pressure and the redox
39
state between the wall rock and fluid provoke the alteration and precipitation of the sulphide
mineralization in the open spaces forming veins and as disseminated minerals (Figure 9.2).
The veins typically contain banded and drusy textures with intervals of massive sulphide
mineralization that are usually less than a metre wide. Between major veins, the sulphide
mineralization occurs in veinlets measuring on the order of millimetres to several centimetres
in width, as well as occurring as disseminations in the altered rock (Figure 9.3).
Figure 9.2
Example of a Massive Sulphide-Filled Vein, Pulacayo Deposit
Figure 9.3
Example of Veinlet and Disseminated Mineralization
Comprising Tetrahedrite, Sphalerite, Galena and Quartz, Pulacayo Deposit
40
The principal mineralized structure at Pulacayo is known as “Veta Tajo”, which was
historically the main silver producer in the Pulacayo mine. The Veta Tajo is part of a larger
structural system that is oriented approximately east-west and dips 75° to 90° south. The
width of this vein varies from less than 1 m to several m. The structure is filled with quartz,
barite, pyrite, sphalerite, galena and silver sulpho-salts (Figure 9.4).
Figure 9.4
Example of a Quartz-Galena-Sphalerite-Filled Vein, Pulacayo Deposit
As described above, two distinct types of rocks are present: volcanic rocks belonging to the
dome complex and sedimentary rocks. The dome complex is interpreted to be intruding the
pile of sediments and sits above them, forming topographic hills (Figure 9.5). The
mineralized structures are believed to behave differently with each rock type. In the
sedimentary rocks there are usually one to three well define narrow mineralized structures,
with no or very little disseminated mineralization present in between. These veins are very
continuous in the sedimentary rocks, but when they enter into the volcanic rocks they change
their character and spread out into multiple veins and veinlets forming almost a true
stockwork. The width of the mineralized zone can reach up to 120 m. The contact between
the dacitic-andesitic volcanic rocks and the sedimentary rocks is typically found about 500 m
below the surface.
41
Figure 9.5
Longitudinal View of the Stratigraphic Sequence, Pulacayo Deposit
The high grade parts of Veta Tajo have been mined out as a single vein in both lithologies
leaving behind all the lower grade veinlets, secondary veins, stockwork and dissemination
that occur in the volcanic rocks. The deepest level of mining in Veta Tajo is -825 m from
surface, however, sulphide mineralization is known to continue below this level. The mine
ceased operation at this level because of the high cost of extraction in those days and also due
to water problems.
Apparently the veins exhibit variable characteristics from meso-thermal at depth to
epithermal closer to the surface. Fluid inclusion studies in sphalerite found temperatures of
formation that vary from 180ºC to 235ºC. Measurements of salinity vary between 6.4 and
10.90% in equivalent weight of NaCl (Villalpando & Ueno, 1987, at Villalpando et al. 1993).
Technical information on the Veta Tajo System has been gathered sporadically over the
years, but a coordinated scientific approach to the geology, mineralogy and metallogeny is
necessary to understand the mineralizing system. Furthermore the Veta Tajo System is not
the only sulphide mineralized zone known to be present within the dome complex. Other
sub-parallel structural systems containing indications of sulphide mineralization have been
found along the northern contact of the dome complex. As well, breccia bodies up to 100 m
in diameter containing galena and manganese stockwork zones with anomalous silver values
are also present in that area.
42
10.0
EXPLORATION
Apogee Minerals Bolivia S.A. commenced an exploration program in Pulacayo in January,
2006, after the company signed a JV agreement with ASC.
Since then, Apogee has carried out detail geological mapping and sampling at surface and in
the old underground workings, followed up by a topographic survey, geophysical survey, and
diamond drilling. The Apogee exploration work also covered the Paca prospect, located
10 km to the north of Pulacayo. The Paca deposit is included within the limits of the
exploration licenses. In brief, exploration drilling at the Paca deposit has been carried out on
a nominal 50 m by 50 m pattern that has outlined silver-zinc-lead mineralization along a
strike length of 500 m and to a depth of approximately 175 m in Zona Principale, the largest
of the three modeled domains. A mineral resource estimate was prepared in March, 2007,
the details of which are presented in Pressacco and Gowans (2007). This mineral resource
estimate suggests that 18,416,000 tonnes of material at an average grade of 43 g/t Au, 1.16%
Zn and 0.68% Pb are present in the Inferred Resource category which could conceptually be
exploited by means of open pit mining methods. The location of the Paca deposit relative to
the Pulacayo deposit has been presented in Figure 4.3, above.
10.1
TOPOGRAPHIC SURVEY
A topography survey on the Pulacayo-Paca areas was carried out under contract by Eliezer
Geodesia y Topografia which is independent of Apogee and is based in La Paz, Bolivia. The
survey covered an area of 24 km2 using the WGS84, Zone 19 South Datum. The coordinates
of the reference point known as the GCP CM-43 were obtained from the IGM (Instituto
Geografico Militar), (Figure 10.1). The equipment used by the contractor company included
four Total Stations LEICA, models TCR 407, TC 703, TC 605L, and TC 600.
As part of the field work, Eliezer Geodesia y Topografia also picked up the collars of those
completed drill-holes and established 12 lines for an Induced Polarization survey, of which 7
were located in the Pulacayo area and 5 in the Paca area to the north. The stations were
spaced at 50 m intervals along each line.
The topographic map for the Pulacayo-Paca area was constructed with topographic contours
at two metre intervals, with less than 0.5 m of error. The new topographic map was used as
the base map to establish road access, geological mapping and surface sampling as well as
for locating drill collars.
43
Figure 10.1
Topographic Survey Crew, Pulacayo Project
10.2
GEOLOGICAL MAPPING AND SAMPLING
ASC completed a 1:5,000 geological map of Pulacayo in 2003; however, this map only
covered a portion of the area of interest. This map was initially used by Apogee’s geologists
as the geological reference until they completed their own map at a 1:1,000 scale that
covered all the exploration licenses, including both the Pulacayo and Paca areas. The results
of the geologic mapping program have been presented in Figure 7.4 above.
COMIBOL provided ASC all of the old underground mine plan maps of the Pulacayo mine.
Using this information, ASC reconstructed a 3D model for the underground mine. Recently
Apogee modified the mine 3D model to its new topographic map as described in detail in
Chapter 17.4, below.
Apogee carried out a surface sampling campaign at Pulacayo in 2005. The sampling
consisted mostly of rock chip samples taken from outcrops. The objective was to characterize
the alteration patterns and locate the presence of sulphide mineralization both at surface and
in the accessible zones of the underground mine. A total of 549 samples were collected from
the following areas: Andesita, Ramales, Paisano, Veta Tajo and Veta Cuatro. Veta Tajo and
Veta Cuatro are the historical veins mined at Pulacayo, and are oriented approximately eastwest. The Andesitas and Ramales areas are located to the east the Tajo Vein System and the
Paisano area is located to the south of Tajo Vein System. Table 10.1 shows the maximum
assay values obtained in the rock chip samples collected from the various areas. Micon
recommends that the rock chip sample results that were collected in 2005 from the Veta Tajo
and Veta Cuatro areas be integrated into the drill-hole/sampling database.
44
Table 10.1
Summary Table of Rock Chip Sampling Completed by Apogee
10.3
No. of Samples
Location
5
121
196
43
184
Andesita
Ramales 1,2 and 3
Paisano Hill
Veta Tajo
Veta Cuatro
Maximum Values
Ag
Pb
Zn
(g/t)
(%)
(%)
300
2.38
0.5
809
7.84
0.3
325
4.43
0.05
58
1.45
2.36
13.1
0.07
1.13
GEOPHYSICAL SURVEY
An Induced Polarization geophysical survey was carried out between November and
December, 2007 over the Pulacayo and Paca areas by Fractal S.R.L (Fractal), a geophysical
consultant company independent of Apogee. The survey used a dipole-dipole electrode
configuration with readings taken every 50 m.
Seven geophysical lines oriented north-south and separated by 400 m with stations every 50
m covered the Pulacayo area (Figure 10.2).
Figure 10.2
Induced Polarization Survey Coverage Area, Pulacayo Project
45
The orientation of the geophysical lines is approximately perpendicular to the east-west strike
of the Tajo Vein system. Another five lines covered the Paca area. The total distance of
survey coverage is 29 line km. The Induced Polarization survey revealed several areas of
anomalous readings. The resistivity values were seen to vary between 8 to 600 ohm/m – the
low values in electric resistivity were interpreted to represent weakly altered rocks while the
high resistivity values were interpreted to represent siliceous bodies.
The chargeability values were seen to vary between 2 and 20 mV/m, with chargeability
values below 7 mV/m being interpreted to represent the background values (Figure 10.3).
Figure 10.3
Induced Polarization Chargeability Results, Pulacayo Project
Red=high chargeability areas, Blue=low chargeability areas.
46
The results of the geophysical survey led Fractal to conclude that an east-west oriented zone
of anomalous readings measuring some 450 m in width is present in the Pulacayo area. The
highest chargeability values are seen in lines LPY4, LPY5 and LPY6, between stations 0 and
-900, and they are also coincident with high resistivity values. This has been interpreted as a
block of rock with some degree of silicification that contains disseminated sulphide
mineralization.
Similarly, high chargeability anomalies are seen to coincide with the location of the Tajo
Vein System, which is located between stations -750 and -900. This is seen particularly well
along the LPY4 and LPY6 lines. Moderately anomalous values in chargeability that are
located at the edges of the main anomalous zone have been interpreted as altered rocks,
which could be related with a mineralized vein system at depth.
47
11.0
DRILLING
Since the Pulacayo mine was closed in 1959, no exploration was carried out until 2002, when
ASC initiated a diamond drilling campaign. In 2006, Apogee Minerals S.A. (Apogee) signed
a Joint Venture agreement with ASC and commenced an initial exploration program that was
completed in May, 2008.
11.1
ASC BOLIVIA LDC (2002-2005)
Between July, 2002 and November, 2003 ASC carried out the first phase of drilling,
consisting of 14 diamond holes totalling 3,095 m in length (PUD001 to PUD017). Eleven
holes were drilled from surface and another three from drill stations located in the
underground mine. The contract drilling company, Leduc Drilling S.R.L., performed the
drilling with two Longyear rigs, models LF-140 and LY-44. Four holes (PUD003, PUD013,
PUD001 and PUD014) did not intercept the target due to technical problems. However, the
results in the other ten holes were encouraging enough to continue drilling.
The second phase of drilling by ASC commenced in February, 2003. This phase had
considered 10 holes, however, the program was terminated after the first two holes were
completed (PUD025 and PUD026) and totalled 554 m in length. Both holes were drilled
from drill stations located in the underground mine by Drilling Bolivia Ltda, the contract
drilling company. ASC re-initiated the phase II drilling program in September, 2003,
completing eight holes totalling 1,302 m in length (PUD018 to PUD024 and PUD027). Six
holes were completed from surface-based locations and another two holes were completed
from drill stations located in the underground mine. The contract drilling company
Maldonado Exploraciones S.R.L. was hired to complete the phase II drilling program and it
used Longyear, model LY-44 and LF-70 drilling rigs.
The drilling contractors encountered serious problem during the phase II program due to the
terrain conditions and the drilling technique. As a result, some of the holes (PUD020,
PUD021 and PUD023) were abandoned before reaching the target depths. Despite the
excellent results obtained in these two phases of drilling, ASC decided not to continue with
the exploration in Pulacayo. As a result, no drilling was carried out in Pulacayo during 2004
or 2005.
11.2
APOGEE (JAN 2006 – MAY 2008)
The first drilling phase undertaken by Apogee took place between January and June, 2006.
The Phase I program consisted of 19 holes (PUD028 to PUD042) totalling 4,148 m in length.
Four of the holes were completed from drill stations located in the underground mine and
another 15 holes were completed from surface-based locations.
The main objective of the Phase I program was to corroborate the previous drilling results
obtained by ASC, during which several new drill-holes were completed to attempt to twin the
results obtained by previous ASC holes. The Phase I program was also successful in
48
demonstrating the presence of significant amounts of disseminated, veinlet, and stockwork
sulphide mineralization located between the high grade veins that were exploited by the old
and narrow underground mine workings.
Between June, 2006 and February, 2007, Apogee decided to prioritize the exploration work
at the Paca project. More than 25,000 m of diamond drilling was completed at the Paca
deposit, resulting in a postponement of drilling at Pulacayo. The results of the exploration
drilling programs at the Paca deposit are described in Pressacco and Gowans (2007). Apogee
re-initiated drilling activities at Pulacayo in November, 2007. During this Phase II drilling
program, Apogee completed 14 holes (PUD043 to PUD056) totalling 3,745 m in length. All
of the Phase II drill-holes were drilled from surface-based locations.
In general, the results of the Phase II drilling program were better than expected by Apogee.
Some intercepts, such as that in hole PUD045, contained grades up to 262.5 g/t Ag, 0.79%
Pb and 2.93% Zn over a core length of 61.00 m. The Phase II drilling program was
successful in demonstrating that the Tajo Vein System was not only a disseminated, veinlet,
and stockwork sulphide mineralized system that measured more than 100 m wide, but also
contained high grade mineralized shoots that were not exploited by the previous operators of
the mine. On the basis of the results of the Phase II drilling program, Apogee believed that
another campaign of drilling was warranted.
The Phase III drilling program took place between January and May, 2008. During this
phase, 54 holes were completed (PUD057 to PUD110) that totalled 14,096 m in length.
Eight of the Phase III drill-holes were completed from drill stations located in the
underground mine, and the rest of the drill-holes were completed from surface-based
locations.
The Leduc Drilling S.R.L Company performed Apogee’s Phase I drilling campaign, while the
Fujita Core Drilling Company carried out the Phase II and III drilling campaigns. Longyear
rigs, models LF44, LM-55, LF-90 and LM-90, were used for the three phases of drilling.
The diameter of the drill core for most of the holes was HQ (63.5 mm) with some exception
where the diameter had to be reduced to NQ (47.6 mm) to traverse the old mine underground
workings.
11.3
APOGEE (JUN 2008 – SEP 2009)
Apogee plans the drilling programs with the help of geological sections. The coordinates of
the collars of the surface-based drill holes are set by the field geologist using a hand-held
GPS unit, with the azimuth and inclination of the hole being set using a compass and a
clinometer. The coordinates of the collars of the underground-based drill holes are set by a
surveying team, with the azimuth and inclination of the holes being set using transits. The
drill-hole deviation is determined at approximately 50 metre intervals using both Tropari and
Reflex survey tools. The core is stored at the drill-site in wooden core-boxes containing
approximately three m of core each. The sides of the core boxes are marked with the hole
identification, box number and the depth intervals of the hole. Every run of core is separated
49
by a wooden tag indicating the depth of the hole. Once the hole is completed, it is sealed and
monumented with cement. A PVC pipe is put in the collar, which has been closed with a
plug. A metallic plate that records the company name , hole identification, easting and
northing coordinates, elevation, final depth and the “start and end” drilling dates is placed
next to the drill-hole collar. The core boxes are transported by 4WD pick-ups to the core
shack that is located in the town of Pulacayo, a distance of about 5 km.
The core was then examined by the supervising geologist, and the depths of geological,
structural, or alteration features were marked. An examination of the distribution of magnetic
intensity of the drill core was conducted using a hand-held magnet. Descriptions of the
lithologies, alteration styles and intensities, structural features, occurrences and orientations
of quartz veins, occurrences of visible gold, and the style, amount and distribution of
sulphide minerals, were then recorded in the diamond drill logs by the geologist.
50
12.0
SAMPLING METHOD AND APPROACH
All of Apogee’s drill-holes at the Pulacayo prospect were completed using equipment to
produce core with an HQ diameter, with the exception of some drill-holes in which drilling
conditions required the reduction to an NQ diameter.
Drill core was collected twice daily by Apogee geologists from the drill site and transported
by truck to the company core yard at the Pulacayo townsite, at average distance of 5 km.
Hole number and box numbers were marked on each core box by the drilling contractor prior
to transportation. Wooden markers were placed in the core boxes after each run (nominally
3.0 m).
Upon arrival at the core yard, company technicians aligned and reconstructed the core
together when possible and marked individual depth marks at one metre intervals on the core
and in the core box walls. Core recovery was measured between core blocks and noted on a
data entry sheet.
The core was then geologically logged and sample intervals were determined by the
geologist. Generally the entire drill-hole was sampled on a 1-m basis; however, occasionally
in the few holes with very bad recoveries, composite 2 – 5 m samples were taken. As well,
sample lengths are adjusted to reflect significant features observed in the core such as
changes in the geology, alteration or mineralization. In general, stockwork and disseminated
mineralization was sampled separately from mineralization occurring as more massive veins.
In the last phase of drilling, only the zones containing obvious mineralization and the
immediately adjacent wall rocks were sampled. Sample numbers were assigned to each
sample interval, the sample interval was marked on the core and the sample number written
on the core box wall. An aluminum tag with the sample information on it (i.e. sample
number, from and to, geologist initials and date), is also affixed in the core box with staples.
Pictures of the core are taken after the boxes are marked, then the core is cut in half using a
diamond saw and returned to the core box. Friable core was cut in half with a knife.
Samples were taken in 1-m lengths and, in certain cases, at shorter lengths, with each sample
put in a polyethylene bag. The sample number was written on the outside of the bag and the
corresponding sample ticket was inserted into the bag. No missing or misidentified sample
bags were reported.
The type of mineralization in Pulacayo is mainly composed of sulpho-salt minerals,
containing silver, and lead and zinc sulphides, with very rare or no native silver or gold.
Therefore, little “nugget effect” is expected.
The overall recovery has been over 90% in most of the cases regardless of the type of rock
(i.e. andesite, dacite, sandstone or conglomerate). Poor recovery was encountered only when
intercepting old underground workings.
51
A total of 1,208 samples were sent to the laboratory of Bondar Clegg, Canada, from the first
phase of drilling by ASC. A total of 15,454 samples were sent to the ALS-Chemex
preparation facility in Oruro, Bolivia by Apogee., which were then analyzed at the ALSChemex facility located in Lima, Peru. Silver, zinc, lead and copper concentrations were
determined using an aqua regia digestion followed by analysis using the ALS method codes
AA46 and AA62 that employed Atomic Absorption Spectroscopy (AAS). For those samples
containing greater than 300 g/t Ag, the gold values were determined using the same digestion
method but using the Au-AA26 analytical method that employs a Fire Assay-Atomic
Absorption finish on a 50 g aliquot.
A total of 1,161 sample pulps from drill-holes PUD043 to PUD056 were analyzed a second
time as replicates through the analytical method ALS AA46 specific element analysis using
aqua regia digest followed by AAS determination (Ag, Zn, Pb, Cu), and fire assay-AAS
method for samples with silver values exceeding 300 ppm.
A total of 133 sample pulps, corresponding to drill-holes PUD045 and PUD063, were
analyzed a second time using the ME-MS-61 method at ALS-CHEMEX, Lima, Peru, for 47
elements.
As part of the Apogee Quality Control protocols, approximately 5% of the total (594
samples) were re-analyzed by a second laboratory, ALS-Chemex in La Serena (Chile), by the
following techniques; ALS Analytical Codes AA46 and AA62 – specific element analysis
using aqua regia digest followed by AAS determination (Ag, Zn, Pb, Cu). A comprehensive
tabulation of significant results obtained from the drilling programs at the Pulacayo project
was presented in Pressacco and Shoemaker (2008).
Density measurements were taken every 10 m in zones where there was no observed
mineralization; one density measurement per sample was determined when the core
contained obvious mineralization (Figure 12.1). Changes in lithology, type or intensity of
alteration, were also considered for density.
Figure 12.1
Bulk Density Determinations, Pulacayo Project, Bolivia
52
13.0
SAMPLE PREPARATION, ANALYSES AND SECURITY
Apogee does not perform any sample preparation or analytical work itself. All such work has
been completed by ALS Chemex (ALS) of Lima, Peru. ALS is a respected international
analytical service which is accredited with NATA and complies with standards of ISO
9001:2000 and ISO 17025:1999. It utilizes standard analytical methodology and employs a
variety of international standards for quality control purposes.
Samples were transported from field projects to the ALS sample preparation facility in
Oruro, Bolivia by Apogee personnel or a reputable commercial carrier. Sample dispatch
forms were utilized to list all samples in each shipment and laboratory personnel crosschecked samples received against this list, reporting any irregularities by fax or email to the
site.
ALS prepared a flow chart describing the sample preparation procedures for Apogee samples
(Figure 13.1).
Figure 13.1
Sample Preparation Flowsheet, Pulacayo Project, Bolivia
53
All samples are weighed upon receipt and prepared using ALS preparation procedure PREP31B which consists of crushing the entire sample to >70% -2 mm, then splitting off 1 kg and
pulverizing to better than 85% passing 75 micron (Figure 13.2). The coarse rejects are
returned to Apogee for storage on site at Pulacayo.
Figure 13.2
Particle Size Analyses of Exploration Samples, Pulacayo Project
Perform particle size analysis (Apogee)
-2mm % Passing
100%
95%
90%
85%
80%
75%
1
26 51 76 101 126 151 176 201 226 251 276 301 326 351 376 401 426 451 476
Real Values
Base line 85%
work orders
Samples from ASC’s drill campaign (2002-2003) were analyzed at ALS’s facility in
Vancouver, BC, Canada; samples from Apogee’s programs (2006- ) were analyzed at ALS’s
facility in Lima, Peru.
All analytical testing is performed utilizing a variety of industrial standard analytical
techniques, including (i) ALS Analytical Code ME-MS41-50 element analysis using aqua
regia digestion followed by ICP-AES analysis, (ii) ALS Analytical Codes AA46 and AA62
specific element analysis using aqua regia digestion followed by AAS determination (Ag, Zn,
Pb, and Cu), and (iii) Fire Assay-AAS finish, for samples with Ag values >300 ppm.
ALS inserts its own quality control samples (reference materials, blanks and duplicates) on
each analytical run, based on the rack sizes associated with the method. The rack size is the
number of samples, including QC samples, included in a batch. The blank is inserted at the
beginning, standards are inserted at random intervals, and duplicates are analyzed at the end
of the batch. All data gathered for quality control samples – blanks, duplicates and reference
materials – are automatically captured, sorted and retained in the QC database.
If any assay for reference materials, duplicates, or blanks falls beyond the control limits
established, it is automatically flagged red by the computer system for serious failures, and
yellow for borderline results.
Apogee has instituted internal QA/QC procedures. Control samples (reference materials,
blanks and duplicates) are routinely inserted in the batches. “Field Blanks” were prepared
from a source of unmineralized quartzite outcropping near Pulacayo, and comprised a
“coarse blank”. A commercial blank was purchased as the “fine blank”. Apogee also
54
purchased “commercial standards” (Certified Reference Material), brand WCM, types PB128 and PB-124.
Field and commercial blanks and standards were inserted at least 1 in every 50 samples to
ensure the presence of enough control samples in each rack. Duplicate samples, comprised
of ¼ core, and duplicate samples of pulp were re-analyzed and both were taken random.
Duplicate samples are given a new, unique number.
Finally and as part of the QCAC procedures, the ALS facility in La Serena, Chile, was used
as a second laboratory for cross laboratory analysis.
The results of all of the standards, as well as the targets and double samples, are monitored
constantly by using Evaluation Software and Quality Control in Reports while preparing the
QA/QC respectively, accompanied by graphs by dates and by drill-holes, with the assistance
of appointed personnel. There were not any significantly abnormal results detected in the
samples.
A full description of the Quality Control results has been provided in Pressacco and
Shoemaker (2008).
55
14.0
DATA VERIFICATION
Micon’s senior geologist, Mr. Reno Pressacco P.Geo., conducted a site visit to the Pulacayo
project area between March 26 and 29, 2008 to examine the general site conditions present
there. A small number drill pads were visited where discussions were undertaken that
examined the drilling procedures that are employed. At the time of Micon’s site visit, one
drill rig was in operation (Figure 14.1). Micon found that the drilling program was being
carried out to the highest standards currently being practiced in the mining industry and
observed that the geological staff was highly motivated and enthusiastic.
Figure 14.1
General View of the Diamond Drilling Operation, Pulacayo Project
Micon continued its data verification by reviewing the drill core logging and sampling
procedures and by comparing the geology and mineralization in several drill-holes against
the descriptions in the drill logs completed by the Apogee geologists. Micon found that the
logging of the drill core was carried out to the highest degree of quality and noted no
significant errors or omissions in the descriptions of the geology and mineralization. Micon
compared the assay results presented in three drill logs against the observed mineralization
for the respective section of drill core and found that the assays correlated well with the
56
visual observations. Micon observed that the sampling procedures are adequate to the
mineralization found at the Pulacayo deposit.
During the course of its inspection of the drill core, Micon noted that a number of the drillholes intersected mined out stopes from the historical mining activities at Pulacayo. In some
cases the stopes are represented by a clay-like material (possibly uncemented back-filled
tailings), loose rock-fill, or voids. In some cases where the drill-holes intersected open
stopes, the drilling contractor was able to continue the drill-hole beyond the stoped area,
while in other cases the drill-hole had to be terminated at the far stope wall.
In conducting spot checks of the density measurements from drill core, Micon noted that the
procedures employed previously consisted of the selection of a small piece of drill core with
the intention of being representative of a longer interval (generally tens of m) of drill core. In
some cases, Micon noted that this resulted in a significant variance locally and attributes this
to sample bias. In other words, the selection of the small piece of drill core was not
representative of the larger interval. Micon recommends that remedial actions be undertaken
wherein additional density measurements be taken on a detailed scale within the zone of
stockworking and veining in support of an accurate estimate of the tonnage of the mineral
resource. As well, Micon also noted that the mineralized intervals contain intervals that are
porous to varying degrees, with little attendant permeability. Micon must point out that
current industry best practice includes employing the wax-seal method for cores containing
porosity.
An audit of the field version of the drill-hole database as at March 29, 2008 was conducted
by selecting approximately 10% of the drill-holes. Micon understands that this field version
of the database is not the most complete version, as the most up-to-date version of the
database is maintained in the La Paz office. Micon understands that the field database is
updated on a periodic basis, approximately once per month, so that the most recent drill-hole
information may not be reflected in the field version of the database. In conducting its audit,
Micon compared the information contained within the paper copies of the selected drill logs
with the information contained in the digital database. The detailed results of this
comparison have been recorded separately and have been supplied to Apogee at the Pulacayo
site. In brief Micon views the findings for the most part as housekeeping items such as:

Disagreement in the drill-hole co-ordinates between the paper logs and the digital
database is common. Micon suspects that this is due to more current information
being utilized in the digital database from such activities as ground-truthing or
detailed survey pick-ups following completion of the drill-holes. Micon recommends
that the most current drill-hole location information be included with the drill logs.
As well, Micon recommends that the drilling method (DD, RC) and size of drill-hole
(HQ, NQ, etc) be recorded on the drill-hole logs.

There are common occurrences where the digital database contained down-hole
deviation information, however, no such information can be found in the paper logs.
57
Micon recommends that the down-hole deviation information be included with the
drill logs.

Micon noted that the digital database contained information in regard to the core
recovery, however, no records of this information could be found with the paper logs.
Micon noted that the core recovery is recorded by the core shack technicians by hand
on pre-printed forms, and recommends that these originals be included with the drill
logs.

Micon noted that the results of the specific gravity tests are included in some of the
drill-hole logs reviewed, however, no digital equivalent could be found. Micon
suspects that this information is contained in a separate digital file and recommends
that the density information be included as either a separate tab in the drill-hole
database, or as a separate column in the assay table.

In examining the assay portion of the database, Micon noted that neither originals nor
copies of the laboratory certificates are contained with the paper drill logs, and are not
located at the field site (assay certificates are stored at the La Paz office). Micon
recommends that either original assay certificates or good quality copies of the assay
certificates be included with the paper drill logs.

Micon noted that the results of the QA/QC samples are currently recorded in the main
assay table, along with the remainder of the assay records. As was discussed at
Pulacayo, this manner of treatment will result in extensive data import errors for
many of the commercial mine modeling softwares, and Micon recommends that such
QA/QC data as blanks, standards, duplicates, and replicates be recorded as separate
tabs in the main database.

As well, it was noted that assay values of less than detection limits were entered in
the assay table using the “<” symbol. In Micon’s experience, this will result in
extensive data import errors for many of the commercial mine modeling softwares,
and recommends that the “<” symbol not be included in the assay table. Alternative
methods include entering any values below detection limits as the detection limits, or
entering them as ½ of the detection limits.
Discussion was also undertaken in respect of how to handle replicate assays (i.e. for silver
values determined by different analytical methods). Micon recommends that the assay table
be amended to include columns for the assay values by different assay methods, and the
inclusion of a set of “Accepted Value” columns in the assay table.
Micon completed its data verification activities by selecting a total of 10 quarter core samples
from mineralized sections of drill-hole PUD045 for check assaying. The core samples were
shipped to Acme Analytical Laboratories in Vancouver, British Columbia where the silver,
zinc and lead contents were determined using similar analytical techniques to those used by
the ALS Chemex laboratory (aqua regia digestion followed by AAS). The numeric data are
presented in Table 14.1.
58
Table 14.1
Comparison of Micon Check-Assay Results from Drill-hole PUD045
Apogee Original
Pb
Ag (ppm)
(%)
191
4.81
5
0.17
34
0.88
6
0.21
23
0.42
SampleID
Hole_ID
From
To
B04239
B04241
B04242
B04243
B04244
PUD045
PUD045
PUD045
PUD045
PUD045
257
258
259
260
261
258
259
260
261
262
B04280
B04281
B04282
B04283
B04285
PUD045
PUD045
PUD045
PUD045
PUD045
293
294
295
296
297
294
295
296
297
298
308
900
707
210
433
B04280
B04281
B04282
B04283
PUD045
PUD045
PUD045
PUD045
293
294
295
296
294
295
296
297
308
900
707
210
0.45
1.53
0.49
0.29
0.32
Zn
(%)
6.67
1.03
2.56
1.57
3.43
1.35
1.06
1.11
8.36
2.83
Micon Check
Ag (ppm)
Pb
(aqua regia)
(%)
279
6.88
5
0.17
7
0.15
6
0.17
15
0.26
304
648
762
607
168
(fire assay)
304
611
693
552
0.39
0.91
0.43
0.16
0.29
Zn
(%)
9.04
0.93
1.43
1.5
3.83
1.51
1.9
3.67
10.23
2.06
Check assay comparisons for silver, zinc and lead are presented graphically in Figures 14.2,
14.3 and 14.4, respectively. These samples were taken with the purpose of independently
confirming the presence of mineralization at the Pulacayo project and the small number does
not represent a valid statistical population to compare against Apogee’s routine analysis.
However, Micon observes that there is a high degree of correlation between the original
assays and its check assays and, on this basis, has reason to believe that the reported assay
values for silver, zinc and lead are reasonable.
Figure 14.2
Comparison of Silver Check-Assay Results, Pulacayo Project
59
Figure 14.3
Comparison of Zinc Check-Assay Results, Pulacayo Project
Figure 14.4
Comparison of Lead Check-Assay Results, Pulacayo Project
60
15.0
15.1
ADJACENT PROPERTIES
SAN CRISTOBAL
Significant quantities of zinc, silver and lead were discovered at the San Cristóbal deposit,
located in the Potosi district of southwestern Bolivia, approximately 100 km southwest of
Uyuni (Figure 15.1). The deposit has been developed towards production and, now under the
ownership of Sumitomo Corporation, is understood to be managed by Golden Minerals
Company, successor to the assets of ASMC after the latter’s emergence from bankruptcy
protection in March, 2009.
A brief description of the San Cristóbal project is given below, based on material previously
published on the Apex Silver website (www.apexsilver.com), as of February 10, 2007.
Figure 15.1
Location of the San Cristobal property
San Cristóbal occupies the central portion of a depression associated with volcanism. The
4-km diameter depression is filled with fine- to coarse-grained volcaniclastic sedimentary
rocks. Disseminated and stockwork silver-lead-zinc mineralization occurs locally both within
the volcaniclastic sediments and in the intrusions themselves. The primary lead mineral is
silver-rich galena, and the primary zinc mineral is sphalerite.
A process flowsheet including crushing, grinding, and flotation was proposed, to produce
lead/silver and zinc concentrates for shipping to smelters for final metal recovery.
15.2
SAN VINCENTE
Pan American Silver currently conducts underground mining and processing operations at its
San Vicente mine, located approximately 150 km south of Uyuni. An expansion of mine
capacity to produce and treat 750 tonnes per day has recently been undertaken. A summary
of the San Vicente mine is presented on the web site maintained by Pan American Silver and
61
is excerpted below1. NB: Micon has not independently verified the information presented on
this web site. The information relating to the San Vicente mine is not necessarily indicative
of the mineralization found at the Pulacayo deposit, but is believed to be similar in style.
“The San Vicente silver-zinc underground mine is Pan American Silver Corp.’s only
mining interest in Bolivia. More than 20 bonanza type silver-zinc veins are known to
occur over an area of 1.5 km on surface and extend to at least 200 m in depth...
San Vicente is a polymetallic vein deposit, located 2.5 km west of a prominent thrust
fault. This north-south striking San Vicente fault forms the eastern limit of the
intermountain Bolivian Altiplano Basin. The lithology of the Project area includes the
fanglomerate facies of the San Vicente formation, which are in contact with Ordovician
shales along the San Vicente fault. The fanglomerate consists of poorly sorted
conglomerate with clastic sub-angular fragments of Palaeozoic sediments cross cut by
quartz veins. The matrix is red in colour and consists of iron bearing sandstone.
Mineralization in the district …consists of vein mineralization in pre-existing faults,
dissemination in brecciated conglomerates in the San Vicente fault, and mineralization
in dacitic dykes. The existing mine was designed and built to extract [material] from
steeply dipping narrow veins using conventional shrinkage stoping. The discovery of
the Litoral Ramal Dos vein has provided a wide and high-grade addition to the mine
resource base. This vein is amenable to longhole mining …and will permit a higher
mining recovery of the wider ore zones than could be achieved through shrinkage
mining.”
Figure 15.2
Schematic diagram showing San Vicente property
(Source: www.panamericansilver.com)
1
http://www.panamericansilver.com/operation/bolivia214.php (as of Sep 25, 2009)
62
16.0
16.1
MINERAL PROCESSING AND METALLURGICAL TESTING
METALLURGICAL TESTWORK
There are two bodies of testwork on the Pulacayo resource:

A preliminary metallurgical and petrographical analysis was carried out by Resource
Development Inc (RDi), Colorado, in March, 2003, on a very high grade sample.

Additional metallurgical tests were undertaken for Apogee by Universidad Técnica
de Oruro (UTO), Bolivia, in November, 2009, on high, medium, and low grade
samples.
This section will discuss both bodies of work, and will cover sample preparations, the
recovery of zinc, lead, silver, and the effects of an ultra-fine size fraction in the ore on the
recovery and flowsheet design.
The head grades of the samples tested are shown in Table 16.1 below.
Table 16.1
Head Assays of the Pulacayo Metallurgical Composites
Description
Ag (g/t)
Pb (wt%)
Zn (wt%)
Cu (wt%)
Low Grade
46
0.79
1.24
.03
Med Grade
181
0.69
2.46
0.07
High Grade
268
1.58
2.71
0.67
High/High Grade*
*Taken from RDi Report-March, 2003
519
2.2
3.96
nr
Note that the RDi sample from 2003 is very high grade (519 g/t silver) and is not typical of a
mill feed grade for this resource. This sample, which plots at the top of the grade-recovery
curve, is not representative of the potential mill-feed. For that reason a second set of tests
were requested by Apogee on samples that would have more typical mill feed grades.
In August, 2009, three sample composites (low, medium, and high grade) were prepared
from drill core and sent to UTO for testing. Intervals were selected that would match what
would be produced by a Pulacayo mining schedule that would produce a range of grades
typically feeding the mill, from 100 g/t to 300 g/t silver. UTO tests were completed in
November, 2009 with a medium and low grade report and a separate, high grade report
issued in February, 2010.
16.1.1
RDi Preliminary Metallurgical Results, March 2003
The RDi results are valuable as preliminary tests. As summarized in the RDi report:
63
“Resource Development Inc. (RDi) completed the preliminary metallurgical testwork
on the composite sample of Pulacayo drill cores. The preliminary objective of the
testwork was to determine the response of silver [and sulphide minerals] in the ore to
various processing options. The bench-scale testwork includes in-place bulk density
determination, sample characterization, mineralogy, leaching, gravity concentration and
flotation.
The highlights of the study indicated the following:

The composite sample assayed 518.9 g/t Ag, 2.20% Pb and 3.96% Zn.

The head analyses of the sample were significantly higher than the values
calculated from the drill core data.

The host rock contained significant amount of clay material which resulted
in problems in settling of tailings and flotation pulp rheology.

The in-place bulk density ranged from 2.2 mt/m3 to 3.4 mt/m3.

The mineralogy of the sample was significantly different from the San
Cristobal deposit. The predominant sulphide-bearing minerals were pyrite,
marcasite, sphalerite and galena. Native silver was not detected in the
samples.

Silver minerals in the ore were not amenable to leaching, even at the fine
grind size. The extractions were generally less than 20%.

The sample was not amenable to gravity concentration for silver recovery.

The sample responded well to the flotation process and reagent suite
developed for the San Cristobal deposit. The locked-cycle flotation process
recovered 88.8% of lead and 63.4% of silver in the third-cleaner
concentrate assaying 62.2% Pb, 4.46% Zn, 10,891 g/t Ag and 0.15 g/t Au.
The zinc third-cleaner concentrate , assaying 61.5% Zn, 3,303 g/t Ag and
0.11 g/t Au recovered 87.6% of zinc, 31.3% of silver and 4% of gold. The
results are summarized in Table 16.1.

The overall silver recovery in the flotation process was 94.7%. The gold
recovery was poor at 7.7%.

The impurity analyses of third-cleaner lead and zinc concentrates indicate
that penalty will be assessed on these concentrates since several impurities
are higher than the norm for the smelter contracts. These impurities should
be carefully monitored in any subsequent testwork.

The recoveries for lead, zinc and silver need to be viewed with caution as
the results would change substantially at lower head grades due to a
constant tail effect.”
Results from the 2003 locked-cycle tests are presented in Table 16.2.
64
Table 16.2
High/High Grade Locked-Cycle Flotation Tests
(from Rdi Report, Mar 2003)
Description
% weight
g/t Ag
Lead concentrate
3.10
Zinc concentrate
5.00
%Zn
%Zn
Dist
%Ag Dist
%Pb
%Pb Dist
10891
63.4
62.20
88.8
4.46
3.9
3303
31.3
0.90
2.1
61.50
87.6
Tails (not reported)
Back calculated head
16.1.2
UTO Metallurgical Testwork, August 2009
16.1.2.1
Sample Preparation
Drill core assays were supplied by Apogee to Micon, which can be found is section 6. From
these assays, core intervals were chosen that would result in a low, medium and high grade
composite. The half-core supplied was spilt once more to quartered-core for the intervals
chosen, that would produce 100 kg of composite for each grade. Figure 16.1 shows the
Pulacayo core shack area where the core was chosen, split and packaged into 5 gallon
buckets for shipment 200 km north to UTO labs in Oruro.
Figure 16.1
Sample Preparation at Pulacayo Core Shack
65
The UTO laboratory in Oruro was toured by Michael Godard, Senior Metallurgist for Micon,
on July 27, 2009, prior to the composite preparation. The laboratory is part of a technical
school attached to the Facultad Nacional de Ingenieria. The laboratory was not being used
by the school at the time of the tour, but appeared well equipped to handle any mineral
processing tests required. Processing the composites at UTO, and in Bolivia in general, had
the advantage of being able to use the UTO’s engineering professors and mineral processing
professionals.
An adjacent assay laboratory, Spectro Lab, was also toured, but was mostly equipped for
solution determinations. Spectro Lab was not used to assay the metallurgical products of the
met-lab for this reason. ALS Laboratory group was selected instead, since they had been
contracted in the past to assay Apogee’s drill core, and had the required QA/QC already in
place. Consequentially ALS’s laboratory in Lima, Peru performed the assay determinations
on UTO’s metallurgical products.
Figure 16.2
Bench Flotation Cells at Universidad Técnica de Oruro
66
The 100 kg of each composite grade was shipped to UTO, where it was crushed and blended.
Three head samples of each were taken and sent to ALS for fire and ICP assays, the results of
which are shown in Table 16.1.
It was decided to focus on metallurgical tests for the medium grade ore to expedite
preliminary information. A report on the low and medium grade sample was received along
with a high grade report and are included in Appendix IV. Four open circuit flotation tests
were done followed by a closed circuit, locked-cycle test. The four open circuit tests were
done first to determine the best test parameters, such as reagent dosages, before the more
definitive locked cycle test. The summary of the locked-cycle tests are given in Table 16.3.
Table 16.3
Locked-Cycle Flotation Tests Assay Results
LOW GRADE
Description
% weight
g/t Ag
%Ag Dist
%Pb
%Pb Dist
Lead concentrate
0.87
3390
67.5
50.50
62.1
Zinc concentrate
1.55
318
8.5
1.24
2.0
Tails
97.58
11
24.0
0.26
35.9
100.00
45
100.0
0.71
100.0
Taken from: Prueba 3 Circuito Cerrado LB, Tabla 28, pagina 59 Informe UTO 1
NSS for re‐assay (CD >1000 g/t)
MEDIUM GRADE
Description
% weight
g/t Ag
%Ag Dist
%Pb
%Pb Dist
Lead concentrate
1.20
6220
33.7
51.00
74.3
Zinc concentrate
3.70
2990
49.7
0.85
3.8
Tails
95.10
39
16.7
0.19
21.9
100.00
222
100.0
0.83
100.0
Taken from: Prueba 3 Circuito Cerrado LM, Tabla 8, pagina 24 Informe UTO 1
HIGH GRADE
Description
% weight
g/t Ag
%Ag Dist
%Pb
%Pb Dist
Lead concentrate
2.14
9670
69.9
53.40
77.0
Zinc concentrate
3.69
1080
13.5
1.00
2.5
Tails
94.17
52
16.7
0.32
20.5
100.00
296
100.0
1.48
100.0
Taken from: Prueba 2 Circuito Cerrado LA, Tabla 7, pagina 21 Informe UTO %Zn
%Zn Dist
19.65
53.30
0.21
1.15
14.9
72.0
17.6
100.00
%Zn
%Zn Dist
3.72
58.30
0.43
2.61
1.7
82.6
15.7
100.00
%Zn
%Zn Dist
5.52
59.00
0.50
2.77
4.3
78.7
17.1
100.00
%Cu
1.15
0.20
0.01
0.02
%Sb
0.64
0.08
0.007
0.01
g/t Hg
6.64
10.50
0.89
1.09
g/t As
1555
571
825
827
g/t Cd
>1000
>1000
0
n/a
%Cu
1.79
1.29
0.02
0.09
%Sb
g/t Hg
1.65
2.81
0.77
11.30
0.02
0.31
0.067347 0.74663
g/t As
3940
457
1145
1153
g/t Cd
261
>1000
156
n/a
%Cu
2.22
0.43
0.02
0.08
%Sb
g/t Hg
3.96
2.79
0.30
13.71
0.01275
0.31
0.107931 0.857532
g/t As
2530
502
1865
1829
g/t Cd
346
>1000
12
n/a
The back calculated lead and zinc head grades matched the composite head grades in Table
16.1, which indicates the metals balance for this test is reasonably accurate. For comparison,
the 2003 RDi locked-cycle test is summarized in Table 16.2.
A “grade vs. recovery” relationship can be developed. The relationship between silver head
grades and recovery are shown in Figure 16.4 and Figure 16.5 (see page 72).
67
Table 16.4
Locked Cycle Test Results-No Desliming Prior to Flotation
High/High Grade Locked‐Cycle Flotation Test (from Rdi Report, Mar 2003)
Description
% weight g/t Ag %Ag Dist
%Pb
%Pb Dist
Lead concentrate
3.10
10891
63.4
62.20
88.8
Zinc concentrate
5.00
3303
31.3
0.90
2.1
Tails (not reported)
head
100.00
519
2.20
High Grade Locked‐Cycle Flotation Test
Description
%Pb
%Pb Dist
Lead concentrate
2.39
6620
56.0
52.40
78.8
Zinc concentrate
3.76
2010
26.7
1.25
3.0
Tails
93.85
52
17.3
0.31
18.3
head
100.00
283
100.0
1.59
99.99
Medium Grade Locked‐Cycle Flotation Test
Description
%Pb
%Pb Dist
Lead concentrate
1.20
6220
33.7
51.00
74.3
Zinc concentrate
3.70
2990
49.7
0.85
3.8
Tails
95.10
39
16.7
0.19
21.9
head
100.00
222
100.0
0.83
99.99
Low Grade Locked‐Cycle Flotation Test
Description
%Pb
%Pb Dist
Lead concentrate
0.53
2600
30.1
46.70
34.7
Zinc concentrate
1.28
749
20.9
4.32
7.7
Tails
98.19
23
49.1
0.42
57.6
head
100.00
46
100.0
0.72
100.00
%Zn
4.46
61.50
%Zn Dist
3.9
87.6
3.96
%Zn
7.94
57.00
0.33
2.64
%Zn Dist
7.2
81.1
11.7
100.0
%Zn
3.72
58.30
0.43
2.61
%Zn Dist
1.7
82.6
15.7
100.0
%Zn
9.72
44.80
0.51
1.13
%Zn Dist
4.6
51.0
44.5
100.0
Ultra-fines generated from the Pulacayo ore appear to be the reason for the low silver
recovery in the medium and low grade sample.
16.1.2.2
The Effect of Fines on the Pulacayo Resource
As noted in the RDi report, the sample contained a significant amount of clay material which
created problems in flotation pulp rheology. The 2009 testwork also identified issues with
the clays in the Pulacayo ore.
For this discussion, fines are defined as any particle with a diameter smaller than 44 microns,
and clays as particles with diameters smaller than 2 microns. Also, “desliming” is defined as
removing the clay fraction from the slurry.
Clay causes problems by:

Significantly reducing the separation efficiencies between mineral and gangue in
flotation. Clay particles preferentially follow the water into either the concentrate or
to the tailings resulting in lower concentrate grade and lower recovery.
68

Blinding the filter cloth and retaining water in the filter cake when dewatering the
concentrates.

