Composition, formation, and leaching behaviour of ... · 3.2.4 Chemical analysis (XRF, EMPA,...

1

Composition, formation, and leaching behaviour of supergene,

polymetallic ores from the Sanyati deposit (Zimbabwe):

A case study

vorgelegt von

Diplom-Geologin

Martina Frei

aus München

von der Fakultät VI

der Technischen Universität Berlin

zur Erlangung des akademischen Grades

Doktor der Naturwissenschaften

- Dr. rer. nat. -

genehmigte Dissertation

Promotionsausschuss:

Vorsitzender: Prof. Dr. W. Dominik Berichter: Prof. Dr. K. Germann Berichter: Prof. Dr. G. Franz Berichter: Prof. Dr.-Ing. Dr. rer.nat. h.c.mult. F.-W. Wellmer Tag der wissenschaftlichen Aussprache: 17.8.2005

Berlin 2005

D83

2

"Mineralogy is the key to applying SX/EW techniques,

and the original ore mineralogy of a deposit determines

the ultimate recovery from the leach process."

(WALLIS & CHLUMSKY 1999)

3

Table of contents

page

Abstract 6

Zusammenfassung 7

1. Introduction 8

1.1 Object and aims of this study 8

1.2 Hydrometallurgical base metal production from leachable deposits 11

1.2.1 Supergene deposits, recovery techniques, economic importance: an overview 11

1.2.2 HL-SX-EW, a hydrometallurgical route for copper recovery 19

1.2.2.1 Pyro- vs hydrometallurgical route – advantages and restrictions 19

1.2.2.2 HL-SX-EW process for copper recovery 20

2. The Sanyati copper ore deposit 24

2.1 Geographical situation 24

2.2 Geological and morphological setting 24

2.2.1 Regional structural situation 24

2.2.2 Lithostratigraphy 27

2.2.3 Morphological situation 30

2.3 Distribution and composition of the primary mineralisation 31

2.4 Distribution and composition of the supergene mineralisation 33

2.4.1 Terminology of supergene mineralisations 33

2.4.2 The Sanyati supergene mineralisation 34

2.5 Mining and benefication activities at the Sanyati mine 35

2.5.1 Historical development 35

2.5.2 Mining process 36

2.6 HL-SX-EW at the Sanyati mine 37

3 Field- and laboratory work 40

3.1 Fieldwork 40

3.1.1 Sampling of orebodies and run-of-mine ore dump 40

3.1.2 Sampling of host rock 41

3.1.3 Sampling of the heap leach pad 42

3.2 Analytical and experimental methods 43

3.2.1 Sample preparation 43

3.2.2 Leaching experiments 44

3.2.2.1 Leaching experiments of the water-soluble fraction (V0) 45

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page

3.2.2.2 Leaching experiments of the H2SO4-soluble fraction (VR, V15, V7, V1, V6) 46

3.2.2.3 Partial extraction experiments (V12 - V14) 48

3.2.2.4 Experimental adsorption of base metals to goethite (experiment V8 - V11) 48

3.2.3 Phase analysis (Pol. Mic., XRD, SEM) 49

3.2.4 Chemical analysis (XRF, EMPA, LA-ICPMS, ICP OES, F-AAS, and G-AAS) 50

4 Weathering products and processes at Sanyati 56

4.1 Weathering products of host rocks with (proximal) and without (distal) the influence of sulfide decay

56

4.1.1 Mineralogical and geochemical composition of fresh host rock 56

4.1.2 Mineralogical characteristics of weathered host rock distal and proximal to the zone of sulfide decay

57

4.1.3 Geochemical characteristics of weathered host rocks distal and proximal to the zone of sulfide decay

58

4.1.3.1 Geochemical changes during weathering processes 58

4.1.3.2 Metal signature 64

4.2 Formation of supergene ore in the oxidation zone 66

4.2.1 Breakdown reactions of primary sulfide ore 66

4.2.2 Composition of the secondary sulfide ore 72

4.2.3 Composition of the supergene ore 73

4.2.3.1 Mineralogical characteristics 73

4.2.3.2 Geochemical characteristics 79

4.2.4 Composition of groundwater in the open pits 87

4.2.5 Neoformation of sulfates in the open pits 87

4.3 Summary Chapter 4 88

5 Leaching products and processes on the heap leach pads 91

5.1 Characteristics of the run-of-mine ore (ROM) 91

5.2 Composition of the leach pad ore (LPO) 94

5.3 Composition of the acid solution used for leaching 99

5.4 Neoformation of phases during leaching 100


6 Composition of goethite and hematite in run-of-mine and leach pad ore 107

6.1 Chrystal chemistry of goethite and hematite 107

6.2 Boxwork texture and chemistry (Microprobe analysis) 110

6.2.1 Development, preservation, classification and geochemistry of relictic decay textures of sulfides

110

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page

6.2.2 Element distribution in goethite-rich and hematite-rich zones of boxwork textures

126

6.2.3 Concentrations in run-of-mine ore (ROM) and leach pad ore (LPO) 133

6.3 Trace element chemistry of goethite- and hematite-rich zones (Laser ablation analysis)

134

6.4 Comparison of element distributions of goethite and hematite in three base metal and lead deposits

136


7 Laboratory leaching behaviour of the supergene ore 140

7.1 Leaching experiments with ROM and LPO (H2SO4-soluble fraction) 140

7.1.1 Percolation experiments 140

7.1.2 Leaching experiments under idealized conditions 145

7.1.3 Metal production rates and rate equations 148

7.2 Extraction experiments on goethite- and hematite-rich supergene ores 151

7.3 Adsorption experiments of Cu onto goethite 153


8 Formation of “invisible” base metal concentrations in supergene goethite and hematite and their consequences for the leaching process

157

8.1 Development of the oxidation zone in Sanyati and the composition of the supergene ores

157

8.2 Distribution of base metals in sulfide decay textures - colloform textures as a proxy for the "invisible" base metal contents in goethite and hematite

159

8.3 Fixation of base metals by goethite and hematite 162

8.3.1 Adsorption to goethite and hematite 163

8.3.2 Lattice incorporation of base metals in goethite and hematite 170

8.3.3 Jarosites/plumbojarosite 170

8.4 Consequences of the base metal retention on the extraction success of heap leaching

171

9 Conclusions and general perspective 175

10 References 180

11 Lists of figures and tables 199

12 Acknowledgements 206

13 Curriculum vitæ 207

14 Appendix Index 208

Abstract

6

Abstract

From the supergene ore of the polymetallic Sanyati deposit in north-western Zimbabwe

copper is won in a heap leaching - solvent extraction - electrowinning (HL-SX-EW)

process. However, the copper recovery is below expectations. Therefore, the composition,

formation, and leaching behaviour of the supergene ore was studied using mineralogical

(optical microscopy, SEM, and XRD), geochemical (XRF, EMPA, AAS, ICP-OES, and

LA-ICP-MS), and experimental methods in order to unravel the processes responsible for

the unsatisfactory copper recovery.

The supergene orebodies of Sanyati developed in a warm humid climate since the Miocene

(LISTER 1987) and are now forming inselbergs since the Pliocene erosion cycle. The

orebodies represent an immature and therefore very heterogeneous oxidation zone with

rudimentarily developed secondary sulfide ore lenses developed at its base.

Beside physical (unfavourable grain size distributions) and technical (a high compaction of

the heap leach pad) aspects, the textural and mineralogical characteristics of the ores lead

to metal retention during the heap leaching process. Significant amounts of copper (and

other base metals) are retained by the formation of ironoxides and -oxyhydroxides

(primarily goethite and hematite), which are ubiquitous in the supergene ore. These

"invisible" base metal contents are significantly higher compared to those reported from

other deposits (SCOTT 1986; SCOTT 1992). The observed distributions of base metals (and

misc. other elements) in goethite- and hematite-rich decay textures of sulfide minerals

demonstrate that they do not contain a geochemical fingerprint of their precursor sulfide

phase. Goethite-rich areas of the decay texture are generally enriched in Cu, Zn, and As, as

well as in selected trace elements (Ga, Ge, Se, Ag, Cd, and Sb), and are depleted in Pb

compared to hematite-rich areas. The base metal contents of goethite and hematite are in

the same range in run-of-mine ore and ore from the heap leach pad that has been leached

for several years. Extraction experiments on run-of-mine ore showed that 19 % of Cu and

6 % of Zn are adsorbed to the surfaces of limonite phases. The remaining 81 % of Cu and

94 % of Zn are fixed in the lattices of limonite phases (predominantly goethite and

hematite). Leaching and adsorption experiments showed that these phases are partly not

dissolved under the conditions used on the heap leach pad (H2SO4, pH 1.5 - 2). From the

dissolved part, limonite phases reprecipitate again and coprecipitate base metals that are

therefore lost for the output of the leaching process.

Zusammenfassung

7

Zusammenfassung

Aus den supergene Erzen der polymetallischen Lagerstätte Sanyati (NW Zimbabwe) wird

im heap leaching - solvent extraction - electrowinning (HL-SX-EW) Verfahren

hochwertiges Kupfer gewonnen. Das Kupferausbringen entspricht jedoch nicht den

Erwartungen. Um genauere Erkenntnis für die Gründe des geringen Ausbringens zu

erlangen, wurde die Zusammensetzung, die Bildung und das Laugungsverhalten der

supergenen Erze mit mineralogischen (Polarisationsmikroskopie, SEM und XRD),

geochemischen (XRF, EMPA, AAS, ICP-OES und LA-ICP-MS) sowie experimentellen

Methoden untersucht.

Die Bildung der supergenen Erzkörper erfolgte seit dem Miozän (LISTER 1987). Rezent

stehen diese Erzkörper als Inselberge in einer seit dem Pliozän gebildeten Erosions-

morphologie. Die Verwitterungsprofile der Erzkörper sind nur rudimentär zoniert. Die

Mineralisation der Oxidationszone ist sehr heterogen und eine Zementationszone ist nur

reliktisch als vereinzelte Erzlinsen ausgebildet.

Neben physikalischen (ungünstige Korngrößenverteilung) und technischen Gründen (zu

hohe Erzkompaktion auf den Laugungsbetten), führen mineralogische und texturelle

Eigenschaften zu einer Fixierung von Metallen im Erz. Wertmetalle wie Cu und Zn, aber

auch Pb und As sind an Limonitphasen (hauptsächlich Goethit und Hämatit) gebunden, die

Hauptmineralbestandteile der supergenen Erze sind. Diese "unsichtbaren" Metallgehalte in

Hämatit und Goethit sind in Sanyati deutlich höher als die in vergleichbaren Lagerstätten

beschriebenen (SCOTT 1986; SCOTT 1992). Die Elementverteilungen in goethit- und

hämatitreichen Abbautexturen von Sulfidmineralen enthalten keinen geochemischen

"Fingerabdruck" der Ausgangssulfide. Cu, Zn und As (sowie die Spurenelemente Ga, Ge,

Se, Ag, Cd und Sb) sind im allgemeinen in goethitreichen Bereichen der Abbautexturen

angereichert, während Pb-Gehalte in hämatitreichen Bereichen erhöht sind. Die

Buntmetallgehalte in Goethit und Hämatit sind in Roherz sowie gelaugtem Erz etwa

gleich. Extraktionsexperimente mit Roherz ergeben, daß 19 % des Cu und 6 % des Zn

adsorptiv an die Limonitphasen gebunden sind. Der restliche Anteil ist im Gitter der

Limonitphasen fixiert. Laugungs- und Adsorptionsexperimente zeigen, daß ein Teil der

Limonitphasen unter den Bedingungen der technogenen Laugung gelöst wird. Davon wird

ein Teil des mobilisierten Cu und Zn durch Copräzipitation mit Goethit und Hämatit

jedoch wieder fixiert. Aus diesen Gründen geht ein signifikanter Teil der Wertmetalle dem

Ausbringen verloren.

1. Introduction

8

1 Introduction

1.1 Object and aims of this study

Because of its mechanical and electrical properties, copper has a wide range of industrial

applications and the annual demand for copper is projected to double within the next two

decades (GRASSMANN et al. 2001). Bearing in mind that copper is a non-renewable raw

material, the sustained exploitation and development of copper ore deposits is therefore of

increasing importance in order to match industrial demands in the future. An analysis of

250 copper producing ore deposits, however, concluded that current mining, extraction,

and processing techniques lead to a combined loss of one third of the annually mined

copper ore, equivalent to an annual copper loss of approximately 4.2 mio t (SIEVERS et al.

2001). Insufficient extraction of copper from the ore is one of the most important factors

contributing to this copper loss.

This observation might be exemplified by the Sanyati mine, a supergene, polymetallic ore

deposit in north-western Zimbabwe with proven reserves of 5.8 mio t supergene ore at 1.1

wt% Cu and 1.2 wt% Zn (CHADWICK 1996). The deposit is mined by means of a heap

leach – solvent extraction – electrowinning (HL-SX-EW) process. However, since its

commissioning the copper recovery has achieved only 50 % of the targeted recovery

estimated from average copper concentrations (MAGOMBEDZE & SANDVIK 2002). It has

been assumed that the main loss of copper occurs during the technogene decay of the ore

on the heap leach pads, but the fundamental processes leading to the observed

unsatisfactory copper recovery remained unknown.

In heap leaching processes recoveries usually vary between 70 to 80 % (WELLMER 2002)

and might be as high as 90 % in well designed heap leach operations (William B. Dresher;

Copper Development Ass. Inc.; pers. comm.). This indicates that a significant copper loss

occurs even under optimal processing conditions. The aim of this study is to investigate the

different reasons for copper retention in leached ores. Because of its unusual low

recoveries, the Sanyati mine is perfectly suited to study copper and zinc retention in

supergene ores in detail. Detailed knowledge of the factors contributing to the low

recoveries in Sanyati is potentially useful also with regard to similar deposits worldwide.

1. Introduction

9

In 1997, the German Federal Institute of Geosciences and Natural Resources, Hannover

(BGR), commenced a project named “The Sanyati Polymetallic Ore Deposits in

Zimbabwe” within the framework of the BGR 2000 strategic research plan. An

introduction into the general aspects of the Sanyati project is given by OBERTHŰR (1999).

The major objectives of the project were (1) to develop models of the genesis of the

primary sulfide ores underground, and (2) to investigate and quantify supergene element

redistribution in the weathered oxidized ores found near the surface. This contribution

summarizes the results of a collaboration project between the Federal Institute of

Geosciences and Natural Resources, Hannover (BGR), and the Department of Economic

Geology and Applied Geochemisty, Institute of Applied Geosciences and Geoengineering,

Technical University Berlin (TUB), carried out within the frame of the BGR 2000 Sanyati

project.

The aim of this case study is to improve the understanding of the factors leading to (1) the

formation of supergene deposits, (2) the observed elemental distributions in these deposits,

and (3) the enrichment-depletion trends developed by both natural and technogene decay.

A central focus of the project is laid on the investigation of the residence, mobility, and

fixation of Cu and Zn (and miscellaneous other metals) in the oxidation zone of the ore

deposit as well as on the heap leach pad with the aim to elucidate the fate of these metals

during the natural and technogene decay of the supergene ores. These factors are of crucial

importance for the long-term operation of mines for both economic and ecologic reasons.

In supergene ores a redistribution of elements and formation of new phases takes place

under the acidic conditions of the sulfide decay. A part of the (base) metals is bound in

secondary minerals (e.g. malachite). Furthermore, these metals can be fixed in the rock-

forming minerals of the oxidation zone, particularly the ubiquitous secondary ironoxides

and -oxyhydroxides (mainly goethite and hematite). The secondary iron phases may act as

potential sinks for metals (CORNELL & SCHWERTMANN 1996), which can prohibit their

exploitation or encapsulate them from release into the environment. Attention has long

been paid to adsorption processes onto goethite and hematite surfaces, but a significant

proportion of metals may also be present in the crystal lattice (e.g. XIE & DUNLOP 1998;

GERTH & BRÜMMER 1990), although some metals may be incorporated by occlusion of

adsorbed metals during crystal growth and diffusion into micropore-like defects of the

1. Introduction

10

crystal structure (GERTH & BRÜMMER 1990).

In this study the supergene polymetallic ores won for the HL-SX-EW process at the

Sanyati mine and their host rocks were examined with special respect to the role of

goethite and hematite for the uptake and fixation of Cu, Zn, and a number of miscellaneous

metals prior to and during the acid leaching process using petrological (optical microscopy

and SEM), mineralogical (XRD), geochemical (XRF, EMPA, AAS, ICP-OES, and LA-

ICP-MS), and experimental methods. This approach offered the opportunity to study the

mineralogical, textural and geochemical changes the sulfide ores suffered during natural

decay under acidic conditions and under the acidic conditions of technological treatment

(HL process) the ores were subsequently subjected to for several years. This information is

important for an improved understanding of the physico-chemical changes affecting

supergene ores used in HL-SX-EW processes in general.

1. Introduction

11

1.2 Hydrometallurgical base metal production from leachable deposits

1.2.1 Supergene deposits, recovery techniques, economic importance: an overview

Supergene deposits

Supergene base metal deposits are the products of weathered precursor ores and rocks.

These source rocks can be volcanogenic massive sulfide ores, porphyry copper ores,

skarns, copper bearing sulfide veins or pyritic shales. The supergene enrichments in the

giant porphyry copper ores are of outstanding importance to copper economics today.

Metals like Cu, Zn, Pb, Ag, and Au are released from disintegrating sulfides and

transported downwards by meteoric water (POHL 1992). They can be precipitated either in

the oxidation zone or are often enriched under reducing conditions in the supergene

enrichment zone below the oxidation zone (Fig. 1.1).

Fig. 1.1: Sketch of the development of a supergene ore deposit (from POHL 1992).

Sufficient permeability is required for the development of a supergene ore deposit. In the

oxidation zone the dissolved minerals can react with carbonate, sulfate, silicate, phosphate,

oxygen and water compounds and precipitate as stable minerals, e.g. for copper: malachite,

azurite (carbonates), cuprite, tenorite (oxides), brochantite (sulfate), chrysocolla (silicate),

phosphates, and native Cu (NICKEL & DANIELS 1986). Below the descending watertable

secondary sulfides, e.g. chalcocite and covellite, are formed. Depending on weathering

conditions and time the maturity of the weathering profiles, which is documented by

1. Introduction

12

zonation, varies strongly (BLAIN & ANDREW 1977). Pioneers in the study of supergene ore

deposits were SCHNEIDERHÖHN (1924) and SMIRNOV (1954), and numerous studies have

been published in the following, including the relatively recent compilations by NICKEL &

DANIELS (1986) and THORNBER & TAYLOR (1992).

In the oxidation zone boxwork structures and impregnations occur. They mainly consist of

goethite and hematite, but also of jarosite, which can all be summarized as "limonite"

(NICKEL & DANIELS 1986). The distribution of limonite provides valuable information

about the history of the oxidation zone. Classical appraisal methods of leached iron caps

were introduced by LOCKE (1926), BLANCHARD (1968), LOGHRY (1972) and CHAO &

THEOBALD (1976). Continuous variations in mineralogy and modal abundance of limonite

can be used to interpret a relatively continuous descent of the watertable (ALPERS &

BRIMHALL 1989), while gaps in the profile indicate sudden drops of the watertable (e.g.

ANDERSON 1982) or multi-stage leaching (e.g. CUMMINGS 1982). With increasing maturity,

the vertical zonation of a weathering profile becomes more pronounced (BLAIN & ANDREW

1977). Mass balance calculations were used to reconstruct palæotopography, -hydrology,

and average rates of erosion and uplift on porphyry copper ores (ALPERS & BRIMHALL

1989).

The iron phases of limonite are capable of trapping and thus fixing metals via adsorption

(PEACOCK & SHERMANN 2004; ZHU 2002; PARKMANN et al. 1999; KOONER 1992, 1993;

BALISTRIERI & MURRAY 1982; FORBES et al. 1976), occlusion (GERTH & BRÜMMER 1990),

and lattice incorporation (XIE 1995; CORNELL & SCHWERTMANN 1996; GERTH &

BRÜMMER 1990).

Recovery techniques

In the history of metallurgy copper takes a prime position. It is believed to be the first of

the seven metals of antiquity (Cu, Au, Ag, Fe, Hg, Sn, and Pb) and was first produced by

mankind around 4000 BC. Zn was the eighth metal to be discovered 1200 AD by the

Hindus and a limited amount was produced soon (HABASHI 2003).

Smelting of supergene ores took place in primitive charcoal-driven furnaces. Reasons for

the early availability of these metals are the occurrence of some of them in their native

state (e.g. Cu), their reducibility below 800°C (e.g. Cu), and their easy attainability by

1. Introduction

13

burning carbonaceous material due to their relatively low melting points, often even

lowered by impurities (HABASHI 2003).

The first pyrometallurgical recovery methods with primitive smelter techniques are the

basis for the oldest sector of extractive metallurgy. First hydrometallurgical attempts,

involving aqueous solutions, were made by the Chinese, winning copper from blue vitriol

as early as 150 BC (DICINOSKI 2000). The technique was then used in the 16th century on

copper-bearing pyrite which was piled up in heaps to extract copper and precipitate it by

cementation with iron (HABASHI 2003). Up to the 1970´s cementation has been the

prefered hydrometallurgical recovery technique, until solvent extraction and

electrowinning offered the perspective to achieve a product quality by hydrometallurgical

means that could compete with smelter-won products (DICINOSKI 2000). Typical ores

treated by hydrometallurgy are those of Cu, Au, U, and Al.

The development of the copper production since 1800 is graphically displayed in Fig. 1.2.

Production rose in the 19th century with the electrification. The strong increase in

consumption during the last decade was caused by the "wiring of the world for the

internet" (FALVEY 2003).

Fig. 1.2: Copper production from 1800-2000 (from FALVEY 2003).

Today 13.9 mio t Cu/y (U.S.BUREAU OF MINES 2004) are produced of which 50 % is used

for electrical, 20 % for building, 25 % for engineering and 5 % for other purposes

(HABASHI 2003). Approximately 20 % of this copper is produced by a hydrometallurgical

route (AME 2004).

1. Introduction

14

In the case of Zn (8.5 mio t Zn/y; U.S.BUREAU OF MINES 2004) most is used for

galvanisation (47 %), die casting (14 %), production of alloys (mainly brass and bronze; 19

%), sheets and wires (8 %), zinc compounds (8 %), and other purposes (4 %). For Zn,

primarily produced from sulfides, the hydrometallurgical route has proved to be more

economical and substituted smelting techniques entirely. By the turn of the 20th century the

zinc production from "galmei" had accounted for a significant share of the world

production, but at present only 2 % of the world zinc production is won from supergene

ores (WELLMER 2002). However, the interest of mining companies in "zinc oxide" deposits

is increasing again (BONI 2003). Especially the development of SX-EW processes for "zinc

oxide" ores has renewed commercial interest (LARGE 2001, BORG et al. 2003) and a share

of more than 10 % of the annual zinc metal production is predicted for the future (BONI

2003, LARGE 2001).

Hydrometallurgy

Hydrometallurgy is used for a variety of raw materials (Table 1.1) to either produce a high

quality metal ready for the market (e.g. Cu and Ni) or metals of lower purity to be purified

by further treatment (e.g. Cu and Au). It can also serve as a first step of producing

compounds for further pyro- or electrometallurgical treatment (e.g. Al). Sometimes, e.g. in

ilmenite ores, the undesirable compound is leached and the remaining ore treated further

(HABASHI 2003).

Table 1.1: Raw materials used for leaching (from HABASHI 2003).

Raw materials Examples Metals Native: Au, Ag, Cu, Pt

By reduction of oxides: Cu, Ni, Co Oxides Bauxite, laterites, supergene copper ores, uranium ores, zinc ores,

manganese ores and nodules Complex ores Chromite, niobite, tantalite, pyrochlore, ilmenite, wolframite, scheelite Sulfides Mainly of the primary metals: Cu, Pb, Zn, Ni Selenides, tellurides Anodic slimes of copper electrolysis Arsenides Co and Ni arsenide ores and speiss Phosphates Phosphate rock, monazite sand Silicates Clay, nepheline-syenite, beryllium ores, serpentinite Chlorides and sulfates Halite, potash, pyrite cinder Carbonates Trona Borates Borax

To treat such diverse chemical compounds a plethora of leaching agents is used including

water, acids, bases, aqueous salt solutions, as well as oxidizing and reducing agents. Acids

1. Introduction

15

are the most common reagents, with sulfuric acid the most commonly used because of the

relatively low costs involved.

A number of methods are established (Table 1.2) and most of them can be used to treat

copper ores. Leaching in place and heap or dump leaching are the most time consuming

ore treatment techniques but are technically straightforward and require only minimal

equipment. Purification and, if possible, concentration of the solutions can be achieved for

dilute solutions with adsorption, e.g. on activated charcoal. More concentrated solutions

are subjected to ion exchangers or solvent extraction plants. Precipitation methods used to

recover the desired metal are fractional crystallisation, hydrolysis, ion precipitation, and

reduction (HABASHI 2003). Extensive studies on sulfide ores and concentrates are carried

out presently (e.g. demonstration plants from the Placer Dome and BioCop companies),

but still have to prove their success. Selective leaching of complex sulfide concentrates

may be possible, but the required leaching time may be too long for economic use (AKCIL

& CIFTCI 2002).

Table 1.2: Comparison of leaching methods and equipment (from HABASHI 2003).

Method Pressure [kPa]

Temperature[°C]

Agitation Ore size Time of leaching

Equipment Examples

Leaching in-situ Ambient Ambient None Lump Years None Cu, U Heap and dump leaching

Ambient Ambient None Lump Month None Cu, U, Au

Percolation or vat leaching

Ambient Ambient None Sandy Days Vats with false bottom

Cu, U, Au

Agitation or pulp leaching

Ambient < 100 Mechanical, compressed air

Fine Hours tanks, agitators

Cu, Au, ZnO, phosphates

Agitation or pulp leaching

1000-2000

110-200 Mechanical, high-pressure steam, rotation

Fine Hours Autoclaves Bauxite, laterite, Ni sulfide, scheelite

Baking Ambient ~ 200 None Fine Hours Digesters, rotary kilns, pugmills

Anodic slimes, monazite sand, ilmenite

In the case of the Sanyati mine, a heap leaching (HL) process with sulfuric acid was

chosen in combination with a solvent extraction (SX) and an electrowinning (EW) plant on

the spot in order to recover high quality copper metal from the supergene ore. Therefore

this process will be described in more detail in Chapter 1.2.2.

1. Introduction

16

Economic overview

The copper market is currently highly volatile and the low price outlook keeps costs under

pressure. The rapid growth in copper demand in the late 1990s led to uncoordinated

development in the following years. This resulted in four years of surplus production in a

row until 2000, when a deficit was recorded (AME 2004). Through 2002 and 2003 the

prices still declined because of the expansion of lower-cost operations and opening of new

mines but also abnormally low treatment and refining charges. A pause in the decline is

now expected in the course of 2004 and 2005 (AME 2004). Thus, at present, copper

producers have to conquer in an extremely competitive market.

The relative market share of the world leading Cu mining companies is shown in Fig. 1.3.

The largest six of the worlds prime copper producers owe their position to the production

of copper from concentrates and by-products, mainly Au (AME 2004).

Fig. 1.3: a. Relative market share of the currently most important Cu producing

companies, and b. relative share of extraction routes on the global Cu production

(from AME 2004).

The most important by-products from Cu mining are listed in Fig. 1.4. They influence the

profitability of a mine especially in times of strongly fluctuating prices at the stock

exchange. In 1998 about 50 % of the global production for Co, Ni, and Mo, and about a

quarter of the Zn production were won as by-products from Cu mines. Zn followed by Au

leads with regard to the value of the global production of by-products (Fig. 1.4,

GRASSMANN 2003).

1. Introduction

17

Fig. 1.4: Share and value of the global by-production from Cu mining relative to the

global primary metal production in 1998 (from GRASSMANN 2003).

Solvent extraction - electrowinning operations have currently a stable share of 20 % of the

global copper production (Fig. 1.3 b.; AME 2004; WELLMER 2002). The first commercial

hydrometallurgical SX-EW plant produced copper in Ranchers Bluebird mine, Arizona,

USA, in 1968. Until 1976 leach production rose by 55 %, accompanied with an increased

use of the SX-EW technique (SCHLITT 1980). In Fig. 1.5 the contribution of the SX-EW

operations to the total mine production is recorded since 1980. Zaire was the main

producer of SX-EW copper until the late 1980s. The technique had then even driven high-

cost smelter route operations out of business.

Fig. 1.5: SX-EW contribution to the mine production 1980-2001 (from WALLIS &

CHLUMSKY 1999).

0

10

20

30

40

50

60

Co Ni Mo Ag Zn Pb Au Sn[%

]

0

500

1000

1500

2000

2500

3000

3500

[Mio

US$

]

Share of by-product production in Cu minesValue of by-product production

1. Introduction

18

Supergene ore is dominantly won from porphyry ore deposits, followed by skarns and

sediment-hosted deposits. The Cu head grades [wt%] (= Cu content in the ore - dilution

through winning) vary in different ore deposit types between 1.5 wt% in VMS deposits

(volcanogenic massive sulfides) and 0.4 wt% in skarns and Cu-Mo-porphyries

(GRASSMANN 2003). They are generally lower for supergene ore compared to the

corresponding sulfide ore (GRASSMANN 2003).

The annual production of Cu ore with different head grades shows a similar distribution

pattern for sulfide and supergene ore in a comparison of 113 deposits (GRASSMANN 2003).

In 1998, 62 % of the copper production from supergene ore had head grades ≤ 1.0 wt%,

and nearly 80 % of the annual Cu production from supergene ore came from ore with head

grades of up to 0.5 wt% (GRASSMANN 2003).

An advantage of supergene ore exploitation is their surficial development. The

overburden/ore ratio is significantly lower compared to the sulfide ore for all six deposit

types compared. However, a lower output rate has to be taken into account for HL-SX-EW

processes compared to pyrometallurgical production from concentrates. The difference in

output is significant and exceeds 20 % for skarn, Cu-Au-porphyry, and Cu-Mo-porphry

deposits (GRASSMANN 2003).

USA13687

Mexico4627

Peru3045

21760Chile

Zimbabwe66

Iran165 India

167

3457Myan mar

Australia485

Cyprus90

Fig. 1.6: Proven and probable reserves of supergene copper ores [in 1000 t of Cu]

(without "Eastern Countries", DR Congo, and Zambia), which contribute 47.5*

106 t Cu, i.e. 11.5 % of the worlds total (from WELLMER 2002).

1. Introduction

19

The proven and probable global copper reserves in supergene ores are shown in Fig. 1.6

(from WELLMER 2002). Chile and USA share more than 80 % of the SX-EW production

and 70 % of the leachable copper reserves (WELLMER 2002). The average costs of the two

routes are currently closely aligned, but SX-EW routes become more competitive during

times of low copper prices or high treatment charges (AME 2004). In 1997 the production

cost of copper from SX-EW were US$ 0.43 / pound, which is about US$ 0.09 / pound

lower than for copper won by smelting (WALLIS & CHLUMSKY 1999). Thus, the leaching

of metals from supergene deposits shows future perspective because of economic as well

as ecologic reasons, and conservative forecasts indicate that the leaching of copper will

increase to 30 % of the total copper production (WELLMER 2002).

The SX-EW technology can be used beside copper production for U, Ni, Co, Zn, W, and V

(DICINOSKI 2000). Commercial interest in non-sulfide zinc has recently been kindled by

the development of SX-EW techniques for these ores (LARGE 2001). This nourishes the

idea of making high quality Zn a by-product that could also be produced at the Sanyati

mine.

1.2.2 HL-SX-EW, a hydrometallurgical route for copper recovery

1.2.2.1 Pyro- vs hydrometallurgical route – advantages and restrictions

A newly introduced technique starts to compete with the old well established ones. With

time, the new technique may displace the well established one or coexistence is achieved if

each technique has its own merits. In the following, the advantages and restrictions of

pyrometallurgy vs hydrometallurgy are shortly summarized based on the in-depth review

by HABASHI (2003).

For high-grade sulfide ores the pyrometallurgical route has obvious advantages because of

effective energy consumption and limited necessity for benefication, as large ore lumps are

directly transported to the plant and fed to the furnaces. Dust formation is lower during on-

site benefication and winning at the plant compared to low-grade ores. Low-grade ores

though have become increasingly important especially because of the giant sized porphyry

Cu mines (KESLER 1994, Chap. 1.2.1). For these ores benefication (e.g. grinding and

flotation) is necessary and horizontal reverberatory furnaces were developed. However,

1. Introduction

20

these furnaces have the disadvantage of high energy consumption and excessive dust

formation and are therefore environmentally much more critical: Dust and gases from these

furnaces have to be recovered to abate pollution and gain valuable material back. This is

not an issue in the wet on-site treatment in hydrometallurgical processes. The liberation of

SO2 during pyrometallurgical processes enforces a further process step in the production

line. In the better cases, SO2 is used to produce sulfuric acid, otherwise it has to be

prepared for adequate disposal. Treatment of sulfides with hydrometallurgical methods

reduces the need to produce sulfuric acid or dispose it to a large extent. Furthermore,

recycling of the agents for leaching and extraction is possible. Material handling is

generally much easier in hydrometallurgical plants by transfer of the solutions in pipelines.

For an efficient and cost-effective operation via the pyrometallurgical route elaborate heat

economy systems are necessary. In hydrometallurgical processes the temperatures rarely

exceed 100°C, so that heat economy systems are usually not necessary. In the

pyrometallurgical route transport to the smelting plant also has to be taken into account.

Low-grade ores and complex ores are preferably treated by hydrometallurgy as too much

energy would be required to melt the gangue minerals that dilute the low-grade ore and

selective leaching agents can solubilize the valuable materials specifically.

Residual slags from pyrometallurgical plants are often little problematic for disposal. In

contrast to these the often fine materials from hydrometallurgical plants are comparably

difficult to dispose as dry deposition generates dust and wet waste releases metal ions and

contaminates the environment, thus well prepared storage sites are needed.

Generally, the pyrometallurgical route is suitable for large-scale operations which require

large capital investments. For small-scale operations and thus low capital investments on-

site hydrometallurgical processes are more suitable. Their units can be increased without

any economic disadvantage when the need arises.

1.2.2.2 HL-SX-EW process for copper recovery

A simplified flow-sheet for the copper extraction by HL-SX-EW is shown in Fig. 1.7. The

ore is crushed, usually to a grain size < 10 mm and deposited on slightly tilted pads that are

sealed to prevent loss of acid into the underground. Typical materials used are asphalt or

clay. The heaps consist of a number of 3 - 10 m high individual ore layers (so-called

“lifts”). They can reach a total height of up to 100 m and usually hold between 100.000 to

500.000 t of ore.