Slow settling rates in the TSF water-cap column; hence, reclaiming water from the
TSF may be problematic unless it is treated.
It is recommended that the clays in the Pulacayo ore be further studied to determine their
effects on concentrate filtering, reclaim water clarity, and TSF deposition density for future
design.
The test program at UTO called for one open-circuit float test to be done with desliming of
the slurry prior to flotation. Figure 16.3 shows the test procedure, described in the UTO
report, which generated the results of test 4, as given in Table 16.5.
Table 16.5
Metallurgical Balance, Deslimed Prior to Float, Medium Grade Test 4
Products
Pb-Ag Flotation
2º Pb Cleaner
1º Pb Cleaner
Pb Rougher
Zn-Ag Flotation
2º Zn Cleaner
1º Zn Cleaner
Zn Rougher
Underflow
Overflow
Total Tailings
Calculated Head Grade
Weight
%
0.68
0.20
0.85
1.73
2.97
0.37
1.41
4.75
71.22
22.31
93.52
100.00
% Pb
57.00
29.20
6.47
29.04
0.29
0.72
1.14
0.58
0.13
0.35
0.18
0.70
Lead
% Dist.
55.49
8.42
7.81
71.72
1.23
0.38
2.30
3.91
13.22
11.15
24.37
100.00
Silver
g/t Ag
% Dist.
10,000
35.49
7,270
7.65
1,790
7.88
5,666
51.02
679
10.49
338
0.64
2,900
21.35
1,314
32.48
16
5.93
91
10.57
34
16.50
192
100.00
% Zn
5.78
4.80
3.90
4.75
55.40
11.25
13.80
39.60
0.12
1.21
0.38
2.32
Zinc
% Dist.
1.70
0.42
1.42
3.54
70.92
1.78
8.42
81.12
3.69
11.65
15.34
100.00
The cyclone overflow, which is the deslimed fraction product, contained 10.57% of the silver
which was sent to tails. If this silver can be recovered, it could potentially increase silver
recovery by 10%. If this silver clay-fraction can be recovered it would also have to be
cleaned to concentrate grade. Sodium fluosilicate was used to depress the clays in Test 4 and
in the other flotation tests. Lime was also used to control pH for differential flotation and to
disperse the clay to help improve the concentrate grade. The treatment of the silver clayfraction requires further metallurgical tests, before final selection of the process flowsheet
and equipment.
The generation of clays in the grinding circuit needs to be kept to a minimum. For this
reason, cyclone separation efficiencies will be an important consideration when the selection
of grinding circuit equipment is done.
69
Figure 16.3
Open Circuit Float Test Parameters - Desliming Prior to Float -Test 4
70
RDi reported the clay as coming from the 30% mica/illite, and 10% kaolinite in the host rock.
In addition, the OTX Mineralogy report (January, 2003), indentifies tuffaceous clastite as a
clay-generating rock. No further mineralogical testing on the clays is considered necessary.
Particle size distributions (PSD), and assaying of size fractions on the flotation feed done by
UTO shows that the clay content increases with decreasing grade. The medium grade had
22.31% fines in the float feed, and the low grade has 26.65% fines in the float feed. This is
also backed up by the size analyses on the flotation tails showing increasing amounts of clay
(including silver-bearing clay) going to tails with decreasing head grades. As the head grade
decreases, clays have an increasingly detrimental effect on concentrate grade and recovery.
Table 16.6 shows increase silver recovery calculated by adding 75% of the silver in the slime
fraction to concentrate, instead of sending it to tails. This is practical using fine recovery
equipment and using SG differentials between oxides clays and metal clays with centrifuges,
spirals, and/or desliming cones, etc. These recovery numbers were used in Table 16.7 to
forecast the concentrate grades and recovery used in the economic assessment.
Table 16.6
Locked Cycle Test Results - Desliming Prior to Flotation
High Grade Locked‐Cycle Flotation Tests (deslimed)
Description
2.14
9670
69.89
Lead concentrate
3.69
1080
13.46
Zinc concentrate
75.43
32
8.16
Float tails
18.74
134
8.49
COF to tails
94.17
52
16.65
Total tails
100.00
296
100
Medium Grade Locked‐Cycle Flotation Tests (deslimed)
Description
0.8
12250
47.49
Lead concentrate
3.56
1460
25.12
Zinc concentrate
75.59
48
17.52
Float tails
20.05
102
9.87
COF to tails
95.63
59
27.39
Total tails
100.00
207
100
Low Grade Locked‐Cycle Flotation Tests (deslimed)
Description
0.87
3390
67.53
Lead concentrate
1.55
318
8.45
Zinc concentrate
71.76
11
18.1
Float tails
25.82
10
5.92
COF to tails
97.58
11
24.03
Total tails
100
45
100
71
%Pb
%Pb Dist
%Zn
%Zn Dist
53.4
76.98
5.52
4.27
1
2.49
59
78.68
0.23
0.7
11.7
8.84
0.27
1.43
7.37
9.69
0.32
20.54
0.5
17.06
1.48
100.01
2.77
100.01
%Pb
%Pb Dist
%Zn
%Zn Dist
56.93
0.74
61.51
3.55
5
51.4
1.83
83.66
0.24
24.42
0.42
14.51
0.39
10.52
1.25
11.45
0.27
0.74
34.94
100
0.59
2.19
25.95
111.45
%Pb
%Pb Dist
%Zn
%Zn Dist
50.5
62.12
19.65
14.85
1.24
2.03
53.3
72.03
0.27
27.44
0.21
13.12
0.23
8.41
0.2
4.49
0.26
0.71
35.85
100
0.21
1.15
17.61
104.49
The zinc head grade distributions (highlighted above, in yellow) indicate the need for UTO to
further refine the reported metallurgical balance.
Figure 16.4 show the grade recovery curve for straight flotation without desliming prior to
flotation, and Figure 16.5 is the grade recovery curve assuming desliming prior to flotation,
but then adding 75% of the fine silver back to the concentrates to obtain the final silver
recovery.
Figure 16.4
Silver Grade vs % Recovery without Desliming prior to Flotation
100
%
95
90
R
e
c
o
v
e
r
y
85
80
75
70
65
y = 0.0875x + 54.483
R² = 0.826
60
55
50
0
100
200
300
400
500
600
Silver grade g/t
Figure 16.5
Silver Grade vs % Recovery when 75% Silver Recovered from Deslimed Clays
100
%
95
90
R
e
c
o
v
e
r
y
85
80
75
70
65
y = 0.0333x + 77.319
R² = 0.8299
60
55
50
0
100
200
300
Silver grade g/t
72
400
500
600
The final calculation, as shown in Table 16.7, estimates the mass yield, concentrate grades
and recoveries consistent with the predicted mill feed grade.
Table 16.7
Concentrate Grades and Recovery at Forecast Average Head Grade
Product
Mill Feed
Lead concentrate
Zinc concentrate
Tailings
16.1.2.3
Mass
Yield
dmt/d
1800
29
59
1713
Grade
Percent Recovery (%)
Ag g/t
%Pb
%Zn
Ag
Pb
Zn
154.2
6220
873
28.5
1.0
51.0
0.85
0.22
2.0
3.72
53.0
0.19
100.0
63.9
18.6
17.6
100.0
77.6
2.7
19.7
100.0
3.0
87.7
9.3
Deleterious Elements on Smelter Returns
The Pulacayo ore contains four deleterious elements which will decrease the value of the lead
concentrate. The composition of the lead concentrate from the medium grade sample
included arsenic (0.394%), copper (2.80%), antimony (3.44%), and zinc (4.19%). The zinc
concentrate had lower concentrations of these elements. The medium grade lead and silver
concentrates assays were used in the economic model and NSR calculations.
16.2
MINERAL PROCESSING
Mineral Processing options are described in Section 18 of this report.
73
17.0
17.1
MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES
INTRODUCTION
An initial drilling program during the 2008 field season at the Pulacayo project has been
successful in outlining a broad zone of relatively low grade disseminated- and stringer-style
silver-zinc-lead mineralization along a strike length of approximately 750 m and from
surface to a vertical depth of approximately 500 m. This mineralization was the subject of an
initial mineral resource estimate that was prepared prior to the completion of the drilling
program which used all drill hole data up to and including hole PUD-110 (Pressacco and
Shoemaker, 2008). This zone contains occasional higher grade intervals that consist of
relatively narrow sulphide-rich veins, stockwork breccias and breccia zones.
Drilling continued subsequent to the crystallization date of the initial mineral resource
estimate such that an additional 29 drill holes were completed, largely from a drill station
established within the existing underground workings, and targeted an area of higher grade
mineralization with the objective of providing sufficient in-fill information to improve the
confidence of the mineral resource estimate in that area. All of the core for these additional
29 drill holes was logged geologically and, due to budgetary constraints, the core from all
holes was assayed, except for six drill holes (PUD-134 to PUD-139, inclusive).
The initial mineral resource estimate completed in 2008 contemplated three possible
production scenarios including extraction of mineralized material by means of an open pit
mine only, and underground mine only and a combination of open pit and underground
mining. The objective of this updated mineral resource estimate is to provide a global
estimate of the tonnage and average grade of the mineralized material present using a
conceptual operational scenario in which mineralized material will be extracted by means of
underground mining methods, with zinc-silver and lead-silver concentrates being produced
either via an on-site concentrator or by means of a toll-milling agreement with an existing
concentrating facility located elsewhere in the region.
17.2
DESCRIPTION OF THE DATABASE
A digital database containing the additional 29 drill holes was provided by Apogee to Micon
wherein such information as collar location, down-hole survey, lithology, density
measurements and assays was stored in comma delimited format. This drill hole information
was modified slightly so as to be compatible with the format requirements of the GemcomSurpac v6.1.1 mine planning software and was merged into the existing drill hole database.
A number of additional tables were created during the process of developing a grade block
model of the mineralization found at Pulacayo to store such information as composite assays,
zone composites and assorted domain codes. A description of the revised database is
provided in Table 17.1.
In all, the database contains information for 138 drill holes that comprise surface-based drill
holes completed by Apex Silver and recent surface- and underground-based drill holes that
74
were completed by Apogee. A listing of the collar information for the initial 109 drill holes
was provided in Pressacco and Shoemaker (2008) and a listing of the collar information for
the additional 29 drill holes is provided in Appendix II.
Table 17.1
Summary of the Pulacayo Drill Hole Database as at October 14, 2009
Table Name
assay_capped
assay_raw
assay_raw2
collar
comp_1m
ddh_composites
styles
survey
translation
zone_flags
17.3
Data Type
interval
interval
interval
Table Type
time-independent
time-independent
time-independent
interval
interval
time-independent
time-independent
interval
time-independent
Records
3,805
0
17,443
138
4,172
76
5
706
0
154
TOPOGRAPHIC SURFACE
The topography in the Pulacayo area ranges in elevation from approximately 4,100 m to
4,500 m amsl and consists largely of rolling to steep-sided slopes along incised valleys. For
the most part the surficial materials are comprised of in-situ weathered, colluvial and alluvial
deposits with occasional low rock outcroppings.
A detailed topographic survey was carried out by Apogee in 2008 where each two-metre
contour and all important topographical features such as roads and shaft collars were
surveyed in using a total station and a series of reflecting prisms that were held in place by
field crews. In such a manner Apogee has generated a high quality topographical map for an
area that measures approximately 2,600 m in an east-west direction and 1,600 m in a northsouth direction.
17.4
HISTORICAL MINE WORKINGS
The Pulacayo project has a long history of mining as discussed in Chapter 6 above. This
mining activity has resulted in an extensive network of shafts, winzes, level development and
stoping dating back to the early 1800’s. Records relating to this historical mining activity are
available from such sources as private company files and from the offices of COMIBOL to
varying degrees of detail. In 2008, Apogee had conducted an initial review of the records
contained within the COMIBOL offices in La Paz and was successful in locating some level
plans for the upper levels, along with a vertical longitudinal projection depicting the mined
out areas that dates to 1945, corresponding to the end of the mining activities by the
Hochschild group. The mine was nationalized in 1952 and was operated by COMIBOL,
which Micon understands had the primary focus of conducting pillar recovery where
available and exploiting high grade areas of mineralization deeper in the mine. The mine
subsequently closed in 1959 and experienced a renewed period of activity when the
75
“cooperativas” (an informal collection of local individuals) began intermittent mining
activities that focussed on exploiting very narrow (on the order of 0.5 m), high grade
structures. Mining activities by the cooperativas continues to this day.
While still using the metric system of measurement, the entire historical underground
infrastructure was completed using a surveying co-ordinate system that is different from that
employed by Apogee. Digital copies of the level plans were provided to Micon which
proceeded to adjust the location of the workings to Apogee’s project grid using the location
of the shafts shown on the level plans and their corresponding locations on the new, highquality topographic survey as control points. Micon used the Gemcom-Surpac mine
modeling package to carry out this transformation and found that, for the most part, the
adjustment from the historical coordinate system to Apogee’s survey system required only a
simple shift of the northing and easting coordinates. A minor adjustment was required in
elevation, but no rotation was required. Comparison of the final location of the shafts against
the information provided from the detailed topographic surface suggested that the
transformed locations of the shafts are accurate to within 5 m.
It is to be noted that the source data for the model of the levels in the mine are the historical
level plans, which contain little to no information regarding the grade or slope of the levels.
In reality, the grade of the floor of the levels is typically excavated at a slight incline
(typically on the order of +0.5% to +1% for track-based underground mining operations) in
order to allow drainage of water. However, considering the age of the workings in the upper
levels, a strong possibility exists that transportation of the muck was carried out with the
assistance of pit-ponies where the floor of the drift can be established at steeper inclines.
Given the lack of detailed information regarding the inclination of the floor of the drifts, for
the purposes of the initial mineral resource estimate Micon assumed an inclination of zero
(i.e. a flat floor) for all of the levels modeled. As well, for the purposes of the initial mineral
resource estimate, Micon assumed a constant cross-sectional dimension of 3.0 m (width) x
3.7 m (height) for all of the modeled drifts on the basis of the results of examination of the
indicated drift widths on a number of the level plans. Should the project proceed to a more
advanced state, Micon recommends that the precise location and inclination of the levels be
established by detailed survey methods.
As well, it is to be noted that the level plans for three of the upper levels have not been
located (the 4252, 4282 and 4316 metre levels). Micon recommended that efforts continue to
be directed towards location of the records of these levels and integration of their results with
the remainder of the model of the mine workings.
In respect of the mined out areas, the only source of information that could be located was a
vertical longitudinal projection that was found in the files of COMIBOL (Figure 17.1). It can
be seen that sufficient information is contained to determine the location and extent of the
mined out areas relative to the levels, shafts and winzes. However, detailed examination of
the development on many levels reveals that the location of the stoped area cannot be
determined with confidence because of the presence of a number of parallel drifts that are
76
oriented along the strike of the mineralization, each of which could have been used as access
and haulage ways for extracted mineralized material. Because of this uncertainty, and until
further information is found to the contrary, Micon assumed that stoping was carried out for
each of the parallel drifts and that the stopes extended completely up to the next level above.
Figure 17.1
Vertical Longitudinal Projection of the Mined Out Areas as at 1945, Pulacayo Project
A major shortcoming of the longitudinal projection method of presentation is that the width
of the mined out stopes cannot be determined. The width of the modeled stopes was
estimated from a description of the mining presented in Ahlfeld and Schneider-Scherbina
(1964) that describes the widths of the stopes as ranging from 1.1 to a maximum 6 m. For
the purposes of this initial mineral resource estimate, Micon assumed a constant, average
stope width of 3 m for the model of the mined out voids. Micon recommends that should the
project proceed to a more advanced state, the shape and location of the mined out stopes be
determined by appropriate methods to an appropriate degree of accuracy.
Plan and longitudinal images of the resulting digital model of the mined out areas are
presented in Figure 17.2.
17.5
METAL PRICE SELECTION
The prices of zinc, lead and silver are cyclical, responding to the supply and demand
relationship and influenced to a degree by market speculation and technical analyses. The
metal prices have varied widely since the year 2000 and the prices for each of the three
metals have recently retreated from their former high levels. Given the cyclical nature of
metal prices it is not reasonable to utilize the metal price at any one point in time, as it is
certain that the price will change in the future.
77
Figure 17.2
Selected Views of Digital Models of Historical Workings, Pulacayo Project
78
While experience has shown that it is difficult at best to predict what the future metal prices
will be, a reasonable alternative is to utilize an average metal price over a time period rather
than using the metal prices at the close of any particular business day. In this manner a
degree of averaging is applied to the cyclical nature of the metals prices and longer-term
trends in the metal prices begin to be taken into account.
For the purposes of this mineral resource estimate, Micon has chosen to use the average
silver price of the 36-month period ending August 31, 2009, resulting in a value of $13.81/oz
(source: Kitco web site) and representing the trading range of the metal for that period.
Since the selection of this average metal price, in the period September 1, 2009 to March 31,
2010, the daily price fix of silver in London has varied between a low of $14.74/oz and a
high of $18.84/oz.
The prices of both lead and zinc have gone through a trough in late 2008 and early 2009 and
are now significantly above their lows. In the absence of a more formal metal price forecast,
Micon believes that an appropriate method of selection of metal prices for lead and zinc is to
examine the longest term forward contract price for each metal that is available on the
London Metal Exchange, as it is believed that these forward prices are the best reflection as
to where the industry as a whole believes that the metal prices will be during the period under
consideration. As of August 2009, the London Metal Exchange 27-month forward seller
price for zinc was $0.864/lb) and the 15-month forward seller price for lead was $0.859/lb).
17.6
DOMAIN MODELING
Based upon its experience in the preparation of the initial mineral resource estimate, Micon
concluded that the relationship of the mineralization to host lithology indicated that the
composition of the hosting lithologic units bears little to no influence upon the concentration
or distribution of the mineralization.
Examination of the metal distributions intersected in the drill hole assay data reveals that the
distribution of the metals varies widely from one sample to the next, such that a potential
economic return of any given sample can be achieved by any of the three metals in any given
sample. Consequently, for the updated mineral resource estimate, Micon proceeded to
prepare a domain model that attempted to represent the distribution of the mineralization that
exceeded estimated operational costs only.
For the purposes of this updated mineral resource estimate, Micon judged that the most
appropriate method to deal with the polymetallic nature of the mineralization was to apply a
Net Smelter Return (NSR) to the assay data. This method recognizes that more than one
metal can contribute to a potential revenue stream and proceeds to derive a factor that
accounts for such items as recovery to concentrate, metal prices payable fraction, penalties,
treatment and refining charges and freight. In this manner, a set of factors are derived that
convert the in-situ grades to net revenue for each metal. The revenue for each metal is
summed to arrive at a NSR value for a given sample.
79
Given that the exact values of many of these input parameters are not known at such an early
point in the project’s development, estimates were derived on the basis of the best available
information from a variety of sources including initial test work results and Micon’s
experience with current smelter terms for zinc and lead concentrates in the region. A
summary of these factors is provided in Table 17.2, however, due to confidentiality reasons,
details of the smelter terms cannot be disclosed.
Table 17.2
Summary of the Input Values and NSR Factors, Pulacayo Project
Item
Metal Price
Recovery to Concentrate
NSR Factor
Silver
$13.81/oz
31.3% to zinc conc
63.4% to lead conc
0.33 per g Ag
Zinc
$0.86/lb
87.6% to zinc conc
3.9% to lead conc
15.29 per % Zn
Lead
$0.86/lb
2.1% to zinc conc
88.8% to lead conc
13.87 per % Pb
The NSR value was then calculated for each sample within the assay database. For purposes
of construction of a domain model of potentially economic mineralization a nominal NSR
value of $40/t was applied as a modeling constraint on in-situ block values..
The NSR value was displayed on the drill hole traces and was used to establish the outline of
the mineralized zone on cross-sections that were spaced at 50 metre centres (viewing
windows of +/- 25 m). The locations of the mineralized contacts were “snapped” to the
observed location in the individual drill holes such that the sectional interpretations
“wobbled” in three dimensional space, to either side of the section plane. In all,
interpretation was carried out on 19 cross-sections along a strike length of 950 m and to a
depth of approximately 450 m, and the resulting “wobbly polylines” were then linked
together to form a three-dimensional solid of the mineralized zone (Figures 17.3 and 17.4).
Examination of drill core revealed that a cap of oxidized material was present throughout the
project area. Based upon visual observations, Micon views the effect of this oxidation as
altering original silver-zinc-lead-bearing hypogene minerals to their oxide, carbonate or
sulphate equivalents. Given that no metallurgical test work has been completed on this
material, on the basis of its experience, Micon believes that the presence of the oxidation in
the mineralized zone will have a negative effect upon the metal recoveries and the quality of
the resulting concentrates. To that end, a model of the oxide-sulphide transition was created
from drill hole data and was used to code the block model accordingly.
During the course of preparing the cross sectional interpretation of the NSR domain, a
number of intervals were noted in the drill holes for which no assay information was
available. In some cases these non-sampled intervals fell inside the NSR domain model, and
so were likely to result in an estimation error on a local basis. Micon elected to adopt a
conservative approach and assumed that the non-sampled intervals that lay within the NSR
domain contained zero metal values. In all, two drill holes were affected (PUD024: 227.40234.60 m and DDH PUD074: 345.82-351.00 m, 351.71-354.70 m, 355.24-358.00 m).
80
Figure 17.3
Plan and Longitudinal Views of the Nominal $40/t NSR Solid
81
Figure 17.4
Cross Section 740300E Showing the Outline of the Nominal $40/t NSR Domain Model
82
17.7
TREND ANALYSIS
An analysis of the trends of the various components of the mineralization such as silver, zinc
and lead grades was conducted to assist in the understanding of the spatial distribution and
any zonation of these items within the limit of the mineralized domain.
In order to prepare longitudinal views of the metal distribution, the composite silver, zinc
and lead grades contained within the three-dimensional model of the mineralized zone were
extracted from the database using the Composite by Geology function of the GemcomSurpac software. The resulting data points were projected in longitudinal view for treatment
and analysis. Longitudinal views of the contoured silver, zinc, lead grades, along with the
contoured NSR values are presented in Figures 17.5, 17.6, 17.7, and 17.78 respectively.
Figure 17.5
Contoured Silver Values for the Nominal $40/t NSR Domain Model, Pulacayo Project
83
Figure 17.6
Contoured Zinc Values for the Nominal $40/t NSR Domain Model, Pulacayo Project
Figure 17.7
Contoured Lead Values for the $40/t NSR Domain Model, Pulacayo Project
84
Figure 17.8
Contoured NSR Values for the Nominal $40/t NSR Domain Model, Pulacayo Project
Although a number of small-scale trends are evident in these images at varying orientations,
the overall trends for the distribution of silver, zinc and lead are generally parallel to the
strike of the mineralizing system (i.e. azimuth 100°) with a horizontal plunge (i.e. no plunge
or rake within the plane of the mineralized system).
17.8
GRADE CAPPING
Examination of the drill core and the raw assay results indicates that high grade samples are
present within the data set that typically are associated with narrow veins/veinlets of semimassive to massive sulphides. These veinlets typically have limited vertical and lateral
continuity, consequently Micon elected to limit the influence of the high grade values by
capping of the grades.
All of the raw samples contained within the mineralized domain were coded and extracted
from the database for examination. The descriptive statistics of these samples are provided
in Table 17.3 and frequency histograms are presented in Figures 17.9, 17.10, 17.11 and
17.12.
85
Table 17.3
Summary Statistics for Raw Samples Contained within the Mineralized Domain Model
Item
Arithmetic Mean
108.81
AgCap
1,8000
97.99
Length-Weighted Mean
100.55
92.94
0.76
0.76
1.70
1.67
0.03
0.03
Standard Error
6.34
3.79
0.03
0.02
0.04
0.03
0.00
0.00
Median
19.00
19.00
0.36
0.36
1.20
1.20
0.01
0.01
Mode
6.00
6.00
0.00
0.00
0.00
0.00
0.01
0.01
390.77
233.77
1.59
1.47
2.21
1.99
0.11
0.08
3.59
2.39
1.94
1.81
1.22
1.12
3.25
2.55
3.89
2.52
2.09
1.94
1.30
1.19
3.53
2.66
Standard Deviation
Coefficient of
Variation-Arithmetic
Coefficient of
Variation-Weighted
Sample Variance
Ag(g/t)
Pb (%)
Zn (%)
0.82
PbCap
15
0.81
1.82
ZnCap
11.5
1.78
Cu
(%)
0.04
CuCap
1.0
0.03
152,704.21
54,650.40
2.54
2.15
4.90
3.97
0.01
0.01
Kurtosis
342.71
24.46
73.27
35.11
18.21
7.63
368.81
56.75
Skewness
15.37
4.59
6.84
5.15
3.51
2.51
15.50
6.63
10,000.00
1,800.00
28.70
15.00
23.20
11.50
3.29
1.00
Minimum
0.00
0.00
0.00
0.00
0.00
0.00
0.01
0.00
Maximum
10,000.00
1,800.00
28.70
15.00
23.20
11.50
3.29
1.00
Sum
414,028.65
372,854.35
3,126.47
3,087.98
6,916.02
6,777.57
120.52
115.14
3,805
3,805
3,805
3,805
3,805
3,805
3,424
3,535
Range
Count
Figure 17.9
Silver Frequency Histogram for Samples within the Mineralized Domain, Pulacayo Project
Upper Tail Histogram of Raw Silver Assays Within the Mineralized Domain Model
Pulacayo Project (n=3,805)
100
90
80
Frequency
70
60
50
40
Grade Cap = ~1,800 g/t Ag
(18 Samples)
30
20
10
0
Ag (g/t)
86
Figure 17.10
Zinc Frequency Histogram for Samples within the Mineralized Domain, Pulacayo Project
Upper Tail Histogram of the Zinc Raw Assays Within the Mineralized Domain Model
(n=3,805)
500
450
400
Frequency
350
300
250
200
150
100
Grade Cap = 11.5% Zn
(32 Samples)
50
0
Zn (%)
Figure 17.11
Lead Frequency Histogram for Samples within the Mineralized Domain, Pulacayo Project
Upper Tail Histogram of Lead Raw Assays Within the Mineralized Domain Model
(n=3,805)
200
180
160
Frequency
140
120
100
80
60
Grade Cap = 15% Pb
(13 Samples)
40
20
0
Pb (%)
87
Figure 17.12
Copper Frequency Histogram for Samples within the Mineralized Domain, Pulacayo Project
Upper Tail Histogram of the Copper Raw Assays Within the Mineralized Domain Model
(n=3,424)
1000
900
800
Frequency
700
600
500
400
300
Grade Cap = 1.0% Cu
5 Samples
200
100
0
Cu (%)
Based upon the distribution of the silver, zinc and lead grades, Micon believes that 1,800 g/t
Ag, 11.5% Zn and 15% Pb are appropriate capping values for this mineralized domain. The
descriptive statistics of the capped samples are provided in Table 17.3.
17.9
COMPOSITING METHODS
The selection of an appropriate composite length for samples contained within the
mineralized domain model began with an examination of the distribution of the sample
lengths within the domain model (Figure 17.13). The sample lengths ranged from a
minimum of 0.2 m to a maximum of 68 m in length, with many samples being 1.0 m in
length. Consequently, Micon elected to utilize a composite length of 1.0 m in consideration
of the relationship between composite length and block size.
All samples were composited to an equal length of 1.0 m using the down-hole compositing
function of the Surpac-Gemcom mine modeling software. In this function, compositing
begins at the point in a drill hole at which the zone of interest is encountered and continues
down the length of the hole until the end of the zone of interest is reached. As often happens,
the thickness of the mineralized zone encountered by any given drill hole is not an equal
multiple of the composite length. In these cases, if the remaining length was 75% or greater
of the composite length (in this case 0.75 m), the composite was accepted as part of the data
set. The remaining sample lengths less than 75% of the composite length were retained for
consideration so as to provide as accurate a grade estimate for the footwall margins of the
88
domain model as possible. A comparison of the descriptive statistics for the capped and
uncapped sample values for the composited data is presented in Table 17.4.
Figure 17.13
Sample Length Histogram for Samples within the Mineralized Domain, Pulacayo Project
Histogram of Sample Lengths for Samples Within the Mineralized Domain Model
3500
3000
Frequency
2500
2000
1500
1000
500
0
0.25
0.5
0.75
1
1.25
1.5
1.75
2
2.25
2.5
2.75
3
More
Sample Length (m)
Table 17.4
Summary Statistics for 1.0 m Composite Samples Contained within the Mineralized Domain Model
Item
Mean
Standard Error
Median
Mode
Standard Deviation
Coefficient of
Variation
Sample Variance
Kurtosis
Skewness
Range
Minimum
Maximum
Sum
Count
Ag_gt
102.16
5.01
19.00
0.00
323.50
AgCap
1,800
94.39
3.37
19.00
0.00
217.71
Pb %
0.77
0.02
0.36
0.00
1.40
PbCap
15
0.77
0.02
0.36
0.00
1.33
Zn %
1.72
0.03
1.15
0.00
2.04
ZnCap
11.5
1.69
0.03
1.15
0.00
1.87
Cu %
0.04
0.00
0.01
0.01
0.11
CuCap
1.0
0.03
0.00
0.01
0.01
0.08
3.17
104,651.59
325.98
14.07
9,808.80
0.00
9,808.80
426,212.63
4,172
2.31
47,396.19
24.15
4.51
1,800.00
0.00
1,800.00
393,807.36
4,172
1.81
1.95
59.00
6.03
26.70
0.00
26.70
3,221.77
4,172
1.73
1.76
32.93
4.91
15.00
0.00
15.00
3,194.19
4,172
1.18
4.17
17.13
3.32
23.20
0.00
23.20
7,192.83
4,172
1.10
3.49
7.65
2.45
11.50
0.00
11.50
7,067.10
4,172
2.98
0.01
335.68
14.44
3.23
0.00
3.23
127.88
3,574
2.44
0.01
58.96
6.59
1.00
0.00
1.00
118.27
3,759
89
17.10 BULK DENSITY
Apogee collected information regarding the bulk density (specific gravity) on a systematic
basis for all of the recent drilling programs. The density of the core was determined by the
core technicians using the Archimedes method on selected samples of core. The resulting
information was transferred to an Excel spreadsheet where the host lithology of the measured
sample was paired with the specific gravity determination. The resulting spreadsheet was
provided to Micon which proceeded to extract those specific gravity measurements that were
contained within the $40/t NSR domain model and determined the average densities of the
resulting data set. In all, 2,744 specific gravity measurements were included within the $40/t
NSR mineralized domain. A frequency histogram displaying the distribution of the specific
gravity measurements is presented in Figure 17.14.
It can be seen that the average specific gravity for the mineralization contained within the
$40/t NSR domain model is 2.40 t/m3.
Figure 17.14
Specific Gravity Histogram for Samples Within the Mineralized Domain, Pulacayo Project
Histogram of Bulk Densities for All Samples Contained Within the Mineralized Domain Model
(n=2,744)
350
Mean Density = 2.40 tonnes/m3
300
Frequency
250
200
150
100
50
0
1.1
1.2
1.3
1.4
1.5
1.6
1.7
1.8
1.9
2
2.1
2.2
2.3
2.4
2.5
2.6
2.7
2.8
2.9
3
3.1
3.2
3.3
3.4
3.5
3.6
3.7
3.8
3.9
4
More
Density
17.11 VARIOGRAPHY
The analysis of the variographic parameters of the mineralization found in the mineralized
domain at the Pulacayo deposit began with the construction of down-hole variograms using
the capped, 1.0 metre composited sample data with the objective of determining the global
nugget (C0) for silver, zinc and lead. Considering the low average grade of copper that is
found within the mineralized domain, the copper grades were not modeled.
The down-hole variogram results were confirmed by construction of omni-directional
variograms, and good model fits were obtained. An evaluation of any anisotropies that may
90
be present in the data was successful in the generation of variograms for the three principal
directions for each of the three metals. A summary of the variographic parameters is
presented in Table 17.5, and the variograms are presented in Appendix III.
Table 17.5
Summary of Variographic Parameters for 1.0 m Composite Samples, Pulacayo Project
Item
Variogram Type
Nugget (Downhole)
Sill (C1-Downhole)
Range (m)
Nugget (OmniDirectional)
Sill (C1-OmniDirectional)
Range (m)
Major Axis:
Orientation
Angular Tolerance
Sill (C1)
Range (m)
Semi Major Axis:
Orientation
Angular Tolerance
Sill (C1)
Range (m)
Minor Axis:
Orientation
Angular Tolerance
Sill (C1)
Range (m)
Major Axis (Pass 2, Short
Range)
Semi-Major Axis
Minor Axis
Major/Semi-Major Ratio
Major/Minor Ratio
Number of Points
Range for Pass 1 (Long
Range)
Minimum Number of Points
Maximum Number of Points
Search Ellipse Type
Silver (D2)
Spherical
NUGGET:
33,052
21,086
21
Lead (D4)
Spherical
Zinc (D6)
Spherical
1.05
0.71
7
2.10
1.38
18
0.88
0.77
6
1.95
1.55
17
-30°  280°
45°
24,058
64
-40°  280°
30°
1.08
67
-50°  280°
45°
1.73
59
+60°  280°
45°
23,788
20
+50°  280°
30°
1.82
24
+40°  270°
45°
1.56
57
29,727
26,893
22
ANISOTROPIES:
0°  010°
0°  010°
45°
30°
4,590
0.33
4
4
SEARCH ELLIPSE:
65m@280°(-30°) 65m@280°(-40°)
60m@280°(-50°)
20m@280°(+60°)
5m@010°(0°)
3.2
13
4,221
140m
25m@280°(+50°)
5m@010°(0°)
2.6
13
4,221
140m
60m@280°(+40°)
5m@010°(0°)
1.0
12
4,221
140m
2
8
Quadrant
2
8
Quadrant
2
8
Quadrant
0°  010°
45°
0.28
3
17.12 BLOCK MODEL CONSTRUCTION
A simple, upright, whole-block model with the long axis of the blocks measuring 10 m
(strike) x 10 m (height) x 2 m (width) and oriented along an azimuth 100° was constructed
91
using the Gemcom-Surpac version 6.1.1 mine planning software package using the
parameters presented in Table 17.6. A number of attributes were also created to store such
information as metal grades by the various interpolation methods, distances to and number of
informing samples, domain codes, oxidation state, mined out status, and resource
classification codes. The block dimensions were selected primarily in an attempt to have
relevance to the selection of underground mining methods. These block dimensions may
require revision at a later date as new information permits the identification of appropriate
mining methods, or should the project scope become better defined.
Table 17.6
Summary of Block Model Parameters, Pulacayo Project
Type
Minimum Coordinates
Maximum Coordinates
User Block Size
Min. Block Size
Rotation
Y (Northing)
7744100
7745100
2
2
10.000
Attribute Name
zn_kvar
ag_avgdist
Type
Real
Real
Decimals
3
1
Background
-99
0
ag_cap_id2
ag_cap_nn
ag_cap_ok
ag_id2_nosample
Real
Real
Real
Integer
2
2
2
-
0
0
0
0
ag_id2_nosample
_pass2
ag_id2_nsr
Integer
-
0
Real
-
0
ag_id2_pass_no
ag_kvar
ag_nearest
Integer
Real
Real
2
1
0
0
0
ag_ok_nosample
Integer
-
0
block_nsr_id2
classification
density
lith_code
mined_out
oxidation_code
pb_avgdist
Real
Integer
Real
Integer
Integer
Character
Real
2
1
0
0
2.28
100
1
sulf
0
pb_cap_id2
pb_cap_nn
pb_cap_ok
pb_id2_nsr
Real
Real
Real
Real
2
2
2
-
0
0
0
0
92
X (Easting)
739700
740900
10
10
0.000
Z (Elevation)
3900
4600
10
10
0.000
Description
Average Distance of Informing Samples,
Silver
Silver by Inverse Distance, Squared
Silver by Nearest Neighbour
Silver by Ordinary Kriging
Number of Informing Samples, Silver,
Inverse Distance
Number of Informing Samples, Long
Range Pass
Silver NSR from ID2 Grade (Ag_gt *
0.37)
1=Short Range, 2=Long Range
Kriging Variance, Silver
Distance to Nearest Informing Sample,
Silver
Number of Informing Samples, Silver
OK
Ag_id2_nsr + Zn_id2_nsr + Pb_id2_nsr
1=Measured, 2=Indicated, 3=Inferred
Rock=2.28, Air=0
403=$14 NSR Domain
0=Mined Out (Void), 1=In Situ
OX=oxidized, SULF=unoxidized
Lead
Lead by Inverse Distance, Squared
Lead by Nearest Neighbour
Lead by Ordinary Kriging
Lead NSR from ID2 Grade (Pb_pct *
22.64)
Attribute Name
pb_kvar
pb_nearest
Type
Real
Real
Decimals
3
1
Background
-99
0
pb_nosample_pas
s1
pb_nosample_pas
s2
pb_pass_no
zn_avgdist
Integer
-
0
Integer
-
0
Integer
Real
1
0
0
zn_cap_id2
zn_cap_nn
zn_cap_ok
zn_id2_nsr
Real
Real
Real
Real
2
2
2
-
0
0
0
0
zn_id2_pass_no
zn_nearest
Integer
Real
1
0
0
zn_nosample_pas
s1
zn_nosample_pas
s2
Integer
-
0
Integer
-
0
Description
Lead
Number of Informing Samples for Pass
1, Lead
2, Lead
Zinc
Zinc by Inverse Distance, Squared
Zinc by Nearest Neighbour
Zinc by Ordinary Kriging
Zinc NSR from ID2 Grade (Zn_pct *
14.07)
Zinc
1, Zinc
2, Zinc
Metal grades were interpolated into the individual blocks for the mineralized domain initially
using the variogram ranges and parameters presented in Table 17.5 above, after which it was
apparent that the density of the drill hole information was not sufficient to provide a full fill
of all the blocks. Consequently, a two-pass approach was taken in order to achieve a filling
of most of the blocks contained within the domain models. In this approach a first pass
interpolation is carried out using a long range of 140 m for the search ellipse in order to
provide as complete a filling of the blocks as possible. This is followed by a shorter range
pass that uses the search ellipse parameters derived from the variographic analysis that reinterpolates and overwrites the grades of the blocks that are located closer to the informing
samples. The interpolation was carried out using Ordinary Kriging (OK), Inverse Distance
Squared (ID2) and Nearest Neighbour (NN) interpolation methods for silver, zinc and lead.
During the course of these interpolation runs, such additional information as the pass
number, distance to the nearest informing sample, average distance of informing samples,
number of informing samples per block and the kriging variance was also recorded for each
block. Hard domain boundaries were used in which only data contained within the $40/t
NSR domain model were allowed to be used to estimate the grades of the blocks, and only
those blocks within the domain limits were allowed to receive grade estimates.
Subsequent to interpolation of the block densities and metal grades, the densities of those
blocks that fell within the model of the mined out stopes was set to zero on the assumption
that all mined stopes do not contain any backfill.
93
17.13 BLOCK MODEL VALIDATION
Validation efforts for the mineral resource estimate for the Pulacayo deposit began with a
comparison of the average block grades for the capped and uncapped metal values against the
respective informing composite samples. As well, the volumes reported from the block
model were compared to the volumes of the solid model of the $40/t NSR mineralized
domain. The reconciliation is presented in Table 17.7. It can be seen that there is a good
correlation for the average block grades estimated using the three interpolation methods, and
between the average estimated block grades with the informing composite samples. As well,
there is a good fit between the reported volumes for the mineralized domain model, with the
block model reporting slightly less volume in comparison to the original solid model. It is to
be noted that this reconciliation report compares the volume and grades inside the
mineralized domain model against the informing data and is not corrected for the mined out
material. In contrast, the tonnage reported in Table 17.8 takes into account the ‘zero’ density
of the mined-out stopes.
Table 17.7
Block Model Validation Results, Pulacayo Project
Volume
Tonnes
5,261,400
11,832,000
5,256,544
Ag Nocap
Ag Cap
Pb Nocap
Pb Cap
Zn Nocap
Block– Model - Inverse Distance, Power 2
95.04
86.82
0.80
0.79
1.62
Block– Model - Ordinary Kriging
95.18
87.28
0.80
0.80
1.63
Block– Model - Nearest Neighbour
102.18
92.87
0.82
0.81
1.66
1m Composites
102.16
94.39
0.77
0.77
1.72
Solid Volume
Block model is reporting +4,856 m3 (<1% difference in volumes)
Zn Cap
1.60
1.61
1.64
1.69
17.14 MINERAL RESOURCE CLASSIFICATION CRITERIA
The mineral resources in this report were estimated in accordance with the definitions
contained in the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on
Mineral Resources and Reserves Definitions and Guidelines that were prepared by the CIM
Standing Committee on Reserve Definitions and adopted by the CIM Council on December
11, 2005.
The mineralized material was classified into either the Indicated or Inferred mineral resource
category after consideration of the following factors:

Interpreted continuity of the mineralization as a function of the existing drill hole
density

The ranges derived from the variographic analysis of the silver grades and resulting
search ellipse parameters
94