1. Introduction

21

Fig. 1.7: General flow-sheet for the extraction of Cu by a heap leaching - solvent

extraction - electrowinning process (from WELLMER 2002).

Depending on the gangue supergene ores are either leached with sulfuric acid, or - in the

case of carbonate gangue - ammoniumcarbonate solution is used (SUTILL 2000). Supergene

copper ore is most commonly leached with sulfuric acid (5 - 30 kg H2SO4/m3) (BISWAS &

DAVENPORT 1994) according to the reaction:

CuO + H2SO4 ⇒ Cu2+ + SO42- + H2O (1.1)

The acid can be applied by sprinkler or trickle systems installed on the heap leach pads. To

achieve the best possible metal output in the "pregnant" solution (= solution after the

passage through the heap leach pad) the irrigation rate (acid distributed to the total area of

the heap) has to be optimised. The application rate, which refers to the area actually

leached and excludes resting areas, should be optimised with regard to the solution entry

into the heap without forming puddles (surface application rate) and the percolation

through the heap. For the latter, low application rates are preferable, because in a system of

blocks with low permeability that are separated by channels and fissures with high

permeability, the leaching of blocks without significant shortcutting of solution flow

should be permitted (JACKSON 1980). In experiments it was shown that sprinkler systems

yielded 15 % more copper to the pregnant solution than trickle system. Sprinkler systems

allow to keep application rates low by spreading the solvent over a greater area compared

to trickle systems (JACKSON 1980). Periodical leaching is proposed by BURGER (1985) and

MORRIS (1993). Leaching of the upper horizons of a heap leach pad takes 1 - 4 month, but

several years are needed for a complete heap leach pad (BISWAS & DAVENPORT 1994).

Coarse crushing

Fine crushing

Heap leaching Solvent extractionloading

Solvent extractionstripping

Electrowinning

Cathode CopperRun-of-mine ore

H2SO4 solution (<0.2 kg Cu/m3)Stripped solvent

Depl. electrolyte (35 kg Cu/m3)

loaded solvent(3-10 kg Cu/m3)

pregnant solution(1-5 kg Cu/m3

1-10 kg H2SO4 /m3)

enriched electrolyte(40-50 kg Cu/m3

~170 kg H2SO4/m3)

Coarse crushing

Fine crushing



Electrowinning


H2SO4 solution (<0.2 kg Cu/m3)Stripped solvent

Depl. electrolyte (35 kg Cu/m3)

loaded solvent(3-10 kg Cu/m3)

pregnant solution(1-5 kg Cu/m3

1-10 kg H2SO4 /m3)

enriched electrolyte(40-50 kg Cu/m3

~170 kg H2SO4/m3)

1. Introduction

22

In the extraction stage the desired metal ions are removed from the aqueous phase by

agitation with an immiscible organic solvent. A variety of extractants are available

including ethers, alcohols, ketones, oximes, oxines, organic acids, phenols, esters, and

amines (HABASHI 2003). The two immiscible phases are then allowed to separate. In the

following stripping stage the loaded organic is agitated with a small volume of

concentrated acid, which concentrates the metals in a relatively pure form. Generally, the

organic solution as well as the acids used can be recycled. Usually, the organic extractant

is diluted with a cheap organic solvent (the so-called “diluent”) to improve its physical

properties (e.g. viscosity and density). Most frequently used are hydrocarbons (e.g.

kerosene) or substituted hydrocarbons (e.g. chloroform). For copper a typical extractant is

hydroxyphenyloxime and petroleum is used as a diluent (KORODOSKY 1992). If secondary

sulfides make up a significant part of the ore oxidation with atmospheric oxygen or iron

(Fe3+) as oxidants is necessary.

Depending on the type of extractant the extraction mechanism can be an ion pair transfer,

anionic or cationic ion exchange, or chelate extraction (HABASHI 2003). In a chelate

reaction an electrically neutral chelate is formed, that is insoluble in the aqueous and

readily soluble in the organic phase, following the general formula:

2XH + M2+ ↔ X2M(ring) + 2H+ (1.2)

Stripping is achieved by a concentrated acid to reverse the equilibrium. The resulting

electrolyte is low in contamination and copper can be won from it using an inert lead anode

with a tension of 2 Volt (BISWAS & DAVENPORT 1994). Copper is won from the electrolyte

following the reactions:

Cathode reaction: Cu2+ + 2e- ⇒ Cuo (1.3)

Anode reaction: H2O + SO42- ⇒ H2SO4 + ½ O2↑ + 2e- (1.4)

Common Cu loads of the acid are stated in the flow-sheet (Fig. 1.7). Most of the acid is

recycled in two circulation systems.

While copper recovery from the heap leach pads is usually restricted to 70 % to 80 % of

the ore´s copper content, the leached copper is recovered almost completely from the

pregnant solution (98 %) in the SX and EW process (WELLMER 2002). The leaching

efficiency can be lowered if ore degradation (i.e. decay of ore blocs and grains because of

the acid attack) occurs within the heap leach pad. Consequences of ore degradation can be

lower permeabilities, increased saturation, sealing of zones in the heap leach pad, plugging

1. Introduction

23

of pipes, and higher solvent levels. Weathering of the surface leads to sealing of the pad

top and development of puddles which impair leaching efficiency. Channelling in the

dump can also be a problem, which increases with the height of lifts as zones of

preferential flow interconnect (SMITH 2002).

An advantage of SX-EW is that "Fe poisoning" is kept low. High iron concentrations, as

e.g. produced in direct leaching solutions, lead to a partial waste of current during

electrolysis by reducing ferric iron at the cathode and oxidizing ferrous iron at the anode.

Furthermore, ferric iron corrodes cathode loops. In SX-EW the iron content is already

reduced before the electrolysis circulation (SMITH 2002). Alternatively, purification of the

acid copper leach solution can be achieved by adding phosphate compounds (CRUZ et al.

1980).

Generally, a supergene ore suitable for HL-SX-EW should have the following

characteristics (BISWAS & DAVENPORT 1994):

- copper content should be between 0.25 - 1.00 wt%;

- copper should reside in phases that are dissolved in an adequate time;

- ores should have a permeability and porosity that allows sufficient contact of ore minerals

and acid;

- grain size should be adjusted so that ore lumps are leached thoroughly and fines prevent

the acid from rushing around the lumps; and

- for acid leaching the host rock should have a low acid neutralisation potential.

The ores from the Sanyati deposit are characterized by a high copper content (1.1 wt% Cu)

and the copper phases are adequate for leaching (see Chap. 2.4.2.). Permeability and

porosity are assumed to be high (see Chap. 2.5.2) and therefore only minimal benefication

is performed. Carbonates at the base of the oxidation zone (see Chap. 2.3) are expected to

be avoided during mining in order to allow effective acid leaching. Therefore, the Sanyati

ores can be regarded as suitable for HL-SX-EW. However, despite all these positive ore

characteristics, the Cu recovery is below expectations. The reasons for this insufficient Cu-

recovery are investigated in the following Chapters, starting with an overview of the

geological framework in which the ore developed.

2. The Sanyati copper ore deposit

24

2 The Sanyati copper ore deposit

2.1 Geographical situation

The Sanyati ore deposit lies in north-western Zimbabwe about 235 km west of Harare. The

mine at Copper Queen is located 6 km west of the Munyati river (Fig. 2.1). The deposits at

Copper King are not yet mined. A broad unpaved road connects the Sanyati mine to the

asphalt road from Chinhoyi to Harare.

Fig. 2.1: Topographical map of the Sanyati region (from Zimbabwe Sheet SE-35-8 Copper

Queen, ed. 3, published by the Surveyor-General, Zimbabwe 1977)

The morphology of the region is dominated by NS striking ranges of hills, which reach a

height of 965 m ASL. The land is partly cleared from its natural shrub and tree vegetation

for agricultural reasons. On the oxidation zones of the ore deposit the diversity of plants,

especially trees, is reduced.

The annual precipitation of 600 - 900 mm concentrates on the period from November to

March. The water supply for the mine and the surrounding settlements is guaranteed

throughout the year by the rivers Munyati and Mupfure, which combine north of Sanyati to

the Sanyati river.

2.2 Geological and morphological setting

2.2.1 Regional structural situation

The Archean Zimbabwe Craton is surrounded by orogenic belts: The Early to


25

Midproterozoic Magondi Belt in the north-west, the Archean Limpopo Belt in the south,

and the Neoproterozoic (Pan-African) Moçambique and Zambesi Belts in the east and

north. The polymetallic mineralisation of Sanyati lies in the south-west of the Magondi

Belt, which flanks the Zimbabwe Craton (Fig. 2.2).

Fig. 2.2: Simplified geological map of Zimbabwe (after TRELOAR 1988) showing the

Archean craton and the surrounding mobile belts. The Magondi Belt is positioned

to the NW of the craton. D = Dett Inlier, SAN = Sanyati mine.

The Magondi Belt is a north-east to north striking fold and thrust belt, with a maximum

exposed length along strike of 250 km and a width of 150 km. It has discordant borders to

the Neoproterozoic Makuti Group in the north-west and the Phanerozoic sediments of the

Karoo Group in the west and south-west. In the east the rocks of the Magondi Supergroup

thrust south-east onto the Archean craton. Especially in the south-east the thin-skinned

thrust belt is well developed (TRELOAR 1988). According to STAGMAN (1978) the rocks of

the Magondi Belt can be correlated with similar rocks occurring in a drill hole in

Bulawayo, while TRELOAR (1988) connects the rocks of the Dett Inlier (Fig. 2.2) to the

Magondi Belt. This would extend the length of the Magondi Belt to at least 450 km.

TRELOAR & KRAMERS (1989) correlate the Magondi Belt to the Kheis Belt, which lies in a

similar geotectonic position west of the Kapvaal Craton in South Africa. This connection is

adopted by GASCOMBE et al. (2000) in a classification of orogenies by absolute protolith

ages of 2400 - 1600 Ga.

The structural style of the Magondi Belt is a typical thin-skinned fold and thrust belt. It is

especially well developed in the southern part where the Sanyati deposit is located. On its


26

eastern margin it is defined by thrusts, in the central parts upright isoclinal folds dominate

(TRELOAR 1988).

The Early to Midproterozoic metasediments and metavolcanic rocks of the Magondi

Supergroup, which comprise the Sijarira, Piriwiri, Lomagundi and Deweras Groups (Table

2.1), are discordant on the Archean basement. The deposition age of the metasediments

and metavolcanics was determined with Rb-Sr whole rock analysis on mafic lavas at the

base of the Deweras Group to 2170 ± 100 Ma (TRELOAR 1988). In the Magondi orogeny

the Magondi Supergroup was deformed and metamorphosed at greenschist facies

conditions in the south and up to granulite facies conditions in the north. Peak

metamorphism was dated by MUNYANYIWA et al. (1995) to 1933 ± 4 Ma and 1959 ± 3 Ma

(zirkon single grain analysis, Pb-Pb evaporation method) and 1932 ± 27 and 1960 ± 39 Ma

(zirkon single grain analysis, conventional U-Pb method). Retrograde metamorphism was

dated by TRELOAR (1988) on Piriwiri phyllites of the southern area to 1753 ± 65 Ma and

1659 ± 50 Ma (K-Ar method on white mica).

Fig. 2.3: Simplified geological map of the Magondi Belt (from TRELOAR 1988). The Sanyati

mine (SAN) is located in the SW of the thrust belt close to the Copper Queen

Dome (CQD).

The intrusion of the granitoids of the Kariba region are dated by TRELOAR (1988) to 1980 ±

80 Ma (Rb-Sr whole rock analysis). These intrusions are syn- to posttectonic, because


27

some are undeformed while others have developed a strong foliation (TRELOAR 1988).

Recent datings of the Urungwe granite west of Karoi (U-Pb-SHRIMP analysis on zirkon)

show an age of 2135 ± 125 Ma (MCCOURT et al. 2000). It is not yet known if this age is

also valid for the neighbouring granitoids of the Sanyati region. However, the Copper

Queen Dome granitoid (Fig. 2.4) is included into the Urungwe Intrusive Belt by NEWHAM

(1986).

Pb-Pb ages of galena from the Sanyati ore deposit in the Piriwiri Group suggest an age of

2100 Ma for the sulfide mineralizing event (TRELOAR 1988). For this sulfide mineralisation

HÖHNDORF & VETTER (1999) calculated a Pb-Pb-model age of 2122 ± 14 Ma.

2.2.2 Lithostratigraphy

The Magondi Supergroup is divided into four groups (TRELOAR 1988); their lithologies are

summarized in Table 2.1. In the eastern part of the belt at the contact to the craton the

Deweras Group is developed. To the west it is superimposed by the Lomagundi Group,

which is itself overlapped by the Piriwiri Group (Fig. 2.3).

Table 2.1: Stratigraphy of the Magondi Belt and related basement and cover units

(TRELOAR 1988).

Group Formation Dominat lithologies Karoo Sandstone and basalts Makuti Vuti

Tsororo Rukwesa

Pink feldspathic psammites, amphibolite layers Pelitic schists, calc-silicate and amphibolite layers Mica schists with amphibolite and calc-silicate layers

Sijarira

Rhythmic series of quartzitic sandstones, arkoses, quartzites and dolomites with occasional shales and greywackes

Piriwiri

Copper Queen Chenjiri Umfuli

Phyllites Graphitic shales, some pyritiferous Phyllites and greywackes

Lomagundi

Sakurgwe Nyagiri Mcheka

Greywackes Argillites and striped slates, minor sandstones Orthoquartzites, dolomites, calcareous grits

M A G O N D I

S U P E R G R O U P

Deweras

U. Arenaceous Volcanic L. Arenaceous

Arkosic quartzites Basic lavas and tuffs Basal conglomerate, quartzite

Chipisa Chiroti Chitumbi Kariba

Silicic biotite paragneisses, some garnetiferous

Early Proterozoic paragneisses

Escarpment Urungwe

Biotite-hornblende and hornblende paragneisses quartzo-feldspathic granite gneisses

Archean Zimbabwe Craton basement granite-greenstone terrain


28

The mineralisation of Sanyati is in the Piriwiri Group, which is subdivided by LEYSHON

(1969) into three lithostratigraphic formations, from NW to SW the Copper Queen

Formation, the Chenjiri Formation and the Umfuli Formation.

During the Magondi orogeny the metasediments were folded with NE-SW to N-S trending

fold axis in the southern part of the belt with NE convergence. In the central parts upright

isoclinal folds dominate the fold and thrust belt (TRELOAR 1988). One the local scale of the

deposit steeply dipping en échelon folds are developed (Fig. 2.7). The folding has resulted

in steep dipping orebodies.

During the Magondi orogeny syn- to posttectonic intrusions took place, e.g. the

granodiorites of the Copper Queen Dome (Fig. 2.3, Fig. 2.4). This intrusive consists of a

marginal gneiss surrounding a central granitoid. At the contact to the metasediments

amphibolites, tremolite skarns, chlorite schists, and biotite- and garnet-bearing schists are

developed (BAHNEMANN 1961). These rocks are described by NEWHAM (1986) as the

“Urungwe Calc-argillite Formation” (Fig. 2.4).

Fig. 2.4: Simplified geological map of the Sanyati area (from NEWHAM 1986) with the

position of the orebodies (termed cupriferous gossans).


29

The Copper Queen Formation is a metasedimentary series dominated by schists,

micaschists, and phyllites. Metacarbonates, which show a metasomatic overprint, are

modally subordinate, but of striking importance as hosts of the mineralisation

(BAHNEMANN 1961). A detailed petrographic description of the rocks is found in LEYSHON

(1969) and BAHNEMANN (1961). The most frequently encountered rocks associated with

the mineralisation are shortly described in the following:

Phyllites and quartz-phyllites:

The dominant rocks around the mineralisation are fine-grained phyllites with a green-grey

colour. Frequently they are banded on the mm to cm scale. The major minerals are quartz,

chlorite, biotite, and plagioclase. Beside these white mica, garnet, and opaque phases

occur. The quartz content varies considerably so that phyllites alternate with quartz-

phyllites. Depending on the mica content the rocks show a gradual change in schistosity. A

description of thin sections is found in BURGATH (1999).

Metacarbonates and calc-silicates:

The metacarbonates are fine-grained and massive. They are slightly weathered and

therefore of beige colour. Major minerals are Fe-bearing dolomite and chlorite. The

metacarbonates show a strong metasomatic overprint and tremolite-rich rocks are often

found adjacent to them.

The calc-silicates are medium to coarse grained, massive, and of dark green colour. The

major minerals are dolomite, clinoamphibole (cummingtonite), and opaque phases (e.g.

magnetite). Subordinate are chlorite and talc (BURGATH 1999). According to BURGATH

(1999) ore and dolomite phases are replaced by clinoamphibole.

In summary, the rocks exposed in the surrounding of the Sanyati mineralisation are

dominantly regionally metamorphosed, siliciclastic metasediments that are locally

overprinted by a contact metamorphic event. The modally subordinate lenses of

metacarbonates are in direct contact to the primary and supergene mineralisation. All

structural features are NE-SW to N-S striking and relatively steeply dipping.


30

2.2.3 Morphological situation

The morphological development of Zimbabwe as well as the surrounding African countries

is subdivided into a series of erosion cycles that are summarized in Table 2.2 (LISTER

1987; KING 1967).

Table 2.2: Erosion cycles in Zimbabwe (LISTER 1987).

Name Geological Age Quaternary End Pliocene to present-day Pliocene Pliocene Post-African Miocene African Mid-Cretaceous to end Oligocene Post-Gondwana Early Cretaceous Gondwana Mid- to end Jurassic Intra-Karoo Triassic Pre-Karoo Late Palaeozoic

The morphology of the Sanyati region developed in the Post-African and Pliocene cycles.

It lies within the Zambesi valley province in the Sanyati-Sengwa basin south-east of lake

Kariba (Fig. 2.5 A, Fig. 2.6). The drainage basins of the Sanyati, Ume, Sasame, and

Sengwa rivers merge here to form a large embayment into the older Central Axis province

(Fig. 2.5). The margins of the Sanyati-Sengwa basin are erosional and not tectonically

controlled (LISTER 1987). The Piriwiri Group outcrops in this Pliocene basin, which forms

a moderately regular terrain of north-east trending ridges and valleys.

Fig. 2.5: Morphological provinces of Zimbabwe (from LISTER 1987).

The hills Copper Queen (965 m) and Copper King (913 m) are relicts of the Miocene Post-

African erosion surface which prevails to the east in the Central Axis province (Fig. 2.5 B).

In the Sanyati region this erosion surface is only preserved as inselbergs associated to


31

resistant rocks (e.g. the quartzitic Nyamapudzi hills, 970 m, and the gneissic Chiringoma

hills, 968 m; LISTER 1987; LEYSHON 1969).

Fig. 2.6: Palæomorphological map of NW Zimbabwe (from LISTER 1987).

The lithology of the hills of Copper Queen and Copper King is similar to the surrounding

area; the metasediments of the Piriwiri group prevail. Therefore the weathering cap of the

ore mineralisation obviously plays a role in increasing their resistance to erosion.

With regard to the morphological development of the Sanyati deposit this means that the

oxidation of the sulfide ore began in the Miocene during the Post-African cycle. The

supergene mineralisation was preserved during the Pliocene erosion cycle due to the

inselberg position, while the surrounding weathering surface was mainly eroded.

2.3 Distribution and composition of the primary mineralisation

The mineralisations of the Sanyati deposit are intercalated in siliciclastic and carbonate

metasediments of the Piriwiri Group. The ore lenses can be followed along a NE-SW strike

for at least 25 km (Fig. 2.4). Their occurrence is almost exclusively linked to

metacarbonates (BAHNEMANN 1961). Occasionally massive, “hornfelsic” quartz-biotite

rocks, coarse-grained amphibolites (tremolite-rich), and subordinate chlorite schists, biotite


32

schists, and garnet-rich rocks occur at the contact to the ore (BAHNEMANN 1961).

At the surface the steep dipping supergene ore lenses are fingerprinted as morphologic hills

(see above). Underground they vary strongly in thickness and length and some orebodies

are tectonically dublicated by folding (e.g. Copper-Queen).

Fig. 2.7: View to the southern wall of Copper-Queen open pit in the Sanyati deposit. The

orebodies are deformed and exhibited to the surface in steep dipping folds. At the

base sulfide ore is exposed.

The primary sulfides are massive as well as disseminated. In massive ores "patches of

massive pyrrhotite alternate with massive chalcopyrite and sphalerite, and arsenopyrite is

virtually absent in one part of the orebody and appears in high concentrations in another

part" (BAHNEMANN 1961).

The major constituents (in order of decreasing abundance) as well as the minor phases, as

reported by BAHNEMANN (1961) and OBERTHÜR & KOCH (1999), are listed in Table 2.3.

The high diversity of sulfidic and arsenidic minerals make it a complex, polymetallic

mineralisation, with the dominant metals Fe, Zn, Cu, Pb, and Co.

The genesis of the primary mineralisation is still discussed controversially in the literature.

A contact metamorphic origin is assumed by CULLIS (1926). MCCANN (1928), JACOBSEN

(1955), and BAHNEMANN (1961) interpret the mineralisation as a replacement or skarn

deposit, while SIMPSON (1988, cited by SARGENT 1992), MASTER (1991), and DUANE et al.

(1997) propose a synsedimentary, either volcanogenic-exhalative or SEDEX-type deposit.

However, although no conclusive model for the genesis of the primary mineralisation has


33

been established yet, the information available on the mineralogy, the lithologic

association of the host rocks and the tectonic situation provide a framework that can be

used to elucidate the genesis of the secondary, supergene mineralisation.

Table 2.3: Paragenesis of the primary ore of the Sanyati deposit (BAHNEMANN 1961;

OBERTHÜR & KOCH 1999) *: ordered with decreasing abundance, **:

alphabetical order.

Major components * Minor components ** pyrrhotite Fe1-xS bismuth Bi sphalerite ZnS bismutite Bi2(CO3)O2 chalcopyrite CuFeS2 cassiterite SnO2 galena PbS chalcocite Cu2S cobaltite CoAsS Co-bearing arsenopyrite FeAsS covellite CuS cubanite CuFe2S3 ilmenite/rutile FeTiO3/TiO2 mackinawite (Fe,Ni)9S8 magnetite Fe3O4 marcasite FeS2 monazite CePO4 pyrite FeS2 valleriite 4(Fe,Cu)S·(Mg,Al)OH2

2.4 Distribution and composition of the supergene mineralisation

2.4.1 Terminology of supergene mineralisations

As the terminology for iron-rich weathering zones is used inconsistently in the literature,

the terms used in this study are defined here:

The oxidation zone of an ore deposit is “an area of mineral deposits modified by surface

waters, e.g. sulfides altered to oxides and carbonates” (JACKSON 1997). Below the

oxidation zone at the watertable supergene secondary sulfide ores are precipitated in the

supergene enrichment zone. At the base of the profile primary sulfide ore (“protore”) is

preserved.

Oxidation zones are classified differently by many authors (e.g. SMIRNOV 1954; BLAIN &

ANDREW 1977; THRONBER & TAYLOR 1992). The differentiation of an oxidation zone

reflects its maturity; e.g. BLAIN & ANDREW (1977) subdivided nine profile types.

Generally, mature profiles have developed a base metal depleted zone close to the surface,

while no or less differentiation is observed in younger, immature profiles.


34

The term gossan is used for iron-bearing weathering products above a sulfide deposit,

often implying a base metal depletion, e.g. iron cap (JACKSON 1997, WHITTEN & BROOKS

1981). Here the term is used according to a classification by TAYLOR et al. (1980) and

WILMSHURST & FISCHER (1982) adopted by NICKEL & DANIELS (1986) and THRONBER &

TAYLOR (1992). Gossan is defined as a weathering product of sulfidic rocks without regard

of its economic value (THRONBER & TAYLOR 1992). Barren gossans do not contain

considerable amounts of economic metal values, either because their protore was

dominated by pyrite and pyrrhotite (NICKEL & DANIELS 1986), or because these metals

have been dissolved and removed from the profile (JACKSON 1997). NICKEL & DANIELS

(1986) defined the fertile gossan as the metal-rich equivalent. The differentiation of these

two gossan types is arbitrary and depends on the economic situation and the development

of metallurgical techniques.

With respect to their relative depth the ores of an oxidation zone are supergene (VISSER

1980) and development occurred under conditions of descending surface waters (JACKSON

1997) in an oxidic milieu. Therefore, the terms supergene and oxide could both be

adequately used to describe the ore. As the transition between ore and proximal host rocks

in Sanyati is continuous, limonite-rich ores with Fe2O3 > 30 wt% are defined as supergene

ore.

The term limonite is used according to NICKEL & DANIELS (1986) as a collective term

including ironoxides, -oxyhydroxides (mainly goethite and hematite) and jarosites.

Oxidation and the subsequent leaching of sulfides lead to the formation of boxworks

(JACKSON 1997). These relic patterns of intersecting blades or plates of limonite are

characteristic for various sulfide minerals (WHITTEN & BROOKS 1981).

To distinguish between alteration by regional weathering and the alteration of host rock

because of the acidic decay of the sulfide ore, the first is defined as distal host rock, while

the second is termed proximal host rock. As the "core stones" investigated in this study

are not of granitic origin ("Core stone: an ellipsoidal joint block of granite formed by

subsurface weathering in the same manner as a tor" (JACKSON 1997)) the term is used in

quotation marks.

2.4.2 The Sanyati supergene mineralisation

Due to weathering an oxidation zone with a depth of up to 35 m has formed above the


35

primary mineralisation at Sanyati (Fig. 2.8). The oxidation zone is in direct vicinity of the

sulfide zone without a distinct intercalated supergene enrichment zone, although locally

secondary sulfide enrichment occurs (see Chap. 4.2.2). The supergene ore is strongly

heterogeneous: porous and earthy textures in general predominate massive parts. The

colour varies from yellow and green to red, brown and black. The mineral assemblage,

examined intensively by VETTER et al. (1999), comprises a multitude of carbonates,

silicates, sulfates, arsenates and oxides/oxyhydroxids. The most frequent phases are listed

in Table 2.4. A table of all minerals is found in Table 4.6.

Table 2.4: Frequently occurring Fe and nonferrous metal phases in the supergene ore of

the Sanyati mine, according to VETTER et al. (1999).

Silicates Carbonate Arsenates Fe-phases chrysocolla malachite olivenite goethite hemimorphite cupro-adamite hematite adamite clinoclase duftite

2.5 Mining and benefication activities at the Sanyati mine

2.5.1 Historical development

There are indications that the ore deposit Sanyati was already mined in pre-colonial times

(VETTER et al. 1999). However, knowledge of the deposit and mining efforts of several

companies with focus on the sulfide ore is recorded only since 1900 (BAHNEMANN 1961;

CHADWICK 1996; SARGENT 1992). In 1989, after five years of field work focussing on the

supergene ore by the state-owned Zimbabwe Mining Development Corporation (ZMDC),

the Joined Venture “Munyati Mining Company” formed with Reunion Mining Ltd (U.K.)

in order to exploit the Sanyati deposit.

In addition to 14 mio t of sulfide ore with 1.25 wt% Cu, 3.22 wt% Zn, and 0.96 wt% Pb the

Munyati Mining Company had estimated 5.8 mio t of supergene ore with 1.1 wt% Cu, 1.2

wt% Zn, 150 mg/kg Co and 0.8 wt% Mn (MOBBS 1994).

In 1994 the development of an hydrometallurgical plant started with an investment volume

of US $ 14 mio. The first copper sheets were stripped in fall 1995 and by the end of that

year Reunion Mining Ltd. was traded at the London stock exchange. In 1995 Reunion

planned to start the "CoZiMa-Project", a pilot plant to recover cobalt, zinc and manganese


36

(MOBBS 1995). In summer 1999 Reunion Mining Ltd. was bought by the Anglo American

Mining Corporation. The Sanyati mine remained in the ownership of several individual

share holders and the ZMDC (R. Mushangwe, pers. comm.).

2.5.2 Mining process

In the period from 1995 to 1999 ore was excavated from seven open pits: Copper-Queen,

Copper-Queen-Beacon, F-Body-North, F-Body-South, J-Body, J-Lines-North, and J-

Lines-South. The location of the mined orebodies is depicted in Fig. 2.8. The overburden /

ore ratio was 1.2 : 1 (MOBBS 1994).

Fig. 2.8: Map of the supergene ore deposits at Sanyati. The seven orebodies mined are

underlined. Detail on the basis of the map: Sanyati Joined Venture: "Ore zone

location and schematic plant layout."- Reunion Mining (Zimbabwe) Ltd.


37

The ore is won by blasting and transported by articulated dump trucks either to a run-of-

mine ore dump or directly onto the heap leach pad. The grain size spectrum includes 25 %

< 1 mm and 10 % > 200 mm (Lee John, pers. comm. 1999). Even blocks up to a size of

1000 mm diameter were placed on the heap leach pad. Benefication was only performed

on blocks > 1000 mm which were crushed with a pneumatic hammer.

In 1999 future plans were to out-source the quarrying activities and to develop further

orebodies in the north. The development of the Copper-King mineralisations as well as the

winning of further elements like Zn and Co, as originally planned, was no longer included

in plans for the near future (R. Mushangwe, pers. comm. 1999).

2.6 HL-SX-EW at the Sanyati mine

A schematic flow-sheet for the HL-SX-EW operation at the Sanyati mine is presented in

Fig. 2.9. The heap leach pad covers an area of approximately 50,000 m2 (Fig. 2.10). The

slightly tilted ground surface (1.1°) is covered with a thick plastic base. An approximately

30 cm thick sand layer separates the ore from the plastic and guarantees the permeability

necessary for the pregnant solution to flow to the exit of the heap leach pad. In the older

heap leach pad areas the ore is piled up to three lifts and reached a height of 7 m in 1999.

In the following years 2 mio t of ore were put on the heap leach pad, so that a height of

nearly 20 m is reached in some areas (MAGOMBEZE & SANDVIK 2002). Other heap leach

pad areas are far lower with a maximum height of 1.5 – 2 m piled in only one lift. The ore

has a bulk density of 1600 kg/m3 (MAGOMBEZE & SANDVIK 2002).

Fig. 2.9: Flow-sheet of the HL-SX-EW plant at the Sanyati mine.



Electrowinning


H2SO4 solutionstripped solvent depl. electrolyte

loaded solventpregnant solution enriched electrolyte

Sulfur plantH2SO3

H2S

O4,

pH

1.5



Electrowinning


H2SO4 solutionstripped solvent depl. electrolyte

loaded solventpregnant solution enriched electrolyte

Sulfur plantH2SO3

H2S

O4,

pH

1.5


38

Fig. 2.10: a. Overview of the heap leach pad; b. Scetch map of the heap leach pad with

trenches for profile sampling; c. Older trickle system for acid dispersion on the

heap leach pad; d. New sprinkler system for acid dispersion on the heap leach

pad.

The acid for the leaching solution is produced by burning sulfur and directing the SO2 into

the leaching solution where it reacts quickly with H2O via H2SO3 to H2SO4. This

technique, first used at Sanyati, has reduced the operating costs significantly (CHADWICK

1996). The ore is leached with diluted H2SO4, at a pH of 1.5 (~6.5 g H2SO4/l) applied at a

rate of 12 l/h/m2 (MAGOMBEZE & SANDVIK 2002). First plans were to disperse the acid by

“trickle leach systems”, which can still be seen in some parts of the heap leach pad (Fig.

2.10 c.). However, “sprinkler systems” were installed subsequently and are currently the

prefered dispersion method (Fig. 2.10 d.).

The metal-rich “pregnant solution” is collected in two basins and fed at a rate of 200 m3/h

to the SX plant (MAGOMBEZE & SANDVIK 2002). Two Mixer-Settler stages are used for

copper extraction by complexation with the organic complexant ARCOGA diluted by

petroleum. The organic / aqueous ratio is 1:1. Before electrolysis Mn and other metals are

scrubbed. For electrolysis the copper is transfered into a strong acidic electrolyte and lead

to electrolysis cells. Reduction takes place for about 14 days until the copper sheets can be


39

stripped of the cathodes.

The solvent extraction plant has extracted 95 % of the Cu delivered to it (MAGOMBEZE &

SANDVIK 2002) and thus functioned satisfactorily. However, the overall copper recovery

since the start of the production is only ~ 50 % of the target (MAGOMBEZE & SANDVIK

2002). It is assumed that the main loss of copper takes place in the heap leaching stage. To

elucidate this further the field campaign focussed on the sampling of supergene ore in the

pits, the run-of-mine ore dump, and on the heap leach pads.

3. Field- and laboratory work

40

3 Field- and laboratory work

3.1 Fieldwork

A fieldwork and sampling campaign was carried out during the southern hemisphere spring

of 1999 (15.9.1999 to 16.10.1999) in collaboration with the BGR. The focus of the

fieldwork campaign was the sampling of profiles and single samples of the supergene ore,

the host rock and the ore on the heap leach pad. Sample lists are given in the Appendix B,

Tables B1-B4.

3.1.1 Sampling of orebodies and run-of-mine ore dump

Samples were taken in all open pits and accessible orebodies not yet mined (e.g. Fig. 3.1).

Except Copper-King, which lies 10.5 km to the south-west of Copper-Queen, all sampled

outcrops are in close vicinity (Fig. 2.8). With one exception (J-Body) the outcrops are

situated in morphologically high positions, i.e. in the centre or on the side of hills. If

adequate, samples were taken in profiles vertical to the steep dipping succession of

mineralized beds. The profiles were taken in varying positions to the "surface-before-

mining" from approximately five to seventy meters. A detailed description of the profiles

sampled as well as exact sample locations are given in FREI & GERMANN (2001b).

Fig. 3.1: a. Profile in F-Body-S open pit (P-FBS) (base of picture = 2 m) b. Chrysocolla-

rich ore lens in F-Body-N open pit (base of picture = 5 m).

Additionally to the 94 samples from the oxidation zone (Table 3.1), secondary sulfide ore

lenses (3 samples) and top parts of the sulfide zone (5 samples) were accessible in the

Copper-Queen open pit.


41

Table 3.1: Locations and numbers of supergene ore samples.