The relationship to the model of the depth of oxidation

The proximity of the mineralization to the mined out stope models.
Those blocks which received interpolated grades below the model of the oxidation surface
that were within the silver variogram ranges were classified as Indicated mineral resources
(i.e. those blocks informed with the short-range pass), while the remaining blocks were
classified into the Inferred mineral resource category.
All material that was contained within the model of the mined out stopes was considered to
have been at a density of zero (i.e. the stopes were assumed to be filled with air), resulting in
no tonnages being ascribed to the blocks that are contained therein. In addition, in an attempt
to address the confidence level relating to the information regarding the presence and shape
of the mined out areas, an envelope of 10 m beyond the modeled stope limits was created
around each of the mined out areas, and all mineralized material between the modeled stope
outlines and this envelope was assigned to the Inferred mineral resource category. As well,
due to the lack of information regarding the three upper levels as discussed above, and in
particular a lack of information relating to the stope indicated on the 4,252-m level, a solid
model with the cross-sectional shape of this stope as indicated on the longitudinal section
was created and was projected across the entire width of the mineralized zone. All material
inside this solid was then assigned to the Inferred mineral resource category.
In addition, the eastern limits of the mineralized domain have been assigned to the Inferred
resource category due to the limited amount of drill hole information in this area.
17.15 RESPONSIBILITY FOR THE ESTIMATE
The estimate of the mineral resources present at the Pulacayo deposit was prepared by Reno
Pressacco, M.Sc.(A)., P.Geo., who is independent of Apogee.
17.16 MINERAL RESOURCE ESTIMATE
The mineral resources for the Pulacayo deposit are reported in Table 17.8. Underground
mineral resources include all blocks that are below the oxidized surface, that are not flagged
as mined out, and that are contained within the $40/t NSR domain model (Figure 17.15).
The report is prepared using the capped, ordinary kriged average grade estimates for silver,
zinc and lead. The estimated uncapped average grades for silver, lead and zinc are also
provided for comparison purposes in Table 17.9.
There is a degree of uncertainty to the estimation of mineral reserves and mineral resources
and corresponding grades being mined or dedicated to future production. The estimating of
mineralization is a somewhat subjective process and the accuracy of estimates is a function
of the accuracy, quantity and quality of available data, the accuracy of statistical
computations, and the assumptions used and judgments made in interpreting engineering and
geological information. There is significant uncertainty in any mineral resource/mineral
reserve estimate, and the actual deposits encountered and the economic viability of mining a
95
deposit may differ significantly from these estimates. Until mineral reserves or mineral
resources are actually mined and processed, the quantity of mineral resources/mineral
reserves and their respective grades must be considered as estimates only. In addition, the
quantity of mineral reserves and mineral resources may vary depending on, among other
things, metal prices. Fluctuation in metal or commodity prices, results of additional drilling,
metallurgical testing, receipt of new information, and production and the evaluation of mine
plans subsequent to the date of any mineral resource estimate may require revision of such
estimate.
Micon has considered the mineral resource estimates in light of known environmental,
permitting, legal, title, taxation, socio-economic, and other relevant issues and believes that
the mineral resources will not be materially affected by these items.
Given the early stage of development of the Pulacayo deposit, few recent studies have been
completed that examine whether the mineral resources may be materially affected by mining,
infrastructure, marketing, political or other relevant factors. Preliminary metallurgical testing
has been completed which indicates favourable recoveries of silver, zinc and lead. The
effective date of this estimate is October 14, 2009.
Table 17.8
Classification
Indicated
Inferred
Tonnes
4,892,000
6,026,000
Ag (g/t)
79.96
98.26
Pb (%)
0.79
0.78
Zn (%)
1.64
1.68
(2) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources
may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues.
(3) The quantity and grade of reported inferred resources in this estimation are conceptual in nature and there has been insufficient
exploration to define these inferred resources as an indicated or measured mineral resource. And it is uncertain if further
exploration will result in upgrading them to an indicated or measured mineral resource category.
Table 17.9
Comparison of Capped vs Uncapped Grades, Pulacayo Deposit
Classification
Tonnes
Indicated (2)
Inferred (3)
4,892,000
6,026,000
Ag Ok
Nocap
90.24
105.42
Ag Ok
Cap
79.96
98.26
96
Pb Ok
Nocap
0.79
0.79
Pb Ok
Cap
0.79
0.78
Zn Ok
Nocap
1.67
1.70
Zn Ok
Cap
1.64
1.68
Figure 17.15
Longitudinal and Isometric Views of the Mineral Resources, Pulacayo Project
97
18.0
OTHER RELEVANT DATA AND INFORMATION
This section of the Technical Report summarises the results of the mining, processing and
other technical work carried out by Micon and which supports the preliminary economic
assessment of the project.
18.1
MINING
Micon has been asked to conduct an underground mining study to determine an appropriate
underground mining method, a corresponding production rate and to evaluate the economics
of mining the Pulacayo deposit. The underground study is at a scoping level to an accepted
level of accuracy of +-30%. General contingency is applied to mining capital in the DCF
model, where an overall contingency of 30% has been assumed.
No detailed geotechnical assessment is available or has been made by Micon. However, from
examination of a selection of drill-core Micon has assumed that the ground conditions will
range from moderate to good in areas of un-mined ground. From visual inspection, during a
tour through the accessible areas of the 4,129 m level, it can be assumed that the ground
conditions will be potentially de-graded where areas of alteration intersect historical
workings. In these areas, any ground support that was previously installed is likely to consist
of wooden props which may have degraded with time. Based on this evidence, Micon has
assumed that ground control will normally be achieved through rock-bolting. In areas of
significant alteration, bolting, limited shotcreting and possibly mesh may be required. The
San Leon adit is self-draining and, therefore, so are the workings above this level. Below the
4,129-m level, the mine is flooded.
18.1.1
Mining Method and Design
Sub-level open stoping (SLOS) with backfill is the mining method which Micon considers
most suitable for underground mining at Pulacayo. The average value of the resource justifies
the use of backfill as opposed to leaving pillars in-situ. SLOS mining with backfill also gives
a reduced risk of surface subsidence. SLOS is a more productive method, even in relatively
narrow stopes, when compared to cut and fill mining.
Longhole stoping will use vertical long holes drilled either upwards or downwards into the
stopes. The sub-level spacing will be 25 m, which leaves an ‘unsupported’ and drillable stope
height of 20.5 m, once the 4.5-m high access drive has been mined and supported. This
spacing fits logically with the existing development. A combination of both up-holes and
down-holes drilled blind and to break-through will be required for production.
Due to the presence of remnant mining voids, the prudent use of backfill and maximum
drilling flexibility will be required to maximize recovery of the mineable resource. A
comprehensive survey of the position and nature of the existing voids will need to be made in
advance of mine planning. A suitably long development lead will be required to allow for
surveying of these voids and to give mine planners sufficient time to carry out short term
98
planning based on this information. There will be an abundant supply of development waste
rock generated during the development of the mine infrastructure. This can be used as either
cemented or un-cemented rock-fill; to fill pre-existing voids or enable the mining of
secondary stopes prior to the tailings backfill system being commissioned.
Sub-levels will be mined in an overhand method to allow the required strength of the fill to
be minimized and allow un-cemented fill to be used where possible. As several production
horizons will be mined simultaneously, there will a requirement to leave a small number of
sill pillars. These will be recovered with the aid of backfill.
18.1.2
Mine Development and Production Schedule
The mine is accessible through the San Leon Adit (4,130m RL), which has a nominal arched
profile of 2.2 m (high) by 2.0 m (wide) and at a minimum drainage gradient. It is Micon’s
opinion that the most efficient route of access and truck haulage is through the adit. This is
on account of the undulations of the terrain making alternative access points either more
inaccessible or more expensive to develop. However, to minimize the environmental impact
to the village of Pulacayo, it is proposed that a new portal and a 710 m long adit be
developed. This adit will start close to the new processing plant and intersect the San Leon
adit at a point approximately 380 m in-bye from its existing south portal. From this point, in
a north-east direction the San Leon adit will be slashed and re-supported to a final dimension
of 4.5 m (high) by 4.0 m (wide) and for a distance of 420 m (see Figure 18.1). It is intended
that the enlarged part of the San Leon adit and the new adit are used as the route for ore and
waste haulage out of the mine. It will also be used as the main access for men and materials.
It is understood that the northern route out of the San Leon adit is trafficable for pedestrians
and suitably small vehicles. Micon has planned that this will become the second means of
egress out of the mine in the event of an emergency. Limited rehabilitation and making safe
may be required to make this possible. The new adit will be designed on minimum drainage
gradient.
The San Leon adit currently houses a potable water supply pipe which supplies the village of
Pulacayo, provision will need to be made to re-route this line during the mine development
works.
Once the San Leon adit has been slashed, two new inclined ramps and two decline ramps are
planned. They will access the ore above and below the 4,130 m level, respectively. The
incline ramps will be developed from the enlarged San Leon adit, starting from the FW to the
south of the ore body. The decline ramps will be developed from the enlarged San Leon adit,
starting from the HW to the north of the ore body. Inclines and declines will be driven at a
maximum gradient of 14% and will have the profile of the enlarged San Leon adit. This
layout has been designed in order to optimize the development schedule and to prevent the
main infrastructure from intersecting the mined out stopes. The incline ramps will be
developed by a contractor between years -3 to -1, as well as several of the production levels.
The remaining infrastructure will be developed by the operator of the mine.
99
Figure 18.1
Plan View of the Existing Development Working and the Planned Mine Infrastructure
It is planned that ventilation air will be exhausted through multiple vent raises to surface.
Main fans will be located on the surface end of each raise. Intake air will be drawn in through
the north and south ends of the San Leon adit. This will ensure that both means of egress are
situated in intake air. Several 2.4-m diameter raise bored ventilation raises have been
budgeted for, they will be tied into the workings at the east and west end of the
mineralisation on each level (see Figure 18.2).
Micon has reviewed the potential for mining at rates of between 1,000 and 1,800 t/d. The
base case mine production rate has been selected to be 1,800 t/d and it has been assumed that
the mine will work 360 production days per year, which equates to a production of 648,000
tonnes per annum. Micon considers that this production rate is the maximum achievable from
the known resource.
100
Figure 18.2
Isometric View Looking North and Showing the Resource Model the Planned Development and the
Mined Out Areas
The silver equivalent value of each block (Ag(g/t) Eq.) was determined by application of the
following formula to the undiluted block grades:
Ag(g/t) Eq.=Ag(g/t)+([Pb% x 2204.622 x $0.983/lb] + [Zn% x 2204.622 x $1.05/lb])
/{$14.66/31.1034}
where:
Silver price ($/oz)
Lead price ($/lb)
Zinc price ($/lb)
Troy ounce
Metric tonne
$14.66
$0.983
$1.05
31.1034 g
2204.622 lb
Due to the preliminary nature of the study and the uncertainty in the geometry, size and
position of the mined out voids, stope outlines were not designed.
The mineable portion of the mineral resource was determined at a number of silver
equivalent cut-off grade values, ranging from 125 g/t to 275 g/t Ag Eq (Table 18.1 and
Figure 18.3). Appropriate mining dilution and recovery factors were estimated by
examination of the resulting block geometry and continuity on a level by level basis. A
tonnage-weighted average for each value was then calculated. For silver equivalent cut-off
grades of 225 g/t and above, the total mining dilution was estimated to be 24 % (at a grade of
30% of in-situ) and mining recovery to be 72%. This results in a grade factor of 83%. For
cut-off grades of 200 g/t and below, greater continuity of the payable blocks results in a
slightly improved grade factor.
101
Table 18.1
Mineral Resources above Silver Equivalent Cut off Values of 125 to 275 g/t Ag Eq.
Mineral Resource with mining dilution and recovery factors applied
COG
Resource
Ag Ok Cap
Pb Ok Cap
Zn Ok Cap
(g/t Ag Eq)
(t 000)
(g/t)
(% Pb)
(% Zn)
Inferred
4,098
125.0
114.8
0.8
1.7
3,496
150.0
129.2
0.9
1.8
2,924
175.0
145.9
1.0
1.9
2,456
200.0
162.2
1.0
1.9
1,910
225.0
179.6
1.1
2.0
1,605
250.0
199.4
1.2
2.0
1,348
275.0
220.6
1.3
2.1
Indicated
125.0
150.0
175.0
200.0
225.0
250.0
275.0
3,006
2,530
2,116
1,793
1,414
1,186
1,015
101.4
114.1
129.0
143.4
157.5
174.3
189.8
0.9
0.9
1.0
1.0
1.1
1.1
1.2
1.8
1.9
2.0
2.1
2.1
2.2
2.2
Ag Eq
(g/t)
236.0
256.9
280.7
304.5
328.1
353.2
379.3
228.1
249.5
272.0
292.8
312.2
334.5
354.4
4,500
450
4,000
400
3,500
350
3,000
300
2,500
250
2,000
200
1,500
150
Tonnes
1,000
100
Ag Eq 500
50
0
0
125
150
175
200
225
250
Cut‐off Grade (g/t Ag Eq.)
102
275
Silver Equivalent Grade (g/t)
Tonnes (000)
Figure 18.3
Grade-tonnage Curve for Mineral Resource vs Silver Equivalent Cut-off Grade
These results were then compared using the cash flow model to determine the optimum cutoff for each processing scenario. Details of this are provided in Section 18.5.7. For the base
case, a cut-off grade of 200 g/t Ag Eq. was selected, giving a LOM production forecast as
shown in Table 18.2.
Table 18.2
Base Case LOM Production at a Cut off Value of 200 g/t Ag Eq.
Class
Resource
t'000
1,793
2,456
Indicated
Inferred
Ag
g/t
143.4
162.2
Pb
%
1.0
1.0
Zn
%
2.1
1.9
Ag Metal
kg
257,000
398,300
Pb Metal
t’000
18.83
25.30
Zn Metal
t’000
36.94
47.40
A production schedule was calculated using the average grade of the mineable portion of
each category of mineral resource, and apportioning the production pro-rata with estimated
tonnage in each of the resource categories. An allowance was made for a three year ramp-up
in production (pre-production Years -2 and -1, and production Year 1), after which it is
assumed a steady rate of 1,800 t/d will be maintained. The production schedule can be seen
in Table 18.3. The small amount of ore arising from development in years -2 and -1 will be
stockpiled. Production is initially planned to be from the levels above 4,130 m, to allow
adequate time for de-watering of the lower levels, also as the upper levels contain fewer
mined out voids and therefore offer reduced mine planning and operational complexity.
However, the ore below 4,130 m does have a higher unit value. Therefore, future studies may
benefit from optimisation of the development and production schedule, once more
information is obtained on the position and nature of the mined out voids and the condition of
the flooded levels.
Table 18.3
Base Case LOM Production Schedule
Indicated (t 000)
Silver g/t
Zinc %
Lead %
Inferred (t 000)
Silver g/t
Zinc %
Lead %
Yr-3
-
-
Yr-2
14
143.4
2.1
1.0
19
162.2
1.9
1.0
Yr-1
82
143.4
2.1
1.0
112
162.2
1.9
1.0
Yr1
178
143.4
2.1
1.0
243
162.2
1.9
1.0
Yr2
273
143.4
2.1
1.0
375
162.2
1.9
1.0
Yr3
273
143.4
2.1
1.0
375
162.2
1.9
1.0
Yr4
273
143.4
2.1
1.0
375
162.2
1.9
1.0
Yr5
273
143.4
2.1
1.0
375
162.2
1.9
1.0
TOTAL
1,793
143.4
2.1
1.0
2,456
162.2
1.9
1.0
*Tonnage and grade after application of dilution and recovery factors.
18.1.3
Mining Equipment
The mobile mining fleet was estimated based on a combination of first principles calculations
(e.g. average haul profiles) and experience of similar operations from the Micon database.
103
The mine is to be trackless and will use a fully mechanised fleet, while employing the
minimum of manual mining practices. The mobile equipment list is shown in Table 18.4.
Table 18.4
Mobile Mine Equipment List
Item
Jumbo (2-boom)
Scissor Lift
Haul Truck (20-t)
LHD (10-t)
LH Drill
Service Vehicle
Grader
Utility Vehicles
Automatic Bolter
Robo-Shotcreter
Remix truck
Light Duty Vehicle
Total
Units
3
2
4
3
2
1
1
3
1
1
1
8
30
Availability and utilisation factors of 85% and 83% (respectively) have been used, operator
efficiency has been assumed to be 90%. Contingencies have been applied to machine
productivities, which accounts for altitude de-rating and also for general operational
inefficiencies.
For production and development, 10 t capacity LHD’s will muck out the stopes and headings
and will either directly load the trucks or re-muck the ore. Ore-passes have not been
designed; however, future optimisation may indicate that they provide a net benefit.
Underground haul trucks of 20 t capacity have been chosen to transport the ore and waste out
of the mine. Once again, this size of equipment is chosen to match the production rate and
yet minimize the cross-sectional area of the development, in an effort to reduce spans and
minimize the waste development cost. This is especially true in the spiral declines where a
short radius of curvature would make larger equipment impractical to operate.
Conveyors were considered as an alternative to trucking the ore out of the main adit to the
mill. This may have environmental benefits, but trucking has been determined to be a more
cost effective system at the chosen production rate.
Twin boom jumbos will carry out all drilling requirements in development headings. Longhole drilling is carried out by trackless and fully mechanised units, capable of fan and parallel
drilling.
104
18.2
PROCESSING
Two process options were considered for this Preliminary Assessment, (1) constructing a
new process plant on site and (2) toll milling. The following sections present the design
criteria, process description, and tailings management for each option.
18.2.1
Pulacayo Process Plant Option
18.2.1.1
Process Design Criteria - Pulacayo Plant Option
The design criteria were based on a mining rate of 1,800 dry metric tonnes per day (t/d), for
an annual rate of 648,000 tonnes per year (t/y). An availability of 67% was chosen for the
crushing circuit, and 90% was chosen for the grinding circuit. The design milling rates using
the given mining rate and availabilities, calculates to a crushing rate of 112 t/h, and grinding
rate of 83 t/h. The design criteria summary can be seen in Table 18.5.
The fine ore stockpile between the crushing and grinding circuits is sized for two days worth
of production, or 3,600 tonnes. This will ensure no disruptions downstream of the stockpile
during extended mine or crusher outages.
Table 18.5
Pulacayo Process Design Criteria
Description
Mill Feed Solids SG
Mine Rate
Crusher Availability
Crusher Rate
Stockpile
Mill Availability
Mill Rate
Flotation Solids
Flotation SG
Concentrate Moisture
Tailings %Solids To TSF
Final Pulp SG In TSF
Process Value
3.1
1,800 t/d
67%
2,687 t/d
~2 days
90%
2,000 t/d
25 wt%
1.2
8 wt%
15 wt%
2.0
Note
112 t/h
3000 t
83 t/h
A surface jaw crusher is needed prior to the grinding circuit as indicated by the mine plan,
which will not have any underground crushing. Crusher availability is based on two, 8-hour
crusher-operating shifts, per 24 hours. Grinding availability is based on industry wide
standards for the grinding flowsheet selected.
The comminution circuit equipment sizes were calculated using the Work Index test results
on the low and medium grade samples from UTO, shown in Table 18.6.
105
Table 18.6
Pulacayo Bond Work Index (kWh/st)
Sample
To 65 Mesh
Low Grade
65 To 100 Mesh
100 To 150 Mesh
9.823
10.775
11.376
Medium Grade
10.206
11.091
12.844
High Grade
10.799
12.434
14.387
Flotation slurry densities are lower than typical due to the un-typical, high fines content in
the flotation feed.
For the concentrate dewatering circuit, a maximum of 8 wt% concentrate moisture is
necessary due to shipping requirements. The dewatering equipment selected consists of a
thickener followed by a pressure filter, as opposed to more capital intensive equipment
consisting of a disk filter, propane fired dryer combination. Pressure filters are able to
produce concentrate moistures of 8 wt%, but more laboratory tests on concentrate filtering
are needed to confirm pressure filter final moistures.
EPCM applied the above design criteria in the preparation of its equipment selection and cost
estimates, which were then reviewed by Micon.
18.2.1.2
Process Description - Pulacayo Plant Option
The location of the new mill relative to the Pulacayo townsite can be seen on Figure 18.4.
Figure 18.4
Town of Pulacayo as Viewed From the Mill Site
106
This location (which is the same as that of the old mill) was selected primarily due to its
proximity to the San Leon adit. A new portal will be established close to the new mill and a
new adit will be driven to connect with the San Leon adit, minimising haulage distance and
ensuring that haul trucks need not travel through the townsite to deliver ore to the crushing
plant.
An alternative location north of Pulacayo peak was also considered, as it would be situated
between the Pulacayo mine and the Paco mine. However, this location was dropped due to
the cost of rehabilitating the adit to the north, with an estimated cost of $4.5M.
A conventional flowsheet was selected for the Pulacayo mill based on metallurgical test
results and on typical Bolivian lead, zinc mills in the area. The flowsheet consists of a
crushing circuit followed by Semi Autogenous Grinding (SAG), differential lead/zinc
flotation cells, concentrate dewatering, ending with the tailings solids deposition in the
Tailings Storage Facility (TSF). Process water will be reclaimed from the TSF and pumped
back to the mill’s Process Water Tank for reuse. A schematic of the flowsheet is presented in
Figure 18.5.
The list of equipment and equipment sizes selected by EPCM can be found in Appendix V.
Some items of primary equipment selected are a jaw crusher (21”x 36”, 100 hp), SAG mill
(14’x7’, 1000 hp), ball mill (12’x 16’, 1500 hp), lead rougher cells (four 300 ft3, 30 hp each),
zinc rougher cells (eight 500 ft3, 50 hp each), lead thickener (24’ diameter), and zinc
thickener (36’ diameter).
The reagent area will consist of a lime slaking station, and mix-and-storage tanks for zinc
sulphate, sodium cyanide, two float collectors, one float frother, copper sulphate, and sodium
silicate.
A tailings pump station and pipeline will transport the tailings slurry 1.2 km to the TSF,
where the solids will be deposited and process water will be reclaimed and pumped back to
the process water holding tank next to the mill. A flocculent mix-and-storage skid will be
purchased to treat the high clay content in the tailings prior to the deposition of the solids in
the TSF, that will ensure process water clarity in the recycled process water system. This
treatment may decrease the deposition density in the TSF, and increase the impoundment
volume required. Further tests are needed to determine the deposition density when treating
the clays.
As discussed in the metallurgical section (Section 16.2), further refinement of the flowsheet
and equipment selection will be necessary at future design phases to ensure the most efficient
recovery of the metal values in the fines.
107
Figure 18.5
Pulacayo Flowsheet (from EPCM report)
108
18.2.1.3
Tailings Storage Facility - Pulacayo Plant Option
This Pulacayo Preliminary Assessment includes the construction of a TSF to contain the
tailings produced by the 1,800 t/d mill. The area selected is close to and downhill of the old
mill area in the Viejo creek area, a distance of 1 km southeast of the old plant site (Figure
18.6).
Figure 18.6
Diagramatic location of Tailings Storage Facility (NTS)
The containment dam will be constructed in two stages to reduce capital costs. Stage 1 will
be built to an elevation of 4,020 m, and will contain 2 Mt of solids. At a mining rate of
1,800 t/d, this equates to almost 3.0 years worth of production. After the completion of Stage
2 to an elevation of 4,030 m, an additional 5.5 Mt will be contained, for a total capacity of
7.5 Mt.
Stage 1 capital cost is estimated at $7.5 million, and Stage 2 at $6.2 million. The itemized
cost breakdown can be found in Table 18.12, and in the TSF Budget Estimate Report in
Appendix V. For this Preliminary Assessment, only Stage 1 capital costs are included in the
economic model in Section 18.5. Stage 2 capital costs will be appropriated from an
operating capital budget later in the mine life, subject to exploration success and/or metal
price expectations at that time.
A mine backfill plant will be constructed when backfilling is scheduled to begin after year 3.
Equipment costs, estimated at $150,000, are reflected in the cash flow as a sustaining capital
109
expense. Due to the high fines and clays expected in the tailings which may comprise up to
30% of the tailings solids, a de-sliming component to the backfill system has to be
incorporated. Typical de-sliming equipment includes a cyclopac followed by a de-sliming
cone, along with a holding tank and pump station. It has not yet been determined if cement
addition will be necessary for mine backfill.
Stage 1 of the TSF construction will consist of a starter dam constructed 5 m high with a
slope factor of 2.5 to 1, as shown in Figure 18.7. The starter dam will be fully lined with the
liner extending upstream 30 m, and anchored in 3 to 5 m intervals.
Figure 18.7
Starter Dam Design (from EPCM report)
There is provision in the study for water diversion channels around the TSF in order to
maintain the desired water cap level during periods of high rainfall and storm events. No
water behind the dam (process water) will be released to the environment, but will be
recycled and reused in the mill.
The process water will be reclaimed from the TSF by a pontoon structure supporting a
vertical pump that will discharge to a holding tank adjacent to the TSF. A horizontal pump
will transport the water from the holding tank to a process water tank adjacent to the mill (see
Figure 18.8). The elevation difference between the holding tank and the process water tank
is expected to be 90 m.
110
Figure 18.8
Conceptual Plan View of the Tailings Storage Facility
(Not to scale. North is up)
Legend:
Dark blue: Pipelines
Red line: Maximum extent of impoundment
18.2.2
Toll Milling Option
18.2.2.1
Process Design Criteria - Toll Milling Option
Light blue shading: Proposed dam area
A study for a toll milling option was requested by Apogee, to process the Pulacayo ore at an
offsite mill. Two mills were visited, one at Potosi, and one at Porco, both owned by
Glencore International. The decision to use the Don Diego mill at Potosi was made based on
the milling rates required and type of metallurgy expected from the Pulacayo ore. The
Pulacayo ore produces high fines, as described in Section 16, which requires a specialized
flowsheet to process it. For example as shown in the Don Diego flowsheet, Figure 18.9, the
fine fraction is floated off with the bulk concentrate. Two concentrate thickeners are then
used in series in an attempt to capture the fines carried from the first thickener overflow to
feed the second thickener. The second thickener would produce a high clay content, lower
grade concentrate, which would be difficult to dewater.
111
Milling rates at Don Diego were reported to be 60 t/h, whereas the Pulacayo milling design
rate is 83 t/h. The difference in milling rates may be made up by the softer Pulacayo ore.
The Don Diego feed was described as medium hard with a work index of around 14 kWh/st.
The low and medium grade Pulacayo ore is softer, with a work index of 12.1 kWh/st, which
would permit increased crushing/grinding rates in the Don Diego mill.
Downstream of grinding, the flotation and dewatering equipment at Don Diego mill is
smaller than that selected for the Pulacayo mill. However, for the purposes of this
Preliminary Assessment, and with the objective of comparing toll milling and the
construction of a dedicated mill at Pulacayo, in Micon’s opinion the Don Diego mill provides
the best comparison of any alternative mill in the region.
At the time of the site visit in July 2009, the Don Diego mill was being run sporadically,
depending on outsourced feed from a dozen or so small private miners in the area.
Figure 18.9
Don Diego Mill, Crushing and Fine Ore Bin
112
Figure 18.10
Don Diego Plant Flowsheet
113
18.2.2.2
Process Description - Toll Milling Option
The toll milling option consists of mining and crushing the Pulacayo ore on site, then
transporting it to the Don Diego mill located 300 km to the northeast.
The transport would be in two stages. Stage one is to load the crushed ore into trucks at
Pulacayo and deliver it 20 km to the rail head at Uyuni. Stage two is to rail transport the ore
from Uyuni 208 km to the Don Diego mill. The Don Diego mill is located 40 km east of the
mining town of Potosi.
A crushing circuit and load-out facility would be required at the Pulacayo site. The crushing
circuit would be the same as the crushing circuit described in the Pulacayo flowsheet above.
The load-out would consist of an apron feeder under the stockpile similar to the one
described above, but instead of feeding the SAG mill, the conveyors would be arranged to
load trucks. The rail head at Uyuni already has ore loading facilities, as does the load-in at
the Don Diego mill.
A tour of the Don Diego mill was arranged. The mill was described as having gone through
a refurbishment in the last couple years. This was evident in new instrumentation and PLC
control equipment, and new rougher flotation cells. In general the mill was found to be in
good condition.
Discussion with staff at the Don Diego mill indicated that there is sufficient permitted
capacity at the site to accommodate the tailings from toll milling the Pulacayo ore.
The transport of concentrate from Don Diego to the port of Antofagasta in northern Chile
was also discussed. Details and costs associated with this, and other transportation costs can
be found in Section 18.5.
18.3
18.3.1
INFRASTRUCTURE
Power Supply
The current electrical power supply at Pulacayo is not adequate for mining and milling at
1,800 t/d. In past production, power was supplied from the Lindara-Kilpani substation via a
44 kV line, 60 km from Pulacayo. Since then this line has been dismantled. A new
substation and power transmission line is required. Three options were studied which can be
found in EPCM’s Power Supply/Power Demand report in Appendix V.
The option chosen is to tie into the San Cristobal-Punutuma 220 kV transmission line. The
closest point to this line is 10 km from Pulacayo. The price for a sub-station off the San
Cristobal line is approximately $2 million. A new 10 km, 34.5 kV transmission line from the
substation to Pulacayo is estimated to cost $165,000. These capital costs were included in
the process capital cost estimate in Section 18.5.
114
Total installed power requirements are estimated to be 5 MVA for the mine, mill, and general
administration.
18.3.2
Water Supply
The current fresh water supply for the town of Pulacayo is from the Yana reservoir located
47 km north of Pulacayo. It is currently transported via an eight inch, cast iron, flanged
pipeline, underneath the mountain through the main mine adit to the town. Cursory water
demand calculations show this line is adequate for makeup water for the daily mining and
milling, since the majority of water for milling will be reclaimed process water from the TSF.
The initial water fill of the TSF would require a larger diameter pipeline if the initial fill had
to be done in less the 28 days. This should not be necessary given proper TSF dam
construction planning, and scheduling the initial fill accordingly.
The dependability and availability of the existing line is questionable. During the visit the
line developed a leak and had to be shut down for a day to be repaired, cutting off fresh water
to the town. As a design exercise, the installation of a new eight inch diameter pipeline was
estimated. A new eight inch, 47 km, HDPE-PN6 pipeline is estimated to cost $1.41 million.
Since the pipeline already exists, the new pipeline capital cost has not been included in the
Section 18.5 capital cost estimates.
18.3.3
Ancillary Buildings
The town of Pulacayo has housing infrastructure, with the current Apogee offices occupying
the old hospital. New buildings to be constructed include the concentrator building and an
administration building that would be comprised of offices, assay laboratory, and a mine/mill
dry.
Labour camp accommodations for construction and for later production crews are not part of
this Preliminary Assessment. It is not known how much of the existing Pulacayo housing
infrastructure can be used for labour accommodation.
18.3.4
Roads
The main access from Pulacayo to Uyuni is a 20 km gravel road. It is well maintained and
can be used for the trucking of ore from Pulacayo to the Uyuni railhead for the toll milling
option.
There is currently a major highway upgrade underway on a 200 km stretch of State Highway
701, between the towns of Uyuni and Potosi. During the visit, three major bridges were
observed being constructed with road widening, culvert installations, and with the
elimination of switchbacks. Some paving along this highway will also be done. The
construction to upgrade this highway will be completed over the next few years. These
highway improvements will enhance the logistics of transporting ore from Pulacayo to the
115
Don Diego mill at Potosi. It is recommended the transport cost elements for the toll milling
option be re-calculated after the highway is open to transport trucks.
Additional road construction and/or upgrades will be needed for the proposed 10 km
electrical transmission line (already included in the capital cost estimate), and for the fresh
water pipeline to the Yana reservoir, 47 km to the north (not included in the capital cost
estimate as explained in Section 18.3.2).
18.4
ENVIRONMENTAL AND SOCIAL ASPECTS
Section 18.4 of this report has been prepared by Jenifer Hill, R.P.Bio., Senior Environmental
Consultant with Micon.
18.4.1
Environmental Conditions
Pulacayo is located in the high plains climate of the Andes. Winter temperatures are as low
as -15oC and average 14oC during the rest of the year with sub-humid conditions occurring in
the summer and dry conditions for the rest of the year. The rainy season is from December
to March and is typically in the form of electrical storms. Mean annual precipitation is
410 mm; relative humidity averages 45%; and winds are dominantly from the southwest
from July to September with average velocities of 2.8 m/s (MINCO S.R.L. 2008).
The countryside is mainly desert with various plants and animals typical of the high plains.
No protected species are known to occur in the project area. The landscape varies from
3,700 to 4,600 m amsl from high to low mountains, foothills, and valley floors. Soils are
generally thin with low fertility in the steeper mountains and become somewhat thicker lower
in the valley. Soil texture is sandy with gravels and cobbles throughout the profile resulting
in poor water retention. The valley floors have thicker silty, sandy soils of class IV arable
lands with limitations due to climate and soil quality. The valley floor of the Rio Negro has
been covered by tailings up to a metre in depth due to past mining activities at Pulacayo.
This has significantly affected the productive capacity of these lands.
The project area has had historical mining since the deposit was discovered in 1833. The
largest production occurred from 1951 to 1959. Due to the past mining operations, there are
many sources of existing contamination throughout the property from waste rock and
tailings. From the baseline work completed by MINCO (2007 and 2008), contamination
from acidic drainage, heavy metals and sediments appear to be along the Rio Negro.
Tailings were released from the operations and allowed to flow down the river basin resulting
in covering of vegetation with sulphidic tailings in the lower river valley. The Rio Negro has
water quality that is low in pH and elevated in various heavy metals including lead, zinc,
manganese, iron, copper, and mercury. Waste rock piles were sampled and continue to
contain acid generating materials.
Air quality samples contained high levels of arsenic and zinc at two sample points within the
village of Pulacayo, likely originating from piles of metallic waste located around the village.
116
Densely populated areas are sources of groundwater contamination where there is no sewage
treatment. Although there are primitive sewage systems in the communities, the plumbing of
most households do not connect to the main sewage pipelines and generally flow into the
smaller creeks.
18.4.2
Social Conditions
The Pulacayo community is represented by a Civic Committee elected from the town council
and includes four members, President, Relations Secretary, Housing Secretary, and First
Committee member. The Civic Committee’s principal function is to hear all the complaints
and demands from the members and organizations in the community, and to take these to the
town council to make decisions through voting. The highest legal authority in the town is the
mayor, who represents the prefecture, however, the President of the Civic Committee calls
the town council meetings, and only when the President is absent does the mayor preside
over social issues.
The government authorities in the community include:










Mayor, the highest executive authority.
Civil Registrar.
Administrator of COMIBOL (the mining corporation of Bolivia).
Defence Counsel for Children and Adolescents.
National Health Office.
Health Station.
Military Exercise Station.
Teacher.
Town Council.
COSEU (electrical service provider).
Organizations in Pulacayo include:









Civic Committee.
Mining Cooperative.
Renters.
Bakery.
Mothers’ Club.
Women’s Centre.
Refinery Club.
Tourism.
Other small groups of former workers.
Currently there are 13 local people working on the project including 11 labourers and 2
cleaners for the geologists’ and engineers’ camp.
117
The closest indigenous community to Pulacayo is Uyuni, 22 km away. The towns of
Huanchaca and Yana, located 15 and 20 km to the north, are no longer populated.
18.4.3
Impact Assessment, Mitigation, and Management
Key potential areas of social and environmental effects are expected on the residents of
Pulacayo, surrounding farming land, surface and ground water quality, and air quality.
Potential social effects on Pulacayo include increased income from direct employment,
increased demand for local services and suppliers, and an influx of workers. Potential
adverse social effects and pressures on housing and infrastructure will need to be effectively
managed. A community development plan should be developed in concert with the
community to help ensure economic benefits are realized in the community. Social and
health effects should be considered for any small miners who are still active in the Pulacayo
area.
Depending on the final locations chosen for the mill, tailings impoundment, and ancillary
facilities, there may be effects on agricultural land. The company will need to consult and
negotiate early and effectively to manage any economic relocation and land acquisition
requirements. It is assumed that no homes will be affected; however, this will need to be
confirmed as project planning proceeds to the next stage. Involuntary resettlement should be
avoided.
The project area is affected by historical mine workings that are causing acidic drainage and
metal contamination to enter the surrounding air and water. Project development needs to
take these into consideration. New development must consider whether operations can be
isolated from historical contaminants, be integrated with historical works to clean up some of
the contamination, or historical contaminants can be remediated prior to new development.
Regardless, definition and documentation of historical contamination is important so that the
company can manage its risks. In addition to the actual environmental impacts the project
may cause, public perception is also a risk.
The waste rock has been shown to be highly acid generating and should not be used for
construction. Dumps should be located in areas where the dump drainage can be captured
and managed during operations and capped after closure to minimize water infiltrating the
dump. Alternately, waste rock could be disposed of underground in a manner that minimizes
acidic drainage.
Based on the ore type and proposed process, it is expected that the tailings will also be
potentially acid generating. Engineering in the next stage of development needs to take this
into consideration. It is expected that during operations, all water from the tailings pond and
seepage collection pond will be recycled to the process plant or evaporated. The
impoundment should be designed to minimize water infiltration through the tailings postclosure.
118
18.4.4
Permitting Process
The exploration phase of the project is permitted by the state government under a
Dispensation Certificate (SRNMA-CD-012/07) that does not require preparation of an
environmental impact assessment. A baseline study and impact assessment will be required
for approval of mine exploitation activities. Public consultation is also required as part of the
permitting process. Baseline social and environmental studies were initiated in 2007. It is
estimated that the impact assessment, consultation, and permitting will take six to twelve
months and could be completed during the feasibility phase of the project. This estimate is
for planning purposes. As with most mining projects worldwide, the permitting process is
subject to outside influences and there is a risk for delays.
Key applicable legislation for the mining industry in Bolivia includes:




The Bolivian Constitution.
Environmental Law (No. 1333; 1992).
Mining Law (No. 1777).
Environmental Rules for Mining Activities (D.S. 24782).
The
following
additional
laws
are
industry.com/sia/marcoreg/Ley/Ley.html):











often
applicable
(http://www.bolivia-
Forest Law.
Law of Biodiversity Conservation.
Water Law.
Energy Law
Hydrocarbon Law.
Mining Code.
Regulatory System Law.
Land Law.
Municipal Law.
Health Legislation.
Law for Medicine.
Applicable regulations under the Environmental Law No. 1333 include (: http://www.boliviaindustry.com/sia/marcoreg/Ley/Ley.html ):






Environmental Management.
Environmental Prevention and Control.
Atmospheric Contamination.
Hydraulic Contamination.
Hazardous Materials Handling.
Waste Management.
119
Micon understands that, as announced on January 8, 2010, Apogee has received notice from
the Bolivian Ministry of Environment and Water ("MEW") accepting Apogee’s
recommendations for the project set out in its submission entitled, "Environmental Form:
Mineral Extraction, Milling and Construction of Tailings Facilities – Pulacayo", submitted to
MEW in November, 2009. The submission forms part of Apogee’s requirements to have the
project categorized for development. The report was prepared on the Company's behalf by
Medio Ambiente Mineria e Industria, a well respected independent Bolivian environmental
engineering consulting firm. Apogee reports that MEW has categorized the Pulacayo Project
as "Category 1" and provided the terms of reference for the Baseline Environmental Impact
Study, required to obtain certain environmental licences and permits necessary for project
development.
18.4.5
International Financing
Additional environmental and social standards must be met if the project requires financing
from international institutions that adhere to the Equator Principles. This includes meeting
the IFC Social and Environmental Performance Standards, IFC Environmental, Health, and
Safety Guidelines for Mining, and International Conventions on such issues as heritage
resources, indigenous communities, human rights, and climate change. In addition to
meeting Bolivian permitting requirements, Apogee would require an internationally
acceptable environmental and social impact assessment, management plans, and action plan.
Security, health and safety management planning would also need to meet international
standards. Any resettlement and/or land acquisition requirements for the project require
appropriate consultation, management, and documentation. Archaeological and cultural
heritage resources would need to be assessed and documented.
If the project is equity funded, it is still recommended that these environmental and social
aspects be addressed and brought to international best practices standards to help minimize
development and investor risks.
18.4.6
Consultation
Apogee has a company liaison for the community of Pulacayo, Sra. Rosario Calderon.
Apogee’s mining project is supported by the community. Based on surveys completed at the
end of 2008 and updated in July 2009, there are many expectations for the project such as:








Generating more jobs.
Improve the town.
Welcome to Pulacayo.
That the company will participate in the activities and celebrations of the town;
That mining will be reactivated.
That they will be concerned and look after the children.
That Pulacayo will return to what it once was.
That the company will fulfill its promises.
120



That it will bring more people.
Good drill results.
That the company will ask more questions from the community.
In general, the feeling is that the town of Pulacayo has been forgotten by the authorities, and
this is why there are hopes for the project.
18.4.7
Environmental and Social Capital and Operating Costs
There are no reclamation bond requirements in Bolivia. The mine must prepare a closure
plan near to the time of closing, but the funding for this plan is not required. Nonetheless,
Apogee should plan a fund in the event that the company requests outside funding for project
development or in the event that Bolivia follows the lead of other countries and amends its
regulations to include this requirement. A closure plan and financial assurance would be
required if the project is financed by an institution following the Equator Principles.
For scoping study purposes, Micon has been provided with an estimate of $2 million for the
eventual cost of mine closure and rehabilitation, based on recent experience of local
engineering firm EPCM Consoltores S.R.L. In line with its recommendations given above,
Micon has provided in the cash flow model for establishing a bond in this amount at the
commencement of production.
It is assumed that baseline social and environmental baseline studies and permitting costs are
already sunk for the purposes of the cash flow model.
Operating costs for environmental and social aspects should include employment of an
Environmental Coordinator, one Social Responsibility Coordinator, one Health and Safety
Manager and three additional technical staff to support these three positions. Annually, it is
recommended that $50,000 be budgeted for environmental consultant and laboratory costs,
and an additional $50,000 to support social and community development programs.
18.5
PROJECT ECONOMICS
The economics of the Pulacayo project have been assessed under two scenarios:

In the first scenario, which forms the base case for this report, a processing facility is
built on-site to treat the material mined from underground, and concentrates of zinc
and lead are produced for shipment to port. Silver credits are obtained for both
products.

Alternatively, no processing facility is constructed, and instead the mine ships
crushed, unbeneficiated ROM material to the Don Diego process plant, where it is toll
treated. This scenario is treated as a sensitivity case, the cash flow from which is then
compared to the base case.
121
In each case, production rates and all other assumptions are kept the same to allow the
relative value of each scenario to be determined.
The net present value of the base case is determined on the basis of a discounted cash flow
model. The model is prepared in constant United States dollars (US$) of 2010 value, and the
present values are stated in mid-2010 terms. A real discount rate of 8%/y was selected for the
base case, taking into account medium-term risk-free interest rates of around 3% in real
terms, and a historical premium for equity of around 5%.
18.5.1
Macro-economic Assumptions
18.5.1.1
Product Prices
In the cash flow analysis, the prices used are:

The 36 month trailing average price of silver, as of 30 April, 2010. This equates to
US$14.78 per ounce.
The 27-month forward prices for lead and zinc, as published by the London Metal
Exchange at the end of April, 2010. These prices are US$1.108/lb zinc, and
US$1.035/lb lead.
18.5.1.2
Royalties and Taxation
A royalty equivalent to 4% of the NSR value of the concentrates has been assumed payable.
The model uses a simplified computation of taxation on corporate income at the rate of 25%.
It is assumed that construction capital allowances are available to offset income tax until
fully recouped, sustaining capital is assumed claimable at the rate of 25% on a declining
balance basis, and other cash taxes are assumed allowable against the amount of income tax.
18.5.2
Production Schedules
The base case LOM processing schedule is given in Table 18.7. This schedule is based on the
mining production schedule (Table 18.2) which uses a 200 g/t Ag Eq cut-off for mill feed. As
discussed in Section 18.5.7, this was determined to be the optimal cut-off for on-site milling.
Table 18.7
Base Case LOM Processing Schedule
Year
-1
1
2
3
4
648
648
648
648
Treated*
t 000
1.98
1.98
1.98
1.98
Zinc
%
1.04
1.04
1.04
1.04
Lead
%
154.2
154.2
Silver
g/t
154.2 154.2
*Tonnage and grade after application of dilution and recovery factors.
122
5
648
1.98
1.04
154.2
6
648
1.98
1.04
154.2
7
361
1.98
1.04
154.2
Total
4,429
1.98
1.04
154.2
18.5.3
Revenue
For both the on-site and toll milling options, the Net Smelter Return (NSR) from sale of the
concentrates is assumed to be the same. Micon has applied its experience to derive generic
off-take terms which it believes to be reasonable at this stage of project development. The
parameters in terms of which the NSR is calculated are given in Table 18.8.
Table 18.8
NSR Parameters
Metal
or unit
Zn%
Zn
Pb
Ag
Zn
Pb
Ag
US$/t
Value
53.0
87.7
2.7
18.6
84.9
n/a
62.5
175.0
Smelting Charge
Pb%
Zn
Pb
Ag
Zn
Pb
Ag
US$/t
51.0
3.0
77.6
63.9
n/a
94.1
94.2
180.0
Concentrate transport costs
US$/t
Description
Zinc Concentrate
Payability of Concentrate
Smelting Charge
Lead Concentrate
Payability of Concentrate
36.03
Comment
Net of min. deduct 3.0 oz/t
Deduct 3.0% Pb in conc
Min. deduct 50 g/t
Dry basis (or $33.15/wmt)
Figure 18.10 shows the resulting split of NSR value between the payable metals.
123
Figure 18.11
NSR Value of Payable Metals
Silver
47%
Gold
0%
Lead
17%
Zinc
36%
18.5.4
Capital Costs
Micon prepared an estimate of mining capital costs for equipment, pre-production and
ongoing development. Micon has also reviewed and adopted the capital cost estimates for the
base case processing, tailings storage, infrastructure prepared by EPCM Consoltores S.R.L.
The estimated cost of environmental remediation was also prepared by EPCM, based on its
local experience and familiarity with Bolivian legislatory requirements. Initial and sustaining
capital costs for the project base case are summarised in Table 18.9.
Table 18.9
Summary of Base Case Capital Expenditure
Capital Cost Summary
Initial
US$ (M)
18.48
27.77
3.00
2.53
2.00
15.63
69.41
Mining
Processing
Tailings
Infrastructure & indirect
Environmental & Social
Contingency
Total
18.5.4.1
Sustaining
US$ (M)
15.06
0.15
4.35
2.22
5.87
27.65
Exploration & Engineering Studies
The preliminary assessment cash flow does not reflect the cost of further exploration or
engineering studies required to bring the project to a construction decision.
124
18.5.4.2
Mining
Mining capital costs are shown as in Table 18.10. A 30% contingency ($10.0 million) is
added to this base estimate in the cash flow model.
Development prior to production commencing is capitalised and is to be carried out by a
contractor. After this time development costs are considered as part of the operating costs.
Raise-boring, de-watering and the drilling of service boreholes is considered as a sustaining
capital cost and will be carried out by a contractor.
Table 18.10
Mining Capital Costs
Description
Unit Cost
$
15,000
2,000
2,000
1,500
1,500
1,500
1,500
1,500
1,500
1,500
250,000
120
80
3,500
Contractor mob/de-mob
New Adit on 4129m
Slash old Adit on 4129m
Access Ramp Upper East
Access Ramp Upper West
Access Ramp Lower East
Access Ramp Lower West
Level 4255
Level 4180
Level 4105
De-watering
Vent Raises bored
Boreholes (inter-level)
Workshops
S/T Excavation Capital
Mechanical & Electrical
Mobile
Start-up Capital
Sustaining Capital
Total Mine Capital
18.5.4.3
Total Cost
$M
0.03
1.42
0.84
1.08
1.66
0.00
0.00
1.44
1.82
0.00
0.50
0.11
0.07
0.53
9.49
4.63
4.36
18.48
15.06
33.54
Processing
Process plant construction capital expenditure for the base case (on-site mill) is estimated to
total $27.7 million, excluding a 30% contingency ($8.3 million) added within the cash flow
model. A breakdown of this amount is given in Table 18.11.
125
Table 18.11
Process Capital Expenditure
Process Plant Area
Civil Works
Site Preparation
Concrete
Stockpile and Ore Bin
Buildings
Concentrator
Steel
Offices and Admin.
Mill Equipment and Materials
Crushing Plant and Stacker
Grinding and Cyclones
Lead Flotation
Conditioners
Zinc Flotation
Dewatering
Regent Mixing and Storage
Lime Circuit
Piping and Fittings
Tailings Pipeline
Power for Construction
Mechanical Installation
Electrical Equipment
Electrical Materials
PLC and Instrumentation
Electrical Installation
Services
Process Water
Fresh Water
Compressed/Blower Air
Connection to Electrical Grid
Safety and Environment
General and Administration.
Engineering
Management
General Expenses
Working Capital
Commissioning
Grand Total Process Plant Capital
$ 000
2,370
510
1,745
115
2,615
1,600
600
415
14,965
805
5,255
755
90
1,040
1,300
210
185
550
250
200
1,000
1,405
330
1,090
500
3,875
565
630
480
2,200
220
3,720
570
400
1,250
1,000
500
27,765
For the base case, provision is also made for the construction of a back-fill plant in Year 3, at
a cost of $0.15 million.
For the alternative (toll-milling) scenario, the construction capital estimate reduces to
$6.2 million, to which a 30% contingency is added. A breakdown of this is given in Table
18.12.
126
Table 18.12
Toll Milling-Pulacayo Site Capital Expenditure
Description
Civil Work
Crushing Plant
Product Loadout
Buildings
Pipes And Fittings
Power During Construction
Mechanical Installation
Electrical Equipment
Electrical Materials
Electrical Installation
TSF (Mine and Admin included)
Engineering and Management
Working Capital and Commissioning
Subtotal
30% Contingency
Total
18.5.4.4
$ 000
698
805
100
523
255
40
200
281
66
218
1,639
666
750
6,241
1,872
8,113
Infrastructure
Site infrastructure and indirect capital costs are summarised in Table 18.13.
Table 18.13
General and Administrative Capital Expenditure
Description
Communications, IT, computers, software, etc
Office costs
Site costs
Subtotal
30% Contingency
Owner's costs
Total
$ 000
200
150
500
850
255
1,681
2,786
For the base case, additional capital is required for the construction of a tailings
impoundment and associated equipment for tailings disposal and water reclaim. The preproduction investment required in this area is estimated to be $3.0 million. Another
$2.74 million is incurred as sustaining capital during Year 1, bringing the initial phase of
dam construction to a total of $5.74 million (Table 18.14). Subsequent phases of dam
construction add a further $1.61 million to sustaining capital, expended during the remainder
of the operating period. All the above amounts are given before the addition of a 30%
contingency.
127
Table 18.14
Tailings Storage Facilities-Capital Expenses (from EPCM Report)
m3
Initial Capital
TSF Construction-Stage 1
Quantity
Unit
Cost
Cost ($)
($ 000)
82,000
3.95
324
m3
107,078
3.95
m3
615,674
7
m2
31,145
5.5
171
m3
960
37
36
M
3,500
50
175
GL
1
25000
25
Unit
Diversion channel: strip, clear
& cut
Starter dam: clear loose
material
Dam: fill and compact using
borrow material
Liner: 80mil HDPE
(elev. 4000-4020 m)
Ditto (elev. 4020-4035 m)
Drainage: geotextile and slotted
corrugated pipes.
3x1m diversion channel on
compacted ground
Seepage collection and decant
pond
Spillway
Process Water Reclaim from
TSF comprising:
pontoon structure
Pumps, 1 pr (duty)
Pumps, 1 pr (standby)
electrical
piping
construction
Sub-total
Contingency
Total
18.5.4.5
Sustaining capital
TSF Construction-Stage 2
Quantity
Unit
Cost
Cost ($)
($ 000)
423
25,000
3.95
98.75
4,310
180.000
8
5,100
450
5.5
37
28.05
16.65
250
115
28.75
1,440.00
280
5
80
80
50
15
50
5,743
1,723
7,466
30%
1,612
484
2,096
Environmental and Social
A provision of $2.0 million has been made for contributions to a reclamation bond in respect
of the Pulacayo mine site. This expenditure is assumed to be incurred at the commencement
of the operating period.
18.5.5
Operating Costs
18.5.5.1
Mining
Operating costs for mining have been factored from similar operations in the Micon database.
Costs have been localised, based on local rates for labour, utilities and consumables. Table
18.15 shows the LOM average operating costs for mining, factored for each of the cut-off
grades used in the evaluation of the optimum extraction strategy.
128
Table 18.15
Average Unit Operating Costs for Mining ($/t mined)
Cut-off Grade (g/t Ag Eq)
Labour
Def Drilling
Development
Production & Backfill
Mucking
Haulage to plant
Services
Maintenance
Energy
Mining Total
18.5.5.2
125
150
175
2.4
0.4
0.6
6.0
0.5
0.6
1.1
5.6
2.7
19.9
2.5
0.4
0.7
6.1
0.5
0.6
1.2
5.9
2.8
20.7
2.6
0.5
0.7
6.4
0.5
0.6
1.4
6.2
2.8
21.7
200
(Base)
2.7
0.5
0.8
6.6
0.5
0.6
1.5
6.5
2.9
22.6
225
250
275
2.8
0.6
0.9
6.9
0.5
0.6
1.7
7.2
2.9
24.1
2.9
0.7
1.0
7.2
0.5
0.6
1.9
7.5
3.0
25.3
3.0
0.8
1.1
7.4
0.5
0.6
2.1
7.9
3.0
26.4
Processing
Process operating costs for the base case are given in Tables 18.16 (Labour, including
burden) and 18.17 (Total cash operating costs). For the alternative (toll-milling) scenario,
operating costs are given in Table 18.18. Costs for the transport of concentrate are not
included here – these are considered as part of the NSR calculation, and are held constant in
both the base case and toll milling scenarios.
Table 18.16
Process Plant Labour Costs (from EPCM Report)
Position
Plant Superintendent
Metallurgist
Secretary
Shift Supervisors
Sampler
Chief workshop
Chief Instrumentation
Met Lab assayer
Mechanics/Welders
Electrical
Instrumentation
Crusher Operators
Grinding Operators
Floatation Operators
Dewatering
Reagent
Tailings
Total Personal
Average cost
No.
1
1
1
3
2
1
1
1
4
2
1
2
3
3
3
2
3
34
US$/t
129
Monthly Cost
($)
3,600
2,400
1,200
5,400
2,000
2,400
2,400
1,800
4,800
2,400
1,200
2,000
3,000
3,000
3,000
2,000
3,000
45,600
Annual Cost
($)
43,200
28,800
14,400
64,800
24,000
28,800
28,800
21,600
57,600
28,800
14,400
24,000
36,000
36,000
36,000
24,000
36,000
547,200
0.84
Table 18.17
Process Plant Cash Operating Costs
Item
Lime
Zinc sulphate
Sodium cyanide
AF 242
MIBC
Copper sulphate
Z-11
Silicate
Flocculent
Mill balls
Spare parts (including liners)
Power ($/kWh x kWh/t ore
Miscellaneous
Subtotal Consumables
Labour (from above)
Grand Total Processing Opex
$/kg
0.12
1.60
2.70
7.50
5.00
5.00
3.00
2.50
approx
1.70
estimate
0.06
g/t ore
9200
400
130
30
40
400
70
350
estimate
600
$/t processed
1.10
0.64
0.35
0.23
0.20
2.00
0.21
0.88
0.50
1.02
2.00
1.80
1.00
11.93
0.84
12.77
30
Table 18.18
Processing Costs - Toll Milling at Don Diego Mill
18.5.5.3
Process Operating Costs: Toll Milling Option
Method
Distance
(km)
Transport charges
from Pulacayo to Uyuni station
Uyuni station to Don Diego mill
truck
rail
20
208
Unit Cost
($/t)
13.24
2.21
11.03
Don Diego Mill operating expenses
Labour Costs (same as Pulacayo Mill Labour)
Don Diego mill rental
Mill consumables
21.00
1.00
12.00
8.00
Total Milling + Transport
34.24
General and Administrative
Table 18.19 summarises the annual provision for General and Administrative operating
expenses. Over the LOM period, this equates to a cost of $2.08/t milled.
Table 18.19
General and Administrative Costs
Annual Cost
($)
Labour
Other overheads
Total
350,000
1,000,000
1,350,000
130
18.5.5.4
Environmental and Social
The annual provision for Environmental and Social expenses is given in Table 18.20.
Table 18.20
Environmental and Social Operating Costs
Annual Cost
($)
Labour (3 staff plus 3 support personnel)
Environmental Consultant and lab costs
Social / community development programs
Total
50,000
50,000
50,000
150,000
Micon considers the operating cost estimates to be reasonable and has not added a
contingency to these amounts in its base case economic assessment.
18.5.6
Project Schedule
The overall project schedule envisages a 3-year period of mine pre-production development
and ramp-up, followed in the base case by an approximately 5.5-year operating period at
steady state production, with mine closure on exhaustion of the identified resource.
Ore produced during the construction period is stockpiled until sufficient material is on hand
to allow the mill to operate continuously. Mill startup occurs after two years of preproduction mining, with material reclaimed from stockpiles making up the balance of millfeed in Year 1. The mill then operates at full capacity for approximately another 5.5 years.
18.5.7
Cash Flow Forecast
Table 18.21 presents the project base case LOM cash flow summary. The annual production
and cash flow details are provided in Table 18.22.
The main components of the project cash flow for the base case are shown in Figure 18.11.
This preliminary assessment is preliminary in nature; it includes inferred mineral resources
that are considered too speculative geologically to have the economic considerations applied
to them that would enable them to be categorized as mineral reserves, and there is no
certainty that the preliminary assessment will be realized.
On a pre-tax basis, at a discount rate of 8%/y, the base case cash flow evaluates to a net
present value (NPV8) of $50.0 million, and has in internal rate of return (IRR) of 24.0%.
After tax, the NPV and IRR are estimated to be $33.0 million and 19.6%, respectively.
Payback on the undiscounted cash flow after tax occurs in Year 4, and over the life-of mine
(LOM) period the net cash flow before and after tax is $109.5 and $80.5 million,
respectively.
131
The estimated maximum funding required before positive cash flow is $89.5 million,
occurring in Year1.
Table 18.21
Project Base Case - LOM Cash Flow Summary
LOM
($ 000)
178,537
203,519
382,056
15,282
366,774
$/t
treated
42.02
47.90
89.92
3.60
86.32
$/oz
Ag
11.81
13.46
25.27
1.01
24.26
Operating Costs
Mining costs
Processing costs
96,027
54,257
9,902
160,186
22.60
12.77
2.33
37.70
6.35
3.59
0.65
10.59
62,500
35,115
6,396
104,011
206,587
48.62
13.66
133,270
Capital Expenditure
97,060
22.84
6.42
83,268
Pre-tax Cash Flow
109,527
25.78
7.24
50,002
Taxation
29,023
6.83
1.92
17,013
80,504
18.95
5.32
32,988
NSR Silver only
NSR Co-products
NSR value
less Royalty
NPV8 (2010)
($ 000)
115,503
131,665
247,168
9,887
237,281
Figure 18.12
Base Case Cash Flow Summary
80
Net Cash Flow
60
Royalty
40
Taxation
Working Capital
0
Capital
(20)
Opcosts
132
Yr9
Yr8
Yr7
Yr6
Yr5
Yr4
Cum C/Flow
Yr3
(80)
Yr2
Cum DCF
Yr1
(60)
Yr‐1
Net Revenue
Yr‐2
(40)
Yr‐3
USD million
20
Table 18.22
Project Base Case Production and Cash Flow Projection
Production Forecast Mine Production
Indicated Resource
Inferred Resource
LG Resource
TOTAL ORE (kt) mined (MII)
Processing Plant Production
Zinc
Lead
Silver
LOM
TOTAL
2,456
1,793
‐
4,249
Yr‐3
‐
‐
‐
‐
Yr‐2
19
14
‐
32
Yr‐1
112
82
‐
194
Yr1
243
178
‐
421
Yr2
375
273
‐
648
Yr3
375
273
‐
648
Yr4
375
273
‐
648
Yr5
375
273
‐
648
Yr6
375
273
‐
648
Yr7
209
152
‐
361
Yr8
‐
‐
‐
‐
4,249
1.985
1.038
154.238
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
361
1.985
1.038
154.238
‐
‐
‐
‐
% 90.7
% 80.3
% 82.4
‐
‐
‐
‐
‐
‐
‐
‐
‐
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
‐
‐
‐
t 000
%
%
g/t
Overall Recovery to Conc
Zinc
Lead
Silver
Zinc Concentrate Production
Zinc
Lead
Silver
dry t 000
t 000
t 000
kg
139.6
73.996
1.182
121,584
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
11.9
6.287
0.100
10,330
‐
‐
‐
‐
Lead Concentrate Production
Zinc
Lead
Silver
dry t 000
t 000
t 000
kg
67.2
2.509
34.259
418,451
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
5.7
0.213
2.911
35,552
‐
‐
‐
‐
% 82.1
% 91.0
% 87.1
‐
‐
‐
‐
‐
‐
‐
‐
‐
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
‐
‐
‐
Payable Metal in Conc (imperial)
Zinc
Lead
Silver
000 lbs 138,509
000 lbs 71,086
oz 15,121,357
‐
‐
‐
‐
‐
‐
‐
‐
‐
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
11,768
6,040
1,284,728
‐
‐
‐
NET SMELTER RETURN (US$ 000)
Zinc
Lead
Silver
382,056
US$ 000 137,568
US$ 000 65,951
US$ 000 178,537
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
32,460
11,688
5,603
15,169
‐
‐
‐
‐
Cash Flow Forecast (US$ 000)
LOM
TOTAL
US$ 000 382,056
US$ 000 15,282
US$ 000 366,774
Yr‐3
‐
‐
‐
Yr‐2
‐
‐
‐
Yr‐1
‐
‐
‐
Yr1
58,266
2,331
55,935
Yr2
58,266
2,331
55,935
Yr3
58,266
2,331
55,935
Yr4
58,266
2,331
55,935
Yr5
58,266
2,331
55,935
Yr6
58,266
2,331
55,935
Yr7
32,460
1,298
31,162
Yr8
‐
‐
‐
4,558
4,393
164
‐
‐
Payability of Metal in Conc
Zinc
Lead
Silver
Net Smelter Return
Less Royalties
Net Revenue
Operating Costs
Mining
Processing
G&A
Social & Environmental
US$ 000
22.60
12.77
2.08
0.25
160,186
96,027
54,257
8,852
1,050
‐
‐
‐
‐
‐
760
732
27
‐
‐
19,102
9,519
8,083
1,350
150
24,419
14,645
8,275
1,350
150
24,419
14,645
8,275
1,350
150
24,419
14,645
8,275
1,350
150
24,419
14,645
8,275
1,350
150
24,419
14,645
8,275
1,350
150
13,670
8,159
4,610
752
150
‐
‐
‐
‐
‐
Operating Margin
48.62
206,587
‐
(760) (4,558) 36,833
31,516
31,516
31,516
31,516
31,516
17,491
‐
Capital Costs
Engineering Studies
Mining Capital
Processing Capital
Infrastructure Capital (including Contingency)
22.84
‐
7.89
6.57
8.38
97,060
‐
33,540
27,915
35,605
9,082
‐
6,986
‐
2,096
13,532
‐
4,643
5,553
3,336
46,794
‐
6,854
22,212
17,728
8,598
‐
3,764
‐
4,834
5,731
‐
3,764
‐
1,967
5,926
‐
3,764
150
2,012
5,731
‐
3,764
‐
1,967
139
‐
‐
‐
139
139
‐
‐
‐
139
1,388
‐
‐
‐
1,388
‐
‐
‐
‐
‐
Change in Working Cap
‐
‐
‐
‐
‐
6,192 (156)
‐
‐
‐
‐
10 (6,046)
Pre‐tax c/flow
Tax payable
C/flow after tax
Cumulative C/Flow
Discounted C/Flow (8%)
Cumulative DCF
Max funding reqmt to positive cashflow
25.78
6.83
18.95
109,527 (9,082)
29,023 ‐
80,504 (9,082)
(9,082)
32,988 (9,082)
(9,082)
(89,516) (9,082)
31,377
7,072
24,305
38,079
14,182
11,024
‐
31,377
7,265
24,112
62,191
13,027
24,051
‐
16,093
3,826
12,267
74,458
6,137
30,188
‐
(14,292)
‐
(14,292)
(23,374)
(13,233)
(22,315)
(23,374)
(51,352)
‐
(51,352)
(74,725)
(44,026)
(66,341)
(74,725)
22,043
‐
22,043
(52,683)
17,498
(48,843)
(89,516)
133
25,942 25,590
4,045
‐
25,942 21,545
(26,741) (5,196)
19,068 14,663
(29,775) (15,112)
(58,257) (36,712)
25,785
6,815
18,970
13,774
11,954
(3,157)
(17,742)
6,046
‐
6,046
80,504
2,800
32,988
‐
18.5.8
Sensitivity Studies
18.5.8.1
Base Case Sensitivity
The sensitivity of the base cash pre-tax cash flow to changes in product pricing, operating
costs and capital expenditure is shown in Figure 18.12.
Figure 18.13
Base Case Sensitivity Chart (NPV After tax)
100
80
60
40
20
0
(20)
70 75 80 85 Product Price (21.4 (12.1 (2.9)
6.1
90 95 100 105 110 115 120 125 130 15.1 24.1 33.0 41.9 50.8 59.7 68.5 77.4 86.2
Opcosts
56.9 52.9 49.0 45.0 41.0 37.0 33.0 29.0 25.0 21.0 17.0 13.0
9.0
Capex
57.4 53.3 49.2 45.2 41.1 37.1 33.0 28.9 24.9 20.8 16.7 12.7
8.6
The chart shows that the project is most sensitive to metal pricing (which would equally
apply to grade and recovery), when a 20% adverse change would reduce NPV8 to below
zero. The base case is less sensitive to capital and operating costs, so that even a 30%
increase in either would result in a positive NPV8.
18.5.8.2
Sensitivity to Cut-Off Grade
Micon evaluated the base case (on site milling) using a series of cut-off grades to determine
the optimum grade/tonnage combination for the project. The assessment utilized the silver
equivalent grade calculation described in Section 18.1, the modified mineral resources given
in Table 18.1, above, and the mining operating cost estimates given in Table 18.14, above.
The results are summarised in Figure 18.14, which shows that project NPV and IRR are
maximized when applying a cut-off grade of 200 g/t Ag Eq. Micon therefore selected this
value of the cut-off grade for its base case economic assessment of the project in this study.
134
35.0
35.0%
30.0
30.0%
25.0
25.0%
20.0
20.0%
15.0
15.0%
10.0
10.0%
5.0
5.0%
.0
0.0%
125
150
175
200
225
Cutoff grade (g/t Ag Eq)
NPV
18.5.8.3
Mining $/t
250
IRR (%)
Figure 18.14
NPV versus Cut-Off Grade for On-Site Milling
275
IRR(%)
Toll Milling Option
The study also considered an alternative to the on-site milling of material mined at Pulacayo.
For this purpose, it was assumed that crushed material was taken by road to Uyuni and
thence by rail to the Don Diego mill for toll treatment.
Mining, concentrate production and recoveries were assumed to remain the same, so that the
trade-off is essentially between the capital cost of the milling and tailings storage facilities at
Pulacayo versus the less capital intensive, higher operating cost option of renting the Don
Diego mill.
Table 18.23 and Figure 18.15 show the LOM cash flow summary for the toll milling option
using a cut-off of 200 g/t Ag Eq, the same cut-off used in the base case. Table 18.24 also
shows the annual cash flow forecast at this cut-off grade. As for the base case, Micon
evaluated this option at a number of cut-off grade scenarios, and measured the impact on
NPV of the cut-off grade selection. The results of this exercise are described below.
135
Table 18.23
Toll-Milling Option - LOM Cash Flow Summary
Using 200 g/t Ag Eq cutoff
LOM
($ 000)
178,537
203,519
382,056
15,282
366,774
$/t
treated
42.02
47.90
89.92
3.60
86.32
$/oz
Ag
11.81
13.46
25.27
1.01
24.26
Operating Costs
Mining costs
Processing costs
96,027
145,486
9,902
251,415
22.60
34.24
2.33
59.17
6.35
9.62
0.65
16.63
62,500
94,121
6,396
163,017
115,359
27.15
7.63
74,264
Capital Expenditure
56,374
13.27
3.73
50,396
Pre-tax Cash Flow
58,985
13.88
3.90
23,868
Taxation
15,878
3.74
1.05
9,339
43,107
10.15
2.85
14,529
NSR Silver only
NSR Co-products
NSR value
less Royalty
NPV8 (2010)
($ 000)
115,503
131,665
247,168
9,887
237,281
Figure 18.15
Toll-Milling Option - Cash Flow Summary
80
Net Cash Flow
60
Royalty
40
Taxation
Working Capital
0
Capital
(20)
Opcosts
(40)
Net Revenue
(60)
Cum DCF
136
Yr7
Yr6
Yr5
Yr4
Yr3
Yr2
Yr1
Yr‐1
Yr‐2
(80)
Yr‐3
USD million
20
Cum C/Flow
Table 18.24
Toll Milling Option – Production and Cash Flow Projection (200 g/t Ag Eq cut off)
Production Forecast Mine Production
Indicated Resource
Inferred Resource
LG Resource
TOTAL ORE (kt) mined (MII)
Processing Plant Production
Zinc
Lead
Silver
LOM
TOTAL
2,456
1,793
‐
4,249
Yr‐3
‐
‐
‐
‐
Yr‐2
19
14
‐
32
Yr‐1
112
82
‐
194
Yr1
243
178
‐
421
Yr2
375
273
‐
648
Yr3
375
273
‐
648
Yr4
375
273
‐
648
Yr5
375
273
‐
648
Yr6
375
273
‐
648
Yr7
209
152
‐
361
Yr8
‐
‐
‐
‐
4,249
1.985
1.038
154.238
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
648
1.985
1.038
154.238
361
1.985
1.038
154.238
‐
‐
‐
‐
% 90.7
% 80.3
% 82.4
‐
‐
‐
‐
‐
‐
‐
‐
‐
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
90.7
80.3
82.4
‐
‐
‐
t 000
%
%
g/t
Overall Recovery to Conc
Zinc
Lead
Silver
Zinc Concentrate Production
Zinc
Lead
Silver
dry t 000
t 000
t 000
kg
139.6
73.996
1.182
121,584
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
21.3
11.285
0.180
18,542
11.9
6.287
0.100
10,330
‐
‐
‐
‐
Lead Concentrate Production
Zinc
Lead
Silver
dry t 000
t 000
t 000
kg
67.2
2.509
34.259
418,451
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
10.2
0.383
5.225
63,816
5.7
0.213
2.911
35,552
‐
‐
‐
‐
% 82.1
% 91.0
% 87.1
‐
‐
‐
‐
‐
‐
‐
‐
‐
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
82.1
91.0
87.1
‐
‐
‐
Payable Metal in Conc (imperial)
Zinc
Lead
Silver
000 lbs 138,509
000 lbs 71,086
oz 15,121,357
‐
‐
‐
‐
‐
‐
‐
‐
‐
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
21,123
10,841
2,306,105
11,768
6,040
1,284,728
‐
‐
‐
NET SMELTER RETURN (US$ 000)
Zinc
Lead
Silver
382,056
US$ 000 137,568
US$ 000 65,951
US$ 000 178,537
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
‐
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
58,266
20,980
10,058
27,228
32,460
11,688
5,603
15,169
‐
‐
‐
‐
Cash Flow Forecast (US$ 000)
LOM
TOTAL
US$ 000 382,056
US$ 000 15,282
US$ 000 366,774
Yr‐3
‐
‐
‐
Yr‐2
‐
‐
‐
Yr‐1
‐
‐
‐
Yr1
58,266
2,331
55,935
Yr2
58,266
2,331
55,935
Yr3
58,266
2,331
55,935
Yr4
58,266
2,331
55,935
Yr5
58,266
2,331
55,935
Yr6
58,266
2,331
55,935
Yr7
32,460
1,298
31,162
Yr8
‐
‐
‐
4,393
4,393
‐
‐
‐
Payability of Metal in Conc
Zinc
Lead
Silver
Net Smelter Return
Less Royalties
Net Revenue
Operating Costs
Mining
Processing
G&A
Social & Environmental
US$/t ore
22.60
34.24
2.08
0.25
251,415
96,027
145,486
8,852
1,050
‐
‐
‐
‐
‐
732
732
‐
‐
‐
33,207
9,519
22,188
1,350
150
38,332
14,645
22,188
1,350
150
38,332
14,645
22,188
1,350
150
38,332
14,645
22,188
1,350
150
38,332
14,645
22,188
1,350
150
38,332
14,645
22,188
1,350
150
21,421
8,159
12,361
752
150
‐
‐
‐
‐
‐
Operating Margin
27.15
115,359
‐
(732) (4,393) 22,729
17,603
17,603
17,603
17,603
17,603
9,740
‐
Capital Costs
Engineering Studies
Mining Capital
Processing Capital
Infrastructure Capital (including Contingency)
13.27
‐
7.89
1.47
3.91
56,374
‐
33,540
6,241
16,593
9,082
‐
6,986
‐
2,096
7,721
‐
4,643
1,248
1,830
19,498
‐
6,854
4,993
7,651
4,925
‐
3,764
‐
1,160
4,925
‐
3,764
‐
1,160
4,925
‐
3,764
‐
1,160
4,925
‐
3,764
‐
1,160
31
‐
‐
‐
31
31
‐
‐
‐
31
312
‐
‐
‐
312
‐
‐
‐
‐
‐
Change in Working Cap
‐
‐
‐
‐
‐
10,016 (140) ‐
‐
‐
‐
10 (9,885)
Pre‐tax c/flow
Tax payable
C/flow after tax
Cumulative C/Flow
Discounted C/Flow (8%)
Cumulative DCF
Max funding reqmt to positive cashflow
13.88
3.74
10.15
58,985 (9,082)
15,878 ‐
43,107 (9,082)
(9,082)
14,529 (9,082)
(9,082)
(56,367) (9,082)
17,572
3,924
13,648
25,862
7,374
6,268
‐
9,418
2,058
7,360
33,222
3,682
9,950
‐
(8,453)
‐
(8,453)
(17,535)
(7,827)
(16,909)
(17,535)
(23,891)
‐
(23,891)
(41,426)
(20,483)
(37,392)
(41,426)
7,788 12,819
‐
‐
7,788 12,819
(33,638) (20,819)
6,183 9,422
(31,209) (21,787)
(56,367) (38,422)
137
12,678 12,678 17,572
2,569 3,559 3,768
10,109 9,119 13,804
(10,710) (1,590) 12,214
6,880 5,747 8,055
(14,907) (9,160) (1,105)
(28,313) (19,194) (5,389)
9,885
‐
9,885
43,107
4,579
14,529
‐
Savings in the process plant and tailings dam construction costs result in a reduction of
approximately $33.1 million in capital invested before positive cash flow is achieved, with
$56.4 million required for toll-milling compared to $89.5 million in the base case.
Nevertheless, Figure 18.15 (above) shows that, although payback on the undiscounted cash
flow occurs in Year 4, the LOM net cash flow after tax of $43.1 million is $37.4 million less
than is forecast in the base case ($80.5 million).
Moreover, the toll-milling option does not appear to maximize project value, since its pre-tax
NPV8 of $27.0 million is $23.0 million less than the base case. Similarly, the after-tax NPV8
of $16.4 million for the toll milling option is approximately half that of the base case ($33.0
million)
Figure 18.16 indicates that, compared to the base case, the toll milling option is more
sensitive to changes in operating costs, and less sensitive to capital costs, as is to be expected.
Figure 18.16
Toll-Milling Option - Sensitivity
60
50
40
30
20
10
0
(10)
(20)
(30)
70 75 80 85 90 95 100 105 110 115 120 125 130 Product Price (27.4 (18.4 (11.0 (4.0)
2.8
9.6
16.4 23.3 30.1 36.9 43.7 50.4 57.2
Opcosts
40.9 36.9 32.8 28.7 24.6 20.5 16.4 12.4
8.3
4.2
.1
Capex
30.6 28.2 25.9 23.5 21.2 18.8 16.4 14.1 11.7
9.4
7.0
(4.0) (8.2)
4.7
2.3
Figure 18.17 shows that, when evaluated at each of the cut-off grade intervals, the on-site
milling option consistently provides a superior economic return. At the optimum cut-off
grade of 200 g/t Ag Eq, the NPV of on-site milling is more than twice that of the toll-milling
equivalent. Even at higher cut-offs, when toll milling becomes more attractive, the NPV of
this option does not exceed that of the base case (on-site milling).
138
35
35%
30
30%
25
25%
20
20%
15
15%
10
10%
5
5%
0
0%
125
150
NPV (Mill)
175
200
225
Cut‐off grade (g/t Ag Eq)
NPV (Toll)
IRR (Mill)
250
IRR (%)
Figure 18.17
On-Site Milling versus Toll-Milling Option
275
IRR (Toll)
Nevertheless, because of the reduction of capital, at cut-off grades above 250 g/t Ag Eq, toll
milling appears to offer an improved internal rate of return, with IRR of 19.4% and 20.7%
after tax at 250 g/t and 275 g/t respectively, compared to rates of 18.6% and 17.6%
respectively in the base case.
Toll-milling is, therefore, shown to be economic at higher cut-offs. Nevertheless, on
economic grounds, Micon concludes that the selection of on-site milling for development of
the Pulacayo project is optimal.
139
19.0
INTERPRETATION AND C
CONCLUSIONS
The mineralization at Pulacayo is a typical low sulphidation epithermal deposit containing
precious and base metals associated with volcan
volcanic
ic rocks. The main geological characteristics
of Pulacayo are:
•
The sulphide mineralization is hosted by Tertiary volcanic rocks of intermediate
composition. These rocks form part of a dome complex, which outcrops at surface.
The mineralized body is compos
composed
ed of stockwork, narrow veins and veinlets, and
disseminations in the argillic
argillic-altered rock controlled by an east-west
west oriented normal
fault system. The width of the mineralization varies from 40 m to 120 m.
•
Sedimentary rocks intruded by the dome complex constitute the host rock for a
bonanza type, high grade vein (Veta Tajo), with high silver and base metals content.
The vein structure rarely is wider than 3 m and continues into the overlying
stockwork and disseminated
sseminated zone in the volcanic rocks.
•
The sulphide mineralization extends along strike for 2,700 m and by almost 1,000 m
to depth, of which 450 m are hosted in the volcanic unit and 550 m are hosted in the
sedimentary unit.
•
The mineral assemblage is relatively simple: barite, quartz, pyrite, calcite as gangue
minerals; and galena, sphalerite, tetrahedrite, and other silver sulfo-salts
sulfo
as ore
minerals. There is also minor chalcopyrite and jamesonite. The internal texture of the
veins is generally bbanded
anded and drusy with segments containing almost massive
sulphides. A vertical zonation appears to exist where base metals increase at depth
and silver content is higher at mid levels.
Micon has estimated the mineral resources at Pulacayo to be as shown in Table 19.1. Details
concerning the preparation of this estimate are given in Section 17 of this report. The
effective date of this estimate
ate is October 14, 2009.
Table 19.1
Classification
Indicated
Inferred
Tonnes
4,892,000
6,026,000
Ag (g/t)
79.96
98.26
Pb (%)
0.79
0.78
Zn (%)
1.64
1.68
(2) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources
res
may be materially affected
cted by environmental, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues.
(3) The quantity and grade of reported inferred resources in this estimation are conceptual in nature and there has been insufficient
insu
explorationn to define these inferred resources as an indicated or measured mineral resource. And it is uncertain if further
exploration will result in upgrading them to an indicated or measured mineral resource category.
140
The project base case comprises the development of an underground mine connecting to
existing workings though a new adit portal, extraction using a sub
sub-level
level open-stoping
open
method
with backfill, feeding 1,800 t/d to a new milling and flotation plant on site, for the production
and sale of lead and zinc concentrates containing economically im
important
portant silver values, and
storage of flotation tailings in a new, purpose
purpose-built
built facility adjacent to the new plant.
With respect to the potential underground m
mining:
•
SLOS mining with backfill has been planned to be the primary mining method. This
is intended to deliver maximum productivity at a lower unit operating cost than more
selective methods.
•
Underground access will be gained by developing a new portal and adit to intersect
the San Leon adit at a position to the south of the orebody. The new portal
port will be
situated in close proximity to the processing plant.
•
Trucks will haul the ore from the underground and deliver it to the processing plant.
•
No geotechnical assessment has been made. However, from observation of the drill
core and from limited exposure of rock underground, it has been judged that the
modifying factors are appropriate for the stoping method selected, at this level of
study.
•
Extensive historical mining areas exist underground, these will need to be surveyed in
detail prior to more detailed mine planning.
Preliminary metallurgical testwork has been carried out which suggests that, based on a
conventional flowsheet, an average mill feed grade of 199 g/t Ag, 1.21% Pb and 2.13% Zn
will yield a lead concentrate assaying 47% Pb and containing 70.3% of the silver, 87.4% of
the lead and 3.9% of the zinc. The zinc concentrate will assay 58.3% Zn and recover 13.6%
of the silver, 2.2% of the lead and 85.2% of the zinc
zinc, as summarized
rized in Table 16.7.
16.7
The flowsheet consists of a crushing circuit followed by Semi-Autogenous
Autogenous Grinding (SAG),
differential lead/zinc flotation cells, concentrate dewatering, and tailings solids deposition
into a Tailings Storage Facility (TSF). Process water reclaimed from the TSF will be
pumped back to the mill forr reuse.
The preliminary assessment of this base case shows it to be economic, with an IRR of 24%
and NPV8 of $50.0
.0 million before tax. Payback is in Year 4, leaving almost 3 years further
production in the ‘tail’.
An alternative scenario, with toll
toll-milling
ing of the underground mine production at the Don
Diego mill, is also shown to be potentially economic. This option has a reduced capital
requirement, resulting in an improved IRR before tax of 27.
27.3%,
%, although the NPV8 is lower
at $27.0 million before tax.
141
20.0
RECOMMENDATIONS
Geology and Mineral Resource estimation considerations:

Analysis of duplicate samples as part of the Quality Control program should be
carried out at a laboratory that is a separate corporate entity from the laboratory that
conducted the primary analyses.

This updated mineral resource estimate is prepared with the objective of providing a
global estimate of the tonnage and average grade of the relatively narrow mineralized
material present, that is envisioned to be extracted by means of underground mining
methods. Specifically, detailed modeling of the narrow higher grade vein structures
will be required, should a local estimate of the amount of material amenable to
underground mining methods be required. In support of this local estimate, Micon
recommends that additional information be acquired in the form of in-fill drilling to
confirm the continuity and grades of these narrow, high grade veins between the
existing observation points.

As described above, a small number of drill holes have been completed (PUD-134 to
PUD-139 inclusive) and have been geologically logged, but no samples have been
taken due to budgetary constraints. Micon recommends that sampling and assaying
be completed for these drill holes as funding becomes available.

Given the lack of detailed information regarding the inclination of the floor of the
drifts, for the purposes of this initial mineral resource estimate Micon assumed an
inclination of zero (i.e. a flat floor) for all of the levels modeled. As well, for the
purposes of this initial mineral resource estimate, Micon assumed a constant crosssectional
dimension
of
3.0 m (width) x 3.7 m (height) for all of the modeled drifts on the basis of the results
of examination of the indicated drift widths on a number of the level plans. Should
the project proceed to a more advanced state, Micon recommends that the precise
location and inclination of the development drives be established by detailed survey
methods.

As well, it is to be noted that the level plans for three of the upper levels have not
been located (the 4,252 m, 4,282 m and 4,316 m levels). Micon recommends that
efforts continue to be directed towards location of the records of these levels and
integration of their results with the remainder of the model of the mine workings.

For the purposes of this initial mineral resource estimate, Micon assumed a constant,
average stope width of 3 m for the model of the mined out voids. Micon recommends
that should the project proceed to a more advanced state, the shape of the mined out
stopes be determined by appropriate methods to an appropriate degree of accuracy.
142

The rock chip sample results that were collected in 2005 from the Veta Tajo and Veta
Cuatro areas be integrated into the drill hole/sampling database. In addition, Micon
recommends that a comprehensive program of chip/channel sampling be carried out
of the higher grade mineralization that is found in those portions of the underground
workings that can be safely accessed.

In consideration of the range of specific gravities observed in the sample data, Micon
recommends that should the project proceed to a more advanced state, additional
density measurements should be taken from samples chipped from the walls of the
existing mine workings to assist in filling in the gaps in the spacing of the
information. The information from these new samples should be integrated with the
existing specific gravity database and the density of each block in the model should
be estimated in detail so as to provide a more accurate local estimate of the tonnages.
Care will need to be taken in order to obtain an accurate specific gravity measurement
for samples that are porous.

Should the project proceed to a more advanced state, a program of geotechnical
characterization of the wall rocks should be carried out in support of mine design.
Based partly on this information, a detailed geotechnical study should be carried out
that will provide the basis of more detailed mine planning.

For resources to be stated as reserves, inferred resource material may need to be
upgraded to the measured and indicated categories to enhance economic viability.
The resources targeted should be in the vicinity of existing infrastructure to minimize
additional capital cost.

The contents of the stopes should be determined and, in conjunction with the results
of the survey recommendations made above, further mine planning should be carried
out to define the potential stope interactions. This will enable planned mining access,
dilution and mining recovery to be more closely estimated.

Subsequent levels of study should incorporate a more detailed mine development,
production and backfill schedule, in order to optimize capital costs and cash flow.
With respect to metallurgy:

Complete pressure-filter moisture tests on the lead and zinc concentrates to confirm
concentrate moistures will be less than 8 wt%. This is required for transport by ship.
If this is not attainable, a disk-filter, gas fired dryer or an atmospheric drying pad may
be required.

Review and modify the flowsheet and equipment selection to maximize silver
recovery from the clay fraction.
143

Bench tests should be completed to determine how the clay fraction responds in the
TSF, both for reclaim water clarity and deposition density for TSF volume
calculations.

For the toll milling option, recalculate the ore transport charges from the Pulacayo
mine to the Don Diego mill, after the road improvements to Highway 701 are
completed, which should be around the first quarter of 2011 (see description in
Section 18.3.4). The direct trucking of ore on this route would significantly reduce
the transport charges, since Highway 701 is the most direct route between Pulacayo
and Don Diego.

A detailed inspection of the eight inch, fresh water pipeline from the Yana reservoir
to Pulacayo is required to determine if a new pipeline is needed. A new pipeline was
estimated in this report to cost $1.5 million. At the same time a pipeline profile
should be done on the existing pipeline, including elevation, lengths, and pressure
measurements.
Environmental and Social Considerations:

Waste rock should not be used for construction.

The waste rock and tailings disposal design and water management plans need to
consider the acid generating and metal leaching properties of the waste rock and
tailings.

It is recommended that the impact assessment further document the extent of
historical contamination.

It is recommended that further project design take historical works into consideration
and remediate historical contaminants where possible.

Community consultation should continue and a Community Development Plan be
developed in concert with the community.
With respect to Project Development:

Project exploration and development should proceed together. The base case would
be strengthened by additional mineral resources to extend the LOM further beyond
the payback period.