Orebody Type Sampling type No. of samples (∑= 94)

Copper-Queen open pit profile: P-CQ singles samples

19 1

Copper-Queen-Beacon open pit profile: P-CQB singles samples

6 4

Copper-Queen-SW open pit singles samples 3 F-Body-N open pit profile: P-FBN

singles samples10 9

F-Body-S open pit profile: P-FBS singles samples

8 3

J-Body open pit singles samples 9 J-Lines-N open pit singles samples 5 Copper-Joker unworked orebody singles samples 6 Copper-King unworked orebody singles samples 11

Water samples could be gained at the base of the open pits Copper-Queen and J-Body.

Efflorescences were sampled close to the groundwater in J-Body and independent of the

current watertable in the Copper-Queen open pit.

As the run-of-mine ore dump was nearly empty at the time of sampling (Fig. 3.2) and no

information could be gained on the exact source of the ore, only few samples were taken

(two large samples of the material < 3 cm and ten arbitrary lump samples ~ 5 cm).

However, their representative character might be limited.

Fig. 3.2: a. Run-of-mine ore dump north of Copper-Queen open pit. b. Separated big ore

blocks (Ø 1 m).

3.1.2 Sampling of host rock

The aim of sampling the weathering profile outside the influence of the sulfide decay was

impeded by the removal of most of this weathering horizon in the Miocene and Plicoene


42

erosion cycle (see Chap. 2.2.3). Only in few positions, e.g. in protected areas at the foot of

hills, relicts of the weathering profile could be obtained. In two cases vertical profiles were

taken and different stages of weathering were sampled from "core stone"-like relicts of

phyllites and their surrounding matrix. Termite hills and alluvial material were sampled to

gain average samples of the surrounding area (Table 3.2). Samples of distal host rock were

taken 5 km south of the deposit east of the Copper Queen Dome intrusion. A detailed

description of the sampled locations is given in FREI & GERMANN (2001b).

Table 3.2: Locations and numbers of distal host rock samples.

Location Number of samples (∑=38)

road cuts 2 weathered profile Copper-Joker 3 weathered profile J-Body 8 “core stone”- outcrop J-Body 4 “core stone”- outcrop Copper-Queen 4 termite hill & alluvial material 2 + 1 metacarbonate-outcrop Copper-Queen 2 Copper Queen Dome 12

3.1.3 Sampling of the heap leach pad

The heap leach pad consists of eight areas which were filled chronologically since 1995

(Fig. 2.10 a., b.). In the areas 1 to 4 material from the Copper-Queen open pit dominates,

the others were filled with a mixture from all open pits. No exact data on time of

deposition and source of the ore were available except that area 7 was filled in 1997.

At the time of the field campaign no leached ore had been removed from the heap leach

pad for final disposal. Therefore other means of collecting ores of varying stages of

leaching had to be chosen. To achieve this two trenches were dug to sample profiles

vertical to the surface (Fig. 2.10 b.). For this study four profiles in area 2 (Fig. 2.10 b., Fig.

3.3) at a lateral distance of 2 to 5 m and at 0.5 m intervals (max. depth 2.7 m) were

sampled to account for heterogeneity of the material. Two of these were chosen for intense

analysis. Additionally, larger blocks (~ 15 cm) were sampled in different positions to

examine their degree of leaching. Again, a detailed description of the profiles is given in

FREI & GERMANN (2001b).


43

Fig. 3.3: Photograph and sketch of profile A2-4 taken on the heap leach pad. (Grain size

distribution: FG < 3 cm, MG 3 - 15 cm, CG > 15 cm).

Efflorescences were collected on dry parts of the surface of heap leach pad (Fig. 2.10 c.).

From the acid circulation the "barren" and "pregnant" solutions, i.e. acid precipitated on

the top of the heap leach pad and acid after passage through the ore were sampled.

3.2 Analytical and experimental methods

3.2.1 Sample preparation

Drying

In the field the samples were collected and stored in plastic bags with inherent moisture.

The acid-moist samples from the heap leach area were placed immediately after sampling

into two plastic bags and sealed with adhesive tape for transport. In Berlin all samples were

stored and left to dry slowly at room temperature. Thin white efflorescences developed on

the surfaces of the samples from the heap leach pad.

Wherever drying was necessary in the further handling, e.g. after wet sieving and leaching

experiments, samples were placed in a drying oven at 40°C to prevent dehydration.


44

Crushing

Large rock pieces were formated with a diamond blade saw, e.g. to take samples from the

rim and centre of leached blocks. For chemical and X-ray investigations samples were

crushed and ground in an agate mill to a grain size < 63 µm. Grinding time for most

samples was approximately 10 min.

Thin- and thick section preparation

A total of 84 thin and thick sections and 14 polished mounts were prepared from supergene

ore, secondary sulfide ore, primary sulfide ore, host rock, and ore from the heap leach pad.

The different consistency and hardness of the phases within the samples was frequently

problematic for preparation, e.g. earthy next to silicified areas. To minimize these different

consistencies the samples were soaked several times in an epoxi-resin (ARALDIT 2020 ,

CIBA GEIGY) prior to preparation.

3.2.2 Leaching experiments

Several types of laboratory experiments were carried out to gain information about:

- the mobility of metals with different solutes (water and H2SO4) in run-of-mine ore and

leach pad ore,

- the bonding of metals to different phases in the ore (partial extraction), and

- sorption of Cu minerals to synthetic goethite under very acidic sulfate-rich conditions.

An overview of all types of experiments carried out is given in Table 3.3.

Table 3.3: Overview of the laboratory experiments.

Exp. nr. Experiment type Ore / matrix Solvent Vessel - method System Results in Appendix Table

V0 water extraction leach pad ore H2O PE-beaker (5 and 1 l) - stirred periodically

evaporated / open

C3

VR acidic extraction by percolation

run-of-mine ore

H2SO4 pH 1.5

PE-filter - percolation experiment

evaporated / open

C4

V15 acidic extraction by percolation

leach pad ore H2SO4 pH 1.5

PE-filter - percolation experiment

evaporated / open

C5

V1 acidic extraction by agitation


PE-beaker - stirred permanently

closed C6

V6 acidic extraction by agitation


glasflask - stirred permanently

closed /const. acid

C6


45

Table 3.3 continued:

Exp. nr. Experiment type Ore / matrix Solvent Vessel - method System Results in Appendix Table

V7 acidic extraction by percolation


plexiglas vessel - percolation experiment

closed C6

V12 NH4Ac extraction leach pad ore NH4Acetate 1M pH 4.5

PE-bottle closed C7

V13 NH4Ox extraction in darkness

leach pad ore NH4Oxalate 0.2 M pH 3

PE-bottle in darkness

closed C7

V14 NH4Ox extraction 80°C

leach pad ore NH4Oxalate 0.2 M pH 3

PE-bottle - 80°C closed C7

V8 sorption of malachite on goethite

goethite (syn.), malachite

H2SO4 pH 2.0

PE-bottle closed C8

V9 sorption of chrysocolla on goethite

goethite (syn.), chryscolla

H2SO4 pH 2.0

PE-bottle closed C8

V10 sorption of Cu-sulfate (syn.) on goethite

goethite (syn.), CuSO4 2H20

H2SO4 pH 2.0

PE-bottle closed C8

Note: all experiment were performed at room temperature except when otherwise stated.

All experiments were performed under oxidizing conditions at room temperature and

atmospheric pressure to mimic the situation on the heap leach pads as closely as possible.

The only exception is the extraction experiment V13 that was performed at a temperature

of 80°C. All experimental conditions and results (e.g. experimental run durations,

elemental concentrations, pH, conductivity, XRF data directly related to the experiments

etc.) are reported in Appendix C.

3.2.2.1 Leaching experiments of the water-soluble fraction (V0)

Because leach pad ore (LPO) samples were covered with efflorescences, leaching of the

water-soluble fraction was performed for these samples.

The samples of profile A2-2 were washed during the sieving process (sieve sizes were 10,

2, 0.71, 0.25, and 0.063 mm) and the filtrate was collected and sampled after every fifth

litre used for the wet sieving.

Samples from profile A2-4 were washed with one litre of deionised water and samples

were taken after the sediments had settled. This procedure was repeated nine times.


46

3.2.2.2 Leaching experiments of the H2SO4-soluble fraction

Prior to all acid leaching experiments the samples were sieved and only the grain size

fraction < 1 cm was used. After sieving all samples were washed ten times with one litre

demineralized H2O each time. This proved to successfully remove the water-soluble

fraction sufficiently (see above and FREI & GERMANN 2001b). The sulfuric acid used for

the experiments was adjusted to a pH of 1.5 in order to mimic the acid strength used during

the day-to-day operation of the mine.

Acid percolation experiments in an open system (experiments VR and V15)

A series of acid percolation experiments were performed via open filters using run-of-mine

ore (ROM) or leach pad ore (LPO) (Fig. 3.4 a.). For the experiments either 100 g of ROM

(experiment VR) and LPO (experiment V15; grain size < 1cm) were used and 1 l of acid

was percolated through the sample and collected. Percolation time varied from several

hours to ~1.5 days. After the pH and conductivity were measured (using a WTW 91 pH-

meter and WTW Lf 92 conductivity-meter) a sample of 30 ml was taken. The removed

amount of acid was replaced before the acid was circulated back through the sample. For

each experiment, this procedure was repeated for ten circulations. The acid irrigation spots

were frequently shifted in order to prevent the development of channels. The filters were

regularly rotated by 90° and acid drip velocity was adjusted so that a slight acid surplus

developed on the sample. For further details see FREI & GERMANN (2001b).

Acid percolation experiments in an semi-closed system (experiment V7)

To minimize the influence of evaporation an experiment with leach pad ore (LPO) was

designed using a closed plexiglass vessel to reduce the moisture loss during the experiment

(experiment V7, see Fig. 3.4 b, c). For this experiment 100 g of the LPO sample M1 = A2-

3-100 see Appendix C, Table C1 (fraction < 1 cm ground to a grain size < 63 µm) was

inserted so that all particles were distributed randomly and no visible layering developed.

The experimental procedure was essentially the same as for the open system experiments.

Additionally, the amount of leachate was measured after each run and the evaporation loss

during the experiment was found to be below 5 vol%. Again channeling was prevented by

a slight acid surplus on the surface and frequent shifting of the irrigated area. After the

experiment the leached ore sample was carefully dissected to check if channelised acid


47

flow had occurred, but only minor cavities were found and no interconnecting network of

channels was observed.

Fig. 3.4: a. Experimental set-up with filters for acid percolation experiments (VR, V15); b.

Set-up for the experiments performed in closed plexiglass vessels (V7). c. Close-

up of the sample vessel.

Acid extraction experiments in agitated systems (experiments V1 and V6)

In addition to the acid percolation experiments, an agitated leaching system with "ideal"

conditions was imitated experimentally. For these experiments a PE-beaker (experiment

V1) and a closed glassflask (experiment V6) were used. An aliquote of 50 g of the same

sample as in experiment V7 (M1 = A2-3-100 see Appendix C, Table C1, fraction < 1 cm

ground to a grain size < 63 µm) was stirred permanently using a magnetic stirrer. For

sampling agitation was stopped and 30 ml of the suspension were taken immediately and

always in the same position 5 cm below the solution surface using a microsyringe. The

extracted acid volume was replaced after pH and conductivity were measured. The samples

were filtered for solution analysis and the filter residue was washed with deionised water

and dried at 40°C. Filter residue samples of 2 - 5 successive runs were combined to gain

enough material for XRD and XRF analysis. Suspension sample F1 represents the early

phase, F3 the middle phase and F5 the end of the experiment and were used for XRF and

samples F2 and F4 were used for XRD measurements (see Appendix C, Table C2).


48

3.2.2.3 Partial extraction experiments (V12 - V14)

Four-step partial extraction experiments were conducted on an iron-rich supergene ore

sample (sample 263) to determine the metal mobility as a function of the leaching agent

(Table 3.4). The sample was split into four subsamples that were treated separately with

each of the various extractants so that the stronger extractions always included the fraction

solved in the weaker extractions.

Table 3.4: Conditions and materials used for the partial extraction experiments.

Leach Preferentially dissolved minerals

Solvent pH Solution to sample ratio [ml] : [g]

T [°C]

Time [h]

References

water-soluble fraction

secondary sulfates

deionised H2O

7 50:1 25 1 DOLD (1999, 2001a,b), RIBET et al. (1995)

exchangeable fraction

adsorbed and exchangeable ions

NH4acetate 1 M

4.5 500:1 25 2 DOLD & FONTBOTE (2001), SONDAG (1981)

Fe(III)oxyhydroxides ferrihydrite, jarosite

NH4oxalate 0.2 M, in darkness

3 500:1 25 1 DOLD (1999, 2001b), CARDOSO FONSECA & MARTIN (1986) SCHWERTMANN (1964)

Fe(III)oxides + oxyhydroxides

hematite, goethite, jarosite

NH4oxalate 0.2 M

3 500:1 80 2 DOLD (1999, 2001a,b)

The experiments were carried out in closed PE-bottles kept in a rocking device. The

extraction with deionised water yielded only concentrations close to or below the detection

(e.g. Zn 0.11 mg/l, Mn 0.06 mg/l). A planned fifth leaching step with HCl was not carried

out, because the sample was already completely dissolved in the 80°C NH4oxalate

solution.

3.2.2.4 Experimental adsorption of base metals to goethite (experiments V8 - V11)

In order to investigate the capacity of goethite for base metal fixation a series of adsorption

experiments were performed designed to closely mimic the technogene treatment on the

heap leach pad. For these experiments, goethite was synthesized following the standard

method of SCHWERTMANN & CORNELL (1991). This method produces acicular, partly

needly (elongated along the c axis) crystals with a surface area of approximately 20 m2/g


49

(SCHWERTMANN & CORNELL 1991). Although the surface area was not determined in this

study, numerous studies confirmed the findings of SCHWERTMANN & CORNELL (1991).

Therefore it can be assumed that the surface area of the produced goethite is within the

same order of magnitude (e.g. PARKMAN et al. 1999; MAJZLAN et al. 2003). The crystal

structure of the synthesised goethite was confirmed by XRD. Prior to the experiment the

goethite was washed four times with deionised water until neutral pH was achieved. The

experiments were carried out in PE-flasks to avoid silica contamination. Mineral separates

of chrysocolla and malachite were picked under a binocular from crystalline crusts of

Sanyati ores to optical purity. The separates were ground in an agate mill to < 63 µm and

subsequently analysed optically for pureness under a petrographic microscope and for

chemistry by electron microprobe analysis (EMPA). Additionally, experiments were made

with highly soluble synthetic copper sulfate (CuSO4*2H2O). For the experiments 10 g of

goethite and 1 mM of mineral separate (i.e. 460 mg chrysocolla in experiment V12, 220

mg malachite in experiment V13, and 498 mg CuSO4*2H2O in experiments V10) were

suspended in 1 l H2SO4 at a pH of 2. This low pH is close to the pH of the sulfuric acid

(pH of approximately 1.5) used during the day-to-day operation of the heap leach pads. For

the entire duration of the experiment the PE-flasks were kept in a rocking device.

Aliquotes of 50 ml were sampled at increasing time intervals. Each sampling was

combined with pH and conductivity measurement. All samples were filtered prior to

analysis.

3.2.3 Phase analysis

Microscopic techniques

All prepared thin- and thick sections were examined with a petrographic microscope

(ZEISS Axioskop) in transmitted as well as reflected light were appropriate. Grain sizes

were subdivided as follows: fine grained < 0.1 mm, medium grained 0.1 - 1.0 mm, and

corse grained > 1.0 mm.

Backscattered electron (BSE) pictures were made using Zeiss DSM 962 (at the GFZ

Potsdam) and Cambridge Instruments (at the Department of Earth Sciences, University of

Bristol) scanning electron microscopes with either 15 or 20 kV acceleration voltage. All

pictures were obtained using carbon coated sections or mounts.


50

X-ray diffraction (XRD)

For XRD measurements sample powder < 63 µm was put into a sample holder as

irregularly as possible to avoid prefered orientation and pressed. A PHILIPS PW 1729 X-

ray diffractometer with a control unit PW 1710 was used for the X-ray diffraction

measurements with a copper tube as X-ray source (TU Berlin). The instrument parameters

are listed in Table 3.5. For most samples programme A was used and programme B was

only used for some host rock samples allowing shorter measurement times.

Table 3.5: Instrument parameters for XRD measurements.

Programme A Programme B Voltage 50 kV 50 kV Current 30 mA 30 mA Goniometer speed 0.02 °2θ/s 0.04 °2θ/s Time constant 2.5 s 2.0 s 2 θ range 03°-80° 03°-70° Analysis time 2:53 h 1:10 h

Unit cell parameters and other structural parameters of selected goethites and hematites

were refined using the X´pert Plus software by PHILIPS.

3.2.4 Chemical analysis (XRF, EMPA, LA-ICP-MS, ICP-OES, F-AAS, and G-AAS)

X-ray fluorescence (XRF)

Major and trace elements of solid samples were determined using a PHILIPS PW 1404

XRF-spectrometer at the geochemical laboratory of the TU Berlin.

The majority of the samples were analysed using powder pellets. For calibration 90

international standards were used and the data evaluation was accomplished with the

programme POWDER (CRB/Hardegsen).

For preparation of powder pellets 6.0 g of the sample < 63 µm were homogenized with 1.5

g C-wax (MERCK) and pressed into Al-capsules (Ø 40 mm) with a pressure of 20 t. As the

concentrations of relevant elements often exceeded the limits of calibration the sample

powder was diluted with pure quartz sand (< 63 µm) where necessary. The dilution factors

in most cases were 1 : 3 and sometimes 1 : 10. These diluted pellets made it possible to

measure relatively high trace element concentrations. However, a matrix and grain size

effect had to be accepted. For the elements Si and Al this allowed only semi-quantitative

determinations. For samples with extreme concentrations of trace metals, a

semiquantitative procedure was used (SEMIQUANT/PHILIPS). The concentrations, which


51

are normalised to 100 %, were corrected for loss of ignition (l.o.i.).

A few samples were analysed using fused glass discs. Only samples with low Cu

concentrations allowed the preparation of fused discs, as high Cu concentrations burst the

pellets during cooling. For calibration 85 international standards were used and the data

evaluation was performed with the programme OXIQUANT (PHILIPS).

For preparation of fused glass discs 0.6 g sample and 3.6 g flux SPECTROMELT A 12

(MERCK) were homogenized. The mixture was melted in Pt/Au crucibles at 1200°C for 6

min in a ROTOMELT furnace before quick cooling in Pt/Au molds. All prepared glass

discs were stored in a desicator.

Since the analysis of powder pellets often yielded unsatisfactory totals, even with

consideration of the loss of ignition, 15 samples were analysed at the geochemical

laboratory of the BGR (Hannover) for quality control. At the BGR major and trace

elements were determined employing PHILIPS PW 2400 and 1480 sequence spectrometers

using fused glass discs. For glass disc preparation 1.0 - 0.5 g sample and a flux (2.5 g Li-

metaborate + 2.5 g Li-tetraborate and LiBr) were homogenized and melted at 1200°C for

20 - 25 min. The addition of LiBr prevented the bursting of pellets with a high Cu content.

The total sums of these analysis are 81.2 – 98.5 wt%. Only one sample with an extreme As

content of 19 wt% yielded a total of only 71.8 wt%.

Comparison of the results obtained at the BGR on fused glass discs with those obtained at

the TUB demonstrates that the analysis of diluted powder pellets at the TUB yielded

reliable results (see Appendix A, Fig. A1). However, the powder pellets prepared with a 1 :

3 dilution show the better correlation with the results obtained on fused glass discs

measured at the BGR and were therefore prefered. Therefore only for samples with

concentrations far beyond the calibration limits a 1 : 10 dilution or the semiquantitative

procedure were used. For Fe two samples do not show a good correlation for both

dilutions, a behaviour that might be attributed to a grain size effect. For Sb the analytical

error rises significantly at concentrations below 50 mg/kg. Since no Co data were available

from BGR, the results obtained for the two different powder pellet dilutions have been

compared. Because of the bad correlation, for Co a large error has to be taken into account.

The XRF results obtained with the best method are reported in the Appendix B, Tables B5-

B9 and Appendix C, Table C2. Data below the detection limit are marked e.g. <10.

Detection limits of all elements vary depending on the dilution used, i.e. for elements


52

determined with a 1:3 dilution the detection limits are three times higher compared to those

elements determined from undiluted samples.

Electron-microprobe analysis (EMPA)

For selected samples of the supergene ore, the sulfide ore, and the leach pad ore major and

trace elements were measured on polished sections using wavelength-dispersive electron-

microprobes. Measurements were carried out at the ZELMI of the TUB using a CAMECA

CAMEBAX with SAM´x programmes and at the European Large Scale Geochemical

Facility hosted by the Department of Earth Sciences of the University of Bristol using a

JEOL JXA 8600 SUPERPROBE.

The instrument parameters applied for analysis in both Berlin and Bristol are listed in

Table 3.6. For calibration natural and synthetic international standard materials were used.

The relative error for major element concentrations is approximately 1 – 5 wt% depending

on concentration. Hydrated ironoxides are notorious for non-stoichiometric incorporation

of water (and other volatiles), i.e. their degree of hydratisation is highly variable (CORNELL

& SCHWERTMANN 1996). The non-stoichiometric, variable degree of hydratization of the

ironoxides studied here is revealed by XRF analysis. Therefore, the oxygen content of all

ironoxides and ironoxyhydroxides was measured directly, rather then calculated from

stoichiometry. Although the error for direct oxygen measurements is about an order of

magnitude higher compared to those for the heavier elements routinely determined by

EMPA (in the range of 10 - 20 % relative, KRONZ 1998), this proved to yield a much better

quality check for the non-stoichiometric phases analysed in this study.

Table 3.6: Instrument parameters applied for microprobe measurements in Berlin

(Cameca Camebax) and Bristol (Jeol JXA 8600).

Cameca Camebax Jeol JXA 8600 Acceleration voltage 20 kV 15 kV Sample current 17 nA 20 nA Counting time on peak 10 - 20 s 8 - 10 s Count time background 5 - 10 s 10 s Beam size Spot - 20x20 µm 3 - 15 µm Matrix correction P.A.P.Samx (POUCHOU

& PICHOIRE 1984) ZAF (BENCE & ALBEE 1968)

Elements on Kα-line K, Mg, Ca, Al, Si, As, O, S, Mn, Fe, Co, Cu

K, Mg, Ca, Al, Si, As, O, S, Mn, Fe, Co, Cu, Zn

Elements on Lα-line Zn Pb Elements on Mα-line Pb, Hg


53

For trace element concentrations the relative error is estimated to be higher due to low

count rates relative to the background and thus lies in the range of 10 – 20 % depending on

concentration.

Laser ablation - inductively coupled plasma - mass spectrometry (LA-ICP-MS)

For selected samples the trace element concentrations in ironoxides and -oxyhydroxides

were determined at the European Large Scale Facility hosted by the Department of Earth

Sciences of the University of Bristol by laser ablation - inductively coupled plasma - mass

spectrometry (LA-ICP-MS) using a Nd:YAG 266 nm UV laser (MERCHANTEK) and a

VG Elemental PQ 3+ S-Option quadrupole ICP-MS. All trace element analyses were

carried out using the same polished sections previously used for EMPA analyses.

Depending on crystal size, the laser beam diameter was 10 – 20 µm. Background blanks

were measured on all measured masses for ~ 50 s before laser ablation and total analyses

time was ~ 80 s. Cu and Si determined by electron microprobe were used as internal

standards. The following isotopes were measured and ratioed to the internal standards: 29Si, 59Co, 63Cu, 65Cu, 66Zn, 68Zn, 69Ga, 71Ga, 72Ge, 73Ge, 77Se, 107Ag, 111Cd, 118Sn, 121Sb, and 209Bi. The NIST 610 and NIST 612 standard reference material glasses were used as

primary calibration and secondary quality check standards, respectively, using the

recommended concentrations suggested by PEARCE et al. (1997). Data reduction was

performed off-line on software following the routines of NORMAN et al. (1996) and

LONGERICH et al. (1996). The results are reported in Appendix D8.

For most elements the concentrations are well above the detection limit (calculated from

the mean background count rates + 3 x standard deviation). Because of the high

background count rates and the resulting high detection limits for Ga (43 mg/kg) and Ge

(131 mg/kg) a manual evaluation of the time-resolved signals was necessary. In Fig. 3.5

the time-resolved signals for ablation of the NIST 610 (Ga 438 mg/kg, Ge 426 mg/kg) and

NIST 612 (Ga 36 mg/kg, Ge 34 mg/kg) standard glasses are compared to those of goethite

unknowns from samples 212/2d-A1-1 (Ga) and 212/2d-A2-3 (Ge). Because of the strongly

fractionated nature of the signal during initial ablation, these parts of the time-resolved

signal were omitted for the quantification of elemental abundances. The count rates for Ga

are significantly above the background and the NIST 612 count rates, so that accurate and


54

precise results could be obtained. In contrast, the high background count rates for 72Ge

close to the peak count rates for the NIST 612 standard glass and the unknown lead to

large analytical errors that have to be kept in mind.

Fig. 3.5: Ablation behaviour for 69Ga and 72Ge during analysis of NIST 610 and NIST 612

standard glasses and a goethite unknown. For 69Ga manual evaluation was

necessary because of the fractionated signal at the beginning of ablation. The

high background count rates for 72Ge leads to high detection limits and high

analytical errors for concentrations close to the detection limit.

Emission- and absorption spectrometry (ICP-OES, F-AAS and G-AAS)

The fluid samples from the open pits and the heap leach pad as well as the eluats from the

experiments were measured according to their concentration with either ICP-OES

(PERKIN ELMER 7500), flame atomic absorption spectrometer (PHILIPS PU 9485) or

69Ga

100

1000

10000

100000

0 20000 40000 60000 80000 100000

time [ms]

cps

NIST 610NIST 612sample

72Ge

100

1000

10000

100000

0 20000 40000 60000 80000 100000

time [ms]

cps

NIST 610NIST 612sample


55

graphite furnace atomic absorption spectrometer (PERKIN ELMER 1100/AS40) at the

geochemical laboratory of the Technical University of Berlin. For calibration,

multielement-standard solutions (MERCK) were used. The quality of the measurements is

affirmed by routine participation in inter-laboratory ring tests. The detection limits are

listed in Appendix A, Table A2.

4. Weathering products and processes at Sanyati

56

4 Weathering products and processes at Sanyati

4.1 Weathering products of host rocks with (proximal) and without (distal) the

influence of sulfide decay

To distinguish between the different alteration phenomena in the Sanyati region, the term distal host rock is used for lithologies that have been alterated by weathering without the influence of acids derived from sulfide decay, while proximal host rock is used for the acid influenced alteration of host rocks in direct contact to the orebodies even though the source rock is often similar (compare Chap. 2.4.1). Due to the palæomorphological situation, complete weathering profiles of the Post-African

weathering cycle are not preserved (see Chap. 2.2.3). Therefore the samples described in

the following section are often from deeper, less mature parts of the weathering profile.

4.1.1 Mineralogical and geochemical composition of fresh host rock

In the surrounding of the mineralisation the rocks of the phyllite series dominate. The

biotite-plagioclase-quartz-phyllites vary in composition within this series (LEYSHON 1969)

and are described in more detail in Chap. 2.2.2. Fresh amphibolites were sampled at the

east side of the Copper Queen Dome. In contrast to rocks in immediate contact to the

mineralisations no indications for the presence of metadolomites or calc-silicates were

observed in their surroundings.

The metadolomites associated with the mineralisation fall into the field of Fe-dolomites of

the classification by RÖSLER (1979), assuming that all Fe is lattice bound. Additionally,

these dolomites contain significant amounts of Mn, which is considered as the source for

Mn-oxyhydroxides during weathering, and subordinate Zn. The average dolomite

composition (in mol%) might be given as: (Ca1.052Mg0.818Fe0.101Mn0.023Zn0.001)CO3.

The geochemistry of the fresh phyllites and the amphibolites is discussed in comparison

with their weathered counterparts in the following.


57

4.1.2 Mineralogical characteristics of weathered host rock distal and proximal to the zone

of sulfide decay

Distal host rocks

Outside the zone of sulfide decay the primary paragenesis of the biotite-plagioclase-quartz

phyllites are weathered to clayey and silty unconsolidated sediments with some “core

stones” in which the relict texture is preserved. A comparison of the minerals in the core

centre and the surrounding matrix yields the following mineralogical composition:

Table 4.1: Mineral phases present in "core stones" and the weathered matrix of distal host

rock.

Mineral „Core stone“ Matrix plagioclase + biotite + quartz + + muscovite/sericite + + chlorite + + smectite + illite + kaolinite +

Only in the centre of the “core stones” biotite and plagioclase are preserved. In the matrix

biotite is completely decomposed and chlorite is found in nearly all samples. The loss of

colour indicates the beginning of the biotite decay. Biotite decay is generally described as a

transformation to chlorite/vermiculite (WILSON 1970; ACKER & BRICKER 1992; WHITE &

YEE 1985). During this decay ironoxyhydroxides are precipitated (NAHON 1991) following

the oxidation of Fe2+ to Fe3+ (REBERTUS et al. 1986; BANFIELD & EGGLETON 1988).

KRUMB (1997) examined weathering profiles in acidic rocks and found an additional decay

from chlorite/vermiculite via illite to kaolinite during both acidic and reducing conditions.

These findings indicate that biotite and chlorite decay could have contributed to the illite

and kaolinite content of the weathered rocks in Sanyati.

In contrast to biotite, the more stable phases chlorite and muscovite/sericite are found in

the cores as well as the matrix, while plagioclase is not preserved outside the cores and

obviously decomposed to smectite and kaolinite. Because the plagioclase content is

generally low in the fresh phyllites, muscovite/sericite is the most likely contributor to the

clay formation. In most of the distal samples some limonite is found either in the matrix or


58

on surfaces of quartz grains or fractures.

Proximal host rocks

In the zone of sulfide decay the decomposition of phyllite is generally more intense.

Feldspars are almost completely replaced by clay minerals. The physical consistency of the

rocks varies strongly, with some displaying a hard framework caused by a silification,

while others are weathered to fragile sandy relicts with limonite impregnations. Thus, the

frequency of occurrence of the minerals in the proximal hosts rocks is, compared to the

distal host rock, generally shifted towards the phases of more intense weathering. Another

difference is the occurrence of alunite in the proximal host rocks. This mineral forms at or

below the watertable in conditions often associated with sulfide weathering (STOFFREGEN

et al. 2000).

In proximal amphibolites primary mineral textures are often well preserved, while the

primary mineral assemblage is mainly replaced by olive-green to petrol coloured chlorite-

dominated products. Grain boundaries and fractures are coated or filled, respectively, with

limonite or the rocks are impregnated by it.

4.1.3 Geochemical characteristics of weathered host rocks distal and proximal to the

zone of sulfide decay

4.1.3.1 Geochemical changes during weathering processes

As the phyllite series varies in lithology as well as in weathering degree, homogeneous

outcrops were used to illustrate the weathering characteristics distal to the mineralized

zone: “core stone”-outcrop Copper-Queen, “core stone”-outcrop J-Body, weathered profile

Copper-Joker and remote phyllites from the Copper Queen Dome area (Chap. 3.1.2).

The advanced weathering stage of the distal siliciclastic rocks in the Post-African

weathering cycle is illustrated in a ternary A-CN-K diagram in Fig. 4.1 a. Only in the

centres of the "core stones" and in one phyllite sample from the Copper Queen Dome the

initial composition is preserved. Most of the samples have reached the advanced

weathering stage (blue dashed arrow).


59

Fig. 4.1: Composition of weathered host rocks (phyllite and amphibolite) of the Sanyati in

the ternary A-CN-K diagram of NESBITT & YOUNG (1984; 1989) a.) In weathered

distal phyllites only the centres of the “core stones” (blue filled squares) and a

phyllite from Copper Queen Dome (blue dot) are in the initial weathering stage.

Most samples from the rim and matrix around the “core stones” are weathered

along the “granite trend” (blue arrow) and have reached the illitic stability field.

b.) In acidic altered proximal phyllites and amphibolites nearly all phyllite relicts

(blue squares) have reached the kaolinitic weathering stage along the “granite

trend” (blue arrows). The acid influenced alteration of the amphibolites

represents the mafic weathering trend (red arrow) often up to the chloritic stage.

(The advanced weathering trend follows the dashed arrow. Black dots represent

ideal composition of minerals).

Additionally, the normalised element distributions of Mg, Ca, Sr, Ba, Na, K, Rb, and Cs

can be used to describe the degree of weathering, as the larger alkali and alkali earth

elements tend to be retained better in weathering profiles (NESBITT et al. 1980). In the

Copper Queen "core stone" the distribution pattern for rim and matrix is well developed

(Fig. 4.2). However, the depletion of the mobile elements (especially Na and Ca) in the

matrix has not yet reached the degree of "highly weathered" rocks according to

WRONKIEWICZ & CONDIE (1987). This coincides with the mineralogical evidence that the

distal host rocks have not yet reached the extreme weathering stage of the kaolinite

stability field (Fig. 3.2, FREI & GERMANN 2001b).


60

Fig. 4.2: Normalised element distributions of the "core stone" phyllite samples from

Copper-Queen open pit (CQ) and the surrounding matrix (according to

WRONKIEWICZ & CONDIE 1987). The samples are shown with increasing degree of

weathering from the "core stone" to the matrix. Concentrations are normalized to

the upper continental crust (TAYLOR & MCLENNAN 1985).

The conditions of alteration inside the zone of sulfide decay were obviously more severe

than in the surrounding. The majority of the proximal host rocks, phyllites as well as

amphibolites, are extremely low in alkalies and alkali earth elements (see Fig. 4.1 b.).

While most distal phyllites (Fig. 4.1 a.) are still plotting in the illite stability field, the

phyllites proximal to the mineralisation are altered to the extreme kaolinite state (Fig. 4.1

b. blue dashed arrow). In accordance with this, most amphiboles are decomposed to

chlorite along the “basic” trend (Fig. 4.1 b. red arrow).

In the following the degree of weathering of the host rocks is compared with their ability to

act as a neutraliser of acid solutions. Therefore, a classification was made on the basis of

the chemical index of alteration CIA (NESBITT & YOUNG 1989, 1984, 1982), where the

CIA is a tool to quantify the degree of alteration and is defined as:

CIA= [Al2O3/(Al2O3+CaO+Na2O+K2O)] * 100 (4.1)

using molecular proportions; CaO must be silicate bound. Kaolinite and chlorite yield the

highest values, about 100, mafic rocks (e.g. fresh basalts) lie in the range of 30 - 45, and

acid rocks (e.g. granites) range from 45 - 55 (NESBITT 1982). The majority of the proximal

host rocks from Sanyati (Fig. 4.3) show a high degree of alteration with a CIA > 85 (9

amphibolite and 17 phyllite samples). These are discriminated from proximal host rocks

with a CIA < 75 - 62 for phyllites (3 samples) and < 53 - 26 for amphibolites (5 samples,

Fig. 4.3).