The toll-milling scenario remains attractive while resource tonnage is limited – this
can therefore be viewed as a fall-back scenario in the event that permitting and/or
construction of a mill on the site cannot be completed expeditiously.
144
21.0
REFERENCES
(1992) Geology and Mineral Resources of the Altiplano and CordilleraOccidental, Bolivia.
US Geological Survey Bulletin 1975. Denver, CO, USA. 1992.
(1997) CODIGO DE MINERIA, Gaceta Oficial de Bolivia. La Paz, Bolivia. 17 March,
1997.
(2003) Overview-San Cristobol Project, Apex Silver Ltd. www.apexsilver.com December,
2003.
Ahlfeld, F., and Schneider-Scherbina, A., 1964, Los Yaciamientos Minerales Y De
Hidrocarburos De Bolivia: Ministerio De Minas Y Petroleo, Boletin No. 5 (Especial), p.
297-300.
Andean Silver Corporation, 2002, Informe Final, Programa de Perforación, 1er. Etapa., ASC
Bolivia LDC. Octubre 2002.
Andean Silver Corporation, 2003, Informe Final, Programa de Perforación, 2da. Etapa., ASC
Bolivia LDC. February 2003.
Andean Silver Corporation, 2003, Informe Final, Programa de Perforación, 3era. Etapa.,
ASC Bolivia LDC. Sep-Nov 2003.
Andean Silver Corporation, 2002, Press Release: “Apex Silver Mines Limited Updates
Bolivian Exploration Activities”. Filling-Apex Silver Mines Limited (AMEX:SIL) October
23, 2002.
Andean Silver Corporation, 2003, Informe Final, Programa De Perforacion, 2da. Etapa., ASC
Bolivia LDC. February 2003.
Andean Silver Corporation, 2003, Preliminary Metallurgical Evaluation of Samples From
Pulacayo Prospect, Bolivia. March 17, 2003. Denver, Co.- EUA.
Apex, 2002, Press Release: “Apex Silver Mines Limited Updates Bolivian Exploration
Activities”. Filing-Apex Silver Mines Limited (AMEX:SIL) October 23, 2002.
Codigo de Mineria (1997), Gaceta Oficial de Bolivia. La Paz-Bolivia.
Consultora Técnica Eliezer, 2007, INFORME TECNICO - Georeferenciación de puntos de
Control Horizontal y Vertical - Pulacayo - Octubre 2007; La Paz-Bolivia.
Consultora Técnica Eliezer, 2007, INFORME TECNICO - Levantamiento Topográfico a
detalle del Centro Minero Pulacayo. Diciembre 2007; La Paz-Bolivia.
145
Consultora Técnica Eliezer, 2007, INFORME TECNICO - Replanteo de Líneas Geofísicas
de Pulacayo y Paca - Diciembre 2007; La Paz-Bolivia.
Fractal, 2007, Sondeo de Líneas Geofísicas por Polarización Inducida IP.
GEOBOL, 1969, Estudio Geológico del Yacimiento de Pulacayo-Huanchaca, Servicio
Geológico de Bolivia; La Paz-Bolivia.
Gustavson Associates Inc. y Servicio Geológico de Bolivia, 1992, Compendio de Geología
Económica de Bolivia; Ministerio de Minería y Metalurgia; La Paz-Bolivia.
Hamilton, G. (2004) Don Mario Mine Production, Press Release, Orvana Minerals Corp.
www.orvana.com May 2004.
Head, R.E., 1939, Microscopic Study of Ore and Rock Specimens, Pulacayo Mine;
COMIBOL, La Paz-Bolivia.
Hedenquist, J., Arribas, A., and Gonzalez-Urien, E., 2000, Exploration for Epithermal Gold
Deposits: SEG Reviews, Vol 13, p. 245-277.
Heuschmidt, B., Miranda-Angles, V., Bellot-La Torre, J., Claire-Zapata, M., CazasSaavedra, A. (2000) Sinopsis De La Metalogenia En Bolivia. Servicio Nacional de
Geologia Y Mineria (SERGEOMIN) No 19, Ano 2000. La Paz, Bolivia 2000, pg 39.
Heuschmidt, B., Bellot-La Torre, J., Miranda-Angles, V. (2000) Las
Provincias
Y
Epocas Metalogenicas De Bolivia En Su Marco Geodinamuco (Bolivian Provinces And
Metallogenic Epochs In Its Geodynamic Context). Chapter 9, in Compendio De Geologia
De Bolivia. Servicio Nacional de Geologia Y Mineria (SERGEOMIN) No 1-2, Ano
2002. Soruco, R. S. editor. La Paz, Bolivia. June 2000, pg 188.
Heuschmidt, B., Bello-La Torre, J., Miranda-Angles, V., Claire-Zapata, M. (2002) Las
Areas Prospectivas de Bolivia Para Yacimientos Metaliferos. Boletin Del Servicio
Nacional de Geologia Y Mineria No 30 Ano 2002. Del Servicio Nacional de Geologia Y
Mineria (SERGEOMIN) La Paz, Bolivia 2002, pp 44-47.
Lindgren, W., 1922, A Suggestion for the Terminology of Certain Mineral Deposits:
Economic Geology, v. 17, p. 292-294.
Malhotra, D., 2003, Preliminary Metallurgical Evaluation of Samples from Pulacayo Project,
Bolivia: Unpublished Internal Report by RDI Resource Development, 16 p.
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Base (ALBA). Prepared on behalf of Apogee Minerals Bolivia S.A. (Draft).
146
Pressacco, R., and Shoemaker, S., 2008, Technical Report for the Pulacayo Project, Potosí
District, Quijarro Province, Pulacayo Township, Bolivia: Unpublished. Document available
at www.SEDAR.com, 207 pp.
Pressacco, R., and Gowans, R., 2007, Technical Report on the Mineral Resource Estimate for
the Paca Deposit, Potosí District, Quijarro Province, Thols, Pampa, Huanchaca and Pulacayo
Townships, Bolivia: Unpublished Document available on the SEDAR web site at
www.SEDAR.com, 226 p.
Roaseler, W. T., 1943, An Analysis of the Napoleon Pero Fault and the Tajo Vein in its
Vicinity, COMIBOL; La Paz-Bolivia.
Servicio Geológico de Bolivia, 1988, Modelo Conceptual y Evaluación del Potencial Mineral
del Yacimiento de Metales Preciosos Pulacayo; La Paz-Bolivia.
Servicio Geológico de Bolivia, 1989, Mapa Geológico y Muestreo en Detalle del Yacimiento
de Pulacayo; La Paz-Bolivia.
Shatwell, D. (1998) Gold Metallogenesis of the Andean Region. David Shatwell Pty Ltd.
February, 1998
Sillitoe, R., and Hedenquist, J., 2003, Linkages Between Volcanotectonic Settings, Ore-Fluid
Compositions, and Epithermal Precious-Metal Deposits: Society of Economic Geologists
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Soruco, R. S., 2000, Compendio de Geologia de Bolivia - Revista Technica de Yacimentos
Petroliferous Fiscales Bolivianos. Vol 18, No 1-2, June 2000. Servicio Nacional de
Geologia Y Mineria Yacimentos Petroliferous Fiscales Bolivianos. La Paz, Bolivia, 2000.
Turneaure, F.S., 1945, Report on the Geology of the Pulacayo Mine and the Development of
the Tajo vein, COMIBOL; La Paz-Bolivia.
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147
22.0
SIGNATURES
The effective date of this report is April 30, 2010.
“Reno Pressacco”
___________________________
Reno Pressacco, M.Sc.(A), P.Geo.
Formerly, Senior Geologist
June 25th, 2010
_______________________
“Geraint Harris”
___________________________
Geraint Harris, CEng, MAusIMM
Senior Mining Engineer
June 25th, 2010
_______________________
“Michael Godard”
___________________________
Michael Godard P.Eng.
Senior Metallurgist
June 25th, 2010
_______________________
“Chris Jacobs”
___________________________
Christopher Jacobs, CEng MIMMM
Vice President
June 25th, 2010
_______________________
148
23.0
CERTIFICATES
149
CERTIFICATE OF AUTHOR
Reno Pressacco
As co-author of this report entitled “Technical Report on the Preliminary Assessment of the Pulacayo
Project, Pulacayo Township, Potosí District, Quijarro Province, Bolivia” dated June 25th, 2010, I,
Reno Pressacco, do hereby certify that:
1.
I was employed as a Senior Geologist by, and carried out this assignment for, Micon
International Limited, Suite 900, 390 Bay Street, Toronto, Ontario M5H 2Y2, tel. (416) 3625135, fax (416) 362-5763, e-mail [email protected].
2.
I hold the following academic qualifications:
CET (Geological Engineering)
Cambrian College
1982
B.Sc (Geology)
Lake Superior State College
1984
M.Sc(A). (Mineral Exploration)
McGill University
1986
3.
I am a Qualified Person as defined in the Instrument.
4.
I am a registered Professional Geoscientist with the Association of Professional Geoscientists
of Ontario (Registration Number 0939); as well, I am a member in good standing of other
technical associations and societies, including the Prospectors and Developers Association of
Canada.
5.
I have worked as a geologist in the minerals industry for 28 years. My experience includes
mineral exploration, advanced exploration and mine development, open pit production,
environmental compliance, financial evaluation and mine commissioning with a variety of
deposit types including gold, silver, copper, zinc, lead, uranium, nickel, platinum-group
metals and industrial minerals.
6.
I visited the subject property, reviewed data and drill core on March 26th to March 28th, 2008.
7.
I am responsible for Sections 4 to 15 inclusive and Section 17 of this report.
8.
I am independent of the issuer for which this report is required, other than providing
consulting services.
9.
My prior involvement with the Pulacayo property was as co-author of the “Technical Report
for the Pulacayo Project, Potosí District, Quijarro Province, Pulacayo Township, Bolivia”,
dated December, 2008 and addressed to Apogee Minerals Ltd.
10.
I have read the Instrument and the Technical Report is prepared in compliance with the
Instrument.
11.
As of the date of this certificate, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be
disclosed to make this Technical Report not misleading.
Dated this 25th day of June, 2010
“Reno Pressacco”
Reno Pressacco, P.Geo
150
Geraint Harris
Geraint Harris, do hereby certify that:
1. I am employed by, and carried out this assignment for Micon International Co. Limited,
Suite 10, Keswick Hall, Keswick, Norwich, Norfolk, UK NR4 6TJ
Tel: +44 (1603) 501 501, email: [email protected]
2. I hold the following academic qualifications:
i.
B.Eng. (Honours) Mining Engineering, University of Nottingham, 1995
ii.
M.Sc., Mining Engineering, Mackay School of Mines, Nevada, United States, 1997;
3. I am registered by or belong to the following professional associations: The Irish Institute of
Engineers (Registered Chartered Engineer, CEng); The Australian Institute of Materials, Mining
and Metallurgy, (Member, AusIMM).;
4. I have worked in the minerals industry for 14 years. I do, by reason of education, experience and
professional registration, fulfill the requirements of a Qualified Person as defined in NI 43-101.
My work experience includes 11 years as a mining engineer at gold, base metal and other mines
and 3 years as a consulting mining engineer.
5. I have visited Apogee’s Pulacayo property between the 6th and 7th August, 2009.
6. I am responsible for the preparation of Sections 18.1, 18.5.4.2 and 18.5.5.1 and portions of
Sections 1, 19 and 20 of this report.
7. I am independent of Apogee Minerals Ltd. as defined in Section 1.4 of NI 43-101;
8. I have had no prior involvement with the mineral properties in question;
9. I have read NI 43-101 and the portions of this report for which I am responsible have been
prepared in compliance with the instrument;
10. As of the date of this certificate to the best of my knowledge, information and belief, the technical
report contains all scientific and technical information that is required to be disclosed to make this
report not misleading
“Geraint Harris”
Geraint Harris, CEng, MAusIMM.
151
Michael Godard
Michael Godard, do hereby certify that:
1. I am employed by, and carried out this assignment for Micon International Limited,
205 – 700 West Pender Street, Vancouver, BC, V6C 1G8,
Tel: 604-647-6463, email: [email protected].
2. I hold the following academic qualifications:
Bachelor of Applied Science Degree (Metallurgy) University of British Columbia, May, 1985
3. I am a Professional Engineer registered with the Association of Professional Engineers and
Geoscientists of BC, APEGBC, (registration number 33114);
4. I have worked in the minerals industry for 22 years;
5. I do, by reason of education, experience and professional registration, fulfill the requirements of a
Qualified Person as defined in NI 43-101. My work experience includes over 25 years of
experience in design, commissioning and process engineering within the oil sands extraction,
mineral processing and metals fabrication industries.
6. I have visited Apogee’s Pulacayo property on July 30, 2009, and Glencore’s Potosi mill (toll
milling) on July 29, 2009.
7. I am responsible for the preparation of Section 16, Sections 18.2, 18.3, 18.5.4.3, 18.5.4.4, 18.5.5.2
and parts of Sections 1, 19 and 20.
8. I am independent of Apogee Minerals Ltd. as defined in Section 1.4 of NI 43-101;
9. I have had no prior involvement with the mineral properties in question;
10. I have read NI 43-101 and the portions of this report for which I am responsible have been
11. As of the date of this certificate to the best of my knowledge, information and belief, the technical
report contains all scientific and technical information that is required to be disclosed to make this
report not misleading;
“Michael Godard”
Michael Godard, P.Eng.
152
Christopher A. Jacobs
Christopher Jacobs, do hereby certify that:
1.
I am employed by, and carried out this assignment for:
Micon International Limited, Suite 900 – 390 Bay Street, Toronto, ON, M5H 2Y2
tel. (416) 362-5135
email: [email protected]
2.
I hold the following academic qualifications:
B.Sc. (Hons) Geochemistry, University of Reading, 1980;
M.B.A., Gordon Institute of Business Science, University of Pretoria, 2004.
3.
I am a Chartered Engineer registered with the Engineering Council of the U.K.
(registration number 369178);
Also, I am a professional member in good standing of: The Institute of Materials, Minerals and
Mining; and The Canadian Institute of Mining, Metallurgy and Petroleum (Member);
4.
I have worked in the minerals industry for 28 years; my work experience includes 10 years as
an exploration and mining geologist on gold, platinum, copper/nickel and chromite deposits; 10
years as a technical/operations manager in both open pit and underground mines; 3 years as
strategic (mine) planning manager and the remainder as an independent consultant;
5.
I do, by reason of education, experience and professional registration, fulfill the requirements
of a Qualified Person as defined in NI 43-101;
6.
I have not visited the Pulacayo Property;
7.
I am responsible for the preparation of Sections 2, 3, 18.5 (except sections 18.5.4.2 to 18.5.4.4
and 18.5.5.1 to 18.5.5.2) and portions of Sections 1, 19 and 20 of this report, entitled
“Technical Report on the Preliminary Assessment of the Pulacayo Project, Pulacayo Township,
Potosí District,Quijarro Province, Bolivia” dated June 25th, 2010,. I also take responsibility for
Section 18.4 of this report, which was prepared under my supervision by Ms Jenifer Hill,
R.P.Bio.;
8.
I am independent of Apogee Minerals Ltd., as defined in Section 1.4 of NI 43-101;
9.
I have had no prior involvement with the mineral property in question;
10.
I have read NI 43-101 and the portions of this report for which I am responsible have been
11.
As of the date of this certificate to the best of my knowledge, information and belief, the
sections of this Technical Report for which I am responsible contain all scientific and technical
information that is required to be disclosed to make this report not misleading.
“Christopher A. Jacobs”
Christopher A. Jacobs, CEng MIMMM
153
24.0
APPENDICES
154
Agreement I
APPENDIX I
APOGEE Minerals Bolivia S.A./ASC Bolivia LDC Option Agreement (03/08/2006)
Apogee Minerals Bolivia S.A. and ASC Bolivia LDC (Sucursal Bolivia) executed an Option
Agreement on 03/08/2006 to establish a joint venture agreement whereby Apogee may earn
an a 60% interest in the Pulacayo Cooperative Ltda / ASC Bolivia LDC Joint Venture
Agreement and also to have 60% participation in the Paca Group of mining properties.
The terms to acquire a 60% participation in both areas (Pulacayo and Paca) are as follows:
Properties:
Pulacayo Group – COMIBOL (2,703ha):
Pulacayo (1,031 ha) Porvenir (1,099 ha)
Huanchaca (460 ha) Galería General (76 ha)
Roschild (3 ha) Temeridad (10 ha) Real del Monte (24 ha)
Paca Group (60% interest) – ASC Bolivia LDC (31,175 ha):
Apuradita (750 ha)
Apuradita II (1,275 ha)
Sally (125 ha)
Tatoe (875 ha)
Khullku (2,775 ha)
Lupitaca (675 ha)
Phico Grande (9,400 ha)
Khasa Pampa (4,150 ha)
El Encanto (11,150 ha)
Start Date: September 9, 2005
Term: Three Years
Payments:
a) $1,000 per month starting from signature to COMIBOL.
b) $1,500 per month starting from signature to the Pulacayo Cooperative.
Expenditure Commitments:
a) $250,000 in the first six months (to February 9, 2006)
b) $1.0 million during the term of the Agreement.
Other Commitments:
a) Prepare a Bankable Feasibility Study during the term of the agreement.
This agreement was ratified by both the Boards of the Pulacayo Cooperative and COMIBOL.
155
Agreement II
Pulacayo Cooperative Ltda. / ASC Bolivia LDC Joint Venture Agreement (effective
07/30/2002)
The Pulacayo Mining Cooperative and ASC BOLIVIA LDC (Sucursal Bolivia) executed a
Joint Venture Agreement on 07/30/2002 for the exploration and development of the Pulacayo
Group of tenements. This agreement satisfied an expectation in an agreement signed in 1997
between the Pulacayo Cooperative and COMIBOL (see below) that a “Strategic Partner”
would become involved in exploration of the properties. This Joint Venture Agreement also
provides that a third party could be integrated into the Joint Venture with the permission of
COMIBOL‟s Board.
Properties:
Pulacayo Group – COMIBOL (2,703 ha):
Pulacayo (1,031 ha)
Porvenir (1,099 ha)
Huanchaca (460 ha)
Galería General (76 ha)
Roschild (3 ha)
Temeridad (10 ha)
Real del Monte (24 ha)
Ubina Group – COMIBOL (397 ha):
Santa Bárbara (149 ha),
La Esperanza (148 ha),
Flora (60 ha)
Victoria (40 ha).
Cholita Chaquiri Group – COMIBOL (230 ha):
Cholita (10 ha)
Tolentino (220 ha)
Start Date: July 30, 2002
Term: 5 Years Exploration; Total 23 Years
Payments:
a) $1,000 per month during the exploration period to COMIBOL.
a) $500,000 during First Stage of Exploration.
b) 2.5% Net Smelter Return royalty payable to COMIBOL.
c) 1.5% Net Smelter Return royalty payable to the Pulacayo Cooperative.
Agreement III
156
COMIBOL / Pulacayo Ltda. Lease Agreement (effective 08/01/97)
The Bolivian Mining Corporation (COMIBOL) and the Pulcayo Ltda. Mining Cooperative
executed a Lease Agreement on 08/01/1997, principally to allow the Pulacayo Cooperative to
mine within the historical Pulacayo Mine. The agreement provided for involvement of a third
party (“Strategic Partner”) with a renowned name and capacity in the mining industry with
permission from COMIBOL‟s Board.
Properties:
Pulacayo Group – COMIBOL (2,703 ha):
Pulacayo (1,031 ha)
Porvenir (1,099 ha)
Huanchaca (460 ha)
Galería General (76 ha)
Roschild (3 ha)
Temeridad (10 ha)
Real del Monte (24 ha)
Ubina Group – COMIBOL (397 ha):
Santa Bárbara (149 ha),
La Esperanza (148 ha),
Flora (60 ha)
Victoria (40 ha).
Cholita Chaquiri Group – COMIBOL (230 ha):
Cholita (10 ha)
Tolentino (220 ha)
Start Date: June 1997
Term: 15 Years (June 2012) extended to June 23, 2025 if Strategic Partner is acquired.
Payments:
a) Rent equal to 1% of the net production value.
a) Strategic Partner to pay $1,000 per month “rent” to COMIBOL during a 5 Year
Exploration Period
b) Minimum investment of $300,000 in exploration costs.
157
APPENDIX II
SUMMARY OF DRILL HOLE COLLARS PUD-111 TO PUD-139
Hole Id
PUD111
PUD112
PUD113
PUD114
PUD115
PUD116
PUD117
PUD118
PUD119
PUD120
PUD121
PUD122
PUD123
PUD124
PUD125
PUD126
PUD127
PUD128
PUD129
PUD130
PUD131
PUD132
PUD133
PUD134
PUD135
PUD136
PUD137
PUD138
PUD139
Northing
7744437.23
7744489.75
7744436.30
7744488.91
7744451.44
7744488.94
7744451.44
7744488.95
7744494.25
7744437.14
7744488.65
7744484.97
7744484.97
7744484.97
7744484.97
7744484.97
7744488.37
7744487.61
7744486.14
7744487.61
7744487.20
7744486.47
7744488.43
7744489.75
7744489.75
7744491.97
7744491.97
7744489.97
7744484.91
Easting
739996.03
740406.74
740095.43
740406.27
740199.91
740406.27
740199.91
740406.27
740408.62
739996.02
740408.62
740407.79
740407.79
740407.79
740407.79
740407.79
740410.75
740410.09
740411.89
740410.09
740409.80
740409.29
740410.66
740406.74
740406.74
740406.79
740406.79
740406.79
740407.79
Elevation
4328.68
4135.00
4305.58
4135.21
4318.45
4134.60
4318.45
4134.90
4136.35
4328.60
4135.45
4135.45
4135.45
4135.45
4135.45
4135.45
4135.45
4135.45
4135.45
4135.45
4135.45
4135.45
4135.80
14135.00
14135.45
14136.45
14135.49
14135.45
14135.45
158
Depth
345.00
228.00
360.00
222.00
300.00
240.00
279.00
216.00
147.00
369.00
171.00
204.00
192.00
186.00
180.00
186.00
198.00
180.00
201.00
171.00
183.00
219.00
180.00
180.00
204.00
180.00
171.00
180.00
210.00
Dip
-48.00
-42.00
-55.50
-40.00
-48.20
-47.00
-47.20
-32.00
-10.00
-52.00
-26.00
-40.00
-25.00
-35.00
-40.00
-11.00
-23.00
-32.00
-38.00
-30.00
-37.00
-46.00
-18.00
-25.00
-35.00
-8.00
-20.00
-30.00
-40.00
Azimuth
0.00
353.00
359.10
345.00
0.60
345.00
19.00
345.00
6.00
0.00
6.00
6.00
24.00
24.00
24.00
41.00
41.00
41.00
41.00
35.00
35.00
35.00
37.00
353.00
353.00
338.00
338.00
338.00
338.00
APPENDIX III
VARIOGRAMS
159
160
161
162
163
164
165
166
167
APPENDIX IV
UTO Metallurgical Reports
168
UNIVERSIDAD TÉCNICA DE ORURO
FACULTAD NACIONAL DE INGENIERÍA
CARRERA DE METALURGIA Y CIENCIA DE MATERIALES
EXPERIMENTACION METALURGICA CON MUESTRA
COMPLEJA DE SULFUROS, DENOMINADA “LEY ALTA”
PROVENIENTE DEL SECTOR DE PULACAYO Y
PERTENECIENTE A LA EMPRESA
APOGEE MINERALS BOLIVIA S.A.
LABORATORIO
CONCENTRACIÓN
ACIÓN
CONCENTR
DE M
MINERALES
INERALES
DE
RESUMEN TECNICO
INFORME Nº 11/09
EXPERIMENTACION METALURGICA CON
MUESTRA COMPLEJA DE SULFUROS
DENOMINADA ALTA LEY Y PROVENIENTE
DEL SECTOR DE PULACAYO Y
PERTENECIENTE A LA EMPRESA APOGEE
MINERALS BOLIVIA S.A.
La Empresa Apogee Minerals Bolivia S. A., por intermedio del Ing. Luis Casal, personero
de la empresa, con correos sucesivos a través del internet y visita a nuestro laboratorio de
un consultor extranjero y del vicepresidente de exploración de la empresa, oficializaron la
realización de pruebas metalúrgicas, encomendando para ello al Laboratorio
Concentración de Minerales, de la Carrera de Ingeniería Metalúrgica de la Facultad
Nacional de Ingeniería de la Universidad Técnica de Oruro, la experimentación de las
mismas con una muestra de complejos sufurosos de Zn, Pb y Ag con la finalidad,
principalmente, de recuperar los contenidos de estos elementos valiosos, por flotación
diferencial.
Para tal efecto se recibió la muestra en cantidad suficiente para la realización de todas
las pruebas programadas. La muestra, proviene de yacimiento primario y se denomina:
LA (alta ley). Esta muestra, típica de perforaciones de diamantina (probetas), tiene
tamaños de grano de hasta 4 pulgadas.
NOVIEMBRE, 2009
Oruro, Bolivia
El trabajo de investigación, a solicitud de la empresa, se encaminó a determinar el rango
de recuperación y grado de concentrados de plomo-plata y zinc-plata, en ciclo abierto,
obtenidos por flotación diferencial y a determinar el rango de recuperación y grado de
concentrados de plomo-plata y zinc-plata, en ciclo cerrado, obtenidos por flotación
diferencial; por otro lado, debe efectuarse análisis granulométricos de las colas de las
pruebas de flotación en ciclo abierto y realizar el análisis size by size de las pruebas de
flotación diferencial en ciclo abierto.
Así mismo, se deben determinar los contenidos de lamas (arcillas) en las colas de todas
las pruebas de flotación, a través de pruebas de ciclonaje, realizar pruebas de
sedimentación, a partir de las colas de flotación y determinar el Indice de Trabajo (Work
Index).
Dirección
Ciudadela Universitaria,
Edif. Carrera de Ingeniería Metalúrgica
Teléfono
591-2-5263888
Correo Electrónico
[email protected]
i
Los resultados alcanzados permiten tener confianza en cuanto a lo que podría lograrse en
una operación industrial.
En cuanto a esta muestra se refiere se puede afirmar claramente que la misma es apta de
ser tratada por el proceso de flotación diferencial ya que se logran obtener concentrados
con índices metalúrgicos bastante aceptables; esta situación se ha visto en las pruebas
tanto en circuito abierto como en ciclo cerrado.
.
En circuito abierto, los mejores índices metalúrgicos que se han logrado son, previo
deslame:
-
Radio de enriquecimiento de la plata, en el concentrado de plomo: 28.01
Radio de enriquecimiento de la plata, en el concentrado de zinc: 5.37
-
Radio de enriquecimiento del plomo: 33.83
Radio de enriquecimiento del zinc: 22.69
Radio de concentración del plomo: 42.55
Radio de concentración del zinc: 27.25
Recuperación de la plata, en el concentrado de plomo: 65.96%
Recuperación de la plata, en el concentrado de zinc: 19.73%
Recuperación total de la plata: 85.69%
Ley de la plata, en el concentrado de plomo: 8290 g/t Ag
Ley de la plata, en el concentrado de zinc: 1590 g/t Ag
Ley del plomo en el concentrado final de plomo: 51.76%
Ley de zinc en el concentrado final de zinc: 57.40%
Recuperación de plomo, 79.47%
Recuperación del Zinc: 83,06%
-
Recuperación del Zinc: 81.10%
También se debe mencionar que si bien la presencia de lamas es grande y perjudicial,
alrededor del 19% en peso, se puede realizar la flotación sin deslame, siempre y cuando
este porcentaje no suba.
Los análisis granulométricos de las colas y los análisis size by size permiten afirmar que la
mayor pérdida de los elementos valiosos en las colas se encuentran en tamaños de fina
granulometría, concretamente por debajo de 400 Mallas Tyler, -38 micrones; asimismo, se
debe resaltar que la velocidad de sedimentación de las partículas, sin la adición de
floculante, a partir de las colas de flotación, es lenta porque algo más del 56% en peso de
la muestra que entra al proceso de flotación está por debajo de la malla 400 y gran parte
de esta fracción corresponde a la presencia de lamas; esta velocidad es de 1.372 x 10-5
m/s. Finamente indicar que el Índice de Trabajo, Work Index, para una malla de corte de
100 Mallas Tyler, 150 micrones, es de 12.434 Kwh/tc.
Mientras que en circuito cerrado, sin deslame se han logrado estros resultados:
-
ii
iii
INDICE
Contenido
Pag.
Resumen ………………………………………………………….
i
Índice ………………………………………………………………
iv
1.
Introducción ……………………………………………………….
1
2.
Objetivo ……………………………………………………………
2
3.
Experimentación Metalúrgica …………………………………...
2
4.
Resultados y comentarios ………………………………………
8
4.1.1
Análisis químico del común …………………………………….
8
4.1.2
Flotación diferencial de sulfuros en circuito abierto ………….
8
4.1.3
Flotación en circuito cerrado ……………………………………
19
4.1.4
Análisis granulométrico de las colas de flotación …………….
23
4.1.5
Análisis size by size ……………………………………………...
26
4.1.6
Determinación del contenido de lamas en colas de flotación .
29
4.1.7
Pruebas de sedimentación a partir de las colas de flotación .
30
4.1.7.1
Con cola, non float, sin previo deslame ……………………….
31
4.1.7.2
Con cola, non float, previo deslame ……………………………
33
4.1.8
Ensayo estándar de Bond para determinación del Work
Index ……………………………………………………………….
34
4.1.8.1
Descripción de la muestra ………………………………………
35
4.1.8.2
Ensayo estándar …………………………………………………
35
4.1.9
Comentarios finales para la muestra LM ………………………
39
5.
Conclusiones ……………………………………………………..
41
Anexo ……………………………………………………………...
43
EXPERIMENTACION METALURGICA CON MUESTRA
COMPLEJA DE SULFUROS, DENOMINADA “LEY ALTA”
PROVENIENTE DEL SECTOR DE PULACAYO Y
PERTENECIENTE A LA EMPRESA
1. INTRODUCCION
La Empresa Apogee Minerals Bolivia S. A., por intermedio del Ing. Luis Casal, personero
de la empresa, con correos sucesivos a través del internet y visita a nuestro laboratorio de
un consultor extranjero y del vicepresidente de exploración de la empresa, oficializaron la
realización de pruebas metalúrgicas, encomendando para ello al Laboratorio
Concentración de Minerales, de la Carrera de Ingeniería Metalúrgica de la Facultad
Nacional de Ingeniería de la Universidad Técnica de Oruro, la experimentación de las
mismas con una muestra de complejos sufurosos de Zn, Pb y Ag con la finalidad,
principalmente, de recuperar los contenidos de estos elementos valiosos, por flotación
diferencial.
Para tal efecto se recibió la muestra en cantidad suficiente para la realización de todas
las pruebas programadas. La muestra, proviene de yacimiento primario y se denomina:
LA (alta ley). Esta muestra, típica de perforaciones de diamantina (probetas), tiene
tamaños de grano de hasta 4 pulgadas.
Una observación estereomicroscópica de la muestra, luego de una adecuada limpieza,
permite identificar pirita en forma mayoritaria, también se observan pequeñas cantidades
de sulfuros de plomo y zinc; está presente una gran cantidad de cuarzo-silicatos y
pizarras. Las muestras presentan características de formar lamas (por el contenido de
arcillas), aunque en menor proporción que otras muestras provenientes del mismo lugar
denominadas: LM y LB.
La representatividad de la muestra es responsabilidad de la Empresa; en esta etapa no
participó el Laboratorio Concentración de Minerales de la Carrera de Ingeniería
Metalúrgica.
1
iv
2. OBJETIVOS
Los objetivos del presente trabajo de investigación, a solicitud de la empresa, se
encaminaron a:
-
Determinar el rango de recuperación y grado de concentrados de plomo-plata y
zinc-plata, en ciclo abierto, obtenidos por flotación diferencial.
-
Determinar el rango de recuperación y grado de concentrados de plomo-plata y
zinc-plata, en ciclo cerrado, obtenidos por flotación diferencial.
-
Efectuar análisis granulométricos de las colas de las pruebas de flotación en ciclo
abierto.
-
Efectuar el análisis size by size de las pruebas de flotación diferencial en ciclo
abierto.
-
Determinar el contenido de lamas (arcillas) en las colas de todas las pruebas de
flotación, a través de pruebas de ciclonaje.
-
Realizar pruebas de sedimentación, a partir de las colas de flotación
-
Determinar el Indice de Trabajo (Work Index).
-
Pruebas de flotación diferencial con deslame
Análisis granulométrico de la alimentación a flotación y de colas de flotación
Determinación de lamas de las colas de flotación por ciclonaje
Análisis size by size
Pruebas de sedimentación
Pruebas de determinación del Work Index.
MUESTRA: LA
TRITURACION
PRIMARIA
CERNIDO, ¼”
Como objetivos secundarios deben establecerse las condiciones de operación y consumo
de reactivos en las pruebas de flotación diferencial.
3. EXPERIMENTACIÓN METALÚRGICA
La experimentación metalúrgica para la presente investigación, se llevó a cabo de
acuerdo a lo que se muestran en los flujogramas de las figuras 1, 2, 3, 4 y 5 y el detalle
descriptivo que se anota a continuación:
-
-¼”
+¼”
HOMOGENEIZACION
Y CUARTEO
TRITURACION
SECUNDARIA
Análisis
químico
Análisis
granulométrico
DETERMINACION
DEL WORK INDEX
PRUEBAS DE
FLOTACION
DIFERENCIAL
Figura 1.- Flujograma de la etapa de preparación de las muestras para la
experimentación, Empresa Apogee
Trituración primaria y secundaria de la muestra
Homogeneización, cuarteo y obtención de comunes representativos
para las diferentes pruebas.
Preparación de la muestra para la realización de las pruebas de flotación
diferencial.
Pruebas de flotación diferencial sin deslame
2
3
MUESTRA
CLASIFICACION,
100 Mallas Ty
CICLONAJE
(-)
(+)
Under flow
Over flow
(lamas-arcillas)
MOLIENDA
Regulador de pH
Depresor
ACONDICIONAMIENTO-1
Espumante
Colector
FLOTACION
ROUGHER de Pb-Ag
Espuma Pb-Ag
1ra FLOTACION
CLEANER
Non Float
Regulador de pH
Activador
NF-1ra Limpieza
Espuma Pb
2da FLOTACION
CLEANER
Espuma
Pb-Ag
ACONDICIONAMIENTO 2
FLOTACION
ROUGHER de Zn
NF-2da Limpieza
Non Float
Espuma Zn
1ra FLOTACION
CLEANER de Zn
Espuma Zn
CUARTEO
NF-1ra Limpieza
Análisis
granulométrico
2da FLOTACION
CLEANER de Zn
Espuma de Zn
Colector
Espumante
CICLONAJE
Prueba de
sedimentación
NF-2da Limpieza
Figura 3.- Flujograma de las pruebas experimentales, que se siguieron
con las muestras complejas de la Empresa Apogee, en
CICLO ABIERTO y PREVIO DESLAMADO
4
5
6
7
2000 g (dos veces)
4. RESULTADOS Y COMENTARIOS
CLASIFICACION, 100#
(+)
Tomando en cuenta los objetivos del presente trabajo y el interés de la empresa Apogee
de obtener la mayor información posible respecto a los resultados de los diferentes
trabajos experimentales con la muestra Alta Ley, se presentarán los resultados de
acuerdo a un desarrollo práctico, de tal manera que se efectúe un seguimiento objetivo
del trabajo experimental.
MOLIENDA
(-)
D-2000
pH: 6.8
Cal: 5000 g/t
pH: 9.5
ZnSO4: 250 g/t
FLOTACION NaCN: 75 g/t
ROUGHER T. Acond.: 8 min
DE Pb-Ag AF-242: 35 g/t
MIBC: 25 g/t
T. Acond.: 6 min
T. Flot.: 6 min
La muestra recibida fue cuidadosamente preparada con el propósito de obtener comunes
representativos para la realización de las diferentes pruebas. Estos pasos se pueden
seguir observando el flujograma que se muestra en la figura 1.
agua
4.1.1 ANALISIS QUIMICO DEL COMÚN
Espuma Pb-Ag
D-500
pH: 7.10
Cal: 1000 g/t
pH: 9.8
PRIMERA
Na2SiF6: 200 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 250 g/t
DE Pb-Ag
NaCN: 75 g/t
T. acond.: 7 min
T. Flot.: 4 min
La ley de cabeza ensayada del común representativo, efectuado en triplicado y calculado
el promedio, da el siguiente resultado:
Plata: 268 g/t
Plomo: 1.58%
Zinc: 2.71%
Cobre: 0.067%
Hierro: 5.86%
Prueba 1:
Esta prueba fue llevada a cabo siguiendo los pasos que se muestran en el flujograma de
la figura 2. Se realiza la flotación diferencial, flotando primero el mineral de Pb-Ag, para
ello se efectuó la molienda a -100 Mallas Tyler y en la etapa de acondicionamiento se usó
cal, como regulador de pH; sulfato de zinc como depresor del mineral de zinc y cianuro de
sodio para depresar las piritas que se encuentran en apreciable cantidad en la muestra;
como colector se usó el ditiofosfato Aero Float-242 y como espumante se usó el Metil
Isobutil Carbinol, más conocido como MIBC. El grado de molienda, el colector y el
espumante, fueron elegidos previas pruebas de flotación exploratorias.
D-250
pH: 9.4
Na2SiF6: 100 g/t
T. acond.: 4 min
ZnSO4: 150 g/t
NaCN: 75 g/t
T. acond.: 7 min
T. Flot.: 3 min
SEGUNDA
FLOTACION
CLEANER
DE Pb-Ag
Espuma de Pb-Ag
Non Float
Espuma Zn-Ag
NF-1ra Limp. Ag
Espuma de Pb-Ag
El peso específico real, determinado por el método del picnómetro es de 2.878 g/cm3.
4.1.2 FLOTACION DIFERENCIAL DE SULFUROS EN CIRCUITO ABIERTO
Non Float
pH: 8.8
Cal: 5000 g/t
pH: 11.4
FLOTACION CuSO4: 187.5 g/t
ROUGHER Z-11: 50 g/t
DE Zn-Ag T. Acond.: 5 min
MIBC: 25 g/t
T. Flot.: 12 min
D-1000
pH: 10.9
PRIMERA Cal: 1500 g/t
FLOTACION pH: 11.5
CLEANER
Na2SiF6: 200 g/t
DE Zn-Ag
T. Acond.: 5 min
T. Flot.: 7 min
Espuma de Zn-Ag
NF-1ra Limpieza Zn-Ag
D-500
pH: 10.1
NF-2da Limp. Ag
Cal: 1500 g/t
SEGUNDA pH: 11.2
FLOTACION Na2SiF6: 100 g/t
CLEANER T. Acond.: 5 min
DE Zn-Ag T. Flot.: 5 min
Espuma de Zn-Ag
NF-2da Limpieza Zn-Ag
Figura 6.- Condiciones de operación y consumo de reactivos de la prueba 1
de flotación diferencial, Muestra LA, Apogee.
8
9
Los resultados de esta primera prueba se muestran en la tabla 1.
Prueba 2:
Tabla 1.- Balance metalúrgico de la prueba de flotación diferencial, prueba 1, muestra LA
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
% Dist. g/t Ag % Dist. % Zn % Dist.
Espuma de Pb-Ag
3.26
34.61
71.58
3940
44.69
3.73
4.42
NF–2da Limp.de Pb
0.95
10.6
6.43
1475
4.91
3.67
1.28
NF-1ra Limp. de Pb
3.80
2.41
5.81
382
5.05
2.35
3.25
Espuma rougher- Pb
8.01
16.48
83.82
1959
54.64
3.07
8.95
Espuma de Zn-Ag
3.57
1.46
3.31
2680
33.37 55.10
71.73
NF-2da Limp. de Zn
1.08
2.59
1.77
1495
5.61
23.8
9.34
NF-1ra Limp. de Zn
7.10
0.54
2.44
66
1.63
0.7
1.81
Espuma rougher Zn
11.75
1.01
7.52
992
40.61 19.36
82.87
Non Float
80.24
0.17
8.66
17
4.75
0.28
8.18
Cabeza Calculada
100.00
1.57
100.00
287 100.00
2.75 100.00
Esta prueba también fue llevada a cabo siguiendo los pasos que se muestran en el
flujograma de la figura 7.
Las condiciones de operación y consumo de reactivos se muestran en la figura 6; se
debieron efectuar dos flotaciones, en las mismas condiciones, para tener suficiente
espuma rougher y llevar a la limpieza, especialmente las espumas de Pb-Ag.
Los índices metalúrgicos que se logran, en estas condiciones de operación, son:
-
-
Recuperación del plomo: 71.58%
Recuperación del zinc: 71.73%
10
En esta prueba se incrementaron un tanto los reactivos y los tiempo de acondicionamiento
y flotación. Los resultados de esta segunda prueba se muestran en la tabla 2 y las
condiciones de operación y consumo de reactivos en la figura 7.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
2.74
36.99
64.88
3490
32.91
2.77
2.72
NF–2da Limp.de Pb
1.28
11.80
9.70
1460
6.45
2.94
1.35
NF-1ra Limp. de Pb
5.11
2.96
9.68
569
10.01
2.67
4.90
Espuma rougher- Pb
9.14
14.41
84.26
1571
49.36
2.74
8.98
Espuma de Zn-Ag
3.61
1.69
3.90
3270
40.56 57.20
74.01
NF-2da Limp. de Zn
0.86
2.78
1.54
1255
3.73 22.00
6.82
NF-1ra Limp. de Zn
6.84
0.61
2.67
107
2.52
1.36
3.34
Espuma rougher Zn
11.31
1.12
8.11
1203
46.81 20.74
84.17
Non Float
79.55
0.15
7.63
14
3.83
0.24
6.85
Cabeza Calculada
100.00
1.56
100.00
291 100.00
2.79 100.00
Se observa una mejora en los resultados, aunque la ley del concentrado de plomo es
todavía baja; la ley del concentrado de zinc es buena y con posibilidades de mejorar la
recuperación del mismo. Por otro lado, las colas tienen una distribución baja de los
elementos valiosos y por tanto pueden considerarse descartables.
Para dispersar y deprimir parte de los óxidos, se introdujo el fluosilicato de sodio, solo en