0.01

0.1

1

10

Mg Ca Sr Ba Na K Rb Cs

CQ center

CQ rim

CQ matrix

CQ matrix


61

Fig. 4.3: Chemical index of alteration (CIA) vs acid neutralisation capacity of phyllites

and amphibolites, distal and proximal to the zone of sulfide decay, based on

alkali and alkali earth cations only (ANCA). For comparison, the black line

marks the CIA of kaolinite, the blue line the lowest CIA for fresh basalts and the

red line the lowest CIA for granites.

During weathering, but also during the extreme acidic weathering conditions in an

oxidation zone, the host rock can act as a source of neutralisation. The acid produced by

sulfide decay is partly neutralised by the leaching of mineral surfaces and dissolution of

elements from them. Binding of protons and liberation of cations occur for example during

the decay of silicates, but this is often complex and subdivided in several reaction steps.

The degree to which acids can be neutralised depends mainly on the chemical composition

of a rock and its resistance to weathering. Equilibrium reactions as well as the kinetics of

solution and precipitation processes control these systems.

To characterise the ability of the host rocks to neutralise the acids produced during sulfide

decay a modified approach that was previously employed to describe soil acidification was

used. The ability of soils and rocks to neutralise solutions are quantified by their acid

neutralisation capacity (ANC). The ANC is a capacity factor based on the substrate

chemistry and bears information on the ion reservoir of a rock or soil (VAN BREEMEN et al.

1984) thus giving long-term information about the ion liberations from it. For comparison,

the pH or the alkalinity of a solution in the same system provide information on the

momentary capability of ion liberation for neutralisation. The ANC of a soil is defined as

0

200

400

600

800

1000

1200

1400

1600

1800

0 50 100 150

CIA

AN

CA

d is tal weathered phyllites

proximal moderatelyweathered phyllite

proximal s tronglyweathered phyllite

proximal moderatelyweathered amphibolite

proximal s tronglyweathered amphibolite

proximal chlorite s chis t


62

the difference of its basic cations minus the strongly acidic anions (VAN BREEMEN et al.

1984; HEINRICHS & BÖTTCHER 1994):

ANC = 6[Al2O3] + 2[CaO] + 2[MgO] + 2[K2O] + 2[Na2O] + 4[MnO2] + 2[MnO] +

6[Fe2O3] + 2[FeO] - 2[SO3] - 2[P2O5] - [HCl] (4.2)

Acidification leads to a reduction of the ANC, where basic cations are drained from the

system more readily than acidic components. Weak acids (SiO2, CO2) are not taken into

account as they do not contribute to the ANC at pH values between 2 - 5 (VAN BREEMEN et

al. 1984). For rocks the contributions of S, Cl, and P to the ANC can be neglected, as they

yield only minor amounts of acid (HEINRICHS 1993). Above the zone of neutralisation,

where acid input exceeds the neutralisation potential, Al, Fe, and Mn are set free from the

weathering of the rocks to clay minerals and oxyhydroxides. This contributes to the ANC.

Below the zone of neutralisation they can be neglected because alkali and earth alkali ions

buffer the system (HEINRICHS 1993). Thus for rocks below the zone of neutralisation

equation 4.2 can be simplified to:

ANCA = 2[CaO] + 2[MgO] + 2[K2O] + 2[Na2O] (4.3)

To describe the conditions in an oxidation zone this approach was modified by including

the total Al, Fe, and Mn calculated to their +III and +IV oxidation states, respectively.

Additionally S and P were included even though their contribution is minor. This leads to

the modified equation:

ANCB = 6[Al2O3] + 2[CaO] + 2[MgO] + 2[K2O] + 2[Na2O] + 4[MnO2] + 6[Fe2O3]

- 2[SO3] - 2[P2O5] (4.4)

The neutralisation of acid fluids passing through a rock leads to its decomposition and

mineral dissolution and is therefore connected to its degree of alteration. This relation is

quantified in Fig. 4.3. As expected, most proximal phyllites altered in the acidic milieu of

sulfide decay have a higher degree of alteration than distal phyllites outside the

mineralized zone. Accordingly the remaining ANCA of the proximal phyllites is lower than

the ANCA of the distal phyllites (Fig. 4.3 a.).

The proximal amphibolites show an astonishingly broad range of alteration degrees. They

are negatively correlated to their ANCA (Fig. 4.3 a.). Only two chlorite schists display

relatively high ANCA as well as CIA values. With regard to the silicate buffering system a

general negative correlation of the degree of alteration and the remaining capacity to

neutralise acids is found.


63

However, the conditions during sulfide decay were presumably more extreme, were pH has

been low and oxyhydroxides can contribute to the buffering capacity via e.g.:

[Fe(OH)2+] + [H+] ↔ [Fe(OH)2+] + [H2O] (4.5)

[Fe(OH)2+] + [H+] ↔ [Fe3+] + [H2O] (4.6)

The inclusion of Al, Fe, and Mn according to equation (4.4) leads to higher acid

neutralisation capacities ANCB (Fig. 4.3 b.). For most phyllite and amphibolite samples a

clear negative correlation between ANCB and CIA is found. However, many samples

mostly from the zone of sulfide decay (open symbols) show a high CIA and a high ANCB

retained due to presence of Al, Fe, and Mn phases.

Fig. 4.4: Chemical index of alteration (CIA) vs acid neutralisation capacity of phyllites

and amphibolites distal and proximal to the zone of sulfide decay based on alkali

and alkali earth cations and including Al, Fe, and Mn as cations as well as S, and

P as anions (ANCB). For comparison, the black line marks the CIA of kaolinite,

the blue line the lowest CIA for fresh basalts and the red line the lowest CIA for

granites. The red triangle highlights the samples with high ANCB and CIA plotted

in Fig. 4.5.

In Fig. 4.5 the ANCB of these samples is plotted versus their Al, Fe, and Mn content. In

some samples, e.g. the distal chlorite schists and phyllites, the high ANCB is due to higher

Al from silicates. For the majority of the samples, which are proximal to the supergene ore,

high Fe contents originate from decaying sulfide ore solutions. Significant amounts of Mn

are only observed for one sample (sample 73).

0

500

1000

1500

2000

2500

3000

3500

4000

0 50 100 150

CIA

AN

CB

d is tal weathered phyllites

proximal moderatelyweathered phyllite

proximal s tronglyweathered phyllite

proximal moderatelyweathered amphibolite

proximal s tronglyweathered amphibolite

proximal chlorite s chis t


64

Proton binding by Al from silicates attributes to a permanent neutralisation. In contrast

proton binding by the iron- and manganese-oxyhydroxides is more critical as these

amphoteric phases release the protons as soon as pH conditions change (see eq. 4.5, 4.6).

Fig. 4.5: Correlation of ANCB and Al, Fe, and Mn in intensely altered samples (indicated

with red triangle in Fig. 4.4).

4.1.3.2 Metal signature

Trace element distributions were used to investigate to what degree transported metals are

retained in the surrounding of the mineralisation. In phyllites the metals, e.g. Zn, are bound

to the clay minerals, where they can be sorbed or grow as new Zn-phyllosilicates onto the

clay surfaces if Si is present in the solution (SCHLEGEL et al. 2001). Sauconite, a Zn

smectite, can contain sufficient Zn to be a major ore forming mineral in non-sulfide Zinc

deposits (e.g. Scorpion Zinc deposit, Namibia; BORG et al. 2003; KÄRNER & BORG 2002).

Table 4.2: Cu, Zn, and Pb contents of samples from weathering outcrops distal to the

mineralisation as indicator for range of solutions derived from the ore lenses.

Cu+Zn [mg/kg] Pb [mg/kg] weathered profile Copper-Joker 5830 90 weathered profile J-Body 3300 60 “core stone”-outcrop Copper-Queen 2570 100 road outcrop 830 20 termite hill 340 40 “core stone”-outcrop J-Body 200 20 typical shale background (WEDEPOHL 1978) 135 30

050

100150200250300350400450

0 2000 4000

ANCB [mmol(eq)/kg]

Al 2O

3+Fe

2O3+

MnO

2 [m

mol

/kg]

distal weatheredphyllitesproximal weatheredphyllitesproximal weatheredamphiboliteproximal chloriteschist


65

The absolute contents of the relatively mobile elements Cu and Zn in the samples distal to

the mineralisation give an indication for the range of the solutions originating from the

decaying ore (Table 4.2).

But even in the “core stone” outcrop J-Body the Cu + Zn concentrations slightly exceed

those of typical shales (mean of shales according to WEDEPOHL (1978): 35 mg Cu/kg and

100 mg Zn/kg). Three phyllite samples (280cd, 282cd1, 283cd1) east of the Copper Queen

Dome in 5 km distance to the mineralisations show mean Cu and Zn concentrations typical

for the geological background of the Sanyati region (42 mg/kg and 102 mg/kg,

respectively, Appendix B, Table B7). In contrast, the relatively immobile Pb is only

slightly increased in the rocks surrounding the mineralisation (Table 4.2). The outcrops

with only slightly increased Cu + Zn contents show background values for Pb (~ 30

mg/kg) (WEDEPOHL 1978). Thus the gradient of immobile metal concentrations decreases

strongly with distance to the mineralisation, while mobile base metals do not reach

background concentrations in the area close to the open pits.

In order to investigate the systematics of metal distribution in the proximal rocks as a

function of alteration degrees, the rocks were classified in amphibolites and phyllites that

have a correlated ANCB and CIA as shown in Fig. 4.4, and those that are high in ANCB as

well as CIA (Fig. 4.3 indicated with a red triangle). The results are shown in Table 4.3.

Table 4.3: Mean of minor and trace metals in proximal rocks (amphibolites and phyllites).

following the weathering trend beyond the weathering trend amphibolites phyllites amphibolites phyllites

As wt% 0.96 3.30 0.03 0.20 Cu wt% 3.64 0.42 0.92 0.26 Mn wt% 0.54 0.52 0.64 0.14 Pb wt% 1.04 1.18 0.33 0.42 Zn wt% 0.95 1.11 0.43 0.14 Ag mg/kg 29 13 3 8 Bi mg/kg 70 428 51 113 Br mg/kg 8 21 5 5 Cd mg/kg 68 30 9 7 Co mg/kg 46 125 58 85 Cr mg/kg 141 49 77 120 Ga mg/kg 4 21 25 37 Ni mg/kg 27 43 38 20 Sb mg/kg 76 161 38 204 Se mg/kg 36 47 20 28 Sn mg/kg 438 49 61 25


66

For the majority of elements the mean concentrations are higher in the rocks following the

alteration trend. One hypothesis is that these rocks were in contact with the metal- and

acid-loaded solutions released from the decaying sulfides and physico-chemical conditions

allowed scavenging of metals into ironoxides and -oxyhydroxides (mainly goethite and

hematite). For the samples lying beyond this trend, either the initial metal input was low or

the physico-chemical conditions (e.g. pH and redox state) prevented metal scavenging. The

mineral assemblages of the rocks following the alteration trend (i.e. clays, limonite) might

also enhance the metal retention of these rocks by offering more possibilities for

adsorption or incorporation.

MAGOMBEZE & SANDVIK (2002) found "refractory Cu in cupriferrous and zinciferrous

micas". The only sheet silicates investigated with EMPA for trace elements were chlorites

in limonite-rich samples. They contain 73 mg Cu /kg, 1240 mg Zn /kg, and 170 mg Pb /kg

and consequently are only enriched significantly for Zn (see Appendix D, Table D4).

4.2 Formation of supergene ore in the oxidation zone

4.2.1 Breakdown reactions of primary sulfide ore

Because this study focuses on the oxidation zone only a few massive sulfide samples were

taken from the open pits Copper-Queen, F-Body-S and J-Lines-N. In accordance with

OBERTHÜR & KOCH (1999) major ore minerals present in the sulfide ore are pyrrhotite,

chalcopyrite, sphalerite, and less frequently galena; minor phases are pyrite, arsenopyrite,

cubanite, magnetite, and less frequently rutile and covellite.

Pyrrhotite

Pyrrhotite is one of the major constituents in the sulfide ore. For the oxidation of pyrrhotite

NICHOLSON & SCHARER (1994) and DOLD (1999) proposed the following mechanism:

Fe(1-x)S + (2-x/2) O2 + x H2O → (1-x) Fe2+ + SO42- + 2x H+ (4.7)

Fe2+ + 1/4 O2 +2 H+ → Fe3+ + H2O (4.8)

Fe3+ + 3 H2O → Fe(OH)3(S) + 3 H+ (4.9)


67

The production of acid is controlled by the stoichiometry of the pyrrhotite with the

maximum amount of acid produced by the iron-deficient end member. This is considered

to be the main reason for the observed (about an order of 100 times) higher oxidation rates

of pyrrhotite compared to those of pyrite. During oxidation, pyrrhotite is initially replaced

by marcasite and further decay leads to the formation of ironoxyhydroxide rims and

pseudomorphs (JAMBOR & BLOWES 1998). However, the decay of pyrrhotite can also

involve the formation of pyrite (BURNS & FISCHER 1990) and sulphur (AHONEN &

TUOVINEN 1994) via acid-consuming reactions (DOLD 1999).

Chalcopyrite

In the sulfide ore chalcopyrite is in some cases intergrown with cubanite, but cubanite also

occurs as an independent phase. The oxidation of chalcopyrite might take place without

acid production according to the reaction (WALDER & SCHUSTER 1998):

2 CuFeS2 + 4 O2 → 2 Cu2+ + Fe2+ + SO42- (4.8)

However, the oxidation of ferrous iron and ferrihydrate hydrolysis leads to the production

of acid via the reaction (DOLD 1999):

2 CuFeS2 + 17/2 O2 + 5 H2O → 2 Cu2+ + Fe(OH)3 + 4 SO42- + 4 H+ (4.9)

The oxidation rate of chalcopyrite increases with its ferric iron content, but has an

oxidation rate 1 – 2 orders of magnitude below pyrite (RIMSTIDT et al. 1994). PLUMLEE

(1999) categorised chalcopyrite as one of the most resistant sulfides to oxidation. On the

other hand, in natural alteration sequences chalcopyrite is found to be one of the more

reactive sulfides (BOWELL & BRUCE 1995).

Sphalerite

The decay of pure sphalerite might also take place without acid production (WALDER &

SCHUSTER 1998) via the reaction:

ZnS + 2 O2 → Zn2+ + SO42- (4.10)

However, several elements can substitute for Zn (e.g. Fe, Cd etc.). If the Fe-content in

sphalerite is high (which can be substituted up to 20 mol%, RÖSLER 1979), sphalerite will

generate acid in the same fashion as pyrrhotite during hydrolysis (WALDER & SCHUSTER

1998). The sphalerite present in the sulfide ores in Sanyati contains on average 7.3 wt% Fe


68

and hence might act as a considerable source for the production of sulfuric acid.

Galena

Oxidation of galena usually leads to a Pb2+ and SO42- rich solution in equilibrium with

anglesite (PbSO4). Anglesite coatings, which have a low solubility, can protect galena from

further dissolution (JAMBOR & BLOWES 1998).

2 PbS + 4 O2 → Pb2+ + SO42- + PbSO4

(4.11)

If the oxidant is oxygen, no acid is produced during the decay of sphalerite and galena. In

the presence of ferrous iron however, acid is produced according to the following reaction

(DOLD 1999):

2 MeS + 4 Fe3+ + 3 O2 + 2 H2O → 2 Me2+ + 4 Fe2+ + 2 SO42- + 4 H+ (4.12)

Pyrite

Based on textural evidence three types of pyrite can be distinguished in sulfide ores from

Sanyati: (1) one type is idiomorphic, in parts slightly rounded and shows only subordinate

fracturing, and (2) the other type is xenomorphic, strongly fractured, and often altered to

secondary products (Fig. 4.6 a., c.). Additionally, in a sample taken close to the supergene

zone, (3) colloform pyrites were frequently observed (Fig. 4.6 e.).

The types (1) and (2) occur in close vicinity and thus have been exposed to similar

conditions of weathering. However, they show different resistance to decay and can

therefore be distinguished easily. With respect to weathering this means that these two

types will alter at different velocities due to their different surface areas.

Even though pyrite is not frequently observed in Sanyati ores, its principal decay reactions

are shortly mentioned here. Pyrite oxidation takes place in three major steps: 1. oxidation

of sulphur, 2. oxidation of ferrous iron and 3. hydrolysis and precipitation of ferric

minerals, according to the reactions:

FeS2 + 7/2 O2 + H2O → FeSO4 + SO42- + 2 H+ (4.13)

Fe2+ + 1/4 O2 +2 H+ → Fe3+ + H2O (4.14)

FeS2 + 14 Fe3+ + 8 H2O → 15 Fe2+ + 2 SO42- + 16 H+ (4.15)

Arsenopyrite

In analogy to pyrite, arsenopyrite also occurs in two types with distinctively different


69

textures: (1) the first type is idiomorphic and (2) the second type is xenomorphic, strongly

fractured and decomposed to secondary phases (Fig. 4.6 a., b.). The Co contents of the

arsenopyrite vary between 3.7 and 0.7 wt% in two investigated samples, possibly reflecting

the two textural types of arsenopyrite or local variation of Co. However, the Co content in

one investigated sample (sample 34) is similar in both idiomorph and strongly weathered

arsenopyrites. Co from the arsenopyrite is regarded as the source for Co minerals in the

oxidation zone.

The oxidation of arsenopyrite can be described by the following reactions (MOK & WAI

1994; DOLD 1999):

4 FeAsS + 13 O2 + 6 H2O → 4 Fe2+ + 4 SO42- + 4 H2AsO4

- + 4 H+ (4.17)

FeAsS + 7/2 O2 + 6 H2O → Fe(OH)3 + SO42- + 4 H2AsO4

- + 3 H+ (4.18)

The oxidation rate is slightly lower compared to pyrite if oxygen is the oxidant (MOK &

WAI 1994), oxidation by ferric iron leads to similar rates (DOLD 1999).

Bornite, covellite

The occurrence of bornite (Cu5FeS4), which has not been described from Sanyati yet, could

be confirmed by microprobe analysis (Appendix D, Table D3). Bornite is typical in

oxidation zones above chalcocite or covellite and weathers to malachite/azurite (RÖSLER

1979).

Most samples investigated were taken in proximity to the supergene zone, therefore e.g.

chalcopyrite is partly altered to covellite. Covellite and chalcocite, which occur as

secondary phases mainly in the supergene enrichment zone, weather without acid

production in the same fashion as iron-free sphalerite.

Non-sulfide iron phases

Frequent iron phases produced by the decay of the sulfides (e.g. pyrite, chalcopyrite) are

ferrihydrite, schwertmannite, goethite, hematite, and jarosite depending on the pH - Eh

conditions (NORDSTROM et al. 1979; NORDSTROM 1982; SATO 1992; JAMBOR 1994;

BIGHAM et al. 1994; BIGHAM et al. 1996; SCHWERTMANN et al. 1995; NORDSTROM &

ALPERS 1999).


70


71

Fig. 4.6 (previous page): Polished section images of primary and secondary sulfide ores

from the Copper Queen, F-Body-S, J-Lines, and J-Body open pits

(microphotograph width: 400 µm, all taken with parallel nicols, except c. taken

with crossed nicols). See text for detailed explanations. Abbreviations of primary

(P) and secondary (S) sulfide minerals:

Abbreviation Mineral name

Origin Abbreviation Mineral name Origin

asp arsenopyrite P dg digenite S cp chalcopyrite P mal malachite S cc chalcocite S nat cu native copper S cub cubanite P py pyrite P cup cuprite S coll. py colloform pyrite S cv covellite S ten tenorite S

(Microprobe analysis of observed sulfide phases are listed in Appendix D, Table D3)

Jarosite is only stable at a pH < 3 and under oxidizing conditions and its stability is also

dependant on the availability of key elements, e.g. K and S (NORDSTROM et al. 1979;

BIGHAM et al. 1996). Ferrihydrite, schwertmannite and jarosite are metastable with respect

to goethite (BIGHAM et al. 1996).

Once ferric iron is produced it is the primary oxidant of pyrite (NORDSTROM et al. 1979;

MOSES et al. 1987; EHRLICH 1996). Below pH 3 the oxidation for pyrite by ferric iron is 10

– 100 times faster than by oxygen (RITCHIE 1994). Microbiological activity often even

accelerates the iron oxidation.

Below pH 3.5 Fe3+ will remain in solution and above pH 3.5 ironoxyhydroxides will

hydrolize and precipitate via the reaction:

Fe3+ + 3H2O → Fe(OH)3 + 3H+ (4.16)

Magnetite (not shown) is partly forming textures with triple junctions, but also occurs as

indepentant grains surrounded by an alteration rim. No evidence was found for the

presence of hematite in sulfide ore samples.

Sulfide alteration strongly depends on the alteration milieu and different sequences are

reported for decay below oxidation zones (BOWELL & BRUCE 1995; NICKEL & DANIELS

1986), in tailings (JAMBOR 1994), or during laboratory experiments (RIMSTIDT 1994).

According to a typical sulfide alteration sequence below an oxidation zone reported by


72

BOWELL & BRUCE (1995) the sulfides in Sanyati are expected to decay in the following

order: pyrrhotite >> sphalerite > chalcopyrite > bornite = arsenopyrite > galena > pyrite >

chalcocite > covellite. However, galena and arsenopyrite are placed in reversed order

according to BOYLE (1994) and chalcopyrite would be placed further to the end by

PLUMLEE (1999).

4.2.2 Composition of the secondary sulfide ore

Even though a continuous supergene enrichment zone (cementation zone) is not developed

at Sanyati, secondary sulfide ores were found in the open pits Copper-Queen, F-Body-S,

and J-Body. Frequently occurring ore minerals in these samples are: chalcopyrite, pyrite,

arsenopyrite, magnetite, ilmenite, chalcocite (mainly digenite), covellite, cuprite, tenorite,

native copper, as well as an unidentified pinkish ore phase. Chalcopyrite is preserved as

relictic cores with replacement rims by digenite and chalcocite (Fig. 4.6 f., i.). Both pyrite

types (see Chap. 4.2.1) are replaced in numerous stages of decay by digenite and covellite.

Cuprite also occurs as a non-replacing phase (Fig. 4.6 c., d.). Cuprite and tenorite are often

intergrown and malachite is formed beside them during weathering (Fig. 4.6 h.). Small

flakes of native copper were found as inclusions in cuprite (Fig. 4.6 g.). According to

MAGOMBEZE & SANDVIK (2002) cuprite is the major copper mineral in the supergene ore,

a finding that could not be confirmed by our observations.

The unidentified pink ore mineral is unlikely to be bornite as it does not oxidize upon

contact with air. The pink reflection colour observed by ore microscopy is suggestive for

renierite (Cu, Zn)11(Ge,As)2Fe4S16) (Fig. 4.6 i.), also called "orange bornite", but no EMPA

data are available to confirm this assumption. Renierite is a fahlore that typically occurs in

dolomite-hosted Cu-Pb-Zn deposits and hydrothermal polymetallic deposits

(WEBMINERALS 2004).

The secondary sulfide ore lenses, which consist of typical Cu-rich secondary sulfide and

oxide phases, are also mined for leaching. However, it is difficult to estimate the overall

volume of these ores, because they were only scarcely found in-situ during the field

campaign.


73

4.2.3 Composition of the supergene ore

The supergene ore of the oxidation zone is, because of the massive as well as disseminated

character of the primary sulfide ore, strongly intercalated and intermingled with the

proximal host rock so that the observed lithological changes are always transitional. For

this reason all ore samples with a Fe2O3 content > 30 wt% have been defined as “supergene

ore” samples and are described in more detail in the following.

A significant depletion of metal values proximal to the surface was not found, so that the

whole oxidation zone can be classified as a fertile gossan according to NICKEL & DANIELS

(1986) and no further subdivision of the oxidation zone is necessary.

Lateral variations in four lithological profiles from four open pits were examined (FREI &

GERMANN 2001b) and were found to be mineralogically and geochemically highly

variable. No correlation between Fe and metal values was found as individual phases and

metals bound to limonite could not be separated. The main reasons for the observed broad

lateral variations are the changing density of the primary ore (massive and disseminated

parts of the lenses), the type of adjacent rock in direct vicinity (which influences pH and

Eh, and thus the local precipitation regime), as well as the hydrogeological situation with

differing water pathways controlled by steep fracture systems.

4.2.3.1 Mineralogical characteristics

As the supergene ore shows transitional boundaries with the proximal host rock, it is

difficult to distinguish strictly between the mineralogical compositions of both lithologies.

Therefore the major mineral phases for both lithologies are listed together in Table 4.4.

Primary silicates are preserved as well as their weathering products.

A distinction between relict massive and disseminated ore can only be made with regard to

massive ores being mainly composed of residual limonites which are often silicified. In

contrast, disseminated ores retain relicts of the proximal host rock with transported and re-

precipitated ironoxyhydroxides occurring on fracture surfaces and as impregnations.


74

Table 4.4: Major minerals in the oxidation zone of the open pits (Copper-Queen (CQ),

Copper-Queen-Beacon (CQB), F-Body-S (FBS), F-Body-N (FBN)).

Minerals CQ CQB FBS FBN iron phases: goethite + + + + hematite + + + + magnetite + lepidocrocite + plumbojarosite/jarosite + primary silicates: quartz + + + + chlorite + + + + tremolite/actinolite + + cummingtonite/anthophyllite + + + weathering products of prim. silicates: kaolinite/halloysite + + + + illite + + smectite + + + + alunite + +

Goethite is the most frequently observed iron phase, which is in agreement with the trend

observed generally in oxidation zones (NICKEL & DANIELS 1986). Goethite is usually

formed via metastable ferrihydrite (5 Fe2O3 * 9 H2O), and direct precipitation is less often

observed (SCHWERTMANN & MURAD 1983). The metastable phases ferrihydrite or

schwertmannite have not been observed in the supergene ore, most likely because they

have been completely altered to goethite. This suggests that with respect to the types of

hydrous ironoxides, the oxidation zone has reached a certain degree of maturity. Hematite,

which is the second-most frequent iron phase in the supergene ore, can form during

dehydration of goethite or ferrihydrite (NICKEL & DANIELS 1986).

Goethite and hematite are by far the most frequent ironoxides and -oxyhydroxides, even

though minor amounts of e.g. lepidocrocite are present. Therefore the terms hematite and

goethite are used in the following.

ANDREW (1980) and SMITH (1977) reported oxidation zones with a decreasing

goethite/hematite ratio relative to the surface, indicating increasing dehydration of goethite

with time. For selected iron-rich samples the goethite/hematite ratios were calculated using

the modal abundances of goethite and hematite (on a wt% basis) determined by Rietfeld-

refinement. Especially samples from the Copper-Queen profile show extreme variations

and do not define a vertical trend with respect to the surface (Table 4.5).


75

Table 4.5: Goethite/hematite ratio of supergene ore samples from the open pits (Copper-

Queen, Copper-Queen-Beacon, F-Body-S, F-Body-N).

Sample nr. Profile Goethite (wt%) Hematite (wt%) Goethite/hematite Other phases 205b P-CQ 58 42 1.3 - 263 P-CQ 30 70 0.4 - 268d P-CQ 86 8 10.7 Quartz 6 wt% 49 P-FBN 19 9 2.1 Quartz 72wt% 13 P-CQB 32 39 0.8 Quartz 29 wt% 14 P-CQB 41 24 1.7 Quartz 35 wt%

Fig. 4.7: SEM photomicrographs of iron-rich supergene ore (sample 230b). a. & b.:

dominantly hummockey surfaces, c.: flat surface in the center, d.: porous rough

surfaces are preparation artefacts).

Examination of the supergene ore using SEM (in combination with semi-quantitative

chemical analysis using EDS) revealed that the mineral constituents of the supergene ore

show low crystallinity. Their morphology is predominantly flat and hummocky, while

rough and porous surfaces are probably preparation artefacts (Fig. 4.7). All examined


76

samples contain Fe, Cu, Al, and Si, while on the flat, rough surfaces additionally Zn is

detected, and As is detected on flat, hummocky surfaces.

Weathering of sulfides and precipitation of hematite and goethite along cleavage planes of

the decaying sulfide can lead to the development of characteristic boxwork textures

(BLANCHARD 1968; ANDREW 1980; BLAIN & ANDREW 1977; REYNOLDS 1982; NICKEL &

DANIELS 1986). The textural development of the investigated sulfides from the supergene

ore is described and discussed in detail in Chap. 6.

In colloform textures lepidocrocite is found only rarely and goethite is the dominant phase.

This is in good agreement with the fact that the formation of goethite is favoured with

respect to lepidocrocite if the Fe2+ oxidation rate is high (SCHWERTMANN & TAYLOR

1989). Furthermore, Al - Fe substitution and excess SO42- also favours goethite formation

(BIGHAM & NORDSTROM 2000; SCHWERTMANN & TAYLOR 1978). Additionally, the

presence of excess HCO3-, CO3

2-, and CO2 in the solution can also suppress lepidocrocite

formation (SCHWERTMANN 1959; CARLSON & SCHWERTMANN 1990; BIGHAM &

NORDSTROM 2000). Most of these conditions are closely matched in the oxidation zone of

Sanyati, so that frequent precipitation of lepidocrocite is not likely. Magnetite, a stable

phase in this environment (SMIRNOV 1954), is well preserved in the oxidation zone and

only a weak martitisation is observed. Plumbojarosite and also jarosites have been detected

frequently. The latter can not be further subdivided as the XRD patterns of hydronium

jarosite and alkali jarosites are difficult to differentiate (DUTRIZAC & JAMBOR 2000) and

the patterns here contain rarely less than three or four phases, leading to complicated XRD

spectra. Jarosites occur e.g. close to the sulfide zone in Copper-Queen but also 5 m below

the surface in Copper-Queen-SW. Minerals of the jarosite group generally form from Fe3+

bearing solutions with a pH < 3. They are typical for acidic and oxidizing conditions

associated with sulfide weathering (STOFFREGEN et al. 2000). Plumbojarosite forms often

from galena via anglesite and cerrusite. Plumbojarosite in turn weathers readily to goethite

if removed from its pH-Eh-stability region, so that near-surface oxidation zones are usually

poorer in jarosite minerals and in mature oxidation zones jarosite is only found deeper in

the profiles (DUTRIZAC & JAMBOR 2000). However, jarosite is formed readily in naturally

weathered mine tailings after only a few years during oxidation of sulfides. The presence

of plumbojarosite close to the surface as well as in the proximity of the primary sulfide ore


77

provides further evidence for the overall immaturity of the oxidation zone in Sanyati, even

though the assemblage of the hydrous ironoxides indicates a low degree of maturity.

A detailed compilation on the minerals derived from sulfide weathering, including the

numerous Cu-, Zn-, and Pb-phases present, is given by VETTER et al. (1999). Additionally,

the arsenates beudantite, and hidalgoite were detected in this study and the presence of

scorodite was confirmed (Table 4.6). The compilation by VETTER et al. (1999) extended by

the newly detected phases is listed Table 4.6.

Table 4.6: List of the Sanyati supergene ore minerals compiled from VETTER et al. (1999),

FREI & GERMANN (2001b) and this study.

Mineral Formula Frequency Reference sulfides: chalcocite (chalcosine) Cu2S + VETTER et al. (1999) and ref. therein covellite CuS + VETTER et al. (1999) and ref. therein halogenides: matlockite PbFCl +? VETTER et al. (1999) and ref. therein oxides, hydroxides: cuprite Cu2O ++ VETTER et al. (1999) and ref. therein tenorite CuO + VETTER et al. (1999) and ref. therein litharge (lithargite) PbO +? VETTER et al. (1999) and ref. therein magnetite Fe3O4 +++ VETTER et al. (1999) and ref. therein hematite Fe2O3 +++ VETTER et al. (1999) and ref. therein hetaerolite ZnMn2O4 +? VETTER et al. (1999) and ref. therein cassiterite SnO2 ++ VETTER et al. (1999) and ref. therein goethite α-FeOOH +++ VETTER et al. (1999) and ref. therein lepidocrocite γ-FeOOH + FREI & GERMANN (2001b) coronadite Pb(Mn4+,Mn2+)8O16 THIS STUDY carbonates, nitrates: smithsonite Zn[CO3] ++ VETTER et al. (1999) and ref. therein cerussite Pb[CO3] + VETTER et al. (1999) and ref. therein azurite Cu3[(OH)|CO3]2 ++ VETTER et al. (1999) and ref. therein malachite Cu2[(OH)2|CO3] +++ VETTER et al. (1999) and ref. therein rosasite (Cu,Zn)2[(OH)2|CO3] + VETTER et al. (1999) and ref. therein unknown mineral ? Cu hydroxy nitrate + VETTER et al. (1999) and ref. therein sulfates: anglesite Pb[SO4] + VETTER et al. (1999) and ref. therein brochantite Cu4[(OH)6|SO4] ++ VETTER et al. (1999) and ref. therein osarizawaite PbCu(Al,Fe)2[(OH)6|(SO4)2] + VETTER et al. (1999) and ref. therein beaverite PbCu(Fe,Al)2[(OH)6|(SO4)4] ++ VETTER et al. (1999) and ref. therein plumbojarosite PbFe6[(OH)6|(SO4)2]2 ++ FREI & GERMANN (2001b) melanterite Fe[SO4] * 7 H2O + VETTER et al. (1999) and ref. therein bieberite Co[SO4] * 7 H2O + VETTER et al. (1999) and ref. therein goslarite Zn[SO4] * 7 H2O + VETTER et al. (1999) and ref. therein gypsum (selenite) Ca[SO4] * 2 H2O ++ VETTER et al. (1999) and ref. therein alunite KAl3[(OH)6|(SO4)2] + FREI & GERMANN (2001b)


78

Table 4.6 continued:

Mineral Formula Frequency Reference jarosite KFe3[(OH)6|(SO4)2] + FREI & GERMANN (2001b) phosphates: pyromorphite Pb5[Cl|(PO4)3] ++? VETTER et al. (1999) and ref. therein veszelyite (Cu,Zn)3[(OH)3|PO4] +? VETTER et al. (1999) and ref. therein arsenates: olivenite Cu2[OH|AsO4] +++ VETTER et al. (1999) and ref. therein adamite Zn2[OH|AsO4] +++ VETTER et al. (1999) and ref. therein -var. cobalto-adamite (Zn,Co)2[OH|AsO4] + VETTER et al. (1999) and ref. therein -var. cupro-adamite (Zn,Cu)2[OH|AsO4] +++ VETTER et al. (1999) and ref. therein cornwallite Cu5[(OH)2|AsO4]2 ++ VETTER et al. (1999) and ref. therein clinoclase Cu3[(OH)3|AsO4] +++ VETTER et al. (1999) and ref. therein duftite PbCu[OH|AsO4] +++ VETTER et al. (1999) and ref. therein bayldonite PbCu3[O(OH)2|AsO3(OH)2] + VETTER et al. (1999) and ref. therein philipsbornite PbAl3[(OH)6|AsO3(OH)|AsO4] + VETTER et al. (1999) and ref. therein mimetite (mimetesite) Pb5[Cl|(AsO4)3] ++ VETTER et al. (1999) and ref. therein scorodite Fe[AsO4] * 2H2O + VETTER et al. (1999) and ref. therein legrandite Zn2[OH|AsO4] * H2O + VETTER et al. (1999) and ref. therein lavendulane NaCaCu5[Cl|AsO4]4 * 5 H2O +? VETTER et al. (1999) and ref. therein pharmacosiderite KFe4[(OH)4|(AsO4)3] * 6-7 H2O +? VETTER et al. (1999) and ref. therein beudantite PbFe[(OH)6|(As,S)O4]2 ++ FREI & GERMANN (2001b) hidalgoite PbAl3[(OH)6|AsO4|SO4) + FREI & GERMANN (2001b) silicates: hemimorphite Zn4[(OH)2|Si2O7] * H2O +++ VETTER et al. (1999) and ref. therein chrysocolla (Cu,Al)2H2[(OH)4|Si2O5 * nH2O] +++ VETTER et al. (1999) and ref. therein halloysite Al4[(OH)8|Si4O10] * 4 H2O +++ VETTER et al. (1999) and ref. therein

A copper mineralisation was very frequently found in form of a dense turquoise-coloured

precipitate. XRD and mass balance calculations showed that it is a tightly intergrown

mixture of chrysocolla (85.2 wt%) and malachite (14.8 wt%). This kind of intergrowth of

the easier soluble malachite with the less-easier soluble chrysocolla might suppress the

solubility of malachite.