la etapa de las limpiezas, con resultados positivos. En esta prueba, como en el resto de la
pruebas, se debieron efectuar 2 flotaciones en las mismas condiciones con la finalidad de
acumular espumas rougher y afrontar adecuadamente las etapas de limpieza.
-
11
-
2000 g (dos veces)
CLASIFICACION, 100#
(+)
MOLIENDA
(-)
D-2000
pH: 6.9
Cal: 6000 g/t
pH: 9.7
ZnSO4: 300 g/t
MIBC: 20 g/t
T. Acond.: 6 min
T. Flot.: 5 min
Prueba 3, con previo deslame:
Esta prueba fue conducida según los pasos que se muestran en la figura 3; esto es, se
efectuó un deslame previo a la flotación. Los resultados de esta tercera prueba se los
agua
Espuma Pb-Ag
Non Float
pH: 8.6
Cal: 5000 g/t
pH: 11.2
FLOTACION CuSO4: 187.5 g/t
MIBC: 20 g/t
T. Flot.: 13 min
D-500
pH: 8.6
Cal: 250 g/t
pH: 9.5
PRIMERA
Na2SiF6: 200 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 250 g/t
DE Pb-Ag
NaCN: 100 g/t
T. acond.: 7 min
T. Flot.: 4 min
Espuma de Pb-Ag
muestra en la tabla 3 y las condiciones de operación y consumo de reactivos en la figura
8.
Non Float
Espuma Zn-Ag
NF-1ra Limp. Ag
D-1000
pH: 10.9
Cal: 1000 g/t
pH: 11.3
Na2SiF6: 200 g/t
T. Acond.: 6 min
PRIMERA
FLOTACION
CLEANER
DE Zn-Ag
D-250
pH: 7.8
Cal: 250 g/t
pH: 9.4
Na2SiF6: 100 g/t
T. acond.: 5 min
ZnSO4: 200 g/t
NaCN: 100 g/t
T. acond.: 6 min
T. Flot.: 3 min
SEGUNDA
FLOTACION
CLEANER
DE Pb-Ag
T. Flot.: 5 min
D-500
Mejoran los resultados en cuanto a la calidad de los concentrados y recuperaciones de los
elementos valiosos, especialmente el concentrado de plomo y al mejorar este
concentrado se ve que la mayor parte de la plata se concentra en este producto. Las
colas finales tienen distribuciones un tanto mas elevadas que la anterior prueba, esto se
debe a que las lamas, producto del ciclonaje, pasan a formar parte de las colas finales, a
pesar de esto, los índices metalúrgicos pueden considerarse como buenos.
NF-2da Limp. Ag
Espuma de Zn-Ag
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
2.35
51.76
79.47
8290
65.96
6.62
6.14
NF–2da Limp.de Pb
0.38
22.70
5.61
3810
4.88
6.3
0.94
NF-1ra Limp. de Pb
1.74
1.84
2.09
308
1.82
1.98
1.36
Espuma rougher- Pb
4.47
29.85
87.17
4801
72.65
4.78
8.44
Espuma de Zn-Ag
3.67
0.98
2.35
1590
19.73 57.40
83.06
NF-2da Limp. de Zn
0.76
2.95
1.46
626
1.60
12.3
3.67
NF-1ra Limp. de Zn
1.74
0.53
0.60
75
0.44
0.44
0.30
Espuma rougher Zn
6.17
1.09
4.41
1044
21.77 35.77
87.03
Non Float
71.67
0.18
8.42
23
5.58
0.16
4.52
Over flow (lamas)
17.69
0.73
8.43
130
7.78
1.40
9.77
Cola Total
89.36
0.29
16.86
44
13.36
0.41
14.30
Cabeza Calculada
100.00
1.53
100.00
296 100.00
2.53 100.00
Espuma de Zn-Ag
pH: 9.8
Cal: 1000 g/t
pH: 11.1
Na2SiF6: 100 g/t
SEGUNDA FLOTACION T. Acond.: 5 min
CLEANER DE Zn-Ag T. Flot.: 3 min
Espuma de Pb-Ag
Recuperación total de la plata 73.47%
Figura 7.- Condiciones de operación y consumo de reactivos de la
prueba 2 de flotación diferencial, Muestra LA, Apogee.
12
13
2000 g
(tres veces)
CICLONAJE
CLASIFICACION, 100#
(+)
(-)
Over flow
(Lama)
Under flow
D-2000
pH: 7.40
Cal: 3000 g/t
pH: 9.6
ZnSO4: 375 g/t
MIBC: 25 g/t
T. Acond.: 6 min
T. Flot.: 4 min
MOLIENDA
Figura 8.- Condiciones de
operación y consumo de
reactivos de la prueba 3 de
flotación diferencial, Muestra
LA, Apogee.
agua
Espuma Pb-Ag
Non Float
pH: 9.3
Cal: 5000 g/t
pH: 11.2
FLOTACION CuSO4: 212 g/t
MIBC: 25 g/t
T. Flot.: 7 min
D-500
pH: 8.60
Cal: 250 g/t
pH: 9.8
PRIMERA
Na2SiF6: 150 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 200 g/t
DE Pb-Ag
NaCN: 75 g/t
T. acond.: 6 min
T. Flot.: 2.5 min
NF-1ra Limp. Ag
Espuma de Pb-Ag
D-1000
pH: 10.0
Cal: 3000 g/t
pH: 11.4
Na2SiF6: 100 g/t
PRIMERA
T. Acond.: 5 min
FLOTACION
CuSO4: 50 g/t
CLEANER
DE Zn-Ag Z-11: 10 g/t
MIBC: 15 g/t
T. Acond.: 5 min
T. Flot.: 6 min
D-250
pH: 8.4
Cal: 200 g!t
SEGUNDA pH: 9.5
DE Pb-Ag ZnSO4: 100 g/t
NaCN: 50 g/t
T. acond.: 4 min
T. Flot.: 1.5 min
Espuma de Zn-Ag
Espuma de Pb-Ag
NF-2da Limp. Ag
D-500
pH: 10.2
SEGUNDA Cal: 1000 g/t
FLOTACION pH: 11.5
CLEANER Na2SiF6: 50 g/t
T. Flot.: 4 min
Espuma de Zn-Ag
-
Prueba 4,
Non Float
Espuma Zn-Ag
-
Esta prueba se llevó delante de la misma forma que la prueba 2, es decir, siguiendo
específicamente el flujograma de la figura 2. Los resultados de esta segunda prueba se
muestran en la tabla 4.
Las condiciones de operación y consumo de reactivos se detallan en la figura 9.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
2.21
49.54
61.79
5850
40.80
3.18
2.33
NF–2da Limp.de Pb
1.12
22.30
14.06
3400
11.99
3.30
1.22
NF-1ra Limp. de Pb
2.78
2.9
4.55
532
4.67
2.48
2.29
Espuma rougher- Pb
6.10
23.32
80.40
2981
57.46
2.88
5.85
Espuma de Zn-Ag
3.72
1.72
3.62
2430
28.56 58.10
71.88
NF-2da Limp. de Zn
1.27
2.66
1.90
1150
4.60 30.60
12.87
NF-1ra Limp. de Zn
5.88
0.85
2.82
138
2.56
1.56
3.05
Espuma rougher Zn
10.87
1.36
8.34
1041
35.72 24.30
87.80
Non Float
83.03
0.24
11.26
26
6.82
0.23
6.35
Cabeza Calculada
100.00
1.77
100.00
317 100.00
3.01 100.00
14
15
Los índices metalúrgicos declinan en sus valores con relación a al aprueba 3 aunque son
un tanto mejores con relación a la prueba 2 y 1 que también se realizaron sin previo
deslame.
-
-
Esta prueba se realizó con previo deslame, según los pasos que se muestran en la figura
3. Los resultados de esta prueba se los muestra en la tabla 5 y las condiciones de
operación y consumo de reactivos en la figura 10.
Tabla 5.- Balance metalúrgico de la prueba 5 de flotación diferencial, muestra LA
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
1.97
52.12
64.75
7810
43.53
5.25
3.44
NF–2da Limp.de Pb
0.67
46.56
19.76
5020
9.56
4.59
1.03
NF-1ra Limp. de Pb
1.69
3.62
3.86
561
2.68
2.54
1.43
Espuma rougher- Pb
4.33
32.33
88.38
4548
55.77
4.09
5.89
Espuma de Zn-Ag
3.69
1.06
2.47
2890
30.20 60.00
73.63
NF-2da Limp. de Zn
0.51
1.80
0.58
764
1.10
28.4
4.80
NF-1ra Limp. de Zn
2.23
1.04
1.46
206
1.30
3.15
2.34
Espuma rougher Zn
6.42
1.11
4.51
1791
32.60 37.78
80.76
Non Float
70.34
0.16
7.11
18
3.59
0.14
3.28
Over flow (lamas)
18.91
0.83
9.92
150
8.04
1.60
10.07
Cola Total
89.25
0.13
7.11
46
11.63
0.45
13.35
Cabeza Calculada
100.00
1.58
100.00
353 100.00
3.00 100.00
16
17
Los resultados no un tanto similares a la prueba 3, que en función de estos se puede
afirmar que son mejores que los que se obtienen sin deslame, las recuperaciones también
son un tanto mas elevadas que las que se obtienen sin previo deslame.
-
Radio de enriquecimiento del zinc: 20
-
4.1.3. FLOTACION EN CIRCUITO CERRADO
Las pruebas de flotación en ciclo cerrado se desarrollaron sin y con deslame, siguiendo
los pasos que se muestran en los flujogramas de las figuras 4 y 5.
Prueba 1:
Esta prueba fue desarrollada siguiendo el flujograma de la figura 4; los resultados que se
alcanzaron se detallan en el balance metalúrgico de la tabla 6.
Tabla 6.- Balance metalúrgico de la prueba de flotación diferencial en
ciclo cerrado, prueba 1, muestra LA
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
2.39
52.40
78.75
6620
55.99
7.94
7.18
Espuma de Zn-Ag
3.76
1.25
2.95
2010
26.74
57.00
81.10
Non Float
93.85
0.31
18.29
52
17.27
0.33
11.72
Cabeza Calculada
100.00
1.59
100.00
283
100.00
2.64
100.00
18
19
Los resultados son buenos ya que los concentrados alcanzan leyes por encima de 50% y
las recuperaciones son bastante aceptables.
2000 g (tres veces)
(dos veces)
CLASIFICACION, 100#
(+)
(-)
D-2000
pH: 6.60
Cal: 7000 g/t
pH: 9.6
ZnSO4: 500 g/t
MIBC: 20 g/t
T. Acond.: 6 min
T. Flot.: 5 min
MOLIENDA
flotación diferencial en ciclo
cerrado, Muestra LA, Apogee.
agua
Espuma Pb-Ag
Non Float
pH: 9.0
Cal: 6000 g/t
pH: 11.1
MIBC: 20 g/t
T. Flot.: 7 min
D-500
pH: 8.6
Cal: 500 g/t
pH: 9.5
PRIMERA
Na2SiF6: 200 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 500 g/t
DE Pb-Ag
NaCN: 100 g/t
T. acond.: 6 min
T. Flot.: 3 min
Non Float
Espuma Zn-Ag
-
-
NF-1ra Limp. Ag
Espuma de Pb-Ag
D-1000
pH: 8.1
Cal: 2500 g/t
pH: 11.5
Na2SiF6: 200 g/t
PRIMERA
T. Acond.: 5 min
FLOTACION
CuSO4: 30 g/t
CLEANER
MIBC: 10 g/t
T. Acond.: 5 min
T. Flot.: 4 min
D-250
pH: 8.6
Cal: 250 g/t
pH: 9.6
SEGUNDA
NaCN: 50 g/t
Prueba 2:
T. acond.: 6 min
T. Flot.: 2 min
Espuma de Zn-Ag
NF-2da Limp. Ag
Las condiciones de operación y consumo de reactivos, se detallan en el flujograma de la
figura 11. Los índices metalúrgicos que se logran, en estas condiciones de operación,
son:
Espuma de Pb-Ag
D-500
Na2SiF6: 100 g/t
SEGUNDA pH: 7.5
FLOTACION Cal: 3000 g/t
CLEANER pH: 11.6
T. Flot.: 3 min
Espuma de Zn-Ag
Esta prueba se llevó adelante con previo deslame, siguiendo los pasos que se muestran
en el flujograma de la figura 5; los resultados se resumen en el balance metalúrgico de la
tabla 7
ciclo cerrado, prueba 2, muestra LA
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
2.14
53.40
76.98
9670
69.89
5.52
4.27
Espuma de Zn-Ag
3.69
1.00
2.49
1080
13.46
59.00
78.68
Non Float
75.43
0.23
11.70
32
8.16
0.27
7.37
Lamas
18.74
0.70
8.84
134
8.49
1.43
9.69
Total colas
94.17
0.32
20.54
52
16.65
0.50
17.06
Cabeza Calculada
100.00
1.48
100.00
296
100.00
2.77
100.00
De esta forma también se obtienen buenos resultados aunque los índices metalúrgicos
son levemente inferiores a la anterior prueba.
21
20
figura 12.
-
-
4.1.4 ANALISIS GRANULOMETRICO DE LAS COLAS DE FLOTACION
Estos análisis granulométricos se refieren específicamente a las colas (non floats) de las
pruebas de flotación a ciclo abierto.
a) Cola de flotación a ciclo abierto, de la prueba 1:
El resultado es el siguiente:
Tabla 9.- Balance metalúrgico del análisis granulométrico, colas de flotación prueba 1
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
22
Peso
%
9.55
12.96
13.27
8.04
7.54
48.64
100.00
PLOMO
% Pb
% Dist.
0.163
8.77
0.135
9.86
0.119
8.89
0.113
5.12
0.147
6.24
0.223
61.11
100.00
0.18
23
PLATA
g/t Ag % Dist.
28
15.28
15
11.11
13
9.85
13
5.97
18
7.75
18
50.03
17.5 100.00
ZINC
% Zn % Dist.
0.977
36.67
0.473
24.11
0.151
7.87
0.094
2.97
0.080
2.37
0.136
26.01
0.25 100.00
En principio, es importante recordar que las leyes de los elementos valiosos en las colas,
de la prueba 1, son: plomo, 0.17%; plata, 17 g/t y zinc, 0.28%; estas leyes deben
compararse con las obtenidas por cálculo en la tabla 9, casi coinciden. En segundo lugar,
se puede establecer que la mayor pérdida de los elementos valiosos que se produce es
en la granulometría fina, por debajo de 400 Mallas Tyler, 38 micrones.
b) Cola de flotación a ciclo abierto, de la prueba 2:
En la prueba de flotación 3, existen diferencias que hasta el momento no existían, es
decir, que las leyes de las colas de esta prueba de flotación son un tanto más altas que
las obtenidas por cálculo después del análisis granulométrico; por ejemplo, las leyes de
los elementos valiosos en las colas de flotación son: 0.15% Pb, 23 g/t Ag y 0.16% Zn y del
análisis granulométrico se obtienen, 0.13% Pb; 14 g/t Ag y 0.09% Zn. En esta tabla,
también se puede establecer que la mayor pérdida de los elementos valiosos que se
produce es en la granulometría fina, por debajo de 400 Mallas Tyler, esta situación se
repite en todas las pruebas.
d) Cola de flotación a ciclo abierto, de la prueba 4:
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
9.06
11.58
16.01
3.52
6.95
52.87
100.00
PLOMO
% Pb
% Dist.
0.103
6.51
0.100
8.07
0.091
10.15
0.093
2.28
0.115
5.57
0.183
67.42
100.00
0.14
PLATA
g/t Ag % Dist.
15
10.18
10
8.67
11
13.19
12
3.17
18
9.37
14
55.43
13.4 100.00
ZINC
% Zn % Dist.
0.831
32.64
0.394
19.78
0.148
10.27
0.101
1.54
0.092
2.77
0.144
33.00
0.23 100.00
En la prueba de flotación 2, las leyes de los elementos valiosos en las colas, son: plomo,
0.15%; plata, 14 g/t y zinc, 0.24%; estas leyes deben compararse con las obtenidas por
cálculo en la tabla 10, casi coinciden. En esta tabla, también se puede establecer que la
mayor pérdida de los elementos valiosos que se produce es en la granulometría fina, por
debajo de 400 Mallas Tyler.
c) Cola de flotación a ciclo abierto, de la prueba 3:
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
7.19
8.40
12.32
7.38
8.78
55.93
100.00
PLOMO
% Pb
% Dist.
0.157
5.80
0.135
5.83
0.126
7.98
0.099
3.75
0.145
6.54
0.244
70.11
100.00
0.19
PLATA
g/t Ag % Dist.
16
7.62
13
7.23
14
11.43
7
3.42
19
11.04
16
59.26
15.1 100.00
ZINC
% Zn % Dist.
0.340
15.35
0.167
8.81
0.121
9.37
0.100
4.63
0.102
5.62
0.160
56.21
0.16 100.00
Las leyes de los elementos valiosos en las colas de la prueba 4, son: plomo, 0.19%; plata,
15 g/t y zinc, 0.16%; estas leyes deben compararse con las obtenidas por cálculo en la
tabla 12, que en esta prueba también difieren un tanto, asi tenemos, 0.24% Pb; 26 g/t Ag
y 0.23% Zn. En esta tabla, también se puede establecer que la mayor pérdida de los
elementos valiosos que se produce es en la granulometría fina, por debajo de 400 Mallas
Tyler, esta situación se repite en todas las pruebas.
Peso
%
10.16
13.08
17.71
4.43
6.44
48.19
100.00
PLOMO
% Pb
% Dist.
0.083
6.35
0.075
7.38
0.078
10.39
0.085
2.83
0.093
4.51
0.189
68.54
100.00
0.13
PLATA
g/t Ag % Dist.
10
7.31
9
8.47
10
12.74
11
3.50
12
5.56
18
62.41
13.9 100.00
ZINC
% Zn % Dist.
0.168
19.57
0.085
12.74
0.065
13.19
0.051
2.59
0.052
3.84
0.087
48.06
0.09 100.00
e) Cola de flotación a ciclo abierto, de la prueba 5:
25
24
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
9.22
12.87
11.75
10.54
10.23
45.39
100.00
PLOMO
% Pb
% Dist.
0.083
5.50
0.086
7.96
0.094
7.94
0.086
6.52
0.128
9.42
0.192
62.66
100.00
0.14
PLATA
g/t Ag % Dist.
8
6.01
10
10.49
12
11.49
8
6.87
16
13.34
14
51.79
12.3 100.00
ZINC
% Zn % Dist.
0.129
10.67
0.090
10.39
0.064
6.75
0.058
5.48
0.052
4.78
0.152
61.92
0.11 100.00
De estos análisis granulométricos, tablas 9, 10, 11, 12 y 13, se puede establecer que las
mayores pérdidas de los elementos valiosos, plomo, plata y zinc, se producen en los
tamaños de grano que se encuentran por debajo de la malla 400.
4.1.5 ANALISIS SIZE BY SIZE
Figura 14.- Análisis granulométrico de la alimentación a flotación rougher
Para realizar este análisis es necesario contar con el análisis granulométrico de la
alimentación a la flotación rougher, es decir, la muestra molida a -100 Mallas Tyler.
i) Análisis granulométrico de la alimentación
Tabla 14.- Balance metalúrgico del análisis granulométrico, ALIMENTACION,
a las pruebas de flotación a ciclo abierto
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
+150#
8.61
1.165
6.32
286
8.06
2.63
8.24
-150# +200#
11.78
1.4
10.38
318
12.26
2.91
12.47
-200# +270#
7.92
2.13
10.62
433
11.23
3.72
10.71
-270# +325#
8.12
1.295
6.62
302
8.03
2.94
8.68
-325# +400#
7.43
2.790
13.04
490
11.91
3.82
10.31
-400#
56.14
1.5
53.01
264
48.51
2.43
49.60
Cabeza calculada
100.00
100.00
1.59
305.5 100.00
2.75 100.00
Entonces, el d80 es igual a 76 micrones y el d50 es igual a 33 micrones
Este análisis granulométrico, también permitirá calcular las leyes de cada tamaño de
grano de las alimentaciones con las que se efectuó cada prueba de flotación y con la cola
de cada prueba se efectúa el análisis size by size.
ii) Alimentación vs cola de la prueba 1, flotación a ciclo abierto
Este análisis granulométrico, a través del % Peso, permite además calcular el d80 del
producto molido; para ello se tiene el gráfico de la figura 14.
Figura 15.- Análisis size by size, alimentación vs cola de la prueba 1, en ciclo abierto
26
27
Este análisis es más práctico y objetivo a través de un gráfico como el que se muestra en
la figura 15.
La figura 15, muestra que las mayores pérdidas de los elementos valiosos se dan con la
plata y el zinc aunque estos valores, de una manera general, son bajos; asi mismo se
puede evidenciar que estas pérdidas se dan en los granos finos.
v) Alimentación vs cola de la prueba 4, flotación a ciclo abierto
De la misma forma, este análisis es más práctico y objetivo a través de un gráfico como el
que se muestra en la figura 17.
iii) Alimentación vs cola de la prueba 2, flotación a ciclo abierto
De la misma forma, este análisis es más práctico y objetivo a través de un gráfico como el
que se muestra en la figura 16.
Se observa que las pérdidas en las colas, de todos los elementos valiosos, disminuyen
considerablemente y por tanto es una prueba que vale la pena tomarse en cuenta.
En esta prueba, la recuperación por fracciones, de una manera general, mejora
sustancialmente; dentro de esta mejora se puede apreciar que el elemento zinc es el que
ocasiona mayores dificultades, principalmente en las fracciones finas que está por debajo
de 45 micrones.
iv) Alimentación vs cola de la prueba 3, flotación a ciclo abierto
vi) Alimentación vs cola de la prueba 5, flotación a ciclo abierto
En este caso no se puede realizar el análisis porque, previa a la flotación diferencial, se
efectuó el deslamado por ciclonaje, impidiendo llevar adelante el análisis correspondiente
por la eliminación de fracciones granulométricas que no ingresan a la prueba de flotación.
4.1.6 DETERMINACION DEL CONTENIDO DE LAMAS EN COLAS DE FLOTACION
Estas pruebas se llevaron a cabo a través de un ciclonaje en dos etapas. Los resultados
se muestran a continuación
efectuó el deslamado por ciclonaje, impidiendo llevar adelante el análisis correspondiente
por la eliminación de fracciones granulométricas que no ingresan a la prueba de flotación.
29
28
a) Pruebas con colas obtenidas en ciclo abierto
Tabla 15.- Resumen de las pruebas de determinación de lamas-arcillas, ciclo abierto
Cola P-1
Cola P-2
Cola P-3*
Cola P-4
Cola P-5*
Denominación
% Peso
% Peso
% Peso
% Peso
% Peso
Over flow
(arcillas, lamas)
19.24
19.25
9.75
18.84
8.85
Under flow
80.76
80.75
90.25
81.16
91.15
Alimentación
100.00
100.00
100.00
100.00
100.00
*La cola proviene de una prueba de flotación previo ciclonaje de la alimentación
De la tabla 15 se puede establecer que la presencia de lamas en las colas de flotación,
cuando no se procede al deslamado previa flotación, alcanzan a un valor que está
próximo a 19.11% en peso y con previo deslamado a 9.30% en peso.
b) Pruebas con colas obtenidas en ciclo cerrado
Tabla 16.- Resumen de las pruebas de determinación de lamas-arcillas, ciclo cerrado
Cola P-1
Cola P-2*
Denominación
% Peso
% Peso
Over flow (arcillas, lamas)
19.60
9.94
Under flow
80.40
90.06
Alimentación
100.00
100.00
La tabla 16, muestra valores un tanto similares; es decir, sin deslame las colas tienen un
19.60% en peso y con previo deslame, un 9.94% en peso.
4.1.7 PRUEBAS DE SEDIMENTACION A PARTIR DE LAS COLAS DE FLOTACION
Antes de mostrar los resaltados de las pruebas de sedimentación es necesario recordar
algunos conceptos básicos en forma preliminar.
La sedimentación es la técnica más usada en las operaciones de desaguado en el
procesamiento de minerales. Relativamente barata, de alta capacidad de procesamiento,
la cual incluye baja influencia de fuerzas de atrición, proporcionando buenas condiciones
para la floculación de partículas finas.
a) Técnica de espesamiento donde lo que interesa es un producto final de alta
densidad.
b) Técnica de clarificación, donde lo que interesa es obtener un rebalse de líquido
claro espesando los sólidos en un underflow sin importar su concentración.
En el espesamiento la obtención de un underflow, o producto espesado, interesa que su
concentración en sólidos sea lo más alta posible sin importar mucho la turbidez del liquido
que rebalse.
En la clarificación es de importancia que el rebalse del líquido sea lo más claro posible.
Los espesadores continuos consisten de un tanque cilíndrico cuyo diámetro fluctúa entre
2 y 200 metros, con una profundidad del orden de 1 a 7 metros. La pulpa es alimentada
en el centro del estanque vía Feed-well, a una profundidad de aproximadamente 1 metro,
con el fin de no producir turbulencia. El liquido clarificado rebalsa a una canaleta
periférica, mientras el sólido espesado es retirado por una salida central en el fondo del
estanque, ayudado por un sistema de bombeo y un sistema de rastra que gira lentamente
arrastrando el sólido al centro.
El tamaño de un espesador se caracteriza por su diámetro (área) y altura. La capacidad
de une espesador está determinada por su área; en cambio, la habilidad para producir
una pulpa de determinado contenido de sólidos esta dado por la altura. El área debe ser
suficiente para permitir que la partícula con la velocidad de sedimentación más lenta
alcance el fondo del tanque, ya que la velocidad de sedimentación de una partícula es
diferente en cada zona.
En la práctica hay varias técnicas para calcular el área del espesador. Así se tienen: la
técnica de Coe-Clevenger, de Talmadge y Fitch y la técnica de Oltmann.
En esta oportunidad de acudió al método descrito por Tamaldge y Fitch. Se efectuaron
varias pruebas, con cola de flotación sin previo deslame y con cola de flotación previo
deslame antes de la operación de flotación; por otro lado, se efectuaron pruebas sin la
adición de floculante y con la adición de dos concentraciones de floculante, 20 g/t y 30 g/t.
A continuación se presentan los resultados de los mismos.
La sedimentación gravitacional puede separarse en dos grupos o técnicas de
funcionamiento:
30
31
4.1.7.1. CON COLA (NON FLOAT) SIN PREVIO DESLAME
Para esta experimentación, se tomo la cola de la prueba 1 de flotación en ciclo abierto.
Los resultados alcanzados se muestran en las tablas 17 y 18, considerando un flujo de
alimentación de 25% sólidos.
Debe tomarse en cuenta que el método empleado sugiere que al diámetro calculado del
espesador debe incrementarse el 25%, es decir, si el diámetro del espesador es de 31 m,
para una alimentación de 25% sólidos, aproximadamente, y una descarga del 50%
Sólidos, en realidad el diámetro deberá ser 39 metros.
Si bien, el diámetro final es grande es el tamaño de estanque que garantizará la
sedimentación de todas las partículas en suspensión, se debe recordar que son partículas
muy finas y que al rededor de 50% en peso están por debajo de 38 micrones.
También es importante y necesario recordar que los espesadores convencionales tienen
la desventaja de poseer una gran área superficial y por ello es necesario una gran
superficie para su construcción; empero, es posible diseñar, construir y usar espesadores
de alta capacidad. Se caracterizan por una reducción del área de espesamiento desde un
área convencional instalada; el asentamiento de los sólidos en el manto de la pulpa se
desliza hacia abajo a lo largo de las placas inclinadas produciendo un “más rápido y más
efectivo espesamiento” que un descenso vertical.
Tabla 17.- EVALUACION RESULTADOS SEDIMENTACION
Prueba Nº
Apogee 822
Volumen pulpa, cm3
1000
Muestra
Colas de flotación
Peso de la pulpa, g
1191.489
Tipo de floculante
Magnafloc 292
Tamaño de partícula
-100 Mallas Ty
Densidad muestra, g/cm3
Dosificación, g/t
0, 10 y 20
2.801
Fracción vol. inicial
0,106
Fracción vol. final
0,263
Interfase sólido-líquido
Concentración del floculante, g/t
0
10
20
Nº
Tiempo, seg.
Altura H, m
Altura H, m
Altura H, m
1
2
3
4
5
6
7
8
9
10
11
12
13
0
480
600
960
1320
1500
3120
3420
4500
5400
6300
7200
8100
0.360
0.350
0.348
0.343
0.340
0.338
0.322
0.321
0.312
0.305
0.298
0.291
0.284
0.360
0.351
0.348
0.338
0.328
0.323
0.278
0.267
0.231
0.222
0.217
0.213
0.209
0.360
0.348
0.342
0.327
0.311
0.298
0.225
0.218
0.202
0.200
0.195
0.190
0.187
14
15
16
17
18
19
20
21
22
23
24
25
26
27
28
10800
12600
13500
14400
15300
16200
17100
21600
23400
25200
27000
28800
30600
36000
86400
90
150
960
1320
1500
2520
3120
3420
4500
5400
7200
8100
9000
12600
13500
14400
16200
17100
19800
20700
23400
25200
28000
75600
86400
0.179
0.176
0.173
0.172
0.169
0.168
0.166
0.160
0.159
0.157
0.155
0.154
0.153
0.150
0.144
Uno de los más usados actualmente es el espesador Lamella. Utiliza un conjunto de
placas paralelas e inclinadas, las cuales reducen la distancia de sedimentación y al mismo
tiempo aumenta el área efectiva. El área efectiva que ocupa un Lamella es solamente del
20% de un espesador convencional; las bandejas inclinadas permiten el asentamiento de
los sólidos deslizándose por gravedad dentro de la tolva.
4.1.7.2. CON COLA (NON FLOAT) PREVIO DESLAME
Para esta prueba, se tomo la cola de la prueba 3 de flotación en ciclo abierto. Los
resultados alcanzados se muestran en las siguientes tablas:
Prueba Nº
Apogee 849
Volumen pulpa, cm3
1000
Muestra
Colas de flotación
Peso de la pulpa, g
1197.969
Tipo de floculante
Magnafloc 292
-100 Mallas Ty
Dosificación, g/t
0, 10 y 20
2.950
Fracción vol. inicial
0,101
0,253
0
10
20
1
0
0.360
0.360
0.360
33
0.357
0.350
0.342
0.336
0.334
0.318
0.309
0.304
0.287
0.271
0.239
0.223
0.183
0.157
0.154
0.150
0.146
0.144
0.141
0.140
0.137
0.135
0.133
0.125
0.115
0.354
0.350
0.272
0.241
0.226
0.180
0.173
0.171
0.165
0.159
0.152
0.149
0.147
0.142
0.141
0.140
0.138
0.137
0.135
0.134
0.133
0.132
0.130
0.121
0.116
0.336
0.320
0.186
0.175
0.164
0.147
0.140
0.139
0.131
0.127
0.122
0.120
0.118
0.117
0.116
0.116
0.116
0.116
0.116
0.115
0.115
0.115
0.115
0.115
0.115
Tabla 20.- Resumen de resultados con diferentes concentraciones de floculante
Detalle de parámetros
0
10
20
Densidad del mineral seco, !s (g/cm3)
2.95
2.95
2.95
Fracción volumétrica de descarga, "D
0.253
0.253
0.253
Velocidad de sedimentación, Vs ("k) m/s
2.382x10-5
1.492x10-4
1.793x10-4
Área unitaria, m2/TPD
0.980
0.284
0.171
Área total espesador, m2 para 700 TPD
686.00
198.80
119.70
Diámetro del espesador calculado, m
29.55
15.91
12.35
Diámetro del espesador calculado, pies
96.96
52.20
40.50
Como se puede ver, en los resultados, el área superficial del espesador convencional
disminuye considerablemente cuando se trata de colas provenientes de la flotación previo
deslame.
4.1.8 ENSAYO ESTANDAR DE BOND PARA DETERMINACION DEL WORK INDEX
El test estándar de Bond es el método más conocido y utilizado para predecir consumos
de energía en molienda de minerales. Esta predicción de consumo de energía se hace
extensiva a molinos de bolas y de barras.
34
0.200
0.196
0.194
0.192
0.191
0.189
0.188
0.180
0.179
0.177
0.175
0.173
0.171
0.168
0.155
0
10
20
2.801
2.801
2.801
0.263
0.263
0.263
1.372x10-5
2.583x10-5
4.406x10-5
1.977
0.848
0.563
1383.90
593.60
394.10
41.98
27.49
22.40
137.72
90.20
73.49
32
2
3
4
5
6
7
8
9
10
11
12
13
14
15
16
17
18
19
20
21
22
23
24
25
26
0.255
0.238
0.231
0.222
0.215
0.209
0.205
0.189
0.186
0.183
0.182
0.179
0.178
0.174
0.153
El test de laboratorio elaborado por Fred Bond, consiste en una simulación de molienda
continua mediante un método que permite lograr estacionariedad a partir de sucesivos
ensayos “batch”.
La prueba entrega un valor para el índice de trabajo Wi, expresado en kWh/ton corta, el
cual introducido en la ecuación básica de la Tercera Ley de Conminución permite predecir
el consumo de energía de un molino de planta.
En general, se acepta que el error de predicción del consumo energético obtenido con
este ensayo es del orden de ! 20%.
Si bien es suficiente efectuar una prueba a una malla de corte determinado, en esta se
efectuaron pruebas a tres mallas de corte, por encima y por debajo de la principal que es
100 Mallas Tyler. Entonces, las mallas de corte estudiadas son: 65 Mallas Ty, 100 Mallas
Ty y 150 Mallas Ty.
4.1.8.1 DESCRIPCION DE LA MUESTRA
La muestra corresponde a yacimiento primario de polisulfuros con poco contenido de
sulfuros en el que prevalece la pirita. La densidad real de la muestra, determinada por el
método del picnómetro, es de 2.784 g/cc. Por las características mostradas durante la
preparación de la muestra, se observa una mena caracterizada como “media dura”, con
tendencia a formar lamas por la presencia de una importante cantidad de arcillas.
4.1.8.2 ENSAYO ESTANDAR
Siguiendo las recomendaciones efectuadas por el Sr. Bond y usando el equipo estándar
establecido, se efectuaron las pruebas a las mallas de corte preestablecidas. Los
resultados
alcanzados en los ensayos son los siguientes:
a) PARA UNA MALLA DE CORTE DE 65 MALLAS TYLER
MUESTRA: LA
MALLA DE CORTE: P1 = 65 Mallas Tyler (212 micrones)
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
TYLER
A MOLINO
"m
#
(micrones) % Retenido % Acumul.Pas
35
PRODUCTO
MOLIDO
% Retenido
% Acumul.
6
3350
8
2360
10
1700
14
1180
20
850
28
600
35
425
48
300
65
212
100
150
150
106
200
75
-200
-75
Alimentación
0
11.63
16.00
12.38
11.13
8.38
7.25
5.88
4.88
3.38
4.25
4.13
10.75
100.00
100,00
88.38
72.38
60.00
48.88
40.50
33.25
27.38
22.50
19.13
14.88
10.75
0.00
b) PARA UNA MALLA DE CORTE DE 100 MALLAS TYLER
0.00
27.58
17.82
13.30
41.30
100.00
100.00
72.42
54.60
41.30
0.00
RESUMEN RESULTADOS
F80 , ("m) = 2014.531(1)
P80, ("m) =
167.042(2)
Gbpe, (g/rev) = 1.725(3)
Wi, (Kwh/tc) = 10.799(4)
44.5
( 10
10 %
(Gbp ) 0.82 ( p1 ) 0.23 &
)
#
F80 #$
'& P80
*
wi
MUESTRA: LA
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
PRODUCTO
TYLER
A MOLINO
MOLIDO
"m
#
(micrones) % Retenido % Acumul.Pas % Retenido
% Acumul.
6
3350
0
100,00
8
2360
11.63
88.38
10
1700
16.00
72.38
14
1180
12.38
60.00
20
850
11.13
48.88
28
600
8.38
40.50
35
425
7.25
33.25
48
300
5.88
27.38
65
212
4.88
22.50
100
150
3.38
19.13
0.00
100.00
150
106
4.25
14.88
30.93
69.07
200
75
4.13
10.75
16.23
52.84
-200
-75
10.75
0.00
52.84
0.00
Alimentación
100.00
100.00
RESUMEN RESULTADOS
F80 , ("m) = 2014.531(1)
P80, ("m) =
wi
*
(1.725) 0.82 x (212) 0.23
44.5
( 10
)
&
' 167.042
* 10.79931 Kwh / tc
%
10
#
2014.531 $
Wi, (Kwh/tc) = 12.434(4)
(1) y (2), del análisis granulométrico de la alimentación y del producto (por interpolación)
(3), de las pruebas, siguiendo las normas del método sugerido por Bond
(4), por cálculo.
Donde: F80 = Tamaño en micrones bajo el cual está el 80% de la alimentación.
P80 = Tamaño en micrones bajo el cual está el 80% del producto.
P1 = Malla de corte en micrones
Gbpe = gramos por revolución del molino de bolas en estado estacionario.
Wi = Consumo unitario de energía que debería tener un material que se muele
en el molino, kWhh/tc
36
MUESTRA: LA
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
PRODUCTO
TYLER
A MOLINO
MOLIDO
"m
#
(micrones) % Retenido
%
% Retenido
% Acumul.
Acumul.Pas
6
3350
0
100,00
8
2360
11.63
88.38
10
1700
16.00
72.38
14
1180
12.38
60.00
20
850
11.13
48.88
28
600
8.38
40.50
35
425
7.25
33.25
48
300
5.88
27.38
65
212
4.88
22.50
100
150
3.38
19.13
150
106
4.25
14.88
0.00
100.00
200
75
4.13
10.75
29.80
70.20
-200
-75
10.75
0.00
70.20
0.00
Alimentación
100.00
100.00
F80 , ("m) = 2014.531(1)
P80, ("m) =
*
44.5
( 10
10 %
(Gbp ) 0.82 ( p1 ) 0.23 &
)
#
F80 #$
'& P80
44.5
( 10
(1.317) 0.82 x (150) 0.23 &
)
' 121.548
%
10
#
2014.531 $
* 12.43424 Kwh / tc
Puesto que se efectuaron pruebas a diferentes mallas de corte es posible realizar un
gráfico, como el que se muestra a continuación, en el cual se puede ver más
objetivamente la variación del consumo de energía en función, precisamente, de la Malla
de Corte.
4.1.9 COMENTARIOS FINALES PARA LA MUESTRA “LA”
85.194(2)
Gbpe, (g/rev) = 0.977(3)
Wi, (Kwh/tc) = 14.387(4)
*
wi
*
Figura 19.- Influencia de la Malla de Corte (65# = 212 µ, 100# = 150 µ, 150# = 106 µ)
en el Índice de Trabajo con la muestra LA, empresa Apogee
RESUMEN RESULTADOS
wi
wi
37
c) PARA UNA MALLA DE CORTE DE 150 MALLAS TYLER
wi
121.548(2)
Gbpe, (g/rev) = 1.317(3)
*
44.5
( 10
10 %
(Gbp ) 0.82 ( p1 ) 0.23 &
)
#
F80 #$
'& P80
44.5
( 10
(0.977) 0.82 x (106) 0.23 &
)
' 85.194
%
10
#
2014.531 $
* 14.38688 Kwh / tc
De una manera general, se puede decir que la muestra responde favorablemente al
proceso de flotación diferencial aunque los resultados en ciclo abierto no son del todo
satisfactorios. Los resultados obtenidos en ciclo cerrado son mejores tanto con deslame
como sin deslame.
En las pruebas de flotación en ciclo abierto sin deslame no se pudo lograr un buen
concentrado de plomo, este no llega a 50% Pb, tomando en cuenta la prueba 4 como la
mejor en esta serie, y también la recuperación es relativamente baja, 61.79%; el
concentrado de zinc alcanza una ley de 58.10% con una recuperación de 71.88% que no
satisface y la recuperación de la plata llega a 69.39% tomando en cuenta los
concentrados de plomo y zinc.
Las pruebas de flotación en ciclo abierto con previo deslame son las mejores y de ellas se
elije la prueba 3; en esta prueba se logra un concentrado final de plomo con una ley
38
39
aceptable que llega a 51.76% y una recuperación de 79.47%, el concentrado de zinc
también tienen índices alentadores como 57.40% Zn y una recuperación de 83.06% y la
recuperación total de la plata, en ambos concentrados, alcanza 85.69%
5. CONCLUSIONES
Del análisis de resultados obtenidos y de las observaciones durante las pruebas
experimentales, se puede colegir lo siguiente:
En cuanto a las pruebas de flotación en ciclo cerrado se puede mencionar que también
son buenas, así, en la prueba sin deslame, se logra un concentrado de plomo de 52.40%
con una recuperación de 78.75% y el concentrado de zinc alcanza a una ley de 57% Zn
con una recuperación de 81.10%, la plata por su parte tienen una recuperación total de
82.73%. Estos resultados son un tanto similares a los obtenidos cuando se realiza un
deslame previo a la flotación, en efecto, el concentrado de plomo de 53.40% con una
recuperación del 76.8%, mientras que el concentrado de zinc tienen una ley de 59% Zn
con una recuperación de 78.68%; la recuperación total de la plata es de 83.35%.
-
-
Los análisis granulométricos de las colas de flotación en ciclo abierto y los análisis size by
size, muestran que la mayor parte de las pérdidas de los elementos valiosos se
encuentran en los granos más finos y que están por debajo de la malla 400, -38 micrones.
plomo, 61.79% y zinc, 71.88%.
-
En circuito abierto, con previo deslame y considerando la prueba 3 como la
mejor, se han logrado estos resultados: ley de plata en el concentrado de
plomo, 8290 g/t y en el concentrado de zinc, 1590 g/t; ley de plomo en el
concentrado de plomo, 51.76% y el concentrado de zinc alcanza una ley de
57.40% Zn, con las siguientes recuperaciones: plata total, 85.69%; plomo,
79.47% y zinc, 83.06%.
-
En circuito cerrado sin deslame, estos son los resultados: ley de plata en el
concentrado de plomo, 6620 g/t; en el concentrado de zinc, 2010 g/t; ley de
plomo en el concentrado de plomo, 52.40% y la ley del zinc en su concentrado
es de 57.00% y las recuperaciones son: plata, 82.73%; plomo, 78.75% y zinc,
81.10%.
-
En circuito cerrado con deslame, los resultados que se han logrado son: ley de
plata en el concentrado de plomo, 9670 g/t; en el concentrado de zinc, 1080
Los reactivos usados son los que habitualmente se usan y los que se han mencionado en
este tipo de operaciones, salvo el fluosilicato de sodio que permitió dispersar y depresar
una parte de la ganga fina.
El contenido de lamas en las colas de flotación, cuando ésta se realiza sin previo
deslame, está alrededor de 19.11% en peso y cuando se realiza un previo deslamado,
estas son del orden del 9.30% en peso y en circuito cerrado es tos valores alcanza a
19.60% sin deslame y a 9.94% con previo deslame. Por otro lado, la velocidad de
sedimentación de las colas cuando sin deslame y cuando no se usan un floculante es del
orden de 1.372x10-5 m/s y ésta mejora cuando previo a la flotación se efectúa el deslame,
la velocidad es de 2.382x10-5 m/s.
Finalmente, el Work Index, de esta muestra tiene la siguiente secuencia de valores:
MALLA DE CORTE
MALLAS TYLER
MICRONES
g/t; ley de plomo en el concentrado de plomo, 53.40% y la ley del zinc en su
concentrado es de 59.00% y las recuperaciones son: plata, 83.35%; plomo,
76.98% y zinc, 78.68%.
WORK INDEX
Kwh/tc
65
212
10.799
100
150
12.434
150
106
14.387
La muestra es apta de ser tratada por el proceso de flotación diferencial ya que
se logran obtener concentrados con índices metalúrgicos aceptables y
susceptibles de ser mejorados; esta situación se ha visto en las pruebas tanto,
principalmente en las de ciclo cerrado.
.
En circuito abierto se han logrado estos resultados: ley de plata en el
concentrado de plomo, 5850 g/t y en el concentrado de zinc, 2430 g/t; ley de
plomo en el concentrado de plomo, 49.54% y el concentrado de zinc alcanza
una ley de 58.10% Zn, con las siguientes recuperaciones: plata, 69.36%;
-
Estos resultados muestran claramente la posibilidad de, obtener similares o
incluso, mejores resultados en una operación industrial.
-
La presencia de lamas es de consideración y perjudicial, alrededor del 19% en
peso; a pesar de ellos se puede flotar sin previo deslame, prueba1.
41
40
-
-
-
-
No fue posible una mayor recuperación del elemento plomo, es el que menor
recuperación ha arrojado en todas las pruebas de flotación, porque se torna
muy difícil la separación de otros sulfuros como el propio mineral de zinc y
sulfuros de hierro.
Los análisis granulométricos de las colas y los análisis size by size permiten
afirmar que la mayor pérdida de los elementos valiosos en las colas se
encuentran en tamaños de fina granulometría, concretamente por debajo de
400 Mallas Tyler, -38 micrones.
La velocidad de sedimentación de las partículas, a partir de las colas de
flotación, es lenta porque algo más del 50% en peso de la muestra que entra al
proceso de flotación está por debajo de la malla 400 y gran parte de esta