Beudantite and scorodite are locally very frequent and can even occur as major phases, e.g.

in the Copper-Queen-Beacon open pit (FREI & GERMANN 2001b, 2002a). Scorodite, which

precipitates after the decay of arsenopyrite, is a relatively stable phase in oxidation zones

(SMIRNOV 1954) and hydrolyzes only slowly after longer contact with water:

FeAsO4*2 H2O → Fe(OH)3 + H3AsO4 (4.19)

However, acidic conditions can enhance scorodite decay (SMIRNOV 1954). After hydrolysis

the Fe is fixed in ironoxyhydroxides, while As is mobilized and might be subsequently

fixed with base metals or lead. In general, lead-bearing arsenates are more stable compared

to base metal-bearing arsenates (SMIRNOV 1954).


79

Additionally, oxide or hydroxide phases of Mn occur, occasionally leading to a black

appearance of some ores, e.g. in Copper-Queen-SW.

Fig. 4.8: Alteration sequence for oxidation of sulfides and products in the oxidation zone of

Sanyati (restricted to the more frequent supergene phases).

The alteration processes observed in Sanyati and the derived weathering sequences from

the sulfides are summarized schematically in Fig. 4.8 (in a simplified form only taking the

most frequently occurring phases into account). Dolomite decay will control the pH as long

as it is present and can provide carbonate anions e.g. for malachite and azurite formation.

Additionally, it provides Mn for hydroxide precipitation in the oxidation zone.

4.2.3.2 Geochemical characteristics

The Fe2O3 contents of the supergene ore samples vary between 30 and 70 wt%. Supergene

ore samples without any macroscopically visible individual base metal minerals

nevertheless contain (as a mean of all open pits) 1.27 wt% Cu, 1.38 wt% Zn, and 229

mg/kg Co. These concentrations are slightly above the mean values estimated by the

mining company, probably because they have taken a higher portion of host rock into

account which can not be separated with the currently used mining technique.

The median metal contents of the supergene ore samples generally exceed the contents of

the proximal phyllites and amphibolites by factors ranging from 1.2 to 5.3 (Ga 1.2, Cu 1.3,

Bi 1.5, Se 1.5, V 1.5, Ba 1.7, As 1.8, Co 1.8, Ag 2.0, Sb 2.1, Cd 3.3, Mn 3.6, Zn 4.9, and


80

Pb 5.3), i.e. the degree of enrichment is more or less in the same order of magnitude for all

metals investigated (Fig. 4.9).

Fig. 4.9: Metal enrichment factor in supergene ore samples calculated from medians of

supergene ore normalised to proximal host rock. Elements are arranged in order

of increasing enrichment.

The investigated profiles yielded no evidence for vertical Cu concentration gradients

relative to the surface (Fig. 4.10). This suggests that no significant metal depletion of the

upper parts of the oxidation zone took place.

Fig. 4.10: Cu vs Fe2O3 in supergene ore proximal and distal to the surface.

As already shown in Chapter 4.2.3.1 and by FREI & GERMANN (2001b), the metal contents

0

1

2

3

4

5

6

Ga Cu Bi Se V Ba As Co Ag Sb Cd Mn Zn Pb

met

al e

nric

hmen

t fac

tor

0

5000

10000

15000

20000

25000

0 20 40 60 80

Fe2O3 (%)

Cu (m

g/kg

)

distal to the surface

proximal to the surface


81

in the supergene ores are highly variable. This variability is on the one side due to the local

occurrence of individual metal-bearing phases and on the other hand due to physico-

chemical variations in the local limonite forming environment (e.g. variations in factors

like pH and Eh that strongly influence metal adsorption and coprecipitation during

formation of goethite and hematite). In an attempt to distinguish between these two

Fig. 4.11: Frequency distribution diagram of Cu concentrations in supergene ore to

exclude outliers. Black: normal element distribution, white: excluded outliers

presumably because of Cu mineralisations.

scenarios the complete geochemical dataset of the supergene ore was analysed using

univariate statistics. Using frequency distribution diagrams, a homogeneous population of

samples characterised by unimodal, Gaussian-type element distributions was established.

This process excludes samples containing signals related to the presence of individual

minerals that are highly enriched in a specific metal (which results in non-Gaussian,

bimodal frequency distributions). The metal budget of the sample population derived by

this method is assumed to be controlled to a large extent by the fraction bound to limonite

phases, i.e. the fraction of the metal budget that is not easily removable during heap

leaching (Fig. 4.11).

The relative composition of this fraction is graphically displayed in pie-diagrams in Fig.

4.12 for all examined open pits and the mean for all open pits. Clearly, the relative metal

proportions that are assumed to be bound to limonite and are thus not easily dissolved

during heap leaching are very variable between the pits. These findings demonstrate that in

02468

1012141618

0-0.1

0.1-0.

25

0.25-0

.5

0.5-0.

750.7

5-11-1

.25

1.25-1

.5

1.5-1.

75

1.75-2

.0

2.0-4.

04.0

-15

15-m

ore

groups

freq

uenc

y

Cu


82

each individual pit the expected leaching behaviour of the ore might vary considerably. For

example, the part of limonite-bound Cu content in the F-Body-S and F-Body-N open pits

are relatively low compared to all other open pits, pointing to an enhanced potential of Cu

leachability. Both of these ore lenses are nearly excavated, presumably because of these

good leaching characteristics.

Fig. 4.12: Boxplots depicting mean, lower quartile, upper quartile, min. and max. of

selected metals Al, As, Cu, Mn, Pb, Ti, and Zn (wt%) and Ag, Bi, Cd, Co, Sb, Se,

and V (mg/kg) of samples with metal contents assumed to be limonite bound.

For the only partly or not yet excaved open pits Copper-Queen-SW and Copper-Joker, the

data indicate a high degree of Cu bound to the limonites. This observation suggests that the

Cu-recovery should be validated with respect to the profitability of the open pit prior to

mining. Copper-Joker shows additionally high As concentrations. The open pit Copper-

King on the other hand indicates a lower proportion of limonite-bound Cu and very high

Mn concentrations.

The Al-budget of the supergene ore samples is presumably dominated by Al bound to

silicate phases, however a part of the Al can also be bound in goethite. The homovalent

substitution of Al3+ for Fe3+ in goethite is strongly controlled by the conditions under

which oxidation takes place and high Al3+ concentrations are found in lateritic weathering

conditions (SCHULZE 1984). Up to 33 mol% Al3+ can be incorporated in goethite as solid

solution along the goethite-diaspor join (SCHULZE 1984). Because the ion radius of Al3+

0

4

8

12

[wt%

]

Al As Cu Mn Pb Ti Zn

0

200

400

600

[mg/

kg]

Ag Bi Cd Co Sb Se V


83

(0.53 Å in 6-fold coordination) is ~ 18 % smaller than Fe3+ (0.645 Å in high-spin state in

6-fold coordination), leading to a shift of the lattice width of the (111)-plane, the Al

content in goethites can be approximated from the XRD patterns (SCHULZE 1984). The Al

contents of goethites in eleven supergene ore samples (Appendix B, Table B10) from

Sanyati were determined in this fashion. The highest Al content observed was only 5 mol%

and no systematic vertical gradients relative to the surface could be found.

Fig. 4.13: Variation of selected metals bound to limonite phases (i.e. the metal fraction

that is not easy removable during heap leaching) in the Sanyati open pits. Also

given is the mean of all open pits examined.


84

Similar to Al the base metals can be bound to goethite and hematite (CORNELL &

SCHWERTMANN 1996). High concentrations and often bimodal distribution for these

elements were found by FREI & GERMANN (2002b) and therefore the distributions of these

elements was studied in more detail using EMPA and LA-ICP-MS. The results of these

microchemical studies are discussed separately in more detail in Chapter 6. In the

following, the chemical composition of the more abundant minerals formed in the

supergene ore are described briefly.

Hematite: Beside hematite-rich zones, that will be discussed in Chapter 6, euhedral

hematite occurs. The base metal contents of euhedral hematite (with an average major

element composition of 70.1 wt% Fe and 27.3 wt% O) are generally low with 0.79 wt%

Cu, 0.08 wt% Zn, 0.08 wt% Co, and 0.3 wt% Pb the maximal concentrations being

observed.

Manganese phases: Manganese hydroxides in the supergene ore are found as black crusts

or impregnations, but could not be specified in more detail by XRD.

Fig. 4.14: BSE image of a Pb-Mn-mineral, probably coronadite surrounded by goethite-

rich precipitates in the oxidation zone sample (212-2, microphotograph width:

380 µm).

Their Co concentrations determined by EMPA are generally low (≤ 0.12 wt%) and are

possibly associated with relatively large errors due to concentrations close to the detection

limit. In sample 212/2 a massive Mn-Pb phase was detected (Fig. 4.14). The optical and


85

chemical properties (average composition: 27.6 wt% O, 0.01 wt% Al, 37.7 wt% Mn, 1.72

wt% Fe, 0.37 wt% Co, 3.34 wt% Cu, 0.07 wt% Zn, 25.9 wt% Pb, and 96.7 wt% total based

on eight EMPA analyses, Appendix D, Table D5) are indicative of coronadite (ideal

formula Pb[Mn4+,Mn2+]8O16, a mineral typically found in Mn-rich parts of oxidation zones

(WEBMINERALS 2004). Significant substitution for metals on the Mn positions is reported

by SCOTT (1992).

00.10.20.30.40.50.60.7

30 35 40 45

Mn [%]

Co

[%]

0

1

2

3

4

5

30 35 40 45

Mn [%]C

u [%

]

Fig. 4.15: Co and Cu correlation observed in coronadite.

Co is significantly enriched (mean 0.37 wt%) and negatively correlated to Mn (Fig. 4.15).

It is interesting to note that Cu concentrations are also very high (Fig. 4.15), while they are

only 1.08 wt% in the goethite and hematite of the direct surrounding. Zn is low (mean 0.07

wt%), but generally low in the surrounding area. Thus coronadite is capable to selectively

scavenge Co and Cu.

Fig. 4.16: Plumbojarosite in the supergene ore surrounded by goethite-rich precipitates.

Transmitted, linear polarized light. Microphotograph width: 1 mm.


86

Plumbojarosite: Plumbojarosite (with an ideal composition of 29.64 wt% Fe, 18.3 wt% Pb,

11.3 wt% S, 39,6 wt% O, and 1.07 wt% H) was detected in section F1a (with a major

element composition of 26.7 wt% Fe and 18.7 wt% Pb (O not determined) as determined

by EMPA and 212/2 (with a major element composition of 23.9 wt% Fe, 18.8 wt% Pb, and

32.8 wt% O as determined by EMPA; Fig. 4.16).

Copper can be incorporated into plumbojarosite via substitution for Fe and forms a

complete solid solution series along the plumbojarosite - beaverite (PbFe6[(OH)6|(SO4)2]2 -

PbCu(Fe,Al)2[(OH)6|(SO4)4]) join (JAMBOR & DUTRIZAC 1985). In acidic environments

like e.g. acid mine drainage (AMD) systems, plumbojarosite and argentojarosite often play

an important role for the fixation of Pb and Ag liberated by the decay of primary sulfides,

even though they do not occur in great modal abundances (HOCHELLA et al. 1999). Indeed,

the Pb and Ag concentrations in these secondary phases can be so high that they can be

used as ore (DUTRIZAC & JAMBOR 2000). Jarosite s.s. can incorporate up to 2.5 wt% Zn

(SCOTT 1987). The observed Cu contents in Sanyati vary from 1.14 wt% to 0.22 wt% and

the Zn concentrations are low with a maximum of 0.4 wt% Zn.

Frequently observed Cu and Zn phases: The mineral composition of some frequent base

metal phases (olivenite, hemimorphite, chrysocolla, malachite, adamite, low Cu adamite,

Co adamite, and clinoclase) occurring at Sanyati was also investigated with EMPA (Table

4.7).

Fig. 4.17: As, Cu, and Zn contents (wt%) of arsenates from Sanyati compared to their

ideal composition. For abbreviations see Table 4.7. A wide range of adamite

compositions is observed. The highest observed Co content in cobalto-adamite is

0.21 wt%.

0102030405060708090

100

Ad Ad Ad Ad Ad Ad idealAd

Cl idealCl

Ol idealOl

As2O5

ZnO

CuO


87

Generally it can be observed that the compositions of the detected phases are close to their

ideal compositions and that their trace metal concentrations are low. One exception is the

slightly higher Zn concentration observed in chrysocolla. Adamite, cupro-adamite, and

cobalto-adamite form continuous solid solution series (Fig. 4.17).

Table 4.7: Trace metal contents observed in frequently occurring base metal phases in the

supergene ores of Sanyati.

Mineral name Mineral formula Co [wt%]

Cu [wt%]

Zn [wt%]

As [wt%]

Pb [wt%]

olivenite (Ol) Cu2[OH|AsO4] 0.003 0.2 0.01 clinoclase (Cl) Cu3[(OH)3|AsO4] 0.001 0.03 0.09 adamites (Ad) (Zn,Co,Cu)2[OH|AsO4] var. malachite Cu2[(OH)2|CO3] 0.001 0.22 0.005 0.08 chrysocolla (Cu,Al)2H2[(OH)4|Si2O5]*nH2O 0.003 1.45 0.006 0.2 hemimorphite Zn4[(OH)2|Si2O7]*H2O] 0.0001 0.0032 0.03 0.014

4.2.4 Composition of groundwater in the open pits

At the base of the open pits Copper-Queen (in 38 m depth) and J-Body (in 16 m depth) the

groundwater table reaches the surface. The height of the watertable varies seasonally, as

can be seen from the occurrence of efflorescences at different levels (FREI & GERMANN

2001a). The presently continuing weathering reactions are reflected in the chemistry of the

groundwater (pH = 3) with significant concentrations of Cu and Zn from sulfide decay, and

Ca and Mg from the decay of predominantly metadolomite found in close vicinity, as well

as subordinately amphibolites and phyllites (Table 4.8).

Table 4.8: Chemical composition of water samples from the bases of Copper-Queen and J-

Body open pits.

[mg/l] Cu Zn Co Pb Cd As Fe Na K Si Ca Mg Al Mn Copper-Queen 69 290 7.4 1.08 0.84 <0.5 <0.1 4.1 6.5 17 150 115 1 22 J-Body 10 165 0.81 0.84 0.27 <0.5 <0.1 111 25 33 415 419 <0.5 9.3

4.2.5 Neoformation of sulfates in the open pits

Although neo-formed sulfates are ephemeral, i.e. they dissolve easily, they provide clues

on pathways of sulfide oxidation and alteration of associated minerals (JAMBOR et al.

2000a). Most efflorescences in the open pits were found as thin white crusts in direct


88

vicinity of the groundwater. One exception was a several cm large occurrence of light blue

chalcanthite on the east wall of the Copper-Queen open pit.

Efflorescences near the groundwater table are Mg- and Ca-rich sulfates as listed in Table

4.9. Additionally the lead sulfate anglesite was found. The presence of minor amounts of

base metal sulfates is likely, however, they could not be detected due to peak interferences

during XRD analysis.

Table 4.9: Minerals constituting the efflorescences found in the open pits of Sanyati.

Mineral name Formula VI/C Water-bearing sulfates without unfamiliar anions

Occurrence

chalcanthite CuSO4 * 5H2O VI/C.04 chalcanthite group at pit wall pentahydrite MgSO4 * 5H2O VI/C.04 chalcanthite group near groundwater starkeyite MgSO4 * 4H2O VI/C.03 rozenite series near groundwater gypsum CaSO4 *2H2O VI/C.22 bassanite-ardealite series near groundwater anglesite PbSO4 VI/A.09 barite group near groundwater

4.3 Summary Chapter 4

In general it can be concluded that the weathering products as well as the base metal

concentrations vary gradually from the proximal contact zone of the mineralisation to the

distal surrounding host rocks. In the distal host rocks they are indicative for the basal part

of a weathering profile that developed under warm and humid conditions.

In the zone proximal to the mineralization, the weathering was enhanced due to the sulfide

decay. Most phyllites are weathered to the kaolinite stage and most amphibolites are

chloritized almost completely. The acidic sulfate-rich solutions allow alunite precipitation.

In the mineralized zone neutralisation is partly due to the presence of Fe-dolomite, which

is dissolved under these conditions. The minerals of the phyllites and amphibolites weather

(and thus contribute to neutralisation) comparatively slowly and very slowly, respectively,

according to the classification of JAMBOR & BLOWES (1998).

The sulfide ore is the source for the metal enrichment in the oxidation and supergene

enrichment zones. Its distribution is texturally and modally highly variable (BAHNEMANN

1961). This is reflected in the variable composition of the oxidation zone. The sulfides in

Sanyati are expected to decay in the following order: pyrrhotite >> sphalerite >

chalcopyrite > bornite = arsenopyrite > galena > pyrite > chalcocite > covellite (based on

BOWELL & BRUCE 1995).


89

Secondary sulfide ore occurs only locally. Frequent ore minerals are: chalcopyrite, pyrite,

arsenopyrite, magnetite, ilmenite, chalcocite (mainly digenite), covellite, cuprite, tenorite,

native copper and presumable orange bornite. Its mineral composition makes the secondary

sulfide ore lenses desirable for heap leaching. Indeed, the secondary sulfide ore lenses

were added to the run-of-mine ore when detected during operation of the mine, but their

volume and thus their contribution to the run-of-mine ore remains unknown.

In the supergene ore metal values are distributed in a variety of individual mineral phases

(VETTER et al. 1999), but also in the limonites. Goethite is by far the most frequent iron

phase in the limonites, followed by hematite. Jarosite/plumbojarosite is found frequently

and lepidocrocite occurs scarcely.

Several indicators for the maturity of the oxidation zone were examined (see Table 4.10),

and it can be concluded that the oxidation zone is dominantly an immature one, without a

distinct horizontal zonation. Because no depleted zone is present, it can be regarded as a

fertile gossan (NICKEL & DANIELS 1986).

Table 4.10: Indicators for maturity of the oxidation zone.

Indicator for maturity of oxidation zone: Maturity reached ? presence of metastable iron phase (ferrihydrite, schwertmannite) yes decreasing goethite/hematite ratio with respect to the surface no decreasing base metal concentations in limonite-rich samples with respect to the surface

no

jarosite/plumbojarosite distribution with respect to the surface no presence of scorodite, beside beudantite no

An attempt was made to extract a sample group, the metal budget of which is controlled

dominantly by the limonite phases and therefore is not easily liberated during the acid heap

leaching. It was shown that each of the open pits and orebodies that are planned to be

mined has an individual distribution of these metals. This observation suggests that each

orebody should be examined separately with respect to the base metal fraction bound to

limonite in order to assess Cu availability in the heap leaching process.

Beside limonite, coronadite was found to be a scavenger for Cu (max. 4.7 wt%) and Co

(max. 0.65 wt%). EMPA measurements of plumbojarosites reveal Cu contents of up to 1.1

wt% Cu and 0.4 wt% Zn, which is significantly below their saturation-concentration in

plumbojarosite (JAMBOR & DUTRIZAC 1985, SCOTT 1987). Frequent base metal phases e.g.


90

malachite, olivenite etc. also show a composition low in trace metals and thus close to their

ideal stoichiometric compositions.

The present weathering situation at Sanyati is reflected by the element load of the

groundwater and the efflorescences in proximity to water sources. Both are high in Mg and

Ca. These elements are liberated when the host rocks (dolomites and silicates) decay. Cu

and Zn, liberated from the sulfides, are enriched in the groundwater. Anglesite, a low

soluble lead sulfate, does occur in the efflorescences. Thus, sulfide decay continues at

present and no indications for recent interruptions were found.

5. Leaching products and processes on the heap leach pads

91

5 Leaching products and processes on the heap leach pads

In contrast to the natural decay processes discussed in Chapter 4, this Chapter will deal

with the ore subjected to technogene decay. The run-of-mine ore (ROM) is either stored on

an ore dump prior to leaching or is brought directly from the open pits to the heap leach

pad. Since only a limited number of ROM samples from the run-of-mine ore dump was

available, data from the supergene ore samples in the previous Chapter are additionally

used for comparison. To distinguish between ROM and the ore samples collected from the

heap leach pad that have been subject to technogene decay, the latter are called leach pad

ore (LPO) in the following.

5.1 Characteristics of the run-of-mine ore (ROM)

Grain size distribution

The extreme grain size spectrum of the ore has already been described qualitatively in

Chapter 2.5.2. A quantitative estimation of the whole grain size spectrum is beyond the

scope of this study. However, the finer fractions (< 1 cm) were sieved and the resulting

grain size distributions are plotted in Fig. 5.1. The fractions of middle- and fine-sized

pebbles and sand are almost equal. Additionally a clay fraction of 10 % was detected

during dry sieving that increases to up to 25 % when the material is sieved wet, as

agglomerated particles disintegrate.

10

100

0.01 0.1 1 10 100 1000

grain size (mm)

cum

mul

ativ

e %

ROM

LPO 0 m

LPO -1 m

LPO -2 m

Fig. 5.1: Grain size distribution of the fraction < 1 cm of run-of-mine ore (ROM sample

210) and leach pad ore (LPO samples from profile A2-2) after wet sieving.


92

Mineralogy

Samples taken from the run-of-mine ore dump were limonite-rich and did, except for two

minor exceptions, not show any visible base metal mineralisations. In some ROM samples

the texture of the host rock is preserved. Their mineral composition matches that of the

samples from the open pits (Chap. 4.2.3), with the exception that jarosites were found more

frequently in ROM and secondary silicates from alteration were rarer.

Geochemistry

As for the mineralogy, the chemistry of the ROM and the samples from the open pits

matches reasonably. Taking into account the restricted possibilities for sampling, the

supergene ore samples from all open pits and the ROM show reasonable agreement (cf.

Fig. 5.2). The contents obtained for Cu are slightly higher (+ 12 %) in this study compared

to the estimations reported by CHADWICK (1996). This is most probably reflecting a higher

portion of host rock in the estimations by CHADWICK (1996).

Fig. 5.2: Average Cu, Zn, and Fe2O3 contents of the supergene ores from the Sanyati mine:

“Munyati Mining” denotes data reported by CHADWICK (1996; no data for Fe2O3

available); supergene "Ore from all open pits", ROM and LPO (profile A2-4) are

based on own samples.

In Fig. 5.3 the relative contents of selected metals in ROM and (calculated as the mean) of

the supergene ore samples from all open pits are plotted in pie-diagrams. Both lead to

strikingly similar results and provide compelling evidence for the assumption that the

ROM closely reflects the mean metal contents in the supergene ores.

0.0

0.2

0.4

0.6

0.8

1.0

1.2

1.4

Cu Zn

wt%

Munyat i Mining

Ore from all open p it s

Run-of-m ine ore(ROM)Leach pad ore (LP O)

0

10

20

30

40

50

60

70

Fe2O3

wt%


93

Fig. 5.3: Relative contents of selected metals estimated from supergene ore samples of all

Sanyati open pits compared to the mean of samples from the run-of-mine ore

dump.

Fig. 5.4: Cu, Zn, As, Co, and Pb contents in the grain size fractions of ROM (sample 210).

The grain size fractions of the sieved ROM show significant variations in base metal

contents (Fig. 5.4). Because of the restricted amount of sample the coarser fraction bears a

larger statistical error. However, it is interesting to note that the finer fractions are

characterised by high base metal contents in the fraction < 63 µm (Fig. 5.4). If the metals

present are easy leachable (i.e. if they are not bound to phases that retain considerable

amounts of metals during leaching) these fine fractions of an ore have an optimal liberation

size for leaching (BARTLETT 1992).

3000

4000

5000

6000

7000

8000

9000

>2>0,71

>0,25

>0,063

<0,063 m m

mg/

kg

Cu

Zn

As

0

50

100

150

200

250

300

>2 >0,71 >0,25 >0,063 <0,063

Co

0

5000

10000

15000

20000

25000

30000

>2 >0,71 >0,25 >0,063 <0,063

P b


94

5.2 Composition of the leach pad ore (LPO)

Grain size distribution

The grain size spectrum of the material on the heap leach pad ranges from ~ 1 m to clay

size. Obviously it was assumed by the operators that the permeability of the blocks was

large enough to allow sufficient acid percolation. This assumption has to be considered as

an oversimplification because even blocks of about 20 cm diameter in 2 - 3 m depth, that

have been leached for several years in a position with intense acid contact, still show Cu

mineralisations.

The leach pad ore of profile A2-2 was sieved (Fig. 5.1) using the same procedures that

have been employed for the ROM. Dry sieving resulted in similar grain size distributions

observed for ROM. In wet sieving, however, the clay fraction in two samples increases

from 25 % to values as high ~ 45 %. This suggests that the dry LPO material is either

agglomerated or preserved in the original grain size and disintegrates upon fluid contact

and mechanical treatment.

A direct consequence of the presence of a large fine grained ore fraction is the

development of puddles (inhabited by alga and bacteria colonies, Fig. 5.5) on the heap

leach pad because the acid does not drain fast enough. This in turn leads to enhanced acid

evaporation. These “sealed” surfaces are occasionally opened using heavy machinery, but

it must be assumed that the use of heavy machines to rip open the surface in turn leads to

compaction at lower levels.

Fig. 5.5: Puddles (inhabited by algae and bacteria colonies) developed on the heap

leaching pad because of insufficient acid drainage.


95

Mineralogy

Generally, the mineral composition of the leach pad ore does not differ much from the run-

of-mine ore (Table 5.1). Neoformations from the solution are discussed separately in

Chapter 5.4. The most important difference observed is the relative frequency of the

ironoxides and -oxyhydroxides. Goethite is by far the most dominant mineral and

jarosite/plumbojarosite is more frequent, while hematite occurs less often in LPO.

Goethite is thermodynamically slightly less stable with respect to hematite and water

(BERNER 1969). Indeed, under acidic conditions goethite may replace hematite (BROWN

1971). Especially the presence of Al favours goethite precipitation (TAYLOR &

SCHWERTMANN 1978). This explains the relatively higher abundance of goethite to

hematite in LPO.

Jarosites can develop in naturally weathered mine tailings after only a few years if sulfides

oxidize (DUTRIZAC & JAMBOR 2000). Since the acidic and oxidizing conditions are broadly

comparably to those on the heap leach pad, additional formation of jarosites as observed is

likely.

Table 5.1: Minerals present in run-of-mine ore (ROM) and leach pad ore (LPO).

Minerals Run-of-mine ore Leach pad ore Fe-phases: magnetite + hematite + (+) goethite + + jarosite/plumbojarosite + + primary silicates: quartz + + chlorite + + tremolite / actinolite + cummingtonite / anthophyllite + talc + + weathering products: illite (+) smectite (+)

The grain size distributions observed in LPO samples from profile A2-2 show a relatively

low goethite content in the fine fractions. Since goethite can scavenge and fix base metals

this might indicate good characteristics for metal release during leaching. However, in

order to assess the leachability of this ore fraction the base metal-bearing phases have to be

qualified and quantified.


96

Geochemistry

The chemical composition of the leach pad ore has to be determined after the water-soluble

crusts have been removed by washing, as direct analysis would include the metal freight of

the dried acid used for leaching (see Chap. 5.4). The effect of washing (and hence the

amount of metals originating from evaporated leaching acids) is visualised in Fig. 5.6 for

Cu and Zn, where the element concentrations of LPO from profile A2-2 are plotted prior to

(dry sieving) and after washing (wet sieving). The dry sieved fractions show a more or less

homogeneous Cu and Zn distribution for all fractions. This is due to the agglomeration of

finer particles with low Cu and Zn contents, which dilute the coarser fractions. For

comparison, in wet sieved fractions the Cu and Zn content decreases by 82 % for Cu and

78 % for Zn from the > 2 mm to the < 63 µm fraction. Additionally, the mean Zn contents

of the wet sieved samples are 56 % lower than the mean Zn contents of dry sieved samples.

This is attributed to the easily soluble sulfate crusts.

Fig. 5.6: Cu and Zn concentrations of the dry and wet sieved fractions of profile A2-2

(grain sizes reported in mm).

0

10002000

30004000

5000

60007000

8000

>2>0.7

1>0.2

5>0.0

63

<0.063 >2

>0.71

>0.25

>0.063

<0.063

wet dry

[mg/

kg]

0

1000

2000

3000

4000

5000

6000

>2>0.7

1>0.2

5>0.0

63

<0.063 >2

>0.71

>0.25

>0.063

<0.063

wet dry

[mg/

kg]

Zn

Cu


97

For Cu the crusts are not enriched to such an extreme extent, because Cu is extracted from

the system regularly (see Chap. 4.2.5).

The maximal and minimal contents of selected metals in the grain size fractions A (100 - 2

mm) and B (< 63 µm) and their ratio (= max./min.) as an indicator for the depletion are

listed in Table 5.2. Clearly, in the finer fractions the depletion of Cu, Zn, As, Co, and

Fe2O3 is enhanced with factors varying between 2.2 for Fe2O3 and 8 for Co.

A different behaviour is observed for lead (Pb concentrations are widely independent of

the grain size), which suggests that the weakly soluble Pb-phases show only little decay

during leaching, and consequently the depletion factor is with ~ 2.6 rather low.

Table 5.2: Fe, As, Cu, Zn, and Co concentrations of the fractions 100 - 2 mm (A) and

<0.063 mm (B) of the wet sieved samples from profile A2-2. The higher metal

recovery from the finer fraction can be estimated from the max./min. ratio.

A2-2 0 m A2-2 1 m A2-2 2 m A max. B min. max./ min. A max. B min. max./ min. A max. B min. max./ min. [wt%] [wt%] [wt%] [wt%] [wt%] [wt%] Fe2O3 26.4 8.3 3.1 34.8 12.4 2.8 28.2 12.7 2.2 As 1.2 0.3 4.0 1.3 0.4 3.3 1.2 0.5 2.4 Cu 0.6 0.1 6.0 0.4 0.1 4.0 0.6 0.1 6.0 Zn 0.5 0.1 5.0 1.0 0.3 3.3 0.7 0.2 3.5 [mg/kg] [mg/kg] [mg/kg] [mg/kg] [mg/kg] [mg/kg] Co 48 10 4.8 105 13 8 78 17 4.6 Pb 0.4 0.15 2.7 2.1 0.8 2.6 1.3 0.5 2.6

However, an average residual Cu content of 0.6 wt% was determined in LPO of the profile

even after several years of leaching (cf. Fig. 5.2). Thus, on the basis of 1.1 wt% Cu in the

ROM (CHADWICK 1996), the recovery for this fraction is only 50 % and is probably even

lower for the coarser fractions > 100 mm. The average contents of Zn (cf. Fig. 5.2) and

also of Co, As, and Mn (not shown) do not show significant variations before and after

leaching. This observation shows that the metals which are retained in the system (i.e. that

are not extracted from the leaching solutions) are either not mobilized to a great extent

during leaching or are quantitatively reprecipitated in the heap leach pad during

recirculation of the acid. Within the heap leach pad, e.g. in the trenches drawn for

sampling, no signs for precipitations (e.g. hard pans) were found, except for the

efflorescences that develop upon drying on all exposed surfaces of the heap leach pad

(Chap. 5.4). However, not further specified precipitations were reported at 6 - 8 m by

MAGOMBEZE & SANDVIK (2002).


98

Even though the frequency of occurrence of the iron minerals changes from ironoxides to

ironoxyhydroxides and -sulfates, iron is not leached to a large extent (Fig. 5.2). Compared

to the Fe2O3 content of all open pits (52 wt%, cf. Fig. 5.2) on average 42 wt% is retained in

the fractions of leach pad ore examined and it is likely that this value is in reality higher if

the coarse fractions of the leach pad ore are considered.

A comparison of the sample compositions with regard to their vertical position in the heap

leach pad shows that the metal output is dominantly controlled by the grain size and the

initial composition and no systematic vertical trend is developed (Fig. 5.7). The fraction >

2 mm contains between 67 % and 82 % more Cu compared to the fraction < 63 µm. For Zn

the fraction > 2 mm is 66 to 78 % higher compared to the < 63 µm fraction (Fig. 5.7).

Fig. 5.7: Cu and Zn concentrations in LPO samples from profile A2-2 with regard to their

position in the heap leach pad and their grain size.

Of the main elements, Si is enriched in the finer fractions, as it is bound to relict phases

e.g. quartz. Mg and Al distributions are independent of grain size. This shows that quartz

and silicates are not dissolved to a great extent. Thus, the decomposition of the ore

proceeds along the grain boundaries of the rockforming minerals, which are generally

coated by limonite.

0 5000 10000

<0.063

>0.063

>0.25

>0.71

>2.00

-2 m:

<0.063

>0.063

>0.25

>0.71

>2.00

-1 m:

<0.063

>0.063

>0.25

>0.71

>2.00

surface

mg/kg

Cu

0 5000 10000 15000

<0.063

>0.063

>0.25

>0.71

>2.00

-2 m:

<0.063

>0.063

>0.25

>0.71

>2.00

-1 m:

<0.063

>0.063

>0.25

>0.71

>2.00

surface

mg/kg

Z n


99

5.3 Composition of the acid solution used for leaching

The sulfuric acid, which is recirculated from the sprinklers through the heap leach pads to

the SX plant and back onto the heap leach pads (Fig. 2.9), was sampled prior to the passage

through the pads (“acid at sprinkler”) and after the passage through the ore (“acid at heap

drainage”).

In the course of one passage the Cu concentration increases by the factor 3.7 (Fig. 5.8).

The remaining elements (Fig. 5.8) are strongly accumulated in the circulating acid during

numerous passages through the heap, because they are not extracted from the solution.

Especially the base metals Zn (18200 mg/l) and Co (370 mg/l) show high concentrations in

both acid at sprinkler and acid at heap drainage. This is important with regard to a potential

economic use of these metals. Also Mn, a potential metal value, is highly concentrated

(6400 mg/l) in the acid solutions.

Fig. 5.8: H2SO4 (pH ~ 1.5) is used to leach the Cu from the supergene ore. The acid is

recycled after the solvent extraction of Cu and shows high concentrations of Zn

and Co, as well as Mn, Al, Mg, and Ca. These elements which are, unlike Cu, not

extracted do not show any change in concentration after the passage through the

heap leach pad. Elevated Zn, Co, and Mn are of interest for a potential economic

use in the future.

Observed concentrations of the toxic elements Pb, As, and Cd are below 10 mg/l (Fig. 5.8),

indicating that their host phases are not solved to a large extent. Possible sinks for these

elements are plumbojarosite and not easily soluble arsenates as well as coprecipitation in

goethite and hematite. In geological timescales plumbojarosite decomposes readily when

removed from its acidic oxygen and sulfate-bearing stability field and decays to goethite

(DUTRIZAC & JAMBOR 2000). However, under the acidic conditions prevailing on the heap

1

10

100

1000

10000

mg/

l

Fe Na Si Ca M g A l M n

acid at sprinkler

acid at heap drainage

1

10

100

1000

10000

100000

mg/

l

Cu Zn Co Pb Cd A s

acid at sprinkler

acid at heap drainage


100

leach pad plumbojarosite is a stable phase. Elements present in the acid solution derived

from silicate and calcite decay are Na, Si, Ca, Mg, and Al. Iron concentrations are kept as

low as possible, because Fe causes problems in the SX-EW plant (cf. "iron poisoning";

CHADWICK 1996). The Fe concentrations are minimized by control of the pH (see Chap.

1.2.2.2) and do not exceed 32 mg Fe/l in the leaching cirquit. This pH control though

restricts the pH to values above 1.5 and prevents a possible enhancement of Cu-recovery

by leaching with a higher acidity.

At the start of a new acid circulation, the concentration of all elements that are not

extracted increased, but reached a steady state relatively soon (R. MUSHANGWE, pers.

comm. 1999). Because of the lack of experimental solubility data in highly acidic and

sulfate-rich systems a qualitative and quantitative interpretation of the observed metal

loads of the acids is impossible.

5.4 Neoformation of phases during leaching

Because of its high metal load white crusts or “efflorescences” of numerous soluble

sulfates precipitate wherever the acid is subjected to evaporation. Cm-thick crusts were

found on parts of the heap leach pad with a dry surface (Fig. 5.9) and in the emergency

pond where surplus acid is stored and dries up (sample 284). The samples Ausb1 and

Ausb2 (Table 5.3) precipitated in the sampling excavation in area 2 after only a few days

from stagnate solutions.

Fig. 5.9: a. View on dry areas of heap leach pad. b. SEM picture of soluble sulfates

(sample 288).


101

Most of these efflorescences are formed by gunningite and gypsum (Table 5.4).

Additionally, from stagnate acid solutions the Mg, Fe, Zn, Al, and Mn-rich sulfates of the

halotrichite and rozenite group precipitated (Table 5.3, Table 5.4). The mineralogy of

samples nr. Ausb1 and Ausb2 is dominated by Mg and Fe sulfates.

Table 5.3: Mineralogical composition of "efflorescences" sampled on the heap leach pad

and emergency pond.

Halotrichite-group Rozenite-series gunningite

Zn, Mn

gypsum Ca

pickeringite

Mg, Al

halotrichite

Fe, Al

apjohnite

Mn Al

rozenite

Fe

starkeyite

Mg

boyleite

Zn, Mg

alunogen

Al Ja + + Jbb + + Jbw + + + + + Ausb 1 + + + Ausb 2 + + + 284 + + + 288 + + 293 + + 294 + +

Gunningite, a hydrous Zn and Mn sulfate of the kieserite group, is frequently occurring in

the efflorescences. It was previously thought to occur only rarely, but is now increasingly

reported (e.g. GRYBECK 1976; AVDONIN 1984; AVDONIN et al. 1988). Gunningite is also

found on waste dumps of the sulfuric acid industry, where it represents the soluble fraction

of the deposited material (LIN 1997; LIN & QVARFORT 1996). Rozenite and starkeyite are

the most common members of the rozenite group. Alunogen is by far the most common of

the trivalent cation sulfate salts and halotrichite and pickeringite are the most abundant

members of the halotrichite group (JAMBOR et al. 2000a).

Table 5.4: Mineral names and formulas of the "efflorescences" detected on the Sanyati

heap leach pads and emergency pond (classification by STRUNZ & NICKEL 2001).

Mineral name Formula VI/C Water-bearing sulfates without unfamiliar anions

gunningite (Zn,Mn)SO4·(H2O) VI/C.01 kieserite-group starkeyite MgSO4·4(H2O) VI/C.03 rozenite-series rozenite Fe2+SO4·4(H2O) VI/C.03 rozenite-series boyleite (Zn,Mg)SO4·4(H2O) VI/C.03 rozenite-series alunogen Al2(SO4)3·17(H2O) VI/C.04 rhomboclase-meta-alunogen series pickeringite MgAl2(SO4)4·22(H2O) VI/C.06 halotrichite-group halotrichite Fe2+Al2(SO4)4·22(H2O) VI/C.06 halotrichite-group apjohnite MnAl2(SO4)4·22(H2O) VI/C.06 halotrichite-group gypsum CaSO4·2(H2O) VI/C.22 bassanite-ardealite series


102

From tailings in arid climates numerous soluble sulfates have been reported (JAMBOR et al.

2000a). Simple hydrated sulfates are commonly accompanied by more complex phases,

e.g. halotrichite (SHCHERBAKOVA & KORABLEV 1998; JAMBOR et al. 2000b). In sulfidic

tailing impoundments Fe sulfates dominate over Mg sulfates, which form from gangue

minerals. In contrast, the occurrence of e.g. Zn sulfates reflects the composition of the

oxidizing source (AVDONIN et al. 1988). Fe2+ sulfates can scavenge Cu and Zn, while As is

captured by Fe3+ sulfates (GIERÉ et al. 2003).

Even though the sulfate efflorescences on the heap leach pads are readily water-soluble,

they act as temporary “sinks” for metals, and the composition of these precipitates can be

used as a proxy for the composition of the circulating acid used for leaching. The chemical

composition of three representative samples is reported in Table 5.5.

The chemical composition of the efflorescences is dominated by Zn (~ 100 mg/kg), with

additionally Mg, Al, Mn, and in one case Ca being observed at higher concentration levels

(in the range of ~ 10 ml/kg). Gunningite contains Zn and Mn, apjohnite shows extensive

solid solution and can incorporate Mg besides Al and Mn (PAULIS 1991). Al is bound in

alunogen and Ca in gypsum. As expected, both Cu and Fe concentrations are low (Table

5.5).

Table 5.5: Chemical composition of three representative soluble sulfate efflorescence

samples from Sanyati (samples 288g, 288r, 284).

Cu Zn Co Pb Cd As Fe Na K Si Ca Mg Al Mn [g/kg] 288g (area 6) 4.5 107 1.5 <0.1 0.43 <0.5 1.8 1.1 0.1 <0.1 1.7 10.0 16.6 55.4288r (area 6) 1.2 90 1.8 <0.1 0.19 <0.5 0.4 <0.1 0.1 <0.1 0.3 10.5 18.0 29.7284 (em. pond) 1.4 112 2.2 <0.1 0.29 <0.5 1.7 0.3 0.1 <0.1 15.8 11.8 19.5 43.1

A comparison of the relative metal contents of the efflorescences and the acid shows very

similar compositions for Zn, Al, Mn, and Mg (Fig. 5.10). The only exception is an

increased Cu content in sulfates on the heap leach pad, demonstrating that these sulfates

can act as minor temporary sinks for Cu.

In efflorescences sampled from the emergency pond (sample 284), the Cu proportion is

very low as this solution has been Cu-depleted in the SX plant prior to deposition. The


103

high Ca proportion in this sample indicates a higher modal contribution of dolomite

dissolution in an earlier stage of leaching on the heap leach pad (Fig. 5.10). A possible

explanation for this high Ca proportion is that ores from the Copper-Queen open pit

deposited at the beginning of the mining operation contained a relatively high amount of

dolomite. This would have been a disadvantage for leaching as acid is neutralised by

carbonate dissolution.

The Fe proportions are slightly higher in the efflorescences compared to the acid,

indicating that some Fe is retained in the crusts, i.e. in rozenite and halotrichite.

Fig. 5.10: Relative metal compositions of soluble sulfates precipitated on the heap leach

pad, the emergency pond, and the acid used for leaching prior to and after the

passage through the heap leach pad.

The samples taken along the profiles on the heap leach pad are also covered by thin white

crusts. These crusts were dissolved from the leach pad ore with demineralized water (for

the procedure see Chap. 3.2.2.1). They consist of Cu, Zn, and Co sulfates that are all

readily water-soluble as depicted in Fig. 5.11.


104

Fig. 5.11: Concentration of Cu, Zn, and Co in the water-soluble fraction of samples of

profile A2-4. Samples were washed with nine 1 l aliqouts of fresh, demineralized

water and every second aliquot was analysed. After nine washings the water-

soluble metal fraction of the leached ore was removed.

The water-soluble fraction preserves information about the metal load of the acid used on

the heap leach pad. In Fig. 5.12 the Cu, Zn, and Co contents of the water-soluble fraction

are plotted as a function of their distance to the surface in the heap leach pad. All three

metals show a characteristic depth gradient: During the passage through the first meter the

Zn and Co concentrations are not enriched significantly, and the Cu concentrations even

decrease slightly. This decrease can be explained by a concentration through evaporation

prior to the precipitation of efflorescences at the heap top. Below 1 m the concentrations

increase because the acid solves base metals from the ore during the passage through

deeper parts of the heap leach pad.

Fig. 5.12: Depths variation of Cu, Zn, and Co concentrations in the water-soluble fraction

of samples in a vertical profile through the heap leach pad (profile A2-2).

Cu

0.0

0.5

1.0

1.5

2.0

0 1 2 3 4 5 6 7 8 9 10

H 20 litre nr.

[mg/

l]A2-4a 0m

A2-4b 0m

A2-4 -0 .5m

Zn

0

2

4

6

8

10

12

14

16

0 1 2 3 4 5 6 7 8 9 10

H20 lit re nr.

[mg/

l]

A2-4a 0m

A2-4b 0m

A2-4 -0 .5m

Co

0.0

0.1

0.1

0.2

0.2

0.3

0.3

0 1 2 3 4 5 6 7 8 9 10

H20 lit re nr.

[mg/

l]

A2-4a 0m

A2-4b 0m

A2-4 -0 .5m

-2

-1.5

-1

-0.5

0

100 200 300 400 500 600

0 2 4 6 8 10

ZnCuCo

Dep

th [

m]

Zn[mg/l]

Cu, Co[mg/l]


105

The concentrations in profile A2-2 (Fig 5.12) are higher than in A2-4 (see Appendix C,

Table C3), even though both profiles are taken only 8 m apart. Zinc concentrations are the

most variable, with up to 270 mg/l in profile A2-2 and one order of magnitude lower

contents in profile A2-4 (max. 14.2 mg/l).

Metals liberated from decaying ores on mine waste tailings may precipitate locally as

either soluble or relatively insoluble sulfate minerals, with the latter acting as solid-phase

controls on dissolved metal concentrations (JAMBOR et al. 2000a). This observation is also

true in the leaching circulation system at Sanyati, where the efflorescences act as

temporary sinks and control the mobility and fixation of metals in the circulation system.

Upon their dissolution, which takes place when dried areas are leached again or during

rainfall, the solution freight is immediately increased considerably (BIGHAM &

NORDSTROM 2000).


The mineralogy and geochemistry of the ROM samples coincide satisfyingly (with an error

of ~ 10 %) with the mean of the supergene ore samples from all the open pits, even though

only a limited number of samples has been available.

Grain size distributions of wet sieved samples of ROM and LPO showed that the

contribution of the fraction < 63 µm is significant in ROM (25 %) and even larger in LPO

(45 %; Fig. 5.1). One direct consequence of this increased fine fraction on the heap leach

pad is the development of puddles due to decreased permeability of the ore. This in turn

leads to enhanced evaporation of the acid used for leaching.

Generally, the metal contents in ROM decrease with decreasing grain size (Fig. 5.4). In

contrast, the contents of the base metals Pb and As increase again in the fraction < 63 µm.

The mineralogical composition of ROM and LPO is similar with respect to the occurrence

of major phases. However, the modal abundance of goethite and plumbojarosites/jarosites

(most likely due to neoformation) is higher and of hematite lower (most likely due to the

conversion to goethite) in LPO.

On average only 50 % of Cu are leached from the fractions < 100 mm (Fig. 5.2), an effect

that is most likely to be more enhanced in coarser fractions. For the metals which are not

extracted, e.g. Zn, the average contents are close to those of ROM (Fig. 5.2). This shows


106

that these metals are only scarcely leached, due to the saturation of leaching solution with

these elements.

The acid used for leaching and the efflorescences are temporary sinks for the various

metals in the circulation system. In the efflorescences significant metal enrichment is

found for Zn (gunningite) and to a lower extent for Mn (gunningite, apjohnite). Zn (18200

mg/l) is also enriched in the leaching acid together with other metal values, e.g. Co (370

mg/l) and Mn (6400 mg/l). The toxic elements Pb, As, and Cd are generally retained in the

LPO, where they are either not mobilized or reprecipitated as low soluble sulfates (e.g.

plumbojarosite) and arsenates. Cu, as the only element not in steady-state in the system, is

enriched in the acid and regularly extracted from it and purified in the SX plant.

6. Composition of ironoxides and -oxyhydroxides in run-of-mine and leach pad ore

107

6 Composition of goethite and hematite in run-of-mine and leach pad ore

6.1 Chrystal chemistry of goethite and hematite

Isomorphic replacement of metals for Fe3+ has been observed in natural and synthetic

phases, but most of these solid solutions have broad miscibility gaps. The structure of α-

FeOOH is adopted by α-AlOOH (diaspor), α-MnOOH and α-CrOOH and synthetic

isomorphs are ScOOH and GaOOH (CORNELL & SCHWERTMANN 1996). The structure of

α-FeOOH is composed of double chains of FeO3(OH)3 octahedra parallel to the c-axis

(Fig. 6.1). The mean Fe-O bond length in goethite is 2.021 Å, leading to an optimal cation

size of 0.661 Å (assuming an ionic radius of 1.36 Å for oxygen) close to the ionic radius of

Fe3+ in 6-fold coordination of 0.645 Å (high-spin state) as determined by SHANNON (1976).

Other metals can be incorporated into the goethite lattice by substitution for iron on this

octahedrally coordinated position (SMYTH & BISH 1988).

Fig. 6.1: Structure of α-FeOOH. The double chains for the FeO3(OH)3 octahedra run

parallel to the c-axis. (OH groups are indicated while the unmarked corners of

the octahedra represent oxygen (from KLEIN & HURLBUT 1993).

Hematite, the non-hydrated equivalent to goethite, is composed of basal sheets of

octahedra with one octahedron vacant for every two octahedra with Fe3+ in center (KLEIN

& HURLBUT 1993). The mean Fe-O bond length in hematite is 2.030 Å, leading to an

optimal cation size of 0.670 Å (assuming an ionic radius of 1.36 Å for oxygen). Hence, the

optimal cation size for the Fe position between goethite and hematite differs by only 0.009

Å, suggesting that cation substitution for Fe in both minerals should generally follow the


108

same pattern. Furthermore, only very few electron microprobe analyses are indicative for

the presence of hematite, whereas the overwhelming majority of all analyses points to

goethite as the dominant Fe-bearing phase (cf. Fig. 6.3). Therefore, only the extent of

possible cation substitutions for Fe into goethite is discussed in more detail in the

following.

Table 6.1: Cations substituting for Fe3+ in goethite - maximum substitution and unit cell

parameters (compiled from CORNELL & SCHWERTMANN 1996; SHANNON 1976;

SMYTH & BISH 1988).

Ionic radius [Å]

Maximum substitution [mol/mol]

Unit cell edge length b0 [Å]

Cell volume [Å3]

Fe 3+ 0.645 - 9.937 137.4 Al 3+ 0.535 0.33 9.800 133.3 Mn 3+ 0.645 0.15 10.06 137.3 Co 3+ 0.545 0.1 9.119 136.7 Cu 2+ 0.73 0.05 10.013 139.3 Zn 2+ 0.74 0.07 10.07 140.0 Cd 2+ 0.95 0.07 10.067 142.6 Pb 2+ 0.775 0.02 9.977 139.3

Theoretically ions with a 3+ valence and a size of ±18 % of the size of Fe in high-spin state

can be tolerated in goethite (CORNELL & SCHWERTMANN 1996). The relationship between

the orthorhombic cell volume and the ionic radius of the Me3+ cations (Al, Co, Cr, Ga, V,

Mn, Fe, and Sc) is linear for some ions, but deviates due to extra OH- in other cases

(CORNELL & SCHWERTMANN 1996). Of the trivalent transition metals Mn3+, having a

similar ionic radius to that of Fe3+ (0.645 Å), substitutes best with up to 0.15 mol Mn/mol

goethite (Table 6.1).

However, because Mn3+ has a tetragonally distorted coordination sphere on an octahedral

site owing to its four d electrons (Jahn-Teller effect), Mn3+ does not fit as readily into the

goethite structure as Fe3+. The distortion of the Mn3+ coordination sphere is reflected in an

increase in the b0 edge length and a decrease in the c0 edge length.

Me2+ cations uptake is also reported but is usually below 0.1 mol Me/mol goethite, due to

the required coupled substitution to maintain charge balance.

The degree of isomorphic substitution for Fe can be examined by observation of the


109

change in unit cell dimensions from XRD line shifts. The unit cell dimensions of metal

substituted goethites from the open pits in Sanyati were calculated from XRD spectra

according to a procedure originally used for Al substitution by SCHULZE (1984). The

relative deviations of the unit cell edge length of the b0 axis and the unit cell volume v0 of

goethites from the open pits in Sanyati to those of pure goethite are plotted in Fig. 6.2.

Also plotted is pure goethite (= zero deviation) and the relative deviations observed for

goethites showing the highest possible substitutions of Al, Mn, Cu, and Zn for Fe

(according to CORNELL & SCHWERTMANN 1996).

Fig. 6.2: Deviation of the unit cell edge length of the b0 axis and the unit cell volume v0 of

base metal containing goethites from Sanyati to those from pure goethite

(expressed as relative deviation from pure goethite). Also depicted are the

deviations observed for endmember compositions with the maximum observed

substitutions of Al, Mn, Cu, and Zn for Fe (calculated from values given in Table

6.1).

All samples from Sanyati plot along an array typical for highly Cu- and Zn-substituted

goethites (Fig. 6.2). Indeed, the investigated goethites show substitutions close to the

maximal substitutions possible for these two metals. Hence, a significant proportion of Cu

and Zn is incorporated into the crystal lattice of the goethites examined.

Therefore, in order to study the element distributions in their textural context in goethite

and hematite in ROM and LPO, representative samples were examined in detail by EMPA.

A restricted set of samples was additionally analysed by LA-ICP-MS.

-2.0

-1.5

-1.0

-0.5

0.0

0.5

1.0

1.5

2.0

-4.0 -3.0 -2.0 -1.0 0.0 1.0 2.0 3.0 4.0

relative deviation of v 0 [%]

rela

tive

devi

atio

n of

bo [%

]

Fe

Al

Mn CuZn


110

6.2 Boxwork texture and chemistry (microprobe analysis)

6.2.1 Development, preservation, classification and geochemistry of relictic decay

textures of sulfides

Under favourable conditions weathering and decay of sulfide assemblages result in the

development of characteristic decay textures in the evolving supergene ore. These decay

textures might be used to identify the original sulfide mineralogy. Textural classification

and interpretation on the basis of visual observation on boxworks obtained by naked eye

and by handlens were made by BLANCHARD (1968), followed by contributions from BLAIN

& ANDREW (1977), ANDREW (1980), REYNOLDS (1982), NICKEL & DANIELS (1986), and

JEONG & LEE (2003). The importance of textural evidence (based on microscopic

observations) in order to unravel the mineralogical development of weathered sulfide

assemblages was already emphasised by SCHNEIDERHÖHN (1924).

One of the goals of this study was, wherever possible, to identify characteristic textures of

former sulfides and to characterise their mineralogical, chemical, and textural development

using microscopic and microchemical techniques (i.e. ore- and petrographical microscopy,

electron microprobe, and LA-ICP-MS, respectively). In a reconnaissance study high and

bimodally distributed base metal contents of up to 1.8 wt% Cu and 5.2 wt% Zn were

observed in goethite and hematite of the supergene ores of Sanyati (FREI & GERMANN

2002b). In the following, the results of textural and microchemical investigations will be

discussed in more detail. Special emphasis was put on the comparison of textures and

chemistry in run-of-mine ore (ROM) and leach pad ore (LPO).

Conditions that favour the development and preservation of pronounced sulfide decay

textures are coarse grained and euhedral sulfide crystals, and a relatively high pH.

Especially during the early stage of the oxidation process, these conditions lead to

ironoxide and -oxyhydroxide precipitation along cleavages, fractures, and grain boundaries

before the remnants of the sulfide grains are dissolved (NICKEL & DANIELS 1986). In order

to maintain a high pH, lithologies with a high acid neutralisation capacity are needed in the

vicinity of the decaying sulfides. Early formed ironoxide and -oxyhydroxide precipitates

are retained throughout the subsequent development of the oxidation zone (NICKEL &

DANIELS 1986) and key textures developing under ideal conditions have been described for


111

the most common minerals (e.g. BLANCHARD 1968).

In the ores from Sanyati the sulfide grain sizes are predominantly middle to fine grained

and decaying textures are best preserved in those parts of the mineralisation bound to

metadolomite lenses. The metadolomite neutralises the acid produced during the early

weathering stage and buffers the pH at values sufficiently high to prevent the dissolution of

goethite and hematite.

Based on the classifications by BLANCHARD (1968) and NICKEL & DANIELS (1986) five

texture types characteristic for decay of chalcopyrite or sphalerite can be differentiated in

the ores of Sanyati (Table 6.2). Because the relicts of these two minerals are difficult to

distinguish from each other by microscopic techniques (NICKEL & DANIELS 1986) it was

evaluated whether their geochemical fingerprint can be used for distinction. Two replica

textures of iron sulfides (pyrite and pyrrhotite) are preserved in the ores (Table 6.2) and

additionally narrow parallel boxworks with different degrees of limonite mineralisation

between the replacement veinlets and colloform textures were observed (Table 6.2).

Table 6.2: Characteristic sulfide decay textures (see images Fig. 6.4 to Fig. 6.20) observed

in Sanyati ROM and LPO. (For abbreviations see Appendix A, Table A1).

Texture type Texture is typical for No. of examples rectangular boxwork cp or sp 2 triangular boxwork cp or sp 3 euhedral grain boundary cp 3 trellis type texture cp 2 cellular sponge cp or sp 3 open rib boxwork po 3 cubic texture py 4 parallel closed to open boxwork - 3 colloform texture - 5

The limonites formed in the decay textures mainly consist of goethite and hematite (see

Chap. 4.2.3.1). Additionally jarosites and plumbojarosite occur void-filling (cf. Fig. 4.16).

BSE imaging reveals that almost all limonites in the decay textures show a strong zonation

with the most pronounced zonations observed in colloform textues (a laminated, more or

less concentric banded pattern of dark- and light-reflecting cryptocrystalline goethite and

hematite; cf. Fig. 6.20). In contrast to observations by HERBERT (1999), who investigated

samples from AMD, the zonations are observable using optical microscopy as well as by

BSE (back scattered electronmicroscopic) imaging.


112

Fig. 6.3: Chemical variation of all investigated goethite and hematite from boxwork

textures (as determined by EMPA) and its bearing on the observed BSE-reflections.

40

50

60

70

80

0 1 2 3 4

Cu [%]

Fe [%

]

light ROM

dark ROM

light LPO

dark LPO

40

50

60

70

80

25 30 35 40

O [%]

Fe [%

]

light ROM

dark ROM

light LPO

dark LPO

hematite

goethite

40

50

60

70

80

0 2 4 6 8 10

Zn [%]

Fe [

%]

light ROM

dark ROM

light LP O

dark LP O

b.

a.

c.


113

SCOTT (1992) describes such alternating dark and light bands consisting of goethite and

hematite, respectively, while HERBERT (1999) suggests that the light reflecting banding

indicates heavier elements. EMPA measurements of run-of-mine ore (ROM) and leach pad

ore (LPO) confirm that the observed zonations reflect chemical as well as mineralogical

differences: The darker reflecting goethite zones show lower iron and higher oxygen

contents compared to the lighter reflecting hematite zones, most probably due to

incorporation of larger amounts of water. These darker reflecting iron-poor goethite zones

are presumably less crystalline and can incorporate more base metals compared to the light

reflecting iron-richer parts of the boxworks. The Fe contents of all measurements vary

between 72.2 and 40.36 wt% (Fig. 6.3) and are for most samples below those expected for

mixtures between pure goethite (62.85 wt%) and hematite (69.94 wt%). Therefore it is

most unlikely that the bands represent pure geothite or hematite but rather more or less

hydrated ironoxides and -oxyhydroxides. Furthermore, neither optical nor microchemical

evidence for the presence of other phases was found. The observed base metal contents of

the goethite-rich and hematite-rich bands vary significantly (Fig. 6.3). The general trend is

that Cu and Zn contents are higher in the darker reflecting, goethite-rich (i.e. more

hydrated), zones (Fig. 6.3). In the following the observed elemental distributions in dark

and light reflecting zones in all boxwork textures examined are described in more detail.

As light and dark reflecting phases will be discussed separately, abbreviations will be used as follows: The sample area (e.g. 294Z1-d-B1) is abbreviated by a letter or number. The letters and numbers assigned to the sample areas are listed in the Appendix D, Table D2. Additionally "l" refers to the light reflecting zones and "d" to the dark reflecting zones (e.g. Bl and Bd respectively).

Rectangular and triangular boxworks (Fig. 6.4, Fig. 6.6): In a first stage of oxidation

chalcopyrite is generally replaced by a secondary sulfide along fractures and grain

boundaries. Primarily crystallographic planes, probably 110, the principle cleavage plane

in chalcopyrite, are attacked. In a second stage rectangular, triangular, and ovoid cells can

be replaced selectively and give rise to an open boxwork with sinuous, subparallel walls

(NICKEL & DANIELS 1986). In Sanyati most distinctive examples for rectangular

replacement ribs (Fig. 6.4) were found in LPO; triangular textures (Fig. 6.6) were

examined in a sample from Copper-Queen-SW taken at shallow depth.


114

Fig. 6.4: Photomicrographs of rectangular boxwork in supergene ore (sample 294Z1).

a. reflected light microscopy (photomicrograph width: 600 µm). b. BSE image

with analysis points and concentrations for Cu and Zn (photomicrograph width:

250 µm).

48

52

56

60

64

Fe [%

]

Al Ad Bl Bd

0.40.81.21.6

22.42.8

Cu

[%]

Al Ad Bl Bd

0.8

1.21.6

22.4

2.8Z

n [%

]

Al Ad Bl Bd

Fig. 6.5: Boxplots (median, lower and upper quartile) of the Fe, Cu, and Zn contents in

light (l) and dark (d) zones in rectangular boxworks in sample 294Z1.

Abbreviations A and B refer to analysed areas 294Z1-d-B1 and 294Z1-d-A3,

respectively (cf. Table D2 in the Appendix D).

The observed Fe, Cu, and Zn distributions in rectangular boxworks are shown in Fig. 6.5.

The elemental distributions are in agreement with the general trend for all analysed zones

shown in Fig. 6.3 a-c, i.e. dark reflecting zones are characterised by relatively low Fe and

high Cu and Zn contents, whereas light reflecting zones show relatively high Fe and low

Cu and Zn contents. However, the observed concentration ranges vary significantly

between both analysed areas (e.g. the maximum Cu and Zn contents observed in dark


115

reflecting zones range from 1.1 wt% to 2.5 wt% and 1.2 wt% to 2.7 wt%, respectively).

Dark reflecting zones show approximately two-times higher As contents compared to light

reflecting zones (max. 0.8 wt%), while Pb, the most immobile element found, is enriched

in the light reflecting zones (max. 2.2 wt% Pb). This could account for the light colour of

these zones, as heavy elements induce light colours in BSE images. Cobalt distributions

are contrary in the two areas analysed. However, the concentrations are close to the

detection limit (max. 0.1 wt% Co). Al, an element known to be substituted for Fe3+ in the

goethite lattice, shows low abundances (0.2 - 1.0 wt%) and concentrations are generally

higher in the dark reflecting zones.

Fig. 6.6: Photomicrographs of triangular boxwork in supergene ore (LPO sample 294Z1).

a. reflected light microscopy (photomicrograph width: 400 µm). b. BSE image of

the same area with analysis points and concentrations for Cu and Zn

(photomicrograph width: 125 µm).

In triangular boxworks (Fig. 6.6) Fe contents are systematically lower in the dark reflecting

zones (Fig. 6.7) which is in agreement with the overall trend observed (cf. Fig. 6.3).

However, the typical trend of higher Cu contents in the dark reflecting zones is found only

in two analysed triangular boxworks. In all three analysed triangular boxworks Zn contents

are higher in the dark reflecting zones (with absolute concentration differences as high as

~0.8 wt% observed). Like in the rectangular boxworks, the observed Cu and Zn

concentrations show a strong variation between analysed areas (Fig. 6.7).


116

48

52

56

60

64

Fe [%

]

ClCd DlDd El Ed

00.40.81.21.6

2

Cu

[%]

Cl Cd DlDd El Ed

00.40.81.21.6

2

Zn [

%]

Cl Cd DlDd El Ed

Fig. 6.7: Boxplots (median, lower and upper quartile) of the Fe, Cu, and Zn contents in

light (l) and dark (d) zones in triangular boxworks in samples 294Z1 and 4F2.

Abbreviations C, D, and E refer to analysed areas 294Z1-d-A1, 4F2-d-A, and

4F2-d-B, respectively (cf. Table D2 in the Appendix D).

No systematical behaviour is found for As (max. 0.4 wt%) and Pb (max. 2.2 wt%). Cobalt

is usually distributed evenly between light and dark reflecting zones. However, the overall

abundances are close to the detection limit (i.e. 0.1 wt%). Al concentrations are, like in

rectangular textures, low (max. 0.6 wt%) but are systematically higher in the dark

reflecting zones.

Fig. 6.8: Photomicrographs of euhedral grain boundary boxworks in supergene ore.

a. BSE image with analysis points and concentrations for Cu and Zn (LPO

sample 294Z1; photomicrograph width: 400 µm). b. BSE image with analysis

points and concentrations for Cu and Zn (LPO sample 294Z1-d;

photomicrograph width: 250 µm).


117

444852566064

Fe

[%]

Fl Fd Gl Gd Hl Hd

0.51

1.52

2.53

Cu

[%]

Fl Fd GlGd HlHd

0

1

2

3

4

Zn [

%]

Fl Fd Gl Gd Hl Hd

Fig. 6.9: Boxplots (median, lower and upper quartile, min. and max.) of the Fe, Cu, and Zn

contents in light (l) and dark (d) zones in euhedral grain boundary boxworks in

samples 294Z1 and 294R1. Abbreviations F, G, and H refer to analysed areas

294Z1-B, 294Z1-B, and 294R1-d-A2, respectively (cf. Table D2 in Appendix D).

Euhedral grain boundarys: In euhedral textures (Fig. 6.8) Fe contents are systematically

lower in dark reflecting zones but vary strongly between the three analysed boxworks (Fig.

6.9). Copper is enriched in the dark reflecting zones (with concentrations up to 2.6 wt%

Cu), while Zn (with concentrations up to 3.3 wt%) shows this trend in two analysed areas

(Fig. 6.9). As was only detected in two analysed areas (maximal concentrations 0.5 wt%)

with higher contents observed in the dark reflecting zones. Pb is slightly higher in the light

reflecting zones in two analysed areas and significantly higher in one analysed area with

mean contents of 0.8 wt% (dark zones) and 2.1 wt% (light areas). Cobalt (max. 0.1 wt%)

shows no systematic distribution. Aluminium (max. 1.9 wt%) is enriched in the dark zones

in two analysed areas.

Trellis type texture: A trellis type texture, which fingerprints the earliest replacement of

chalcopyrite due to different rates of acid attack along crystallographic planes, possibly

(111), can be retained in the further boxwork development fringing the replacement ribs

(NICKEL & DANIELS 1986). This texture is a distinctive diagnostic feature for boxworks

derived from chalcopyrite (NICKEL & DANIELS 1986). In two supergene ore samples from a

shallow position of the Copper-King outcrop these textures are preserved in boxworks

which are open and have continuous subparallel walls, also typical for chalcopyrite

(BLANCHARD 1968).