fracción corresponde a la presencia de lamas; esta velocidad es de 1.1372 x
10-5 m/s.
Referencias
Informe Nº
11/2009
Trabajo realizado:
Experimentación metalúrgica con
muestra compleja de sulfuros,
denominada “Alta Ley” proveniente
del sector de Pulacayo y
pertenecientes a la empresa Apogee
Minerals Bolivia S.A.
Fecha:
Oruro, NOVIEMBRE de 2009
Autores del Informe:
Ing. Octavio Hinojosa C.
Ing. Cinda E. Beltran O.
Trabajo realizado por:
Ing. Octavio Hinojosa C.
Ing. Cinda E. Beltran O.
Sr. Celestino Mamani R.
Sr. Francisco Sanchez
En:
Laboratorio Concentración de
Minerales de la Facultad Nacional de
Ingeniería – UTO.
Análisis Químico:
Als Chemex
Copiado:
Secretaría Laboratorio Concentración
de Minerales
Dirección:
Telf. 5263888 Laboratorio
Telf. 5261046 Secretaría Carrera
FAX: 5260008
El Índice de Trabajo, Work Index, para una malla de corte de 100 Mallas Tyler,
150 micrones, es de 12.434 Kwh/tc.
42
43
ANEXO
FOTOGRAFIAS DE LAS PRUEBAS EXPERIMENTALES
Fotografía 3.- Se procede al desembalaje de las muestras recibidas
Fotografía 1.- Las muestras de la empresa Apogee se
recibieron en doble embase; este es el embase externo
Fotografía 4.- Se aprecia el embalaje interior de las muestras recibidas
Fotografía 2.- Inspección de cada uno de los recipientes que se
recibió en el Laboratorio Concentración de Minerales de la UTO.
Fotografía 5.- Trituración primaria de las muestras
Fotografía 6.- Otra vista de la trituración primaria de las muestras
Fotografía 9.- Trituración secundaria y cernido
Fotografía 7.- Cernido, antes de la trituración secundaria
Fotografía 10.- Mezclado y homogeneizado de las muestras trituradas
Fotografía 8.- Trituración secundaria en Trituradora de Rollos
Fotografía 11.- Cuarteo, obtención de muestras para la
realización de las diferentes pruebas
Fotografía 12.- Cargado de la muestra al molino de bolas,
operación previa a la flotación diferencial
Fotografía 13.- Descargado de la muestra del molino de bolas
Fotografía 14.- Parte final del descargado de la muestra del molino
Fotografía 18.- Cargado de la pulpa a la celda de flotación
(de dos kilogramos)
Fotografía 15.- Clasificación de la muestra molida, previo
a la flotación diferencial
Fotografía 16.- Cargado, del producto molido, al reactor del
ciclón para eliminar las lamas
Fotografía 17.- Ajuste del equipo de ciclonaje
Fotografía 21.- Flotación rougher de Pb-Ag
Fotografía 22.- Pesaje de reactivos, cal para regular el pH
Fotografía 19.- Acondicionamiento de la pulpa, antes de la flotación Pb-Ag
Fotografía 20.- Control de pH de la pulpa
Fotografía 23.- Adición de reactivos a la etapa de flotación de Zn-Ag
Fotografía 24.- Flotación rougher de Zn-Ag
Fotografía 27.- Flotación cleaner de Zn-Ag
Fotografía 25.- Flotación rougher de Zn-Ag
Fotografía 28.- Concentrados finales; concentrado de Pb-Ag,
Izquierda y concentrado de Zn-Ag, derecha
Fotografía 26.- Flotación cleaner de Pb-Ag
UNIVERSIDAD TÉCNICA DE ORURO
FACULTAD NACIONAL DE INGENIERÍA
CARRERA DE METALURGIA Y CIENCIA DE MATERIALES
Facultad Nacional de Ingeniería
Carrera de Ingeniería Metalúrgica
Laboratorio Concentración de Minerales
LABORATORIO
EXPERIMENTACION METALURGICA CON DOS MUESTRAS
COMPLEJAS DE SULFUROS PROVENIENTES DEL SECTOR
DE PULACAYO Y PERTENECIENTES A LA EMPRESA
CONCENTRACIÓN
ACIÓN
CONCENTR
DE M
MINERALES
INERALES
DE
1. INTRODUCCION
INFORME Nº 11/09
EXPERIMENTACION METALURGICA CON
MUESTRA COMPLEJA DE SULFUROS
DENOMINADA ALTA LEY Y PROVENIENTE
DEL SECTOR DE PULACAYO Y
PERTENECIENTE A LA EMPRESA APOGEE
MINERALS BOLIVIA S.A.
NOVIEMBRE, 2009
Oruro, Bolivia
La Empresa Apogee Minerals Bolivia S. A., por intermedio del Ing. Luis Casal, personero de
la empresa, con correos sucesivos a través del internet y visita a nuestro laboratorio de un
consultor extranjero y del vicepresidente de exploración de la empresa, oficializaron la
realización de pruebas metalúrgicas, encomendando para ello al Laboratorio Concentración
de Minerales, de la Carrera de Ingeniería Metalúrgica de la Facultad Nacional de Ingeniería
de la Universidad Técnica de Oruro, la experimentación de las mismas con dos muestras de
complejos sufurosos de Zn, Pb y Ag con la finalidad, principalmente, de recuperar los
contenidos de estos elementos valiosos, por flotación diferencial.
Para tal efecto se recibieron las muestras en cantidades suficientes para la realización de
todas las pruebas programadas. Las muestras, provienen de yacimiento primario y se
denominan: LM (media ley) y LB (baja ley). Estas muestras, características de perforaciones
de diamantina (probetas), tienen tamaños de grano de hasta 4 pulgadas.
Una observación estereomicroscópica de las muestras, luego de una adecuada limpieza,
permite identificar pirita en forma mayoritaria, también se observan pequeñas cantidades de
sulfuros de plomo y zinc; está presente una gran cantidad de cuarzo-silicatos y pizarras. Las
muestras presentan características de formar lamas (por el contenido de arcillas), la más
notoria en este aspecto es la muestra LB (baja ley).
La representatividad de las muestras es responsabilidad de la Empresa; en esta etapa no
participó el Laboratorio Concentración de Minerales de la Carrera de Ingeniería Metalúrgica.
Dirección
Ciudadela Universitaria,
Edif. Carrera de Ingeniería Metalúrgica
Teléfono
591-2-5263888
Correo Electrónico
[email protected]
1
Informe Nº 09/09
2. OBJETIVOS
Los objetivos del presente trabajo de investigación con las dos muestras, a solicitud de la
empresa, se encaminaron a:
-
Determinar el rango de recuperación y grado de concentrados de plomo-plata y zincplata, en ciclo abierto, obtenidos por flotación diferencial.
-
Determinar el rango de recuperación y grado de concentrados de plomo-plata y zincplata, en ciclo cerrado, obtenidos por flotación diferencial.
-
-
Análisis granulométrico de la alimentación a flotación y de colas de flotación
Determinación de lamas de las colas de flotación por ciclonaje
Análisis size by size
Pruebas de sedimentación
Pruebas de determinación del Work Index.
MUESTRA:
Efectuar análisis granulométricos de las colas de las pruebas de flotación en ciclo
abierto.
-
Efectuar el análisis size by size de las pruebas de flotación diferencial en ciclo abierto.
-
Determinar el contenido de lamas (arcillas) en las colas de todas las pruebas de
flotación, a través de pruebas de ciclonaje.
-
Realizar pruebas de sedimentación, a partir de las colas de flotación
-
Determinar el Indice de Trabajo (Work Index)..
LM
LB
TRITURACION
PRIMARIA
CERNIDO, ¼”
Como objetivos secundarios deben establecerse las condiciones de operación y consumo de
reactivos en las pruebas de flotación diferencial.
La experimentación metalúrgica para la presente investigación, se llevó a cabo de acuerdo a
lo que se muestran en los flujogramas de las figuras 1, 2, 3, 4 y 5 y el detalle descriptivo que
se anota específicamente a continuación:
-
+¼”
HOMOGENEIZACION
Y CUARTEO
TRITURACION
SECUNDARIA
Análisis
químico
PRUEBAS DE
FLOTACION
DIFERENCIAL
3. EXPERIMENTACIÓN METALÚRGICA
-¼”
Análisis
granulométrico
DETERMINACION
DEL WORK INDEX
Figura 1.- Flujograma de la etapa de preparación de las muestras para la
experimentación, Empresa Apogee
Trituración primaria y secundaria de las muestras
Homogeneización, cuarteo y obtención de comunes representativos
para las diferentes pruebas.
Preparación de las muestras para la realización de las pruebas de flotación
diferencial.
Pruebas de flotación diferencial sin deslame
Pruebas de flotación diferencial con deslame
Informe Nº 09/09
2
Informe Nº 09/09
3
MUESTRA (LM o LB)
CLASIFICACION,
100 Mallas Ty
CICLONAJE
(-)
(+)
Under flow
Over flow
(lamas-arcillas)
MOLIENDA
Regulador de pH
Depresor
ACONDICIONAMIENTO-1
Espumante
Colector
FLOTACION
ROUGHER de Pb-Ag
Espuma Pb-Ag
1ra FLOTACION
CLEANER
Non Float
Regulador de pH
Activador
Espuma Pb
NF-1ra Limpieza
2da FLOTACION
CLEANER
Espuma
Pb-Ag
ACONDICIONAMIENTO 2
Colector
Espumante
FLOTACION
ROUGHER de Zn
NF-2da Limpieza
Non Float
Espuma Zn
1ra FLOTACION
CLEANER de Zn
Espuma Zn
2da FLOTACION
CLEANER de Zn
CUARTEO
NF-1ra Limpieza
Análisis
granulométrico
CICLONAJE
Prueba de
sedimentación
Espuma de Zn
NF-2da Limpieza
Figura 3.- Flujograma de las pruebas experimentales, que se siguieron
con las muestras complejas de la Empresa Apogee, en
CICLO ABIERTO y PREVIO DESLAMADO
Informe Nº 09/09
Informe Nº 09/09
4
6
Informe Nº 09/09
Informe Nº 09/09
5
7
4. RESULTADOS Y COMENTARIOS
Tomando en cuenta los objetivos del presente trabajo y el interés de la empresa Apogee de
obtener la mayor información posible respecto a los resultados de los diferentes trabajos de
investigación con las dos muestras, Media Ley y Baja Ley, la presente investigación se dividió
en dos partes; a pesar de que varias de las operaciones unitarias debían realizarse
simultáneamente, para fines prácticos, la explicación se desarrollará por etapas.
4.1 PRIMERA PARTE: MUESTRA “LM”, MEDIA LEY
A continuación se presentarán los resultados de acuerdo a un desarrollo práctico, de tal
manera que se efectúe un seguimiento objetivo del trabajo experimental.
representativos para la realización de las diferentes pruebas. Estos pasos se pueden seguir
observando el flujograma que se muestra en la figura 1.
La ley de cabeza ensayada del común representativo de esta muestra da el siguiente
resultado:
Plata: 181 g/t
Plomo: 0.69%
Zinc: 2.45%
Cobre: 0.068%
4.1.2 FLOTACION DIFERENCIAL DE SULFUROS EN CIRCUITO ABIERTO
Prueba 1:
Esta prueba fue llevada a cabo siguiendo los pasos que se muestran en el flujograma de la
figura 2, pero, en esta oportunidad, con una sola limpieza. Se realiza la flotación diferencial,
flotando primero el mineral de Pb-Ag, para ello se efectuó la molienda a -100 Mallas Tyler y
en la etapa de acondicionamiento se usó cal, como regulador de pH; sulfato de zinc como
depresor del mineral de zinc y cianuro de sodio para depresar las piritas que se encuentran
en apreciable cantidad en la muestra; como colector se usó el ditiofosfato Aero Float-242 y
8
Informe Nº 09/09
como espumante se usó el Metil Isobutil Carbinol, más conocido como MIBC. El grado de
molienda, el colector y el espumante, fueron elegidos previas pruebas de flotación
exploratorias.
Los resultados de esta primera prueba se muestran en la tabla 1.
Tabla 1.- Balance metalúrgico de la prueba de flotación diferencial, prueba 1, muestra LM
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
1,16
38,00
63,10
2890
17,34
2,30
1,13
NF – Limpieza Pb
8,09
0,54
6,28
242
10,16
1,96
6,74
Espuma rougher- Pb
9,24
5,22
69,37
573
27,50
2,00
7,87
Espuma de Zn-Ag
3,56
1,55
7,94
3280
60,71 46,60
70,66
NF-Limpieza Zn
1,75
2,68
6,72
419
3,80
2,47
1,83
Espuma rougher Zn
5,31
1,92
14,66
2340
64,51 32,09
72,50
Non Float
85,45
0,13
15,96
18
7,99
0,54
19,63
Cabeza Calculada
100,00
0,70
100,00
193 100,00
2,35 100,00
Las condiciones de operación y consumo de reactivos se muestran en la figura 6; se debieron
efectuar dos flotaciones, en las mismas condiciones, para tener suficiente espuma rougher y
llevar a la limpieza, especialmente espumas de Pb-Ag.
En esta prueba, la calidad de los productos finales es baja y también son bajas las
recuperaciones; no se ha logrado una adecuada separación de los componentes
mineralógicos, debido, posiblemente, a la falta de reactivo y/o los tiempos de flotación y
acondicionamiento no fueron suficientes.
-
Prueba 2:
Esta prueba también fue llevada a cabo siguiendo los pasos que se muestran en el flujograma
de la figura 2. En esta prueba se incrementaron un tanto los reactivos y los tiempo de
acondicionamiento y flotación, también se incrementaron las etapas de limpieza. Los
resultados de esta segunda prueba se muestran en la tabla 2 y las condiciones de operación
y consumo de reactivos en la figura 7.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0,70
43,00
42,90
8160
28,98 14,75
4,20
NF–2da Limp.de Pb
0,50
13,15
9,37
4010
10,17 17,45
3,55
NF-1ra Limp. de Pb
2,78
1,94
7,69
454
6,41
3,37
3,81
Espuma rougher- Pb
3,98
10,56
59,96
2255
45,56
7,14
11,56
Espuma de Zn-Ag
3,40
3,15
15,26
2250
38,81 53,60
74,11
NF-2da Limp. de Zn
0,55
3,23
2,53
1315
3,67 14,35
3,21
NF-1ra Limp. de Zn
3,88
0,84
4,65
198
3,90
2,27
3,58
Espuma rougher Zn
7,83
2,01
22,44
1167
46,38 25,40
80,91
Non Float
88,19
0,14
17,60
18
8,05
0,21
7,53
Cabeza Calculada
100,00
0,70
100,00
197 100,00
2,46 100,00
Se observa una mejora en los resultados, tanto en calidad como en recuperación, en los
concentrados finales de plomo y zinc, aunque esta recuperación disminuye en el elemento
plata; es importante remarcar que la recuperación total de las espumas rougher de los
elementos plomo y zinc se mantiene.
Ante la presencia de material fino, lamas, que perjudican la calidad de los concentrados
finales, se introdujo el fluosilicato de sodio, solo en la etapa de las limpiezas, con resultados
positivos. Otro aspecto importante que se observa en esta prueba es el hecho de que la plata
se enriquece más en el concentrado de plomo, disminuyendo en consecuencia en el
concentrado de zinc. En esta prueba se debieron efectuar 3 flotaciones e las mismas
condiciones con la finalidad de acumular espumas rougher y afrontar adecuadamente las
etapas de limpieza.
Informe Nº 09/09
9
Informe Nº 09/09
-
10
Informe Nº 09/09
11
-
Recuperación total de la plata 67.79%
Recuperación del zinc: 74,11%
Prueba 3:
De la misma manera que la anterior prueba, ésta fue llevada a cabo siguiendo los pasos que
se muestran en el flujograma de la figura 2. En esta prueba se incrementó un poco más el
consumo de los reactivos. Los resultados de esta tercera prueba se los muestra en la tabla 3
y las condiciones de operación y consumo de reactivos en la figura 8.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0,82
50,00
58,04
9240
36,85
4,27
1,37
NF–2da Limp.de Pb
0,15
6,92
1,48
1950
1,43
4,50
0,26
NF-1ra Limp. de Pb
3,81
2,57
13,88
719
13,34
2,99
4,46
Espuma rougher- Pb
4,77
10,83
73,39
2218
51,62
3,26
6,09
Espuma de Zn-Ag
4,22
1,12
6,71
1875
38,61 50,50
83,55
NF-2da Limp. de Zn
0,95
0,57
0,77
273
1,27
4,78
1,78
NF-1ra Limp. de Zn
6,23
0,28
2,47
65
1,97
0,96
2,34
Espuma rougher Zn
11,40
0,62
9,96
753
41,85 19,63
87,67
Non Float
83,82
0,14
16,65
16
6,54
0,19
6,24
Cabeza Calculada
100,00
0,70
100,00
205 100,00
2,55 100,00
Mejoran los resultados en cuanto a la calidad de los concentrados y recuperaciones de los
elementos valiosos, merced a un incremento de los reactivos, este aspecto puede observarse
en el flujograma de la figura 8. La distribución del plomo en las colas finales es todavía
ligeramente alta, aspecto que debería corregirse a través de una tiempo mayor de flotación
del plomo y/o quizás incrementando un tanto más el colector en esta etapa. En esta prueba
como en la prueba 2, tuvieron que efectuarse tres flotaciones en las mismas condiciones con
la finalidad de contar con suficiente espuma como para afrontar dos etapas de limpieza
especialmente con las espumas de plomo que son las más escasas.
-
Informe Nº 09/09
12
13
Informe Nº 09/09
-
Recuperación de plomo, 58,04%
Esta prueba se llevó adelante con el under flow de la operación de ciclonaje, operación que
se llevó a cabo previa a la flotación diferencial.
La muestra molida a -100 Mallas Tyler fue sometida a un deslamado en un ciclón, la
secuencia de esta prueba se puede seguir en el flujograma de la figura 3. Esta prueba, con
excepción del ciclonaje, fue llevada a cabo en forma similar a la prueba 3 en cuanto al
consumo de reactivos se refiere. Los resultados de esta segunda prueba se los muestra en la
tabla 4.
Informe Nº 09/09
14
Informe Nº 09/09
15
Productos
Espuma de Pb-Ag
NF–2da Limp.de Pb
NF-1ra Limp. de Pb
Espuma rougher- Pb
Espuma de Zn-Ag
NF-2da Limp. de Zn
NF-1ra Limp. de Zn
Espuma rougher Zn
Non Float
Over flow (lamas)
Cola Total
Cabeza Calculada
Peso
%
0,68
0,20
0,85
1,73
2,97
0,37
1,41
4,75
71,22
22,31
93,52
100,00
PLOMO
% Pb
% Dist.
57,00
55,49
29,20
8,42
6,47
7,81
29,04
71,72
0,29
1,23
0,72
0,38
1,14
2,30
0,58
3,91
0,13
13,22
0,35
11,15
0,18
24,37
0,70
100,00
PLATA
g/t Ag % Dist.
10000
35,49
7270
7,65
1790
7,88
5666
51,02
679
10,49
338
0,64
2900
21,35
1314
32,48
16
5,93
91
10,57
34
16,50
192 100,00
ZINC
% Zn % Dist.
5,78
1,70
4,80
0,42
3,90
1,42
4,75
3,54
55,40
70,92
11,25
1,78
13,80
8,42
39,60
81,12
0,12
3,69
1,21
11,65
0,38
15,34
2,32 100,00
Las condiciones de operación y consumo de reactivos en la figura 9.
-
Prueba 5, sin deslame:
16
Informe Nº 09/09
17
Informe Nº 09/09
Esta prueba, se llevó a cabo pretendiendo mejorar los anteriores resultados y tratando de
demostrar que el deslamado, en esta muestra, no es necesario. La prueba de flotación
diferencial se basó bastante en las condiciones en las que se llevó a cabo la prueba 3 y, por
supuesto, siguiendo los pasos que se ven en la figura 2. Los resultados de esta última
prueba se los muestra en la tabla 5 y las condiciones de operación y consumo de reactivos en
la figura 10.
Tabla 5.- Balance metalúrgico de la prueba 5 de flotación diferencial, muestra LM
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
0,98
50,00
59,34 6440
29,54
3,14
1,20
Espuma de Pb-Ag
NF–2da Limp.de Pb
0,54
17,30
11,29 3050
7,69
4,22
0,89
NF-1ra Limp. de Pb
6,27
0,49
3,72
152
4,46
0,42
1,03
Espuma rougher- Pb
7,79
7,88
74,35 1143
41,69
1,03
3,11
3,99
0,71
3,43 2020
37,76 49,90
77,63
Espuma de Zn-Ag
NF-2da Limp. de Zn
1,13
0,49
0,67
777
4,10 13,90
6,10
NF-1ra Limp. de Zn
4,75
0,28
1,61
116
2,58
1,39
2,57
Espuma rougher Zn
9,88
0,48
5,71
962
44,43 22,44
86,31
Non Float
82,33
0,20
19,94
36
13,87
0,33
10,58
Cabeza Calculada
100,00
0,83
100,00
214
100,00
2,57 100,00
Los resultados no son mejores a la prueba 3, pero es posible obtener y mejorar estos
resultados. Lo que no se puede mejorar es una mayor recuperación del plomo de las colas,
siendo este el elemento que en mayor distribución se encuentra en estas colas. Las
condiciones de operación y consumo de reactivos se encuentran detallados en la figura 10.
-
Informe Nº 09/09
18
Informe Nº 09/09
19
-
Las pruebas de flotación en ciclo cerrado se desarrollaron sin deslame y con previo deslame,
siguiendo los pasos que se muestran en los flujogramas de las figuras 4 y 5.
Prueba 1:
Esta prueba fue desarrollada siguiendo los pasos que se muestran en la figura 4; los
resultados que se alcanzaron están detallados en el balance metalúrgico de la tabla 6.
ciclo cerrado, prueba 1, muestra LM
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0,70
52,00
51,97 10250
34,28
4,19
1,16
Espuma de Zn-Ag
4,36
1,15
7,20
1790
37,48 47,90
83,32
Non Float
94,94
0,30
40,84
62
28,24
0,41
15,51
Cabeza Calculada
100,00
0,70
100,00
207 100,00
2,51 100,00
La calidad del concentrado de plomo en cuanto al mismo plomo y plata son aceptables pero
las recuperaciones de estos dos elementos son relativamente bajas, se observa una alta
distribución de los mismos en las colas; por otro lado, el concentrado de zinc no llega a 50%,
pero la recuperación es aceptable aunque también una cantidad apreciable de zinc se
encuentra en las colas.
figura 11. Los índices metalúrgicos que se logran, en estas condiciones de operación, son:
-
20
Informe Nº 09/09
-
21
Informe Nº 09/09
Prueba 2:
Esta prueba se llevó adelante con previo deslame, siguiendo los pasos que se muestran en el
flujograma de la figura 5; los resultados se resumen en el balance metalúrgico de la tabla 7
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0,80
56,93
61,51 12250
47.49
5,00
1,83
Espuma de Zn-Ag
3,56
0,74
3,55
1460
25.12 51,40
83,66
Non Float
75,59
0,24
24,42
48
17.52
0,42
14,51
Lamas
20,05
0,39
10,52
102
9.87
1,25
11,45
Total colas
95,63
0,27
34,94
59
27.39
0,59
25,95
Cabeza Calculada
100,00
0,74
100,00
207 100,00
2,19 100,00
La calidad de los concentrados, tanto de plomo como de zinc, mejora y también mejora la
recuperación de estos elementos valiosos; lo mismo se puede decir de la plata. La
distribución de todos los elementos valiosos en las colas es alta y debe mejorarse este
aspecto ya que otra parte, aunque pequeña, se distribuye en las lamas, de todas formas,
estos resultados son mejores que los obtenidos en la anterior prueba.
figura 12.
-
Radio de concentración del plomo: 125
Informe Nº 09/09
22
Informe Nº 09/09
23
-
Prueba 3:
Esta tercera prueba se llevó a cabo para intentar mejorar los dos anteriores resultados y para
ello se decidió efectuar la misma sin deslame, siguiendo los pasos que se muestran en el
flujograma de la figura 4; los resultados se resumen en el balance metalúrgico de la tabla 8
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
1,20
51,00
74,32
6220
33,66
3,72
1,72
Espuma de Zn-Ag
3,70
0,85
3,80
2990
49,67 58,30
82,60
Non Float
95,10
0,19
21,87
39
16,67
0,43
15,68
Cabeza Calculada
100,00
0,83
100,00
222 100,00
2,61 100,00
La calidad de los concentrados, tanto de plomo como de zinc, mejora y también mejora la
recuperación de estos elementos valiosos; lo mismo se puede decir de la plata. La
distribución de todos los elementos valiosos en las colas disminuye considerablemente.
-
24
Informe Nº 09/09
-
Peso
%
9,65
13,98
10,20
8,36
5,74
52,08
100,00
PLOMO
% Pb
% Dist.
0,112
7,69
0,088
8,76
0,078
5,66
0,102
6,07
0,097
3,96
0,183
67,85
0,14
100,00
PLATA
g/t Ag % Dist.
21
10,95
15
11,33
13
7,16
13
5,88
18
5,58
21
59,10
18,5 100,00
25
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Informe Nº 09/09
Peso
%
6,42
12,00
9,55
9,97
7,10
54,95
100,00
PLOMO
% Pb
% Dist.
0,076
3,50
0,071
6,10
0,077
5,27
0,074
5,29
0,092
4,68
0,191
75,17
0,14
100,00
PLATA
g/t Ag % Dist.
12
4,44
11
7,60
11
6,05
10
5,74
16
6,54
22
69,62
17,4 100,00
ZINC
% Zn % Dist.
0,394
13,28
0,095
5,98
0,171
8,57
0,074
3,87
0,069
2,57
0,228
65,72
0,19 100,00
debajo de 400 Mallas Tyler.
ZINC
% Zn % Dist.
1,825
33,43
1,050
27,88
0,402
7,78
0,102
1,62
0,200
2,18
0,274
27,10
0,53 100,00
En principio, es importante recordar que las leyes de los elementos valiosos en las colas, de
la prueba 1, son: plomo, 0.13%; plata, 18 g/t y zinc, 0.54%; estas leyes deben compararse
con las obtenidas por cálculo en la tabla 9, casi coinciden. En segundo lugar, se puede
establecer que la mayor pérdida de los elementos valiosos que se produce es en la
granulometría fina, por debajo de 400 Mallas Tyler, 38 micrones.
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
7,76
13,02
9,10
10,90
5,65
53,57
100,00
PLOMO
% Pb
% Dist.
0,071
4,10
0,076
7,36
0,074
5,01
0,076
6,16
0,087
3,65
0,185
73,71
0,13
100,00
PLATA
g/t Ag % Dist.
10
5,15
9
7,77
10
6,03
12
8,68
13
4,87
19
67,50
15,1 100,00
ZINC
% Zn % Dist.
0,348
13,53
0,174
11,35
0,104
4,74
0,211
11,52
0,070
1,98
0,212
56,88
0,20 100,00
debajo de 400 Mallas Tyler, esta situación se repite en todas las pruebas.
Informe Nº 09/09
Peso
26
Informe Nº 09/09
PLOMO
PLATA
ZINC
27
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
%
9,37
14,35
18,23
8,61
11,39
38,06
100,00
% Pb
0,070
0,073
0,074
0,090
0,117
0,227
0,138
% Dist.
4,75
7,59
9,78
5,61
9,66
62,61
100,00
g/t Ag
10
10
10
12
20
22
15,9
% Dist.
5,90
9,03
11,48
6,51
14,35
52,73
100,00
% Zn
0,166
0,100
0,075
0,107
0,070
0,155
0,12
% Dist.
12,99
11,98
11,42
7,69
6,66
49,27
100,00
Para realizar este análisis es necesario contar con el análisis granulométrico de la
alimentación a la flotación rougher, es decir, la muestra molida a -100 Mallas Tyler.
i) Análisis granulométrico de la alimentación
Las leyes de los elementos valiosos en las colas de la prueba 4, son: plomo, 0.13%; plata, 16
g/t y zinc, 0.12%; estas leyes deben compararse con las obtenidas por cálculo en la tabla 12,
casi coinciden. En esta tabla, también se puede establecer que la mayor pérdida de los
elementos valiosos que se produce es en la granulometría fina, por debajo de 400 Mallas
Tyler, esta situación se repite en todas las pruebas.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
+150#
11,47
0,296
4,21
123
7,13
2,07
10,07
-150# +200#
4,16
0,866
4,47
268
5,64
3,08
5,44
-200# +270#
4,84
0,608
3,64
232
5,67
2,64
5,41
-270# +325#
18,79
0,695
16,18
208
19,75
2,58
20,55
-325# +400#
6,52
2,520
20,37
495
16,32
4,37
12,09
-400#
54,22
0,761
51,13
166
45,49
2,02
46,44
Cabeza calculada
100,00
0,81
100,00
197,9 100,00
2,36 100,00
Este análisis granulométrico, a través del % Peso, permite además calcular el d80 del producto
molido; para ello se tiene el gráfico de la figura 14.
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
9,92
10,65
11,83
8,74
8,92
49,95
100,00
PLOMO
% Pb
% Dist.
0,115
5,44
0,116
5,89
0,134
7,56
0,124
5,16
0,252
10,71
0,274
65,25
100,00
0,21
PLATA
g/t Ag % Dist.
23
5,43
27
6,85
32
9,02
26
5,41
65
13,81
50
59,49
100,00
42,0
ZINC
% Zn % Dist.
1,160
21,59
0,870
17,38
0,592
13,14
0,405
6,64
0,550
9,20
0,342
32,06
0,53 100,00
Estos resultados muestran que esta prueba no fue conducida adecuadamente porque las
leyes de los elemento valiosos en esta cola son ligeramente elevadas, aspecto que incide
negativamente en la recuperación final de estos elementos. También es bueno remarcar que
las leyes calculadas con las ensayadas en la cola de esta prueba son bastante similares.
De estos análisis granulométricos, tablas 9, 10, 11, 12 y 13, se puede establecer que las
mayores pérdidas de los elementos valiosos, plomo, plata y zinc, se producen en los tamaños
de grano que se encuentran por debajo de la malla 400.
Entonces, el d80 es igual a 53 micrones y el d50 es igual a 41.5 micrones
Este análisis granulométrico, también permitirá calcular las leyes de cada tamaño de grano de
las alimentaciones con las que se efectuó cada prueba de flotación y con la cola de cada
prueba se efectúa el análisis size by size.
ii) Alimentación vs cola de la prueba 1, flotación a ciclo abierto
28
Informe Nº 09/09
29
Informe Nº 09/09
Este análisis es más práctico y objetivo a través de un gráfico como el que se muestra en la
figura 15.
La figura 15, permite observar en forma clara que el elemento plomo no se recupera bien en
los granos gruesos y sobre todo, no se recupera adecuadamente los tamaños de grano que
se encuentran por debajo de la malla 400; solo es posible una adecuada recuperación de los
tamaños de grano que están por encima de 400 Mallas y por debajo de 270 Mallas Ty. Una
situación más pronunciada e inaceptable se da con el elemento zinc; la recuperación de este
elemento en las fracciones gruesas es pésima, especialmente por encima de 200 Mallas Ty.
Esta situación debe ser corregida en siguientes pruebas.
En esta prueba, la recuperación por fracciones, de una manera general, mejora
sustancialmente; dentro de esta mejora se puede apreciar que el elemento plomo es el que
ocasiona mayores dificultades, solo las fracciones finas, las que se encuentran entre -270#
+400#, son adecuadamente recuperados; la recuperación del elemento zinc, mejoró
sustancialmente, especialmente en los granos gruesos; también mejoró la recuperación de la
plata.
iv) Alimentación vs cola de la prueba 3, flotación a ciclo abierto
En este caso, también se realiza el análisis a través del gráfico que se encuentra en la figura
17.
Con el elemento plata, ocurre una situación similar a la del plomo pero en menor proporción.
iii) Alimentación vs cola de la prueba 2, flotación a ciclo abierto
De la misma forma, este análisis es más práctico y objetivo a través de un gráfico como el que
se muestra en la figura 16.
Informe Nº 09/09
30
Informe Nº 09/09
31
Se observa una situación un tanto similar a la prueba 2.
Estas pruebas se llevaron a cabo a través de un ciclonaje en dos etapas. Los resultados se
muestran a continuación
v) Alimentación vs cola de la prueba 4, flotación a ciclo abierto
efectuó el deslamado por ciclonaje, impidiendo llevar adelante el análisis correspondiente por
la eliminación de fracciones granulométricas que no ingresan a la prueba de flotación..
vi) Alimentación vs cola de la prueba 5, flotación a ciclo abierto
En este caso, también se realiza el análisis a través del gráfico que se encuentra en la figura
18.
En este gráfico se puede evidenciar las pérdidas de los elementos valiosos en las colas, estas
se incrementaron, en esta prueba; a pesar de ello se puede afirmar que estos resultados son
susceptibles de ser mejorados.
Cola P-1
Cola P-2
Cola P-3
Cola P-4*
Cola P-5
% Peso
% Peso
% Peso
% Peso
% Peso
Denominación
Over flow
(arcillas, lamas)
Under flow
20,53
21,16
20,13
11,37
79,47
78,84
79,87
88,63
Alimentación
100,00
100,00
100,00
100,00
20,21
79,79
100,00
De la tabla 15 se puede establecer que la presencia de lamas en las colas de flotación,
cuando no se procede al deslamado previa flotación, alcanzan a un valor que está próximo a
20.50% en peso y con previo deslamado a 13.37% en peso.
32
Informe Nº 09/09
Cola P-1
Cola P-2*
Cola P-3
Denominación
% Peso
% Peso
% Peso
21,90
11,63
20,70
Under flow
78,10
88,37
El tamaño de un espesador se caracteriza por su diámetro (área) y altura. La capacidad de
une espesador está determinada por su área; en cambio, la habilidad para producir una pulpa
de determinado contenido de sólidos esta dado por la altura. El área debe ser suficiente para
permitir que la partícula con la velocidad de sedimentación más lenta alcance el fondo del
tanque, ya que la velocidad de sedimentación de una partícula es diferente en cada zona.
79,30
Alimentación
100,00
100,00
100,00
La tabla 16, muestra valores un tanto similares; es decir, sin deslame las colas tienen un
21.3% en peso y con previo deslame, un 11.63% en peso.
Antes de mostrar los resaltados de las pruebas de sedimentación es necesario recordar
algunos conceptos básicos en forma preliminar.
La sedimentación es la técnica más usada en las operaciones de desaguado en el
procesamiento de minerales. Relativamente barata, de alta capacidad de procesamiento, la
cual incluye baja influencia de fuerzas de atrición, proporcionando buenas condiciones para la
floculación de partículas finas.
La sedimentación gravitacional puede separarse en dos grupos o técnicas de funcionamiento:
a) Técnica de espesamiento donde lo que interesa es un producto final de alta
densidad.
b) Técnica de clarificación, donde lo que interesa es obtener un rebalse de líquido
claro espesando los sólidos en un underflow sin importar su concentración.
En el espesamiento la obtención de un underflow, o producto espesado, interesa que su
concentración en sólidos sea lo más alta posible sin importar mucho la turbidez del liquido
que rebalse.
En la clarificación es de importancia que el rebalse del líquido sea lo más claro posible.
Los espesadores continuos consisten de un tanque cilíndrico cuyo diámetro fluctúa entre 2 y
200 metros, con una profundidad del orden de 1 a 7 metros. La pulpa es alimentada en el
centro del estanque vía Feed-well, a una profundidad de aproximadamente 1 metro, con el fin
de no producir turbulencia. El liquido clarificado rebalsa a una canaleta periférica, mientras el
sólido espesado es retirado por una salida central en el fondo del estanque, ayudado por un
sistema de bombeo y un sistema de rastra que gira lentamente arrastrando el sólido al centro.
Informe Nº 09/09
33
Informe Nº 09/09
34
En la práctica hay varias técnicas para calcular el área del espesador. Así se tienen: la
técnica de Coe-Clevenger, de Talmadge y Fitch y la técnica de Oltmann.
En esta oportunidad de acudió al método descrito por Tamaldge y Fitch. Se efectuaron varias
pruebas, con cola de flotación sin previo deslame y con cola de flotación previo deslame
antes de la operación de flotación; por otro lado, se efectuaron pruebas sin la adición de
floculante y con la adición de dos concentraciones de floculante, 20 g/t y 30 g/t. A
continuación se presentan los resultados de los mismos.
Para esta experimentación, se tomo la cola de la prueba 2 de flotación en ciclo abierto. Los
resultados alcanzados se muestran en las tablas 17 y 18, considerando un flujo de
Debe tomarse en cuenta que el método empleado sugiere que al diámetro calculado del
espesador debe incrementarse el 25%, es decir, si el diámetro del espesador es de 31 m,
para una alimentación de 25% sólidos, aproximadamente, y una descarga del 50% Sólidos,
en realidad el diámetro deberá ser 39 metros.
Si bien, el diámetro final es grande es el tamaño de estanque que garantizará la
sedimentación de todas las partículas en suspensión, se debe recordar que son partículas
muy finas y que al rededor de 50% en peso están por debajo de 38 micrones.
También es importante y necesario recordar que los espesadores convencionales tienen la
desventaja de poseer una gran área superficial y por ello es necesario una gran superficie
para su construcción; empero, es posible diseñar, construir y usar espesadores de alta
capacidad. Se caracterizan por una reducción del área de espesamiento desde un área
convencional instalada; el asentamiento de los sólidos en el manto de la pulpa se desliza
hacia abajo a lo largo de las placas inclinadas produciendo un “más rápido y más efectivo
espesamiento” que un descenso vertical.
Prueba Nº
Muestra
Informe Nº 09/09
1000
Apogee 625
Volumen pulpa, cm3
Colas de flotación
Peso de la pulpa, g
1194.589
35
1
2
3
4
5
6
7
8
9
10
11
12
13
14
15
16
17
18
19
20
21
22
23
24
25
0
90
150
480
600
960
1320
1500
2220
2520
3120
3420
4500
5400
6300
7200
8100
9900
10800
12600
13500
14400
16560
21840
86400
0,360
0,358
0,355
0,350
0,349
0,343
0,338
0,336
0,327
0,323
0,316
0,312
0,301
0,290
0,273
0,261
0,250
0,237
0,229
0,221
0,218
0,216
0,213
0,205
0,156
0,360
0,359
0,356
0,344
0,338
0,321
0,309
0,303
0,273
0,272
0,261
0,254
0,241
0,236
0,228
0,225
0,220
0,213
0,211
0,205
0,203
0,201
0,198
0,192
0,160
Para esta prueba, se tomo la cola de la prueba 4 de flotación en ciclo abierto. Los resultados
alcanzados se muestran en las siguientes tablas:
Tipo de floculante
Magnafloc 292
-100 Mallas Ty
Dosificación, g/t
0, 10 y 20
2.742
Fracción vol. iniacial 0,108
0,267
0
20
30
Nº
Tiempo, seg.
Altura H, m
Altura H, m
Altura H, m
Prueba Nº
Apogee 648
Volumen pulpa, cm3
1000
Muestra
Colas de flotación
Peso de la pulpa, g
1192.835
Tipo de floculante
Magnafloc 292
-100 Mallas Ty
Dosificación, g/t
0, 10 y 20
2.831
0,261
0
20
30
Nº
Tiempo, seg.
Altura H, m
Altura H, m
Altura H, m
0,360
0,356
0,355
0,347
0,343
0,333
0,320
0,310
0,288
0,280
0,265
0,256
0,247
0,238
0,229
0,219
0,211
0,203
0,198
0,193
0,192
0,187
0,185
0,179
0,165
1
2
3
4
5
6
7
8
9
10
11
12
13
14
15
16
17
18
19
20
21
22
23
24
25
0
20
30
3
Densidad del mineral seco, !s (g/cm )
2,742
2,742
2,742
0,267
0,267
0,267
1.143x10-5
2.003x10-5
2.873x10-5
2,012
1,064
0,875
1408,40
744,80
612,50
42,35
30,79
27,93
138,93
101,03
91,62
Uno de los más usados actualmente es el espesador Lamella. Utiliza un conjunto de placas
paralelas e inclinadas, las cuales reducen la distancia de sedimentación y al mismo tiempo
aumenta el área efectiva. El área efectiva que ocupa un Lamella es solamente del 20% de un
espesador convencional; las bandejas inclinadas permiten el asentamiento de los sólidos
deslizándose por gravedad dentro de la tolva.
0
90
150
480
600
960
1320
1500
2220
2520
3120
3420
4500
5400
6300
7200
8100
9900
10800
12600
13500
14400
16560
21840
86400
0.360
0.355
0.354
0.344
0.341
0.331
0.323
0.317
0.295
0.285
0.269
0.256
0.234
0.215
0.198
0.185
0.178
0.167
0.162
0.156
0.153
0.152
0.147
0.143
0.118
0.360
0.352
0.342
0.295
0.275
0.234
0.197
0.187
0.170
0.166
0.160
0.158
0.151
0.147
0.143
0.141
0.138
0.134
0.132
0.130
0.129
0.128
0.126
0.124
0.119
0.360
0.343
0.330
0.269
0.257
0.217
0.182
0.175
0.161
0.156
0.150
0.145
0.139
0.136
0.132
0.130
0.129
0.126
0.124
0.123
0.122
0.121
0.120
0.120
0.120
0
20
30
2,831
2,831
2,831
0,261
0,261
0,261
2.312x10-5
1.012x10-4
1.256x10-4
0.993
0.203
0.197
695.10
142.10
137.90
29.75
13.45
13.25
97.60
44.13
43.47
36
Informe Nº 09/09
deslame.
4.1.8 ENSAYO ESTANDAR DE BOND PARA DETERMINACION DEL WORK INDEX
El test estándar de Bond es el método más conocido y utilizado para predecir consumos de
energía en molienda de minerales. Esta predicción de consumo de energía se hace extensiva
a molinos de bolas y de barras.
El test de laboratorio elaborado por Fred Bond, consiste en una simulación de molienda
continua mediante un método que permite lograr estacionariedad a partir de sucesivos
ensayos “batch” .
La prueba entrega un valor para el índice de trabajo Wi, expresado en kWh/ton corta, el cual
introducido en la ecuación básica de la Tercera Ley de Conminución permite predecir el
consumo de energía de un molino de planta.
En general, se acepta que el error de predicción del consumo energético obtenido con este
ensayo es del orden de ! 20%.
37
Informe Nº 09/09
MUESTRA: LM
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
PRODUCTO
TYLER
A MOLINO
MOLIDO
"m
#
% Acumul.
6
3350
0
100,00
8
2360
10,86
89,14
10
1700
12.82
76.32
14
1180
12.49
63.83
20
850
10.58
53.25
28
600
8.19
45.06
35
425
7.39
37.67
48
300
5.94
31.73
65
212
4.82
26.91
0.00
100.00
100
150
3.37
23.54
33.95
66.05
150
106
3.88
19.65
12.98
53.07
200
75
2.90
16.75
9.84
43.22
-200
-75
16.75
0.00
43.22
0.00
Alimentación
100.00
RESUMEN RESULTADOS