118

Fig. 6.10: Photomicrographs of trellis type boxwork in supergene ore (sample 79)

a. reflected light microscopy (photomicrograph width: 2500 µm). b. BSE image


500 µm).

40

44

48

52

56

Fe [%

]

Il Jl

1.2

1.4

1.6

1.8

Cu

[%]

Il Jl

1.4

1.61.8

22.2

2.4

Zn

[%]

Il Jl

Fig. 6.11: Boxplots (median, lower and upper quartile, min and max.) of the Fe, Cu, and

Zn contents in trellis type boxworks in sample 79. Abbreviations I and J refer to

analysed areas 79-d and 79-e, respectively. Abbreviation l refers to the light

zones analysed (cf. Table D2 in the Appendix D).

In the trellis type textures (Fig. 6.10) only dark zones are present. The highest observed Cu

and Zn concentrations are 1.8 wt% and 2.3 wt%, respectively (Fig. 6.11). Average As

concentrations (with a mean of 1.6 wt% and a max. 2.2 wt%, respectively) are very high,

while Pb (with a max. of 0.35 wt%), Co (max. 0.1 wt%), and Al contents (max. 0.12 wt%)

are relatively low.


119

Parallel closed and open boxwork: In the case of the parallel textures (Fig. 6.12) the

sections were evaluated in an order of decreasing degree of dark filling in the voids.

Generally Fe concentrations show the typical trend of lower contents in the dark reflecting

zones. The overall Fe concentrations increase with decreasing degree of void filling (from

57 to 62 wt%)(Fig. 6.13). This can be interpreted as a maturation, where water-rich

goethite dominated fillings slowly alter to hematite-richer crusts. For Cu (min. 0.6 wt%

and max. 2.9 wt%) and Zn (min. below detection limit and max. 3.1 wt%) the typical trend

is retained for most analysed areas and the concentrations decrease with decreasing degree

of void filling in dark and light reflecting zones. This is in agreement with the typical trend

for lower base metal content in hematite-richer light reflecting zones. For As (with

contents from the detection limit to 2.3 wt% ), Pb (with contents varying from 0.1 to 1.7

wt%), Co (max. 0.1 wt%), and Al (with a very low max. concentration of 0.2 wt%) no

systematic trends are found.

Fig. 6.12: Photomicrographs of parallel closed and open boxworks in supergene ore: a.

BSE image of parallel closed boxwork with analysis points and concentrations

for Cu and Zn (sample 26-A3; photomicrograph width: 250 µm). b. BSE image of

parallel open boxwork with analysis points and concentrations for Cu and Zn

(sample 212-2-B-1 photomicrograph width: 400 µm).


120

48

52

56

60

64

Fe

[%]

Kl K d Ll Ld MlMd Nl Nd Ol Od

0.51

1.52

2.53

Cu

[%]

Kl Kd Ll Ld MlMd Nl Nd Ol Od

0

1

2

3

4

Zn [

%]

Kl Kd Ll Ld MlMd Nl Nd Ol Od

Fig. 6.13: Boxplots (mean, lower and upper quartile, min. and max.) of the Fe, Cu, and Zn

contents in light (l) and dark (d) zones in parallel open and closed boxworks in

samples 26, 209c, and 212-2. Abbreviations K, L, M, N, and O refer to analysed

areas 26-A3, 209-c-d-B, 209c-d-A, 212-2-d-B1, and 212-2-B1, respectively (cf.

Table D2 in the Appendix D).

Cellular sponge: Direct replacement of chalcopyrite by ironoxyhydroxides produces a

dense, cellular texture called "cellular sponge" (BLANCHARD 1968). The observed cellular

sponge textures (Fig. 6.14) show a systematic depletion trend in dark reflecting zones for

Fe, but not the systematic enrichment for Cu and Zn (Fig. 6.15). Maximum contents are 2.6

wt% for Cu and 4.4 wt% for Zn and are relatively high compared to other textures. Except

for one example (82-d with As contents of max. 2.9 wt%) the observed As contents are

low. Observed Pb, Co, and Al distributions are not systematic, but with up to 5.5 wt% Pb

contents are very high, while the maximum Co and Al contents are very low with 0.1 wt%

and 0.2 wt%, respectively.


121

Fig. 6.14: Photomicrograph of cellular sponge boxwork in supergene ore (sample 268a-

d): a. reflected light microscopy (photomicrograph width: 800 µm). b. BSE image


250 µm).

404448525660

Fe

[%]

Pl Pd Ql Qd Rl Rd

0.81.21.6

22.42.8

Cu

[%]

Pl Pd QlQd Rl Rd

2

3

4

5

6

Zn [

%]

Pl Pd Ql Qd Rl Rd

Fig. 6.15: Boxplots (mean, lower and upper quartile) of the Fe, Cu, and Zn contents in

light (l) and dark (d) zones in cellular sponge boxworks in samples 82 and 268a.

Abbreviations P, Q, and R refer to analysed areas 82-d-A2, 268a-d-B, and 268a-

B, respectively (cf. Table D2 in the Appendix D).

Open rib texture: This texture is described by BLANCHARD (1968) as boxwork with flaky

or shrivelled limonite crusts, which is typical for pyrrhotite decay (Fig. 6.16). In these

textures the Fe contents are high (with mean values ranging from 59.7 to 62.7 wt%; Fig.

6.17), while Cu and Zn concentrations are below 1.1 wt%. Contents of As, Pb, Co, and Al

are low with max. concentrations of 0.7 wt% for As and Pb, up to 0.1 wt% for Co, and 0.3

wt% for Al being observed.


122

Fig. 6.16: Photomicrograph of open rib boxwork in supergene ore (sample 212/2-d-A2).

BSE image with analysis points and concentrations for Cu and Zn

(photomicrograph width: 500 µm).

58596061626364

Fe

[%]

Sl Tl Ul Vl

0.2

0.4

0.6

0.8

1

Cu

[%]

Sl Tl Ul Vl

0

0.2

0.4

0.6

0.8

Zn [

%]

Sl Tl Ul Vl

Fig. 6.17: Boxplots (mean, lower and upper quartile) of the Fe, Cu, and Zn contents in

light (l) zones in open rib boxworks in samples 212/2, 212/2-d, and 4F2-d.

Abbreviations S, T, U, and V refer to analysed areas 212/2-A5, 212/2-d-A2,

212/2-d-B4, and 4F2-d-C, respectively (cf. Table D2 in the Appendix D).

Cubic boxworks: In dark reflecting zones in cubic textures (Fig. 6.18) the typical trend of

relative Fe depletion and Cu and Zn enrichment is observed, if both dark and light

reflecting zones are developed (Fig. 6.19). The Cu (max. 3.7 wt%) and Zn (max. 4.7 wt%)

contents in these zones are unusually high. For As (max. 1.4 wt%), Pb (max. 2.2 wt%) and

Al (usually below 0.2 wt%) no systematic trend could be found, whereas Co (max. 0.1

wt%) might be slightly enriched in the light reflecting zones, bearing in mind that

concentrations are close to the detection limit.


123

Fig. 6.18: Cubic boxwork in supergene ore (sample 268a-d-C). BSE image with analysis

points and concentrations for Cu and Zn (photomicrograph width: 500 µm).

404448525660

Fe

[%]

Wl WdXd Yl Yd Zl

0

1

2

3

4

Cu

[%]

WlWdXd Yl Yd Zl

012345

Zn [

%]

WlWdXd Yl Yd Zl

Fig. 6.19: Boxplots (mean, lower and upper quartile, min. and max.) of the Fe, Cu, and Zn

contents in light (l) and dark (d) zones in cubic boxworks in samples 268a and

72a. Abbreviations W, X, Y, and Z refer to analysed areas 268a-d-C, 268a-A2,

268a-d-A, and 72a, respectively (cf. Table D2 in the Appendix D).

Colloform textures: Compositional zoning is a commonly observed phenomena in a large

range of minerals from various geologic environments. These zonations are a direct

function of the physico-chemical conditions during crystal growth and can therefore

provide valuable information about the environment during different growth stages. For

example, in magmatic minerals the observed zonations are controlled by extrinsic (e.g.

temperature change, crystal settling) and intrinsic mechanisms (e.g. rate of crystal growth,

solute diffusion through the crystal-melt boundary; e.g. SHORE & FOWLER 1996).


124

In hydrothermal, sedimentary, and ore-forming environments extrinsic factors in aqueous

solutions are changes in P, T, and redox state, while intrinsic factors are reactions at the

solution-mineral boundary (e.g. adsorption). In low temperature, near-surface

environments changes in T and P are usually confined to a fairly narrow range. Pore water

chemistry and flow are the most important extrinsic factors affecting the compositional

zoning (HERBERT 1999). Iron- and manganese-oxides show banding and concentric

accumulation zonations that are formed during their growth under terrestrial, limnic, and

marine conditions and are suggestive for a cyclic precipitation (CORNELL &

SCHWERTMANN 1996). Banding in manganese oxides of stream sediments is related to

changes in solute chemistry, pH, and dissolved oxygen, that occur during seasonal cycles

(ROBINSON 1993). Intrinsic factors may also be autonomous self-organisation during

crystal growth, as has been shown for carbonates by REEDER et al. (1990).

Fig. 6.20: Microphotographs of colloform textures in supergene ores: a. BSE image with

analysis points and concentrations for Cu and Zn (sample 294Z1-A;

photomicrograph width: 150 µm) b. BSE image with analysis points and

concentrations for Cu and Zn (sample 212-2-d-A1; photomicrograph width: 500

µm).


125

444852566064

Fe

[%]

1l 1d 2l 2d 3l 3d 4l 4d 5l 5d 6l 6d

00.5

11.5

22.5

3

Cu

[%]

1l 1d 2l 2d 3l 3d 4l 4d 5l 5d 6l 6d

0

1

2

3

4

Zn [

%]

1l 1d 2l 2d 3l 3d 4l 4d 5l 5d 6l 6d

0

1

2

3

4

As

[%]

1l 1d 2l 2d 4l 4d 5l 5d 6l 6d

Fig. 6.21: Boxplots (median, lower and upper quartile, min and max.) of the Fe, Cu, Zn,

and As contents in light (l) and dark (d) zones in colloform textures in samples

212-2, 294-Z1, 79, and 268d. Abbreviations 1, 2, 3, 4, 5 and 6 refer to analysed

areas 212-2-d-A1, 212-2-A2, 294Z1-A, 294Z1, 79-d-B and 268d-d-A, respectively

(cf. Table D2 in the Appendix D).

In sulfide mine waste deposits compositional zoning has been described by JAMBOR

(1994), LIN (1996), and HERBERT (1999). The precipitation-forming ions are dissolved,

transported over a certain distance and reprecipitated from the solution. Here too, seasonal

variations, e.g. rainfall or snowmelt flushing of sulfide oxidation products leads to

temporal changes in the groundwater composition and effects the chemical signal recorded

in the precipitations (ALPERS et al. 1994; HERBERT 1995).

In all examined colloform textures in supergene ores from Sanyati (Fig. 6.20) the typical,

above described depletion-enrichment trends are found for the distribution of Fe, Cu, and

Zn in dark and light reflecting zones, and only two exceptions could be observed (analysed

areas 6 and 4; Fig. 6.21). The observed Fe contents span a wide range between 46.2 wt%

and 63.1 wt%. Except for one unusually high value (2.86 wt% Cu), Cu contents are below

1.7 wt%, while Zn contents reach a maximal concentration of up to 3.6 wt%. As (Fig. 6.21)

and Pb (not shown) contents are highly variable in all investigated samples and no

systematic distributions could be found. Co concentrations reach 0.1 wt% but are not


126

systematic. When Al is present it is enriched in the dark reflecting zones with max.

contents of up to 3.5 wt% being observed.

6.2.2 Element distribution in goethite-rich and hematite-rich zones of boxwork textures

Al, Si, and S: Al shows no systematic distributions in the studied textures and the observed

concentrations are very low (mean of light reflecting zones 0.2 wt%, mean of dark

reflecting zones 0.4 wt%). This is in agreement with the estimation of 1.5 wt% from the

XRD of selected goethite-rich samples (see Chapter 4.2.3.2). Higher concentrations are

only observed in locally enriched zones. Si is also locally enriched and reaches in one

exceptional case a content of 12 wt%; however, for the majority of the samples contents

are below 1.0 wt%. For comparison Si concentrations of 5.4 mol% (= 1.56 wt%) in

oxyhydroxides have been reported by JOHNSON & NORRISH (1981). HERBERT (1999) found

an inverse correlation between the Al+Si and the Fe+S contents in ironoxyhydroxides of

waste rock dumps. In the samples studied here, this correlation is only found in the

previously leached euhedral textures in sample F (294Z1) and the colloform textures in

sample 3 (294Z1, Fig. 6.22). Furthermore, in these samples Si and Al show a positive

correlation (Fig. 6.23). It is likely that this correlation is also present in the unleached

samples, but it is superimposed by other signatures (e.g. due to adsorption).

Fig. 6.22: Fe+S vs Si+Al in the goethite- and hematite-rich zones of LPO sample 294Z1 A

(=LPO 3 coll) and B (LOP F euhedral).

LPO 3 coll

40

45

50

55

60

65

0 5 10

Si+Al [%]

Fe+S

[%]

light areas

dark areas

LPO F euhedral

40

45

50

55

60

65

0 5 10

Si+Al [%]

Fe+S

[%]

light areas

dark areas


127

Fig. 6.23: Si vs Al in the goethite- and hematite-rich zones of the LPO sample 294Z1 areas

A and B and all other samples measured.

Al and Si concentrations in ironoxyhydroxides are generally assumed to be controlled by

the solution chemistry, i.e. by the Al/Si ratio of the solutions. Due to the low Al

concentrations observed the Al/Si ratio in the investigated goethite-rich and hematite-rich

zones varies strongly (from unity to values up to 50). The most likely Al and Si sources are

phyllosilicates which liberate Al and Si during weathering.

Cu and Zn: Based on the classification by BLANCHARD (1968) and NICKEL & DANIELS

(1986) five of the observed textural types were classified as relicts of Cu- or Zn-rich

sulfides (chalcopyrite or sphalerite), while two were found to be characteristic for Fe-

sulfides (e.g. pyrite and pyrrhotite) and should display low base metal (i.e. Cu and Zn)

contents. In Fig. 6.24 the mean Cu and Zn contents of the dark reflecting goethite-rich

zones are plotted in column diagrams. The results from dark reflecting zones were chosen

because of the generally higher contents of Cu and Zn, and therefore lower analytical

errors compared to the light reflecting zones. Clearly, only the open rib textures,

interpreted as relicts of pyrrhotite, show the low Cu and Zn contents expected for the decay

of Fe sulfides. In contrast the cubic relicts, characteristic for pyrite decay, show Cu and Zn

contents that are on the high end of the observed variations in all relictic textures.

Furthermore, Fig. 6.24 demonstrates that the investigated goethite- and hematite-rich zones

do not show distinctive Cu or Zn enrichments, which might allow a distinction between

chalcopyrite and sphalerite relictic textures. Triangular textures, according to BLANCHARD

0

1

2

3

4

5

6

0 1 2 3 4 5 6Al [%]

Si [%

]

294Z1-B-light

294Z1-B-dark

294Z1-A-light

294Z1-A-dark

all others


128

(1968) the most likely relicts of sphalerite, show the lowest Zn contents. However,

difficulties with respect to the identification of sphalerite relicts were already pointed out

by NICKEL & DANIELS (1986). Overall, the observed Cu and Zn distributions suggest a

slight positive correlation of both elements, rather than a complementary behaviour (Fig.

6.25). These observations suggest that the mobility of base metals is high even in the early

stages of sulfide decay so that no particular elemental fingerprints indicative for the decay

of a specific mineral phase are preserved in the goethite-rich and hematite-rich zones,

irrespective of the observed textural evidence.

Fig. 6.24: Cu vs Zn of all investigated goethite and hematite from boxwork textures (as

determined by EMPA) show a correlated behaviour.

Fig. 6.25: Mean Cu and Zn contents in dark reflecting zones of different types of boxworks

from both ROM and LPO. The solid line represents the gross average of all

determinations.

0.0

0.5

1 .0

1 .5

2 .0

2 .5

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

mea

n C

u in

box

wor

k ty

pes

[%]

Cu

0

1

2

3

4

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

mea

n Z

n in

box

wor

k ty

pes

[%]

rect angular

t riangular

euhedral

t rellis-work

cellular sp.

parallel rib

open rib

cubic

collom orph

boxwork t ype:Zn

0

2

4

6

8

10

0 1 2 3 4

Cu [%]

Zn

[%]

light ROM

dark ROM

light LPO

dark LPO


129

Fig. 6.26 a: Chemical variation (expressed as standard deviations) in light reflecting zones

for boxwork types and cases in each section to compare the variabilty for each

situation. The solid line represents the overall mean for all boxwork types and

sections in ROM and LPO. The variability is always lower in the compilation of

sections.

Fe- light

0

1

2

3

4

5

6

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

294Z

1

4F2 79 82 268

26A

209

212/

2

stan

dard

dev

atio

n

Cu - light

0

0.1

0.2

0.3

0.4

0.5

0.6

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

294Z

1

4F2 79 82 268

26A

209

212/

2

stan

dard

dev

atio

n

Zn- light

0

0.2

0.4

0.6

0.8

1

1.2

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

294Z

1

4F2 79 82 268

26A

209

212/

2

stan

dard

dev

atio

n


130

Fig. 6.26 b: Chemical variation (expressed as standard deviations) in dark reflecting zones

for boxwork types and cases in each section to compare the variabilty for each

situation. The solid line represents the overall mean for all boxwork types and

sections in ROM and LPO. The variability is always lower in the compilation of

sections.

Cu - dark

0

0.2

0.4

0.6

0.8

1

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

294Z

1

4F2 79 82 268

26A

209

212/

2

24A

stan

dard

dev

atio

n

Zn - dark

0

0.2

0.4

0.6

0.8

1

1.2

1.4

1.6

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

294Z

1

4F2 79 82 268

26A

209

212/

2

24A

stan

dard

dev

atio

n

Fe - dark

0

1

2

3

4

5

6

rect

angu

lar

tria

ngul

ar

euhe

dral

trel

lis-w

ork

cellu

lar

sp.

para

llel r

ib

open

rib

cubi

c

collo

mor

ph

294Z

1

4F2 79 82 268

26A

209

212/

2

24A

stan

dard

dev

atio

n


131

If such a high mobility of the solutions is assumed, precipitations that are derived from

these moving solutions should yield base metal contents close to the mean of the

surrounding relicts. The observed base metal contents in the colloform textures provide

evidence for this assumption as their mean base metal contents approximates the overall

mean contents of all investigated textures (Fig. 6.24).

Hence, base metal contents are primarily reflecting the chemical fingerprint of the wider

surrounding and not a preserved chemical fingerprint of single decaying grains even if

these leave characteristic boxworks. This is also shown in Fig. 6.26 a. and Fig. 6.26 b.,

where the overall observed variations of the Fe, Cu, and Zn contents in dark and light

reflecting zones, expressed as standard deviations, for all analysed boxworks are plotted

for the different textural types and for the different samples examined. For all three

elements, the observed variations on a sample scale are less pronounced, i.e. their

variability is lower, compared to the observed variations in the different types of

boxworks.

A further indicator that the chemistry of the goethites is governed by the local environment

is their composition close to frequent base metal minerals (Fig. 6.27). Here Cu, Zn, and As

concentrations vary systematically with the chemistry of the base metal phase nearby, i.e.

goethites in the vicinity of olivenite, a Cu arsenate, incorporate high Cu and As contents.

One exception are the relatively high As contents in goethites close to hemimorphite.

Fig. 6.27: Average element concentrations (determined by EMPA) of goethites that are in

direct proximity (i.e. in grain contact) to frequent base metal minerals.

Abbreviations are: Mal = malachite, Ol = olivenite, Cl = clinochlore, Ad =

adamite, and Hem = hemimorphite.

0

5

10

15

20

25

wt%

Mal Ol Cl Ad Ad Hem

As2O5ZnOCuOCoO*10Fe2O3/10


132

These findings suggest that the mobility of Cu and Zn is high enough to homogenize

original concentration gradients in the primary sulfides (at least on the scale of a sample,

i.e. on a dm scale). Hence, the incorporation of these elements into goethite and hematite is

primarily controlled by physical and chemical factors like permeability, pH, and Eh.

Existing primary concentration gradients, which would fingerprint the primary sulfide, are

erased or blurred.

The average chemical compositions of the homogenizing fluids are retained in the

colloform textures. The chemical signature of the colloform textures might therefore in

general reflect the chemistry of the goethite and hematite on a sample scale. These

observations suggest that the textural context and the chemistry of the colloform textures,

which can be easily investigated by microscopic techniques and by EMPA, might be used

as an approximation for the amount of Cu and Zn fixed in the goethite and hematite in a

given ore sample. Together with additional information, e.g. whole rock geochemistry and

modal abundances, this might offer a new approach to discriminate between the mobile

and immobile fraction of base metals in the supergene ore.

Mn, Co, As, and Pb: Mn contents in goethite and hematite are low due to the presence of

Mn-hydroxides. Co does not show systematic trends in the investigated goethite- and

hematite-rich zones. However, the Co contents are in most cases close to the detection

limit, and therefore the data have to be regarded as only half-quantitative.

As is present in the investigated goethite- and hematite-rich zones from contents below the

detection limit up to 2.57 wt% in the light reflecting hematite-rich zones and 3.10 wt% in

the dark reflecting goethite-rich zones. In analogy to Fig. 6.26 a. and Fig. 6.26 b. the mean

of the standard deviation has been calculated for the As contents of the boxwork types and

for the As contents sorted by the sections (Table 6.3). This serves to compare the variabilty

for each situation. The standard deviation for As is less than half for the sorting by sections

compared to the sorting by types, which means that in the same way as for the base metals

the element distibution is controlled by the local environment. In cases where a systematic

As distribution between light and dark reflecting zones is developed, As is always more

enriched in the dark reflecting zones.

With respect to spatial distribution Pb shows the same trend as observed for As and the

base metals. The only difference found is that higher concentrations are generally observed


133

in the light reflecting zones (max. 5.5 wt% in the light reflecting zones, max. 3.4 wt% in

the dark reflecting zones). Minimal concentrations are 0.1 wt%.

Table 6.3: Mean of the standard deviations of the data sorted by boxwork types and

examined sections.

Zone Al Mn As Pb Mean type light 0.2 0.04 0.47 0.67 Mean section light 0.09 0.03 0.16 0.41 Mean type dark 0.21 0.04 0.49 0.44 Mean section dark 0.21 0.08 0.20 0.25

6.2.3 Concentrations in run-of-mine ore (ROM) and leach pad ore (LPO)

One aim of the microanalytical approach was to investigate if element mobilisation from

goethite and hematite occurs on the heap leach pad. Therefore a sample block leached for

approximately four years (sample 294, Ø 15 cm) was examined in detail. The mean

concentrations in goethite and hematite from the core and the rim of this block are

compared to the ROM samples in Fig. 6.28. The concentrations in the ROM samples are so

highly variable (for Cu min. 0.46 wt% and max. 2.65 wt%; for Zn min. 0.27 wt% and max.

3.86 wt%) that the LPO sample lies somewhere in the middle of this span (Cu light 0.89

wt%, Cu dark 1.81 wt%).

Fig. 6.28: Mean Zn concentrations of the sections of ROM (solid bars) and LPO block 294

(open bars) showing that Zn concentrations in the LPO bloc lie within the

average of the ROM samples.

0

1

2

3

4

5

4F2 79 82268

26a209

212/2 294

mea

n co

nc. [

%]

Z n light reflecting

Z n dark reflecting

Zn

ROM LPO


134

A comparison of the iron phases in the centre and the rim of the leached block 294

(distance 7 cm) shows that weak depletion in the rim is found for Cu and Zn (Fig. 6.29).

The Cu depletion is (-14 wt%) significantly higher than the depletion for Zn (-5 wt%),

since the Cu is undersaturated and Zn is saturated in the acid under the conditions on the

heap leach pad.

Fig. 6.29: Cu and Zn in rim and centre of the leach pad ore sample 294.

6.3 Trace element chemistry of goethite- and hematite-rich zones (Laser ablation

analysis)

In bulk supergene ore samples enrichment relative to the average crustal composition was

also established for Ga, Se, Ag, Cd, Sb, and Bi (Table 6.4). For crystal chemical reasons,

these elements are also believed to be incorporated into the structure of ironoxyhydroxides

(e.g. BERNSTEIN & WAYCHUNAS 1987). Adsorption also plays an important role in the

scavenging of e.g. Se from solutions. In seawater adsorption on ironoxides and -

oxyhydroxides might exert a significant control on the Te and Se budget (HEIN et al. 2003).

Because of their low concentrations, these elements can not be identified by EMPA

analysis.

Table 6.4: Element concentrations in bulk supergene ore samples determined by XRF.

element Ga mg/kg Se mg/kg Ag mg/kg Cd mg/kg Sb mg/kg Pb wt% Bi mg/kg min. (LDL)

<9 <12 <9 5 <18 0.01 <21

max. 87 134 124 517 1200 3.2 571 mean 11 36 19 63 171 0.94 109 median 9 23 9 29 74 0.65 77

average crustal composition [µg/kg]: Ga 18, Se 50, Ag 80, Cd 98 Bi 60 [mg/kg]: Sb 0.2, Pb 8 (Taylor & McLennan 1985)

0

1

2

3

4

center rim center rim

[wt%

]

Cu Zn


135

Therefore the distribution of these elements in goethite- and hematite-rich bands were

investigated in selected samples using LA-ICP-MS. The samples investigated were ROM

212-2-d, 268d, 209c-d, 4F2-d, 79-d and LPO 294Z-d, 294R-d. LA-ICP-MS analyses were

performed in the areas previously analysed by EMPA. In cases were the Laser spot size (~

Fig. 6.30: Observed distributions of Ga, Ge, Se, Ag, Cd, and Sb in goethite- and hematite-

rich zones as determined by LA-ICP-MS. Blue solid diamonds represent dark

reflecting zones and orange solid diamonds represent light reflecting zones. Open

symbols indicate samples with on average unusually high trace metal contents

compared to the other samples (see text for explanation).

05

10152025303540

0 20000 40000Cu [mg/kg]

Ga [m

g/kg

]

0

5

10

15

20

25

30

35

0 20000 40000Cu [mg/kg]

Ge [m

g/kg

]

0200400600800

100012001400

0 20000 40000

Cu [mg/kg]

Se [m

g/kg

]

0

10

20

30

40

50

0 20000 40000Cu [mg/kg]

Ag

[mg/

kg]

0

100

200

300

400

500

600

0 20000 40000Cu [mg/kg]

Cd [m

g/kg

]

0100200300400500600700800

0 20000 40000Cu [mg/kg]

Sb [m

g/kg

]


136

15 to 20 µm) was to large to analyse adjacent light and dark reflecting zones, the nearest

possible similar zone was used for measurement.

In agreement with the element distributions observed by EMPA (Chap. 6.2),

concentrations for the analysed trace elements are generally higher in the dark reflecting

zones.

In analogy to the EMPA investigations, different boxwork types were examined. Again it

was found that the chemistry of goethite and hematite is dominated by the chemistry of the

surrounding material and not by the type of ore texture. High trace element concentrations

found in selected samples are not related to specific boxwork types. These samples are

highlighted with open symbols in Fig. 6.30 (e.g. for Ga samples 212-2, 294R and 294Z; for

Se sample 212-2, and for Cd sample 209 and 79). For Ge, Ag, Sb, and Bi (the last not

shown in Fig. 6.30) no characteristic enrichment could be observed. High trace element

concentrations are found in ROM as well as LPO (e.g. Ga).

6.4 Comparison of element distributions of goethite and hematite in three base

metal and lead deposits

In the literature the database on EMPA analyses for the rock-forming minerals goethite and

hematite suitable for comparison is very limited.

EMPA data of ironoxyhydroxides precipitated from acid mine drainage (AMD) are

available for Fe, Al, Si, and S only (HERBERT 1999; Fig. 6.31).

50

60

70

80

90

100

[wt %

]

Fe AMD Fe San

0

1

2

3

4

5

[wt %

]

Al AMD Al San Si AMD Si San S AMD S San

Fig. 6.31: Comparison of the chemical composition of Sanyati goethites from colloform

textures with those precipitated from acid mine drainage (HERBERT 1999). The

boxplots depict the median, upper and lower quartile, min. and max. value.


137

The median of all four elements are lower in the colloform textures from Sanyati compared

to the AMD data and show a wider compositional spread. One likely reason are the

different physico-chemical milieus and especially the varying element sources. Beside this,

the different aging processes of the minerals may influence element adsorption and

incorporation.

goethite hematite

0

4000

8000

12000

16000

20000

Cu

[mg/

kg]

Tu Vo Do San

0

4000

8000

12000

16000

Cu

[mg/

kg]

Tu Vo Do San

0

20000

40000

60000

Zn [

mg/

kg]

Tu Vo Do San

0

10000

20000

30000

40000

Zn [

mg/

kg]

Tu Vo Do San

0

10000

20000

30000

As

[mg/

kg]

Tu Vo San

0

4000

8000

12000

16000

20000

As

[mg/

kg]

Tu Vo San

Fig. 6.32: Comparison of EMPA analyses of goethite and hematite from base metal

bearing mineralisations in turbidites (Tu), volcaniclastics (Vo) (SCOTT 1992),

dolomitic shale (SCOTT 1986), and Sanyati ores (colloform textures). The

boxplots depict the median, upper and lower quartile, min. and max. value.


138

In Fig. 6.32 goethite and hematite analyses from two Australian deposits (1. the Wagga

Tank polymetallic deposit developed in turbidites, Tu, and volcaniclastics, Vo; SCOTT

1992, and 2. a dolomitic shale-hosted base metal mineralisation, Do; SCOTT 1986), are

compared with those from Sanyati. Clearly, the Sanyati ores generally show extremely

high Cu, Zn, and As concentrations. For Pb (not shown) the medians equal those of the

dolomite hosted mineralisation, which is mined for Pb and Zn. Al concentrations are lower

and Si concentrations significantly higher in both goethite and hematite from Sanyati

compared to those from the Wagga Tagga deposit (no data were available for the dolomitic

shale hosted mineralisation).

This shows that a significant amount of metals is bound in or to the ironoxides and -

oxyhydroxides of the Sanyati ores. These phases are the main constitutents of boxwork

textures which form during the decay of the primary sulfides.


Sulfide decay often leads to the development of characteristic textures formed by

ironoxides and -oxyhydroxides. On the basis of classifications by BLANCHARD (1968) and

NICKEL & DANIELS (1986) these were assigned to base metal sulfides

(chalcopyrite/sphalerite), and iron sulfides (pyrrhotite and pyrite). It was shown that

element distributions in these textures are not controlled by the elements liberated from a

single grain, but by the local environment on a dm3 scale. No geochemical fingerprint of

Cu or Zn is therefore retained in the textures characteristic for chalcopyrite and sphalerite.

The mean chemical composition of the residual iron phases of a local environment is

reflected in the colloform goethite- and hematite-rich zones that precipitate from the

solutions that mobilize and redistribute the metals. These colloform textures reflect the

mean composition on a sample scale.

In all examined ironoxyhydroxide textures light and dark reflecting zones reflect the

mineralogical and geochemical zoning of the precipitations. In the light reflecting zones Fe

contents are generally higher, indicating a higher proportion of hematite. Additionally,

light reflecting zones often show Pb enrichment. In the dark reflecting zones lower Fe and

higher O contents indicate goethite-rich precipitations. In these zones Cu, Zn, and As are

generally enriched. Additionally, the trace elements Ga, Ge, Se, Ag, Cd, and Sb are

generally also enriched in the dark reflecting, goethite-rich zones.


139

A comparison of the ironoxyhydroxides compositions in the centre and the rim of a sample

block shows that Zn is only weakly depleted (-5 %) and Cu with -14 % somewhat more

depleted in the rim (Fig 6.29). This again indicates that base metal contents of the

ironoxyhydroxides are not readily leachable and are therefore "non-available" for the

leaching process. In order to quantify this "non-available" fraction further a series of

laboratory experiments were performed.

7. Laboratory leaching behaviour of the supergene ore

140

7 Laboratory leaching behaviour of the supergene ore

One of the most important findings of the mineralogical and geochemical investigations on

the supergene ores from Sanyati is that goethite and hematite are carriers of considerable

amounts of base metals. The dark reflecting zones of boxwork textures show particularly

characteristic enrichments for Cu and Zn. Investigations on leach pad ore revealed that

especially the goethite-rich zones might play a crucial role for the mobilisation and

retention of base metals.

Therefore, as an attempt to better constrain the bearing of goethite and hematite on the

leaching characteristics of supergene ores, a series of laboratory experiments were

designed using leach pad ore (LPO), run-of-mine ore (ROM), and ore-analogues consisting

of natural and synthetic materials. A detailed description of the materials used as well as

experimental set-ups and protocols is provided in Chapter 3.2.2.

In order to compare the element output from the supergene ore before and after technogene

leaching on the heap leach pad percolation experiments on ROM and LPO were performed

in a first step (Chap. 7.1.1). Additionally, LPO was leached in an "ideal" system to

mobilize the metals still retained in the ore (Chap. 7.1.2). Partial extraction experiments

were performed to elucidate to which phases the metals are bound and to delineate metal

mobilisation as a function of the strength of acid attack (Chap. 7.2). Finally, the adsorption

behaviour of Cu on goethite was evaluated in various chemical environments under acidic,

sulfate-rich conditions (Chap. 7.3). All experimental results, i.e. the initial composition of

the starting materials used, the final composition of the residues recovered after the

experiments as well as the solution compositions are reported in Appendix C.

7.1 Leaching experiments with ROM and LPO (H2SO4-soluble fraction)

7.1.1 Percolation experiments

Small-scale leaching experiments with ROM and LPO were performed to estimate the

output of base metals (experiments VR and V15 in Table 3.3 in Chap. 3.2.2). Two ROM

samples (210 and 211, fraction < 1cm) and four LPO samples (114, 115, 116/2, and 116/4)

were leached 10 times with H2SO4 (pH 1.5). The results for these experiments are reported


141

in Table C2, C4, and C5 in Appendix C. In Fig. 7.1 the relative elemental depletions or

enrichments observed in the sample residues recovered after leaching (calculated from

elemental concentrations in the starting material determined prior to and after the

experiment according to the equation 100*[(final conc. – initial conc.)/initial conc.]) are

displayed graphically. The combined absolute experimental and analytical uncertainties are

estimated to be in the 10 to 15 % range.