F80 , ("m) = 1889.328(1)
Si bien es suficiente efectuar una prueba a una malla de corte determinado, en esta se
efectuaron pruebas a tres mallas de corte, por encima y por debajo de la principal que es 100
Mallas Tyler. Entonces, las mallas de corte estudiadas son: 65 Mallas Ty, 100 Mallas Ty y 150
Mallas Ty.
P80, ("m) =
175,477(2)
Gbpe, (g/rev) = 1.906(3)
Wi, (Kwh/tc) = 10.206(4)
wi
*
(Gbp )
La muestra corresponde a yacimiento primario de polisulfuros con poco contenido de sulfuros
en el que prevalece la pirita. La densidad real de la muestra, determinada por el método del
picnómetro, es de 2.784 g/cc. Por las características mostradas durante la preparación de la
muestra, se observa una mena caracterizada como “media dura”, con tendencia a formar
lamas por la presencia de una importante cantidad de arcillas.
wi
*
0.82
44.5
( 10
10 %
)
( p1 ) &
#
F80 #$
'& P80
0.23
44.5
%
( 10
10
(1.906) 0.82 x (212) 0.23 &
)
#
1889.33 $
' 175.48
* 10.20575 Kwh / tc
Siguiendo las recomendaciones efectuadas por el Sr. Bond y usando el equipo estándar
establecido, se efectuaron las pruebas a las mallas de corte preestablecidas. Los resultados
alcanzados en los ensayos son los siguientes:
Informe Nº 09/09
38
(4), por cálculo.
Informe Nº 09/09
39
en el molino, kWhh/tc
MUESTRA: LM
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
PRODUCTO
TYLER
A MOLINO
MOLIDO
"m
#
% Acumul.
6
3350
0
100,00
8
2360
10,86
89,14
10
1700
12.82
76.32
14
1180
12.49
63.83
20
850
10.58
53.25
28
600
8.19
45.06
35
425
7.39
37.67
48
300
5.94
31.73
65
212
4.82
26.91
100
150
3.37
23.54
0.00
100.00
150
106
3.88
19.65
31.01
68.99
200
75
2.90
16.75
13.91
55.07
-200
-75
16.75
0.00
55.07
0.00
Alimentación
100.00
100.00
MUESTRA: LM
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
PRODUCTO
TYLER
A MOLINO
MOLIDO
"m
#
%
% Retenido
% Acumul.
Acumul.Pas
6
3350
0
100,00
8
2360
10,86
89,14
10
1700
12.82
76.32
14
1180
12.49
63.83
20
850
10.58
53.25
28
600
8.19
45.06
35
425
7.39
37.67
48
300
5.94
31.73
65
212
4.82
26.91
100
150
3.37
23.54
150
106
3.88
19.65
0.00
100.00
200
75
2.90
16.75
28.24
71.76
-200
-75
16.75
0.00
71.76
0.00
Alimentación
100.00
100.00
RESUMEN RESULTADOS
RESUMEN RESULTADOS
F80 , ("m) = 1889.328(1)
F80 , ("m) = 1.889.328(1)
P80, ("m) =
P80, ("m) =
121.626(2)
84.047(2)
Gbpe, (g/rev) = 1.113(3)
Gbpe, (g/rev) = 1.515(3)
Wi, (Kwh/tc) = 12.844(4)
Wi, (Kwh/tc) = 11.091(4)
wi
wi
wi
*
44.5
( 10
10 %
(Gbp ) 0.82 ( p1 ) 0.23 &
)
#
F80 #$
'& P80
*
(1.515) 0.82
44.5
%
( 10
10
x (150) 0.23 &
)
#
1889.33 $
' 121.63
wi
* 11.09085 Kwh / tc
40
Informe Nº 09/09
*
Puesto que se efectuaron pruebas a diferentes mallas de corte es posible realizar un gráfico,
como el que se muestra a continuación, en el cual se puede ver más objetivamente la
variación del consumo de energía en función, precisamente, de la Malla de Corte.
*
44.5
( 10
10 %
(Gbp ) 0.82 ( p1 ) 0.23 &
)
#
F80 #$
'& P80
44.5
%
( 10
10
(1.113) 0.82 x (106) 0.23 &
)
#
1889.328 $
' 84.047
* 12.844 Kwh / tc
41
Informe Nº 09/09
Justamente, esta tendencia se puede observar en la tabla 8; en la que se muestra los
resultados de la flotación en ciclo cerrado. La recuperación del plomo mejora sustancialmente
y una situación similar debe ocurrir con el elemento plata.
En cuanto al elemento zinc, no hay mayores problemas en obtener un concentrado de alta ley
y con una elevada recuperación, esto se ha visto en la mayoría de las pruebas.
size, muestran que la mayor parte de las pérdidas de los elementos valiosos se encuentran
en los granos más finos y que se encuentra por debajo de la malla 400, -38 micrones.
Los reactivos usados son los que habitualmente se usan y los que se han mencionado en
este tipo de operaciones, salvo el fluosilicato de sodio que permitió dispersar y depresar una
parte de la ganga fina.
El contenido de lamas en las colas de flotación, cuando ésta se realiza sin previo deslame,
está alrededor de 20.50% en peso y cuando se realiza un previo deslamado, estas son del
orden del 13.37% en peso. Por otro lado, la velocidad de sedimentación de las colas cuando
no se realiza el deslame y cuando no se usan un floculante es del orden de 1.143 x 10-5 m/s y
esta mejora cuando se efectúa un previo deslame, la velocidad es de 2.312 x 10-5 m/s.
Figura 19.- Influencia de la Malla de Corte (65# = 212 µ, 100# = 150 µ, 150# = 106 µ)
en el Índice de Trabajo con la muestra LM, empresa Apogee
4.1.9 COMENTARIOS FINALES PARA LA MUESTRA LM
Si bien, el contenido de lamas es relativamente elevado, se puede afirmar que no influye en
forma determinante en la tarea de obtener concentrados finales de calidad y recuperaciones
elevadas; esta apreciación es tanto para pruebas de flotación en circuito abierto y para
pruebas en ciclo cerrado como se puede constatar observando los balances metalúrgicos de
la prueba 3, flotación en ciclo abierto, tabla 3 y prueba 3, flotación en ciclo cerrado, tabla 8.
En la tabla 3 se ve que el concentrado de plomo tienen una ley de plomo de 50% y una ley de
plata de 9240 g/t, mientras que el concentrado de zinc tiene una ley de zinc de 50.50% y una
ley de plata de 1875 g/t; la recuperación total de plata alcanza a 75.46%, mientras que la
recuperación de plomo llega a 58.04% y el de zinc a 83.55%.. Se debe hacer notar la
dificultad de lograr una mejor recuperación del plomo en vista de que si se prolonga el tiempo
de flotación rougher de plomo o si se aumenta, más de lo usado, más colector, empieza a
flotar mineral de zinc y pirita que es muy difícil su depresión en las etapas de limpieza.
También se debe remarcar que la distribución de plomo y plata en la mayor parte de las
espumas rougher tienen valores por encima de 70% y 50%, respectivamente, esta situación
implica que finalmente se elevará las recuperación final de estos elementos.
Informe Nº 09/09
42
MALLA DE CORTE
MALLAS TYLER
MICRONES
WORK INDEX
Kwh/tc
65
212
10.206
100
150
11.091
150
106
12.844
4.2 SEGUNDA PARTE: MUESTRA “LB”, BAJA LEY
Informe Nº 09/09
43
A continuación se presentarán los resultados en forma similar a los presentados para la
muestra LM.
representativos para la realización de las diferentes pruebas. Estos pasos se pueden seguir
observando el flujograma que se muestra en la figura 1.
La ley de cabeza ensayada de un común representativo de este mineral da el siguiente
resultado:
Plata: 46 g/t
Plomo: 0.79%
Zinc: 1.24%
Cobre: 0.026%
4.2.2
NF–2da Limp.de Pb
NF-1ra Limp. de Pb
Espuma rougher- Pb
Espuma de Zn-Ag
NF-2da Limp. de Zn
NF-1ra Limp. de Zn
Espuma rougher Zn
Non Float
Cabeza Calculada
1,26
5,78
7,87
1,61
2,85
6,65
11,11
81,02
100,00
5,24
2,04
5,17
2,81
0,69
0,44
0,85
0,26
0,71
9,28
16,58
57,17
6,34
2,77
4,11
13,22
29,61
100,00
278
171
294
578
38
27
110
11
44
7,92
22,35
52,36
20,99
2,45
4,06
27,50
20,15
100,00
2,81
2,22
2,81
43,50
1,05
0,81
7,04
0,16
1,13
3,13
11,33
19,50
61,66
2,64
4,75
69,06
11,44
100,00
En esta prueba la calidad de los productos finales son bajas y también son bajas las
recuperaciones; no se ha logrado una adecuada separación de los componentes
mineralógicos, debido, a la enorme cantidad de lamas presentes en la muestra.
Las condiciones de operación y consumo de reactivos se muestran en la figura 20; se
debieron efectuar dos flotaciones, en las mismas condiciones, para tener suficiente espuma
rougher y lleva r a la limpieza, especialmente espumas de Pb-Ag.
FLOTACION DIFERENCIAL DE SULFUROS
Prueba 1:
figura 2, sin previo deslame. Se realiza la flotación diferencial, flotando primero el mineral de
Pb-Ag, para ello se efectuó la molienda a -100 Mallas Tyler y en la etapa de
acondicionamiento se usó cal, como regulador de pH; sulfato de zinc como depresor del
mineral de zinc y cianuro de sodio para depresar las piritas que se encuentran en apreciable
cantidad en la muestra; como colector se usó el ditiofosfato Aero Float-242 y como
espumante se usó el MIBC. El grado de molienda, el colector y el espumante, en forma similar
a la muestra LM, fueron elegidos previas pruebas de flotación exploratorias.
Los resultados de esta primera prueba se los muestra en la tabla 21.
-
Tabla 21.- Balance metalúrgico de la prueba de flotación diferencial, prueba 1, muestra LB
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0,83
26,9
31,30
1180
22,08
6,91
5,05
44
Informe Nº 09/09
CLASIFICACION, 100#
(+)
D-500
pH: 8.8
Cal: 200 g/t
pH: 9.6
PRIMERA
Na2SiF6: 140 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 150 g/t
DE Pb-Ag
NaCN: 60 g/t
T. acond.: 5 min
T. Flot.: 4 min
LB, Empresa Apogee.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0,72
16,80
21,80
2170
35,30
20,3
12,52
NF–2da Limp.de Pb
0,52
14,45
13,58
1150
13,55 10,10
4,51
NF-1ra Limp. de Pb
2,34
2,08
8,70
155
8,13
2,83
5,63
Espuma rougher- Pb
3,59
6,87
44,08
708
56,98
7,42
22,65
Espuma de Zn-Ag
1,46
1,34
3,51
409
13,42 47,80
59,45
NF-2da Limp. de Zn
0,62
1,30
1,45
96
1,35
3,23
1,72
NF-1ra Limp. de Zn
2,41
0,44
1,90
30
1,62
0,92
1,89
Espuma rougher Zn
4,50
0,85
6,86
162
16,39 16,48
63,05
Non Float
65,27
0,24
28,06
10
14,66
0,09
5,00
Over flow (lamas)
26,65
0,44
21,00
20
11,97
0,41
9,30
Cola Total
91,92
0,30
49,06
13
26,62
0,18
14,30
Cabeza Calculada
100,00
0,56
100,00
45 100,00
1,18 100,00
Non Float
Espuma Zn-Ag
A pesar de que la mayor parte de los índices metalúrgicos mejoraron, los resultados todavía
no satisfacen, los índices metalúrgicos siguen siendo bajos; principalmente en lo que respecta
al elemento plomo.
D-1000
pH: 10.9
Cal: 1500 g/t
pH: 11.2
Na2SiF6: 200 g/t
PRIMERA
T. Acond.: 5 min
FLOTACION
CuSO4: 30 g/t
CLEANER
Z-11:
15 g/t
DE Zn-Ag
MIBC: 15 g/t
T. Acond.: 6 min
T. Flot.: 5 min
D-250
pH: 8.4
Cal: 300 g!t
pH: 9.5
Na2SiF6: 100 g/t
ZnSO4: 100 g/t
NaCN: 30 g/t
T. acond.: 5 min
T. Flot.: 1.5 min
SEGUNDA
FLOTACION
CLEANER
DE Pb-Ag
Espuma de Pb-Ag
Non Float
pH: 9.0
Cal: 5000 g/t
pH: 11.1
MIBC: 20 g/t
T. Flot.: 5 min
NF-1ra Limp. Ag
Espuma de Pb-Ag
NF-2da Limp. Ag
Espuma de Zn-Ag
-
D-500
pH: 10.7
FLOTACION pH: 11.1
T. Flot.: 3 min
Espuma de Zn-Ag
Informe Nº 09/09
figura 3. En esta prueba se incrementaron un tanto los reactivos y los tiempo de
acondicionamiento y flotación, también se incrementaron las etapas de limpieza. Los
resultados de esta segunda prueba se los muestra en la tabla 22 y las condiciones de
operación y consumo de reactivos en la figura 21.
MOLIENDA
agua
Espuma Pb-Ag
Prueba 2:
2000 g (se repite 4 veces)
(-)
D-2000
pH: 6.0
Cal: 5250 g/t
pH: 9.4
ZnSO4: 150 g/t
MIBC: 20 g/t
T. Acond.: 6 min
T. Flot.: 4 min
45
Informe Nº 09/09
46
Informe Nº 09/09
47
-
Prueba 3:
figura 2. En esta prueba se incrementaron un tanto los reactivos y los tiempos de
acondicionamiento y flotación. Los resultados de esta segunda prueba se los muestra en la
tabla 23 y las condiciones de operación y consumo de reactivos en la figura 22.
Los resultados de esta primera prueba se los muestra en la tabla 23.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
1.23
20.35
35.39
1488
40.25
7.45
7.94
NF–2da Limp.de Pb
1.73
5.06
12.37
297
11.29
3.62
5.42
NF-1ra Limp. de Pb
7.72
0.72
7.84
40
6.77
0.95
6.34
Espuma rougher- Pb
10.69
3.69
55.59
249
58.32
2.13
19.70
Espuma de Zn-Ag
1.47
3.52
7.27
673
21.64
48.50
61.43
NF-2da Limp. de Zn
1.91
1.66
4.48
87
3.65
3.91
6.46
NF-1ra Limp. de Zn
7.53
0.37
3.93
16
2.64
0.45
2.93
Espuma rougher Zn
10.91
1.02
15.68
117
27.93
7.52
70.82
Non Float
78.40
0.26
28.73
8
13.75
0.14
9.48
Cabeza Calculada
100.00
0.71
100.00
46
100.00
1.16
100.00
-
48
Informe Nº 09/09
CLASIFICACION, 100#
(+)
(-)
D-2000
pH: 6.4
Cal: 4500 g/t
pH: 9.4
ZnSO4: 175 g/t
MIBC: 20 g/t
T. Acond.: 6 min
T. Flot.: 5 min
D-500
pH: 8.6
Cal: 250 g/t
pH: 9.7
PRIMERA
Na2SiF6: 200 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 250 g/t
DE Pb-Ag
NaCN: 75 g/t
T. acond.: 6 min
T. Flot.: 3 min
Prueba 4:
LB, Empresa Apogee.
figura 3, con previo deslame. En esta prueba se incrementaron un tanto los reactivos y los
tiempos de acondicionamiento y flotación. Los resultados de esta segunda prueba se los
muestra en la tabla 24 y las condiciones de operación y consumo de reactivos en la figura 23.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0.99
38.00
48.57
2230
45.73
16.8
14.29
NF–2da Limp.de Pb
0.32
16.40
6.81
1330
8.86
11.10
3.07
NF-1ra Limp. de Pb
1.29
2.26
3.76
156
4.16
2.53
2.80
Espuma rougher- Pb
2.61
17.63
59.14
1091
58.75
9.03
20.16
Espuma de Zn-Ag
1.43
1.54
2.83
471
13.88
50.30
61.51
NF-2da Limp. de Zn
0.68
1.42
1.25
129
1.82
5.36
3.13
NF-1ra Limp. de Zn
2.06
0.62
1.64
38
1.62
0.87
1.54
Espuma rougher Zn
4.17
1.07
5.72
201
17.32
18.52
66.18
Non Float
68.57
0.24
21.18
9
12.74
0.06
3.52
Over flow (lamas)
24.65
0.44
13.96
22
11.19
0.48
10.13
Cola Total
93.22
0.29
35.14
12
23.93
0.17
13.66
Cabeza Calculada
100.00
0.78
100.00
48
100.00
1.17
100.00
Non Float
Espuma Zn-Ag
D-1000
pH: 10.8
Cal: 1500 g/t
pH: 11.1
Na2SiF6: 200 g/t
PRIMERA
T. Acond.: 5 min
FLOTACION
CuSO4: 40 g/t
CLEANER
MIBC: 20 g/t
T. Acond.: 8 min
T. Flot.: 5 min
D-250
pH: 8.3
FLOTACION pH: 9.5
NaCN: 50 g/t
T. acond.: 5 min
T. Flot.: 2 min
Espuma de Pb-Ag
Non Float
pH: 9.1
Cal: 6000 g/t
pH: 11.3
MIBC: 20 g/t
T. Flot.: 5 min
NF-1ra Limp. Ag
Espuma de Pb-Ag
NF-2da Limp. Ag
Espuma de Zn-Ag
son insatisfactorios, los índices metalúrgicos siguen siendo bajos; principalmente con
elemento plomo. Se continuaran con las próximas pruebas programadas.
-
D-500
pH: 10.7
FLOTACION pH: 11.1
T. Flot.: 3 min
Espuma de Zn-Ag
Informe Nº 09/09
MOLIENDA
agua
Espuma Pb-Ag
-
2000 g (se repite 3 veces)
49
Informe Nº 09/09
50
Informe Nº 09/09
51
-
Prueba 5:
figura 3, con previo deslame y tratando de mejorar la anterior prueba. Los resultados de esta
quinta prueba se los muestra en la tabla 25 y las condiciones de operación y consumo de
reactivos en la figura 24.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0.84
43.00
46.01
1940
40.67
20.70
13.65
NF–2da Limp.de Pb
0.45
13.35
7.69
838
9.46
11.50
4.08
NF-1ra Limp. de Pb
2.08
2.20
5.87
165
8.62
6.88
11.31
Espuma rougher- Pb
3.37
13.81
59.57
695
58.76
10.93
29.04
Espuma de Zn-Ag
1.17
1.12
1.68
331
9.72
52.00
48.01
NF-2da Limp. de Zn
0.63
0.83
0.67
95
1.50
8.47
4.21
NF-1ra Limp. de Zn
2.08
0.50
1.33
27
1.41
0.60
0.99
Espuma rougher Zn
3.88
0.74
3.68
130
12.63
17.37
53.21
Non Float
67.21
0.26
22.37
9
15.17
0.16
8.48
Over flow (lamas)
25.53
0.44
14.38
21
13.45
0.46
9.26
Cola Total
92.75
0.31
36.74
12
28.61
0.24
17.75
Cabeza Calculada
100.00
0.78
100.00
40
100.00
1.27
100.00
no son buenos, los índices metalúrgicos siguen siendo bajos; principalmente con elemento
plomo.
-
52
Informe Nº 09/09
(se repite 4 veces)
CICLONAJE
CLASIFICACION, 100#
(+)
(-)
Under flow
Over flow
(Lama)
D-2000
pH: 7.00
Cal: 2500 g/t
pH: 9.5
ZnSO4: 200 g/t
MIBC: 20 g/t
T. Acond.: 6 min
T. Flot.: 5 min
MOLIENDA
LB, Empresa Apogee.
agua
Espuma Pb-Ag
D-500
pH: 8.10
Cal: 1000 g/t
pH: 9.9
PRIMERA
Na2SiF6: 150 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 150 g/t
DE Pb-Ag
NaCN: 75 g/t
T. acond.: 7 min
T. Flot.: 3 min
-
2000 g
Non Float
pH: 9.1
Cal: 6500 g/t
pH: 11.1
MIBC: 20 g/t
T. Flot.: 6 min
53
Informe Nº 09/09
Las pruebas de flotación en ciclo cerrado se desarrollaron siguiendo los pasos que se
muestran en el flujograma de la figura 4 y en el flujograma de la figura 5.
Non Float
Espuma Zn-Ag
Prueba 1:
NF-1ra Limp. Ag
Espuma de Pb-Ag
D-1000
pH: 9.8
Cal: 2500 g/t
pH: 11.0
Na2SiF6: 200 g/t
PRIMERA
T. Acond.: 5 min
FLOTACION
CuSO4: 30 g/t
CLEANER
MIBC: 15 g/t
T. Acond.: 5 min
T. Flot.: 5 min
D-250
pH: 9.40
Cal: 0 g/t
SEGUNDA pH: 9.40
FLOTACION
CLEANER
DE Pb-Ag
Na2SiF6: 100 g/t
T. Acond.: 4 min
ZnSO4: 150 g/t
NaCN: 50 g/t
T. acond.: 8 min
T. Flot.: 2 min
ciclo cerrado, prueba 1, muestra LB
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0.53
46.70
34.71
2600
30.08
9.72
4.59
Espuma de Zn-Ag
1.28
4.32
7.73
749
20.85 44.80
50.96
Non Float
98.19
0.42
57.56
23
49.06
0.51
44.45
Cabeza Calculada
100.00
0.72
100.00
46 100.00
1.13 100.00
Espuma de Zn-Ag
Espuma de Pb-Ag
Esta prueba fue desarrollada sin deslame; los resultados que se alcanzaron están detallados
en el balance metalúrgico de la tabla 26.
NF-2da Limp. Ag
D-500
pH: 10.8
FLOTACION pH: 11.2
T. Flot.: 3 min
Espuma de Zn-Ag
-
La calidad del concentrado de plomo en cuanto al mismo plomo y plata son aceptables pero
las recuperaciones de estos dos elementos son bajas, se observa una alta distribución de los
mismos en las colas; por otro lado, el concentrado de zinc no llega a 50% y la recuperación
también es baja.Las condiciones de operación y consumo de reactivos, se detallan en el
flujograma de la figura 25. Los índices metalúrgicos que se logran, en estas condiciones de
operación, son:
Informe Nº 09/09
54
Informe Nº 09/09
55
CLASIFICACION, 100#
(+)
(-)
D-2000
pH: 7.2
Cal: 5000 g/t
pH: 9.9
ZnSO4: 175 g/t
MIBC: 20 g/t
T. Acond.: 6 min
T. Flot.: 4 min
MOLIENDA
flotación diferencial en ciclo
cerrado, Muestra LB, Apogee.
agua
Espuma Pb-Ag
Non Float
pH: 8.9
Cal: 5000 g/t
pH: 11.0
MIBC: 20 g/t
T. Flot.: 5 min
D-500
pH: 8.80
Cal: 500 g/t
pH: 9.6
PRIMERA
Na2SiF6: 200 g/t
FLOTACION
T. acond.: 5 min
CLEANER
ZnSO4: 225 g/t
DE Pb-Ag
NaCN: 100 g/t
T. acond.: 6 min
T. Flot.: 4 min
Prueba 2:
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0.80
16.75
18.25
3200
56.24 19.05
13.54
Espuma de Zn-Ag
1.48
1.41
2.83
440
14.26 48.70
63.84
Non Float
70.71
0.64
61.63
19
29.51
0.36
22.62
Over flow (lamas)
27.02
0.47
17.29
21
12.46
0.40
9.60
Total colas
97.72
0.59
78.92
20
41.97
0.37
32.22
Cabeza Calculada
100.00
0.73
100.00
46 100.00
1.13 100.00
NF-1ra Limp. Ag
Espuma de Pb-Ag
D-1000
pH: 10.2
Cal: 2000 g/t
pH: 11.3
Na2SiF6: 200 g/t
PRIMERA
T. Acond.: 5 min
FLOTACION
CuSO4: 50 g/t
CLEANER
MIBC: 10 g/t
T. Acond.: 5 min
T. Flot.: 6 min
D-250
pH: 8.4
Cal: 300 g/t
pH: 9.5
SEGUNDA
NaCN: 75 g/t
Los resultados no son buenos ni en ley menos en recuperación especialmente con el
elemento plomo.
T. acond.: 6 min
T. Flot.: 2 min
Espuma de Zn-Ag
Espuma de Pb-Ag
NF-2da Limp. Ag
D-500
pH: 10.3
Cal: 1500 g/t
pH: 11.2
Na2SiF6: 150 g/t
T. Acond.: 6 min
T. Flot.: 4 min
SEGUNDA
FLOTACION
CLEANER
DE Zn-Ag
Espuma de Zn-Ag
-
Esta prueba fue desarrollada previo deslame, según la figura 5; los resultados que se
alcanzaron están detallados en el balance metalúrgico de la tabla 27.
Non Float
Espuma Zn-Ag
-
(se repite 3 veces) 2000 g
-
Informe Nº 09/09
56
57
Informe Nº 09/09
-
Radio de concentración del plomo: 125
Prueba 3:
Esta prueba fue desarrollada previo deslame, figura 5; los resultados que se alcanzaron están
detallados en el balance metalúrgico de la tabla 28.
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
Espuma de Pb-Ag
0.87
50.50
62.12
3390
67.53 19.65
14.85
Espuma de Zn-Ag
1.55
1.24
2.03
318
8.45 53.30
72.03
Non Float
71.76
0.27
27.44
11
18.10
0.21
13.12
Over flow (lamas)
25.82
0.23
8.41
10
5.92
0.20
4.49
Total colas
97.58
0.26
35.85
11
24.03
0.21
17.61
Cabeza Calculada
100.00
0.71
100.00
45 100.00
1.15 100.00
Los resultados mejoran considerablemente en ley y en recuperación en los tres elementos
valiosos.
figura 27.
-
-
Informe Nº 09/09
58
Informe Nº 09/09
59
-
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
8.92
12.71
8.56
9.28
6.38
54.15
100.00
PLOMO
% Pb
% Dist.
0.211
7.32
0.186
9.20
0.186
6.19
0.154
5.56
0.207
5.14
0.316
66.58
100.00
0.26
PLATA
g/t Ag % Dist.
17
15.39
11
14.19
11
9.55
9
8.48
13
8.42
8
43.97
9.9 100.00
ZINC
% Zn % Dist.
0.557
33.65
0.296
25.47
0.145
8.40
0.072
4.53
0.078
3.37
0.067
24.57
0.15 100.00
-
60
Informe Nº 09/09
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
10.05
13.53
16.40
7.66
10.90
41.45
100.00
PLOMO
% Pb
% Dist.
0.162
7.01
0.169
9.84
0.160
11.29
0.184
6.06
0.205
9.62
0.315
56.18
100.00
0.23
PLATA
g/t Ag % Dist.
9
9.54
9
12.85
9
15.57
11
8.88
12
13.80
9
39.35
9.5 100.00
ZINC
% Zn % Dist.
0.272
32.43
0.099
15.89
0.065
12.64
0.048
4.36
0.044
5.69
0.059
29.00
0.08 100.00
61
Informe Nº 09/09
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
10.17
13.28
10.08
9.66
7.90
48.91
100.00
PLOMO
% Pb
% Dist.
0.179
7.04
0.176
9.04
0.189
7.37
0.178
6.65
0.220
6.72
0.334
63.18
100.00
0.26
PLATA
g/t Ag % Dist.
8
9.66
7
11.03
9
10.77
7
8.03
15
14.06
8
46.44
8.4 100.00
ZINC
% Zn % Dist.
0.466
40.73
0.195
22.26
0.075
6.50
0.053
4.40
0.044
2.99
0.055
23.12
0.12 100.00
En forma similar a lo que ocurrió con la muestra LM, en ésta en esta también se encuentran
las mayores distribuciones de los elementos valiosos por debajo de la malla 400.
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
9.11
10.20
10.54
6.41
8.09
55.65
100.00
PLOMO
% Pb
% Dist.
0.22
7.67
0.235
9.18
0.197
7.95
0.181
4.44
0.208
6.44
0.302
64.32
100.00
0.26
PLATA
g/t Ag % Dist.
12
12.00
12
13.45
9
10.42
9
6.34
10
8.89
8
48.90
9.1 100.00
ZINC
% Zn % Dist.
0.166
11.45
0.100
7.73
0.075
5.99
0.107
5.19
0.070
4.29
0.155
65.34
0.13 100.00
Para realizar este diagnóstico es necesario contar con el análisis granulométrico de la
alimentación a la flotación rougher.
i)
Peso
PLOMO
PLATA
ZINC
Productos
%
% Pb
+150#
6.33
0.432
3.78
38
5.52
1.24
6.89
-150# +200#
8.22
0.636
7.22
54
10.19
1.57
11.30
-200# +270#
12.46
0.674
11.62
52
14.89
1.57
17.14
-270# +325#
3.78
0.883
4.61
70
6.07
2.02
6.70
-325# +400#
7.27
1.470
14.78
98
16.37
2.37
15.14
-400#
61.95
0.677
57.99
33
46.96
0.79
42.83
Cabeza calculada
100.00
100.00
100.00
100.00
0.72
43.5
1.14
Productos
+150#
-150# +200#
-200# +270#
-270# +325#
-325# +400#
-400#
Cabeza calculada
Peso
%
11.28
14.03
12.43
10.83
10.04
41.39
100.00
PLOMO
% Pb
% Dist.
0.164
7.59
0.167
9.61
0.174
8.87
0.155
6.89
0.210
8.64
0.344
58.39
100.00
0.24
PLATA
g/t Ag % Dist.
10
9.74
10
12.11
12
12.88
9
8.42
12
10.40
13
46.45
11.6 100.00
Análisis granulométrico de la alimentación
ZINC
% Zn % Dist.
0.163
28.65
0.085
18.59
0.048
9.30
0.033
5.57
0.032
5.00
0.051
32.89
0.06 100.00
Este análisis granulométrico, a través del % Peso, permite calcular el d80 del producto molido;
para ello se tiene el siguiente gráfico:
Informe Nº 09/09
62
Informe Nº 09/09
63
Este gráfico permite observar en forma clara que el elemento plomo no se recupera bien en
los granos gruesos y tampoco se recupera adecuadamente lo que se encuentra por debajo de
la malla 400; solo es posible una adecuada recuperación los tamaños de grano que están por
encima de 400 Mallas y por debajo de 270 Mallas Ty. Una situación más pronunciada e
inaceptable se da con el elemento zinc; la recuperación de este elemento en las fracciones
gruesas es pésima, especialmente por encima de 200 Mallas Ty. Esta situación debe ser
corregida en siguientes pruebas.
100.00
P es o!P as ante!Ac um ulado,!%
90.00
80.00
70.00
60.00
50.00
40.00
30.00
ii)
20.00
10.00
0.00
0
20
40
60
80
100
Alimentación vs cola de la prueba 2, flotación a ciclo abierto
efectuó el deslamado por ciclonaje, eliminando la posibilidad de llevar adelante el análisis
correspondiente.
120
T a m a ño!de !g ra no,!Mic rone s
iii)
Entonces, el d80 es igual a 64 micrones y el d50 es igual a 30 micrones
Este análisis granulométrico, también permitirá calcular las leyes de cada tamaño de grano de
las alimentaciones con las que se efectuó cada prueba de flotación y con la cola de cada
prueba se efectúa el análisis size by size.
ii)
figura 30.
figura 29.
Las pérdidas de los elementos valiosos, en esta prueba, disminuyen en forma considerable.
iv)
64
Informe Nº 09/09
No se puede efectuar este análisis en virtud de que la prueba de flotación se efectuó previo
deslame de la muestra.
v)
No se puede efectuar este análisis en virtud de que la prueba de flotación se efectuó previo
deslame de la muestra.
Estas pruebas se llevaron a cabo a través de un ciclonaje en dos etapas. Los resultados se
muestran a continuación
Cola P-1
Cola P-2*
Cola P-3
Cola P-4*
Cola P-5*
% Peso
% Peso
% Peso
% Peso
% Peso
24,92
13,82
25,55
13,04
13,55
75,08
86,18
74,45
86,96
86,45
Alimentación
100,00
100,00
100,00
100,00
100,00
De la tabla 25, se conoce que un 25.24% e peso, aproximadamente, son lamas cuando la
muestra que se flota no se deslamo previamente y un 13.47% en peso cuando previamente a
la flotación se deslamó.
Cola P-1
Cola P-2*
Cola P-3*
Denominación
% Peso
% Peso
% Peso
26,83
13,82
14,07
Under flow
73,17
86,18
Prueba Nº
Apogee 655
Volumen pulpa, cm3
1000
Muestra
Colas de flotación
Peso de la pulpa, g
1186.329
Tipo de floculante
Magnafloc 292
-100 Mallas Ty
Dosificación, g/t
0, 10 y 20
2.691
0,271
0
20
30
Nº
Tiempo, seg.
Altura H, m
Altura H, m
Altura H, m
1
2
3
4
5
6
7
8
9
10
11
12
13
14
15
16
17
18
19
20
21
22
23
24
25
0
90
150
480
600
960
1320
1500
2220
2520
3120
3420
4500
5400
6300
7200
8100
9900
10800
12600
13500
14400
16560
21840
86400
0.360
0.359
0.358
0.357
0.355
0.350
0.346
0.344
0.332
0.330
0.326
0.323
0.312
0.302
0.294
0.284
0.276
0.263
0.254
0.244
0.239
0.236
0.221
0.211
0.178
0.360
0.359
0.357
0.354
0.352
0.346
0.340
0.337
0.325
0.320
0.314
0.310
0.298
0.284
0.270
0.262
0.247
0.228
0.224
0.216
0.215
0.214
0.203
0.198
0.185
0.360
0.359
0.356
0.351
0.347
0.342
0.336
0.331
0.319
0.314
0.298
0.293
0.278
0.265
0.251
0.241
0.228
0.211
0.207
0.201
0.199
0.197
0.195
0.194
0.193
85,93
Alimentación
100,00
100,00
100,00
La tabla 26, muestra que el 26.83% en peso corresponde a las lamas cuando no se realiza un
deslame antes de la flotación y un 13.95% cuando esta deslamado.
Informe Nº 09/09
65
Para estas pruebas, se tomo la cola de la prueba 1 de flotación en ciclo abierto. Los
resultados alcanzados se muestran en las siguientes tablas, considerando un flujo de
Denominación
Over flow
(arcillas, lamas)
Under flow
Informe Nº 09/09
66
0
20
30
2,691
2,691
2,691
0,271
0,271
0,271
9.892x10-6
1.423x10-5
1.856x10-5
2,536
1,756
1,501
1775,20
1229,20
1050,70
47,54
39,56
36,58
155,98
129,79
120,00
Informe Nº 09/09
67
Finalmente debe tomarse en cuenta que el método empleado sugiere que al diámetro
calculado del espesador debe incrementarse el 25%, es decir, si el diámetro del espesador es
de 48 m, para una alimentación de 25% sólidos, aproximadamente, y una descarga del 50%
Sólidos, en realidad el diámetro deberá ser 60 metros.
0,266
1.583x10-5
1.538
1076.60
37.02
121.47
0,266
1.982x10-5
1.180
826.00
32.43
106.40
0,266
2.913x10-5
0.844
590.80
27.43
89.98
deslame.
Para estas pruebas, se tomo la cola de la prueba 2 de flotación en ciclo abierto. Los
resultados alcanzados se muestran en las siguientes tablas:
4.2.8 DETERMINACION DEL WORK INDEX
Prueba Nº
Apogee 664
Volumen pulpa, cm3
1000
Muestra
Colas de flotación
Peso de la pulpa, g
1189.655
Tipo de floculante
Magnafloc 292
-100 Mallas Ty
Dosificación, g/t
0, 10 y 20
2.760
0,266
0
20
30
Nº
Tiempo, seg.
Altura H, m
Altura H, m
Altura H, m
1
2
3
4
5
6
7
8
9
10
11
12
13
14
15
16
17
18
19
20
21
22
23
24
25
0
90
150
480
600
960
1320
1500
2220
2520
3120
3420
4500
5400
6300
7200
8100
9900
10800
12600
13500
14400
16560
21840
86400
0.360
0.357
0.355
0.350
0.348
0.343
0.338
0.335
0.324
0.320
0.310
0.306
0.278
0.265
0.252
0.241
0.227
0.210
0.198
0.182
0.176
0.173
0.163
0.152
0.130
0.360
0.356
0.352
0.343
0.339
0.332
0.325
0.322
0.310
0.302
0.292
0.280
0.255
0.238
0.221
0.210
0.193
0.175
0.165
0.155
0.153
0.150
0.145
0.137
0.133
La muestra corresponde a yacimiento primario de polisulfuros con poco contenido de sulfuros
en el que prevalece la pirita. La densidad real de la muestra, determinada por el método del
picnómetro, es de 2.712 g/cc. Por las características mostradas durante la preparación de la
muestra, se observa una mena caracterizada como “blanda”, con tendencia a formar lamas
por la presencia de una importante cantidad de arcillas.
0.360
0.355
0.351
0.338
0.328
0.322
0.311
0.302
0.272
0.260
0.249
0.238
0.211
0.196
0.180
0.168
0.160
0.153
0.149
0.144
0.143
0.141
0.139
0.138
0.137
MUESTRA: LB
GRANULOMETRIAS
TAMAÑO
ALIMENTACION
PRODUCTO
MALLA
TYLER
A MOLINO
MOLIDO
"m
#
% Acumul.
6
3350
0
100,00
8
2360
8.09
91.91
10
1700
9.97
81.94
14
1180
8.58
73.36
20
850
8.13
65.23
28
600
6.65
58.58
35
425
6.10
52.48
48
300
5.51
46.97
65
212
4.66
42.31
0.00
100.00
100
150
3.57
38.74
19.30
80.70
150
106
3.52
35.22
16.77
63.92
200
75
3.82
31.40
10.13
53.80
-200
-75
31.40
0.00
53.80
0.00
Alimentación
100.00
100.00
0
20
30
2,760
2,760
2,760
68
Informe Nº 09/09
RESUMEN RESULTADOS
200
75
-200
-75
Alimentación
F80 , ("m) = 1582.173(1)
P80, ("m) =
69
Informe Nº 09/09
(2)
148.174
Gbpe, (g/rev) = 1.802(3)
3.82
31.40
100.00
(Gbp ) 0.82
wi
*
44.5
( 10
10 %
( p1 ) 0.23 &
)
#
F80 #$
'& P80
44.5
%
( 10
10
(1.802) 0.82 x (212) 0.23 &
)
#
1582.17 $
' 148.174
58.09
0.00
F80 , ("m) = 1582.173(1)
P80, ("m) =
*
16.34
58.09
100.00
RESUMEN RESULTADOS
Wi, (Kwh/tc) = 9.8233(4)
wi
31.40
0.00
115.58(2)
Gbpe, (g/rev) = 1.523(3)
Wi, (Kwh/tc) = 10.775(4)
wi
* 9.8233 Kwh / tc
(4), por cálculo.
wi
en el molino, Kwh/tc
*
*
44.5
( 10
10 %
(Gbp ) 0.82 ( p1 ) 0.23 &
)
#
F80 #$
'& P80
44.5
%
( 10
10
(1.523) 0.82 x (150) 0.23 &
)
#
1582.173 $
' 115.58
* 10.77538 Kwh / tc
MUESTRA: LB
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
PRODUCTO
TYLER
A MOLINO
MOLIDO
"m
#
% Acumul.
6
3350
0
100,00
8
2360
8.09
91.91
10
1700
9.97
81.94
14
1180
8.58
73.36
20
850
8.13
65.23
28
600
6.65
58.58
35
425
6.10
52.48
48
300
5.51
46.97
65
212
4.66
42.31
100
150
3.57
38.74
0.00
100.00
150
106
3.52
35.22
25.57
74.43
Informe Nº 09/09
70
MUESTRA: LB
GRANULOMETRIAS
MALLA
TAMAÑO
ALIMENTACION
PRODUCTO
TYLER
A MOLINO
MOLIDO
"m
#
%
% Retenido
% Acumul.
Acumul.Pas
6
3350
0
100,00
8
2360
8.09
91.91
10
1700
9.97
81.94
14
1180
8.58
73.36
20
850
8.13
65.23
28
600
6.65
58.58
35
425
6.10
52.48
48
300
5.51
46.97
65
212
4.66
42.31
100
150
3.57
38.74
150
106
3.52
35.22
0.00
100.00
200
75
3.82
31.40
21.60
78.40
Informe Nº 09/09
71
-200
-75
Alimentación
31.40
100.00
0.00
78.40
100.00
F80 , ("m) = 1582.173(1)
77.296(2)
Gbpe, (g/rev) = 1.228(3)
Wi, (Kwh/tc) = 11.376(4)
wi
*
(Gbp )
*
wi
0.82
Las pruebas de flotación diferencial, con esta muestra, arrojaron resultados por debajo de las
expectativas, aunque se puede afirmar que para encarar una flotación diferencial a nivel
industrial, necesariamente debe deslamarse la carga, puesto que las lamas perjudican
enormemente en el proceso.
RESUMEN RESULTADOS
P80, ("m) =
0.00
44.5
( 10
10 %
)
( p1 ) &
#
F80 $#
'& P80
0.23
44.5
( 10
%
10
(1.228) 0.82 x (106) 0.23 &
)
#
1582.173 $
' 77.296
* 11.3764 Kwh / tc
Las mejores pruebas tanto en circuito abierto como en ciclo cerrado fueron obtenidas previo
deslamado; en efecto, la tabla 25 muestra los resultados de la prueba de flotación Nº 5, en
circuito abierto y se puede ver un concentrado de plomo con 43% Pb y 1940 g/t Ag y el
concentrado de zinc tiene una ley de 52% Zn, con recuperaciones que están en el orden de
46.01% para el plomo, 50.39% para la plata (sumando las recuperaciones del concentrado de
plomo y del concentrado de zinc) y 48.01% para el zinc; estos valores son bajos, comparando
por ejemplo con los obtenidos con la muestra LM; pero, en un hecho bastante positivo, los
resultados de la prueba 3 de flotación en ciclo cerrado, tabla 28, muestran índices
metalúrgicos bastante más alentadores ya que se logra obtener un concentrado de plomo
con una ley de 50.5%Pb y 3390 g/t Ag , mientras que el concentrado de zinc tiene una ley de
53.30% Zn, en cuanto a las recuperaciones se refiere se puede indicar que el plomo alcanza
a 62.12%, 75.98% para la plata y 72.03% para el zinc.
Puesto que se efectuaron pruebas a diferentes mallas de corte es posible realizar un gráfico,
como el que se muestra a continuación, en el cual se puede ver más objetivamente la
variación del consumo de energía en función, precisamente, de la Malla de Corte.
size, muestran que la mayor parte de las pérdidas de los elementos valiosos, en forma similar
lo que ocurre con la muestra LM, se encuentran en los granos más finos y que están por
debajo de la malla 400, -38 micrones.
Work!Index ,!Mues tra!L B
El contenido de lamas en las colas de flotación, cuando ésta se realiza sin previo deslame,
está alrededor de 25.24% en peso y cuando se realiza un deslamado previo, estas son del
orden del 13.47% en peso. Por otro lado, la velocidad de sedimentación de las colas cuando
no se realiza el deslame y cuando no se usan un floculante es del orden de 9.892 x 10-6 m/s y
esta mejora cuando se efectúa un previo deslame, la velocidad es de 2.312 x 10-5 m/s.
Work!Index,!K wh/tc
12
10
8
6
4
2
MALLA DE CORTE
0
0
50
100
150
200
MICRONES
65
212
9.823
100
150
10.775
150
106
11.376
Ma lla !de !c orte ,!Mic rone s
Figura 31- Influencia de la Malla de Corte (65# = 212 µ, 100# = 150 µ, 150# = 106 µ)
en el Índice de Trabajo con la muestra LM, empresa Apogee
De la figura 31 se puede colegir que a medida que disminuye la malla de corte el molino
requiere de un mayor consumo de energía por tonelada de mineral tratado
WORK INDEX
Kwh/tc
MALLAS TYLER
250
5. CONCLUSIONES
4.2.9 COMENTARIOS FINALES PARA LA MUESTRA LB
Informe Nº 09/09
72
Del análisis de resultados obtenidos y de las observaciones durante las pruebas
experimentales, se puede concluir lo siguiente:
Informe Nº 09/09
-
MUESTRA LB
-
-
-
-
-
Estos resultados muestran claramente la posibilidad de, obtener similares o
incluso, mejores resultados en una operación industrial.
-
Si bien la presencia de lamas es grande y perjudicial, alrededor del 20% en peso,
se ha podido demostrar que nos es necesario un deslame previo a la flotación,
posibilitando una mejor recuperación de todos los elementos valiosos.
-
-
-
recuperación ha arrojado en todas las pruebas de flotación, porque se torna muy
difícil la separación de otros sulfuros como el propio mineral de zinc y sulfuros de
hierro.
afirmar que la mayor pérdida de los elementos valiosos en las colas se encuentran
en tamaños de fina granulometría, concretamente por debajo de 400 Mallas Tyler,
-38 micrones.
La velocidad de sedimentación de las partículas, a partir de las colas de flotación,
es lenta porque algo más del 50% en peso de la muestra que entra al proceso de
flotación está por debajo de la malla 400 y gran parte de esta fracción corresponde
a la presencia de lamas; esta velocidad es de 1.143 x 10-5 m/s.
Informe Nº 09/09
74
-
El Índice de Trabajo, Work Index, para una malla de corte de 100 Mallas Tyler, 150
micrones, es de 11.091 Kwh/tc.
MUESTRA LM
La muestra es apta de ser tratada por el proceso de flotación diferencial ya que se
logran obtener concentrados con índices metalúrgicos bastante aceptables; esta
situación se ha visto en las pruebas tanto en circuito abierto como en ciclo cerrado.
.
En circuito abierto se han logrado estos resultados: ley de plata en el concentrado
de plomo, 9240 g/t y en el concentrado de zinc, 1875 g/t; ley de plomo en el
concentrado de plomo, 50% y el concentrado de zinc alcanza una ley de 50.50%
Zn, con las siguientes recuperaciones: plata, 75.46%; plomo, 58.04% y zinc,
83.55%..
En circuito cerrado, estos resultados son todavía mejores: ley de plata en el
concentrado de plomo, 6220 g/t; en el concentrado de zinc, 2990 g/t; ley de plomo
en el concentrado de plomo, 51.0% y la ley del zinc en su concentrado es de
58.30% y las recuperaciones son: plata, 83.33%; plomo, 74.32% y zinc, 82.60%.
73
Esta muestra también es apta de ser tratada por el proceso de flotación diferencial
aunque los índices metalúrgico que se logran son un tanto menores, comparando
con la muestra LM, y que es preciso un deslame previo; ya que de otra manera es
difícil lograr concentrados finales. esta situación se ha visto solo en la prueba en
ciclo cerrado y no así en circuito abierto.
.
En circuito cerrado se han logrado estos resultados: ley de plata en el concentrado
de plomo, 3390 g/t y en el concentrado de zinc, 318 g/t; ley de plomo en el
concentrado de plomo, 50.50% y el concentrado de zinc alcanza una ley de
53.30% Zn, con las siguientes recuperaciones: plata, 75.98%; plomo, 62.12% y
zinc, 72.03%.
-
Estos resultados, muestran claramente la tendencia de, incluso, obtener mejores
resultados en una operación industrial.
-
La presencia de lamas es enorme y perjudicial, alrededor del 25% en peso, por lo
que se ha demostrado que no es posible una adecuada flotación si no se efectúa
previamente un deslamado.
-
recuperación ha arrojado en todas las pruebas de flotación, porque se torna muy
difícil la separación de otros sulfuros como el propio mineral de zinc y sulfuros de
hierro, esta situación similar también se observó con la muestra LM.
-
afirmar que la mayor pérdida de los elementos valiosos en las colas se encuentran
en tamaños de fina granulometría, concretamente por debajo de 400 Mallas Tyler,
-38 micrones.
-
La velocidad de sedimentación de las partículas, sin la adición de floculante, a
partir de las colas de flotación, es lenta porque algo más del 50% en peso de la
muestra que entra al proceso de flotación está por debajo de la malla 400 y gran
parte de esta fracción corresponde a la presencia de lamas; esta velocidad es de
19.982 x 10-6 m/s.
-
El Indice de Trabajo, Work Index, para una malla de corte de 100 Mallas Tyler, 150
micrones, es de 10.775 Kwh/tc.
Informe Nº 09/09
75
Informe Nº 09/09
76
APPENDIX V
EPCM REPORTS AND ESTIMATES
169

apogee minerals ltd. technical report on the preliminary assessment

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