Both ROM samples appeared very similar when examined visually, and show reasonable

similar chemistry (cf. Appendix C, Table C2). Bearing these similarities in mind, it is

surprising that both samples show a markedly different leaching behaviour (Fig. 7.1).

Sample 210 shows the expected element depletions for Cu (-30 %), Zn (-53 %), Co (-62

%), Mn (-45 %), and Pb (-27 %) whereas sample 211 shows a marked depletion only for

Zn (-34 %) and a slight depletion for Mn (-15 %). The results for both Cu and Co suggest a

slight passive enrichment (+5 and +12 %, respectively) in the residue. However, for both

elements the measured concentration changes are within the combined experimental and

analytical errors. The apparent enrichment of Cu, Co, and Pb in sample 211 is interpreted

as not mobilized within the combined experimental and analytical errors. This suggests

that Cu and Co are fixed in phases that are difficult to dissolve and occur with different

frequency in the two ROM samples used. Fe is significantly enriched in both ROM

samples and, within analytical and experimental error, constant in previously leached LPO.

The enrichment of Fe in previously unleached ROM can be explained with a passive

enrichment due to the preferential dissolution of Fe-free phases. Because major element

concentrations have only been measured in the solids prior to and after the experiments

(and not in the acid used for the percolation experiments), mass-balance calculations for

these elements are not possible. For trace elements, concentrations were determined in both

solids and liquids (Appendix C, Table C4 + C5). However, mass-balance calculations yield

unsatisfactory results that can be explained best with substantial evaporation of the liquid

during the exeriment in an open system (cf. Table 3.3, experimental set-up VR + V15).

In contrast, all previously leached LPO samples show, within uncertainties, coherent

results (Fig. 7.1). The remaining Cu fraction is fixed and can not be mobilized even with

repeated leaching cycles (i.e. 10 cycles in these experiments). Due to the high Zn, Co, and

Mn loads of the recirculated acid used for leaching on the heap leach pads in Sanyati, these


142

Fig. 7.1: Comparison of the observed relative elemental enrichments and depletions in

previously unleached ore (ROM = red) and previously leached ore (LPO =black)

after leaching in percolation experiments (leached with H2SO4 at a pH of 1.5

recirculated 10 times).

Cu

-80

-60

-40

-20

0

20

40

60

80

2 1 0 2 1 1 1 1 4 1 1 5 1 1 6 /2 1 1 6 /4

210 211 114 115 116/2 116/4

depl

eded

%

en

rich

edZn

-80

-60

-40

-20

0

20

40

60

80

2 1 0 2 1 1 1 1 4 1 1 5 1 1 6 /2 1 1 6 /4

210 211 114 115 116/2 116/4

depl

eded

%

en

rich

ed

Co

-80

-60

-40

-20

0

20

40

60

80

2 1 0 2 1 1 1 1 4 1 1 5 1 1 6 /2 1 1 6 /4

210 211 114 115 116/2 116/4

depl

eded

%

en

rich

ed

Mn

-80

-60

-40

-20

0

20

40

60

80

2 1 0 2 1 1 1 1 4 1 1 5 1 1 6 /2 1 1 6 /4

210 211 114 115 116/2 116/4

depl

eded

%

en

rich

ed

Pb

-80

-60

-40

-20

0

20

40

60

80

2 1 0 2 1 1 1 1 4 1 1 5 1 1 6 /2 1 1 6 /4

210 211 114 115 116/2 116/4

depl

eded

%

en

rich

ed

Fe 2O3

-80

-60

-40

-20

0

20

40

60

80

2 1 0 2 1 1 1 1 4 1 1 5 1 1 6 /2 1 1 6 /4

210 211 114 115 116/2 116/4

depl

eded

%

en

rich

ed

ROM LPO ROM LPO

ROM LPO ROM LPO

ROM LPO ROM LPO


143

fractions have not been previously mobilized (even though these ores had a residence time

of several years on the heap leach pad). Consequently, these elements are easily mobilized

during the percolation experiments and are strongly depleted in the residual ore with

average losses of ~ 60 % being observed. These findings are in agreement with the

mineralogical, textural, and geochemical observations and confirm that a significant part of

the Cu budget of the ore is not mobilized even during repeated steps of leaching.

Fig. 7.2: Relative fractions of the total Zn released as a function of the relative fractions of

the total Cu released at each sampling interval during percolation experiments

VR and V15, depicting the relative dissolution rates for these elements. See text

for details.

The dissolution behaviour of Cu and Zn during the percolation experiments is compared in

Fig. 7.2, where the relative fractions of both metals released at each sampling interval are

plotted versus each other. Coupled release of Cu and Zn via the dissolution of a single

phase or more phases where these elements are present in equal proportions should result

in a linear dissolution trend along the 1:1 line. Clearly, within the experimental and

analytical errors, three out of four of the previously technogene leached LPO samples

define linear dissolution trends along the 1:1 line pointing to the preferential dissolution of

both Cu and Zn from phases with equal Cu and Zn contents, most likely goethite and

hematite. In contrast, the previously unleached ROM samples 210 and 211, and one

previously technogene leached LPO sample (114) define dissolution trends that

significantly deviate from the 1:1 line. At the onset of leaching Zn is liberated much faster

than Cu from these samples, followed by linear dissolution trends with Cu:Zn ratios of 1:3

0

20

40

60

80

100

0 20 40 60 80 100

Cu % of max. extraction

Zn %

of m

ax. e

xtra

ctio

n 210

211

114

115

116/2

116/4


144

(sample 211; R2 = 0.888) and 2:1 (samples 210 and 114; R2 = 0.927 and 0.996,

respectively). The non-linear dissolution behaviour points to the presence of rapidly

dissolving Zn-bearing phases at the onset of the experiment. This is followed by the

dissolution of one or more phases with equal Cu and Zn contents. These observations

demonstrate that the leaching progress of each individual ore sample might be influenced

by the presence of modally less abundant mineral phases.

Another important factor controlling the liberation of base metals from ores is the grain

size distribution within the ores. In the case of Sanyati, where the grain size distribution of

the ROM is not modified prior to heap leaching, it is assumed that the permeability of the

heap leach pad is high enough to allow dissolution of metals from larger grains. In order to

simulate the effect of larger grain sizes on leaching, a percolation experiment was

performed using a massive ore block (ø 5 cm, sample 212/4) embedded in a quartz matrix

(see Chap. 3.3.3 for a detailed description of the experimental design). A representative

aliquot of the block employed for the experiment was used separately for chemical

analyses. It was found that the block contained higher Fe (about a factor of ~2) and lower

Si contents compared to the fine grained ore, an observation that can be attributed to the

higher amount of silicate phases in the fine grained ore.

Fig. 7.3: Comparison of the observed relative elemental enrichments and depletions in fine

grained ROM (sample 210) and in an embedded coarse grained, massive ore

block (sample 212/4) after leaching in percolation experiments (leached with

H2SO4 at a pH of 1.5 recirculated 10 times).

In Fig. 7.3 the results for Cu, Zn, and Co of this experiment are compared to the results of

an experiment (sample 210) using ROM with a grain size < 1 cm. Although the

Zn

-80

-40

0

40

80

210 212/4

depl

eted

%

en

rich

ed

Cu

-80

-40

0

40

80

210 212/4

depl

eted

%

en

rich

ed

Co

-80

-40

0

40

80

210 212/4

depl

eted

%

en

rich

ed


145

permeability still allows the dissolution of base metals, the recovery is clearly below the

recovery from the fine grained ore. Only 12 % Cu and 12 % Zn are leached from the

coarse grained ore block, whereas up to 29 % Cu and 53 % Zn are leached from the fine

grained sample. The relative high liberation of Co (42 %) from the bloc compared to Cu

and Zn can best be attributed to a non-ferrous phase (e.g. manganese oxyhydroxide or

cobalto-adamite).

7.1.2 Leaching experiments under idealized conditions

In order to estimate the potential metal output under "ideal" leaching conditions

experiments simulating an agitated leaching process were carried out (experiments V1 and

V6 in Table 3.3 in Chap. 3.2.2). In these experiments fine grained leach pad ore (sample

MM1, < 63 µm) was leached in a surplus of acid and stirred constantly for 164 hrs.

Samples of the solution were taken during the entire duration of the experiment in

increasing time intervalls. The solution samples were filtered to ensure removal of any

suspended matter from the solution. In experiment V6 samples of the suspended finest

fractions were additionally taken with a syringe 5 cm below the solution top at the initial,

intermediate, and final leaching stage of the experiment.

In Fig. 7.4 the concentrations of Fe, Cu, Zn, As, Mn, and Co in the solution after leaching

are plotted as a function of leaching duration. The leaching behaviour and the observed

absolute amounts mobilized (with the exception of As) in both experiments (V1 and V6)

are in very good agreement. The observed leaching curves can be divided into three

different types: (1) the first type (i.e. Fe, Cu, and Zn; Fig. 7.4 a, b, and c) is characterised

by a strong increase in the first 5 h of the experiment followed by a moderate more ore less

continuous increase of the dissolution rates; (2) the second type (i.e. Mn and Co; Fig. 7.4 e

and f) shows contrasting behaviour with no or minor dissolution during the first 92 hrs

leaching time followed by a strong increase of the dissolution rate during the rest of the

experiment; and (3) the third type (i.e. As; Fig. d) is characterised by steep initial slopes

with a peak at ~ 20 hrs leaching time followed by a moderate continuous decrease

indicating subsequent removal (i.e. precipitation) of the elemental species from the

solution. Except for As no saturation of the solution is reached during the entire

experiment.


146

Fig. 7.4: Concentration of Fe, Cu, Zn, As, Mn, and Co in the acid used (H2SO4 at a pH =

1.5) in the permanently stirred leaching experiments V1 and V6 as a function of

leaching time.

In experiment V1 additionally to the bulk sample three samples of filter residues from the

acid sampling procedure at the initial (sample F1), intermediate (sample F2), and final

stage (sample F3) of the experiment were analysed (see Chap. 3.2.2.). The recoveries

Fe

0

50

100

150

200

250

0 50 100 150 200[h]

[mg/

l]

V1

V6

Cu

0

2

4

6

8

10

12

14

0 50 100 150 200[h]

[mg/

l]

V1

V6

Zn

0

5

10

15

20

0 50 100 150 200[h]

[mg/

l]

V1

V6

Co

0

5

10

15

20

25

0 50 100 150 200[h]

[mg/

l]

V1

V6

Mn

0102030405060708090

0 50 100 150 200[h]

[mg/

l]

V1

V6

As

0

5

10

15

20

0 50 100 150 200[h]

[mg/

l]

V1

V6

a. b.

c. d.

e. f.


147

relative to the starting compositions for Cu, Zn, Al, and Fe from the bulk ore and the filter

residues are shown in Fig. 7.5. The most striking observation is that, in contrast to the

percolation experiments, approximately 30 % of the Cu retained after technogene leaching

is liberated additionally. However, significant mobilisation of base metals is not achieved

from the filter residues, which represent the very fine fraction of the bulk sample

suspended in the acid during the experiment. The only element that is subsequently

mobilized from the filter residues is Al, suggesting that a clay mineral is liberated from the

fraction and progressively dissolved during the experiment.

Fig. 7.5: Compositional changes in the bulk ore residue (V1) and three filter residues (fine

fraction sampled; F1, F2, and F3) relative to the starting composition resulting

from leaching under idealized conditions.

The dissolution behaviour of Cu, Zn, Co, and Fe during idealized leaching is compared in

Fig. 7.6. The dissolution trends for Cu, Zn, and Fe are linear, indicating equal dissolution

of all three elements from the liberation of Cu and Zn from iron-bearing phases (i.e.

goethite or hematite). In contrast, the exponential variation of Co (and, not shown, Mn)

Cu

-50

-25

0

25

50

V1 F1 F3 F5

depl

eted

%

en

riche

d

Zn

-50

-25

0

25

50

V1 F1 F3 F5

depl

eted

%

en

riche

d

Al2O3

-50

-25

0

25

50

V1 F1 F3 F5

depl

eted

%

en

riche

d

Fe2O3

-50

-25

0

25

50

V1 F1 F3 F5

depl

eted

%

en

riche

d


148

with Fe indicates that these elements are dissolved from different phases (i.e. a manganese-

oxyhydroxide).

Fig. 7.6: Comparison of relative dissolution rates during idealized leaching (V6): a.

Relative fractions of the total Zn released as a function of the relative fractions of

the total Cu released at each sampling interval. b. Relative fractions of the total

Cu, Zn, and Co as a function of the relative fractions of the total Fe released at

each sampling interval.

7.1.3 Metal production rates and rate equations

In Fig. 7.7 the concentrations of Cu, Zn, and Fe in the solution are plotted as a function of

time for the "idealized" leaching experiment V6 (for the part > 5 h). Adapting the methods

and nomenclature of GLEISNER & HERBERT (2002), metal production rates have been

calculated from the gradients of the linear regressions shown. In a similar fashion, metal

production rates have been calculated for Co and Mn (not shown in Fig. 7.7) for the two

curve segment sections showing markedly different dissolution rates, i.e. before and after

92 h (cf. Fig. 7.4).

Fig. 7.7: Cu, Zn, and Fe concentrations vs time in the "ideal" leaching experiment V6 after

5 hrs. Production rates (reported as mmol/l*h) are highlighted.

R2 = 0.9782

0

20

40

60

80

100

0 20 40 60 80 100

Cu [% of max. extraction]

Zn [%

of m

ax. e

xtra

ctio

n]R2 = 0.995

R2 = 0.9706

0

20

40

60

80

100

0 20 40 60 80 100

Fe [% of max. extraction]

Me

[% o

f max

. ext

ract

ion]

Zn

Cu

Co

Cu

y = 0.0011x + 0.0443R2 = 0.9709

0.00

0.05

0.10

0.15

0.20

0.25

0 100 200[h]

[mm

ol/l]

Zn

y = 0.0014x + 0.048R2 = 0.9881

0.000.050.100.150.200.250.300.35

0 100 200[h]

[mm

ol/l]

Fe

y = 0.0211x + 0.2579R2 = 0.9774

0.00

1.00

2.00

3.00

4.00

5.00

0 100 200[h]

[mm

ol/l]


149

Cu and Zn are dissolved with similar production rates, while the production rates for Fe are

an order of magnitude higher. This observation is in good agreement with the observation

that Cu, Zn, and Fe are equaly strong dissolved from goethite and hematite. Co, that is

most likely hosted by a manganese-oxyhydroxide shows a production rate similar to that of

manganese after 92 h (Table 7.1). Arsenic has a very high initial production rate and is

precipitated (cf. Fig. 7.5) in the further course of the experiment (Table 7.1).

Table 7.1: Metal production rates calculated from leaching experiment V6.

Production rate region <5 h

Production rate region >5 h

[mol/l*s] [mol/l*s] Cu 5.76*10-6 3.24*10-6 Zn 1.01*10-5 4.68*10-6 Fe 2.13*10-4 7.49*10-5 region <92 h region >92 h Co-1 n.d. Co-2 1.58*10-5

Mn-1 9.72*10-6 Mn-2 5.58*10-5

region <20 h region >20 h As-prod 2.37*10-5 As-prec -1.8*10-6

In order to characterise the dominant dissolution mechanism the experimentally

determined dissolution curves were modelled with rate equations as outlined by CORNELL

& SCHWERTMANN (1996) and HABASHI (1969). The correlation coefficient R2 was used as

a criterion to evaluate the likelihood of a given dissolution mechanism. Three equations

which describe different mechanistic types of element mobilisation from minerals were

used (cf. Table 7.2).

Table 7.2: Rate equations used for modelling the dissolution kinetics of experiment V6

(from BROWN et al. 1980; GIOVANOLI & BRÜTSCH 1975, cited in CORNELL &

SCHWERTMANN 1996; and HABASHI 1969).

Rate equation Physical meaning -ln(1-X) random nucleation (first order) [1-(1-X)1/3] phase boundary controlled for a contracting sphere (cube

root) [1-2/3X-(1-X)2/3] three dimensional diffusion (Ginstling-Brounstein)

reaction controlled by diffusion through a non-porous solid product

Because the dissolution curves show a change in dissolution behaviour after 5 h for Cu,


150

Zn, and Fe, and 92 h for Co and Mn (Fig. 7.4) these two segments were treated separately

(Table 7.3). In region 1 (i.e. up to 5 h for Cu, Zn, and Fe, and 92 h for Co and Mn) the

behaviour of Cu, Zn, Fe, Mn and Al is best described (highest R2) by the Ginstling-

Brounstein equation [1-2/3X-(1-X)2/3], which describes dissolution with the build-up of a

diffusive rim around the decaying grain through which the ions have to move by diffusion.

This scenario is plausible especially at the beginning of the experiment where adsorbtive

bound base metals are liberated at a high rate while goethite and hematite do not break

down to a large extent. In contrast, in region 2 (i.e. after 5 h for Cu, Zn, and Fe or 92 h for

Co and Mn) the behaviour of Zn, Fe, Mn and Al is equally well described by random

nucleation, where the surface control is rate-determining (i.e. the reaction rate is controlled

by the surface left at any specific time) or a phase boundary controlled contracting sphere

model (i.e. the geometry of the solid influences dissolution of a particle and the interface

moves inwards at a constant rate (implying isotropic dissolution). This indicates consistent

liberation during the decay of the ironoxides and -oxyhydroxides.

Table 7.3: k and R2 for rate equation modelling of experiment V6.

Region 1 Region 2

k R2 k R2 k=-ln(1-X) Region 1 (< 5 h) Region 2 (> 5 h) Cu 2E-06 0.8167 2E-07 0.984 Zn 9E-06 0.7789 1E-07 0.9962 Fe 7E-07 0.8151 4E-08 0.9956 Region 1 (< 92 h) Region 2 (> 92 h) Mn 3E-06 0.7639 4E-07 0.9777 Al 7E-07 0.7412 9E-08 0.9794

k=[1-(1-X)1/3] Region 1 (< 5 h) Region 2 (> 5 h) Cu 2E-06 0.9075 6E-08 0.984 Zn 4E-07 0.7026 4E-08 0.9962 Fe 2E-08 0.9591 1E-08 0.9956 Region 1 (< 92 h) Region 2 (> 92 h) Mn 1E-06 0.7639 1E-07 0.9777 Al 2E-07 0.7412 3E-08 0.9794

k=[1-2/3X-(1-X)2/3] Region 1 (< 5 h) Region 2 (> 5 h) Cu 1E-13 0.9922 8E-13 0.9909 Zn 2E-12 0.9985 5E-13 0.9839 Fe 1E-13 0.9791 4E-14 0.966 Region 1 (< 92 h) Region 2 (> 92 h) Mn 7E-12 0.9991 3E-12 0.9053 Al 3E-13 0.9925 2E-13 0.9117


151

7.2 Extraction experiments on goethite- and hematite-rich supergene ores

Extraction procedures have long been used to estimate the relative mobility of elements

and, if possible, to unravel to which minerals they are bound (e.g. TESSIER 1979). Because

of the plethora of reactants available, the extraction procedure employed has to be adjusted

as good as possible to the composition of the material investigated. In this study a

mineralogically controlled extraction procedure for supergene porphyry ore waste (DOLD

1999) was adopted and modified according to the mineralogical composition of the

supergene ore at Sanyati (Table 7.4). The emphasis of the extraction experiments was to

gain further insight into the residence of the base metals to the supergene ores, especially

the elements adsorbed or bound to goethite and hematite.

Table 7.4: Experimental conditions and materials used for extraction experiments on

sample 263.

Solvent Liberated fraction Solid/liquid ratio

Conc. [M] / pH

T [°C]

Ref.

NH4-Acetate

sorbed/exchangeable fraction

1:500 1 / pH 4.5 25 DOLD (1999), CARDOSO FONSECA (1986)

NH4-Oxalate- (darkness)

dissolution of jarosite, ferrihydrite etc.

1:500 0.2 / pH 3.0 25 SCHWERTMANN (1964), DOLD (1999)

NH4-Oxalate-(80° C)

dissolution of goethite, hematite

1:500 0.2 / pH 3.0 80 DOLD (1999)

A goethite and hematite-rich sample (263) was washed with demineralized water and three

aliquots were treated using different reactants with particular dissolution capabilities

(Table 7.4). A further experiment with HCl was planned but proved to be not necessary as

hot NH4-Ox-80° dissolved the sample completely. Although the reactants have been

chosen to dissolve particular fractions of the samples, some overlaps between the fractions

dissolved have to be considered.

A fraction of Cu (19 %) is bound to the exchangeable fraction (Fig. 7.8). Mass balance

calculations based on the modal abundance of hematite and goethite indicate that the

amount of Cu adsorbed to the goethite surface is ~ 7.5 mmol/kg.

A comparison of the relative fractions yielded by the different extraction techniques is

shown in Fig. 7.8. The most striking feature of Fig. 7.8 is that the extraction technique that


152

completely dissolves goethite and hematite yields by far the highest extraction rates. Only

35 % of Cu, 15 % of Zn, but 79 % of As can be extracted without the dissolution of

goethite and hematite. These findings provide further compelling evidence for a dominant

role of goethite and hematite in governing the mobilization behaviour of a range of

elements during leaching. Furthermore, these findings clearly demonstrate that Cu and Zn

are not bound to goethite and hematite via simple, easily reversible adsorbtion processes.

The observation that the majority of As (79 %) is already mobilized in the easily

exchangeable fraction is important for environmental considerations. Mn and Co are both

mobilized in NH4-Ox-80° in agreement with the assumption that Co is bound to a

manganese-oxyhxdroxide and is therefore showing a dissolution behaviour simultaneous to

Mn (cf. Fig. 7.4).

Fig. 7.8: Relative fractions of metals and As mobilized from the goethite- and hematite-

rich supergene ore sample 263 by the different extraction techniques used.

In Fig. 7.9 the element concentration of the solutions is normalised to the final solution

concentration and is plotted as a function of the time for the experiments. Systematic

curves are only obtained in the NH4-Ox-80°C experiment (Fig. 7.9 b). This also shows that

distinct mineral phases are dissolved. In the experiments NH4-Ac (not shown) and NH4-

Ox-d (Fig. 7.9 a) random distribution points to the overlapping of desorption and

dissolution processes.

0%

20%

40%

60%

80%

100%

Al Mn Fe Co Cu Zn As

NH4-ox-80°

NH4-ox-d

NH4-ac


153

Fig. 7.9: Extraction of Fe, Cu, Zn, Al, and Si in goethite- and hematite-rich supergene ore

with a.) NH4Ox-d performed in darkness at room temperature and b.) NH4Ox-80°

performed in daylight at 80°C. The data are normalised to concentration at the

end of the experiment. Systematic dissolution is only found in NH4Ox extraction

in daylight at 80°C.

7.3 Adsorption experiments of Cu onto goethite

The extraction experiments indicate that 19 % of the Cu is bound in the exchangeable

fraction, i.e. is adsorbed onto mineral surfaces. The existing experimental evidence

suggests that adsorption of positively charged metals onto ironoxyhydroxides, e.g.

goethite, is usually strongly decreasing with decreasing pH (e.g. CORNELL &

SCHWERTMANN 1996). However, sulfate ions enhance the capability of ironoxyhydroxides

to adsorb positively charged metals (e.g. ALI & DZOMBAK 1996), but it is still relatively

low. However, there are discrepancies between the experimental conditions used in these

studies and the conditions during technogene decay on the heap leach pad. The most

important discrepancy with respect to the adsorption of Cu is that the sulfate

concentrations of the solutions used in the existing experiments are much lower compared

with the solutions employed for technogene leaching. Therefore experiments in an acidic,

sulfate-rich system using natural and synthetic starting materials have been performed.

Three Cu minerals handpicked to optical purity from Sanyati ore samples were dissolved

in the presence of synthetic goethite in sulfuric acid (pH 2). For experimental details see

Chap. 3.2.2.

The temporal variation of the solution chemistry (Cu, Fe, and H+ concentrations) and its

0.0

0.2

0.4

0.6

0.8

1.0

0 2 4 6 8 10

time [h]

0.0

0.2

0.4

0.6

0.8

1.0

0 2 4 6 8 10

time [h]

Fe

Cu

Zn

Al

Si

a. b.

NH4Ox-d NH4Ox-80°C


154

conductivity are shown in Fig. 7.10 for all three experiments. In the initial stage of all three

experiments Fe (i.e. goethite) is partly dissolved and afterwards constantly re-precipitated

during the rest of the experiment. As expected, the dissolution of Fe is pH-sensitive and Fe

is most readily dissolved in the experiment with CuSO4 * 2 H2O where pH values are the

lowest. Due to the mobility of H3O+ the conductivity in this experiment is the highest.

Fig. 7.10: Dissolution of malachite, chrysocolla, and copper sulfate (syn.) in the presence

of fine grained goethite (syn.) at pH 1.5.

From the Cu-bearing phases, 48 wt% of the chrysocolla, 74 wt% of malachite, and 80 wt%

of copper sulfate are dissolved in the initial stage of the experiment. During the

adsorption/precipitation phase of the experiment Cu is systematically removed from the

solution in the synthetic Cu sulfate system and the malachite system. In contrast, in the

chrysocolla system the Cu content of the solution decreases only slightly and stays

approximately constant for the rest of the experiment. The dissolution/adsorption

mechanisms of Cu in the silica-bearing chrysocolla system thus differs significantly from

the malachite and synthetic CuSO4 * 2H2O systems. This behaviour can not be explained

pH

1.8

2

2.2

2.4

2.6

0 50 100 150 200

[h]

cond.

1.8

2.2

2.6

3

3.4

3.8

0 50 100 150 200

[h]

[mS/

cm]

Cu

40

50

60

70

80

90

100

110

0 50 100 150 200

[h]

[mg/

l]

MALCRYCuSO4*2H20

Fe

0

10

20

30

40

50

60

70

0 50 100 150 200

[h][m

g/l]


155

by differences in pH, conductivity or Fe-content of the solution. Sorption in the malachite

and synthetic CuSO4 * 2H2O occurs in an environment dominated by the presence of

sulfate anions. In the malachite system, the sulfate present is solely derived from the acid

because the carbonate readily leaves the solution as CO2 (0.1 mol/l). In the Cu-sulfate

system, additional sulfate (2.6 mol/l) is liberated from the dissolving solid. In total, 49 and

51 wt% of Cu were adsorbed onto goethite in the malachite and the Cu-sulfate

experiments, respectively. The observation is in good agreement with work from ALI &

DZOMBAK (1996) who investigated the sorption of Cu onto goethite in the presence of

sulfate (at lower sulfate concentrations of 0.25 and 1 mM/l) and found enhanced sorption

in the low pH-range.

The contrasting behaviour of Cu in the chrysocolla system is best explained with the

presence of significant amounts of silica (1 mmol/l ~ 28 mg/l). It is suspected that Cu is

stabilized in the solution as silica complexes. Complexation of Cu with polymeric silica at

higher concentrations and pH is known from the literature (e.g. YATES et al. 1998), as well

as complexation with amorphous silica, e.g. Ca and Mg (SANTSCHI & SCHINDLER 1974). A

complexation of Cu would change its properties (e.g. surface charge), and thus might retain

it in the solution. To elucidate the speciation of Cu in silica-bearing solutions (i.e.

speciation with amorphous or polymeric silica) it would be necessary to separate the Cu-

bearing silica by ultrafiltration.

As shown by the different experiments base metal fixation occurs in different types (e.g.

latticebound versus adsorbed). The mechanisms leading to base metal fixation by

ironoxides and -oxyhydroxides will be further discussed in Chapter 8.3.


An experimental comparison of the H2SO4-leaching behaviour of ROM and LPO from

Sanyati shows that ~ 60 % of Zn, Co, and Mn can be mobilized from both unleached and

previously leached ore. In contrast, Cu is only leachable from previously unleached ore

(Fig. 7.1) and can not be mobilized from LPO even during repeated steps of leaching.

In order to constrain the maximal possible metal liberation from the LPO leaching

experiments under "idealized" conditions have been performed. Under these conditions ~

30 % of Cu can be liberated additionally (Fig. 7.5). The most plausible factors contributing

to this enhanced mobilisation is a better degree of liberation of the base metal minerals, the


156

formation of new limonite surfaces accessible for leaching, and limited dissolution of

goethite and hematite. The correlation of Cu, Zn, and Fe points to goethite and hematite as

their common source (Fig. 7.6). In contrast, the main Co hosts are manganese phases (Fig.

7.6).

Rate equation modelling was performed in order to determine the most probable

dissolution mechanism. For Cu, Zn, Fe, Mn, and Al the dissolution mechanism in the first

stage of the experiment (region 1) is best explained by the Ginstling-Brounstein equation,

which describes a three dimensional dissolution of a sphere that is controlled by a non-

porous solid layer through which the ions have to diffuse. In this scenario adsorbtive

processes probably dominate. In the second stage (region 2) the dissolution behaviour is

best explained by simpler reaction models, e.g. a consumed shrinking sphere, indicating

the breakdown of iron phases.

Partial extraction experiments revealed that the overwhelming majority of Cu, Zn, and Co

is lattice bound in iron- and manganese oxyhydroxides. The differing behaviour of As,

where ~ 79 % is bound adsorptively, has important implications for environmental

considerations.

Adsorption of Cu on goethite was experimentally investigated in acidic, sulfate-rich

systems. These experiments demonstrated that significant amounts of Cu can be removed

from the solution in silicate-free systems. In silica-bearing systems the formation of silica

complexes might stabilize Cu in the solution.

8. "Invisible" base metal concentrations in supergene ironoxides and -oxyhydroxides

157

8 Formation of “invisible” base metal concentrations in supergene goethite and hematite and their consequences for the leaching process

In this Chapter the development of copper- and zinc-rich ironoxyhydroxides and -oxides

will be summarized using the supergene ore composition and the reconstruction of the

weathering regime as a background. Subsequently, the mechanisms of metal fixation

processes and the consequences of this metal retention will be discussed.

8.1 Development of the oxidation zone in Sanyati and the composition of the

supergene ores

In the Sanyati region weathering has taken place since the Miocene (LISTER 1987) in a

warm and humid climate. A schematic sketch summarizes the weathering situation (Fig.

8.1).

Fig. 8.1: Schematic sketch depicting the weathering and alteration situation of the

proximal and distal rocks of the Sanyati region.

Pliocene erosion extensively exhumed lower horizons of an originally about 100 m deep

advanced weathering profile. The supergene ore was mainly preserved from erosion due to


158

its higher resistance to weathering and is now forming positive morphological anomalies as

relict mountains.

In the proximal host rock sulfuric acid solutions from the decay of sulfides dominate the

groundwater (Chap. 4.2.4). Therefore the degradation of minerals is much more intense in

the proximal rocks and leads to an extreme depletion of alkali and alkaline earth elements

(Fig. 4.1). As a consequence kaolinite is frequent in the weathering products of proximal

rocks. Additionally alunite occurs, which is formed in a sulfate-bearing acidic milieu and

conditions at or below the watertable (STOFFREGEN et al. 2000).

Amphibolites, which are only found proximal to the mineralisation, are also extremely

weathered to chlorite-dominated products (Table 4.4). Metadolomites were found only in

the lowest parts of the oxidation zone.

The weathering process is still progressing, as can be seen from the composition of the

groundwater at the base of the open pits (69 mg/l Cu, 290 mg/l Zn) and efflorescences of

sulfates (e.g. chalcantite) where metal-bearing solutions have dried.

On the basis of the chemical index of alteration (CIA) as defined by NESBITT & YOUNG

(1989) and the acid neutralisation capacity (ANC), the proximal and distal phyllites and

amphibolites can be classified as: (1) rocks showing a good CIA-ANCB negative

correlation and (2) rocks showing a high CIA as well as ANCB (Fig. 4.3). Metal

concentrations of the host rock are generally higher in the first group (Table 4.3). This

point to adsorption or incorporation to host rock minerals, most likely sheet silicates,

during late-stage solution-rock interaction. Some metal retention on the clay minerals

(kaolinite, illite, and smectite) and sheet silicates (chlorite, mica, and biotite) is very likely.

Especially clays with a low zero-point-of-charge surface adsorb trace metals, e.g. clays

from the kaolinite-montmorillonite group (STUMM & MORGAN 1981; BOWELL & BRUCE

1995). Metal retention on biotite has been observed by ÖHLANDER et al. (2003).

MAGOMBEZE & SANDVIK (2002) also report the occurrence of Cu-bearing chlorites in the

Sanyati host rock. These observed trace element anomalies in exploration targets usually

extend over some distance into the wallrock (TAYLOR & SCOTT 1983).

At Sanyati the oxidation zone is developed in steep dipping host rocks, which leads to a

variable water access to the ore. The metal sources of the oxidation zone were massive as


159

well as disseminated polymetallic sulfide ores with a "patchy" sulfide distribution

(BAHNEMANN 1961), i.e. variability in thickness and composition. Of the ore-forming

sulfides pyrrhotite, pyrite, chalcopyrite, and arsenopyrite are classified as acid producers,

while sphalerite and galena are non-acid producers (DOLD 1999). Variable acid production

influenced the degree of degradation of surrounding ores and silicate minerals. Therefore

the composition of the supergene ore is highly variable, as was shown in profiles across

various supergene ore lenses (FREI & GERMANN 2001b). Several lines of investigation

were followed in order to determine the degree of maturity of the oxidation zone (Table

4.10). They point to a low degree of maturity. Except for the alteration of unstable

ferrihydrite and schwertmannite to goethite, no indications for zoning, i.e. development of

an upper depleted part in the oxidation zone, was found. Thus the complete oxidation zone

can be classified as fertile and is adequate for mining.

The metal values of the supergene ore are bound to a variety of carbonate (e.g. malachite),

silicate (e.g. chrysocolla, hemimorphite), oxide (e.g. cuprite, tenorite), sulfate (e.g.

brochantite), and arsenate (e.g. olivenite, adamite) minerals. However, a significant

fraction is bound to goethite and hematite (see Chap. 6).

To summarize, the formation of the supergene deposit by degradation of acid-producing as

well as non-acid-producing sulfides has taken place since the Miocene. The acidic

solutions, that have dissolved the primary sulfides, are the metal source for the base metal

minerals, the metal-bearing limonites, as well as the metal-enriched proximal host rocks

which form a halo around the mineralisation. Sulfide ore decay is still progressing and the

steep-dipping rocks as well as the "patchy" ore texture lead to a very heterogeneous oxide

ore, with a large fraction of metals bound in a variety of minerals with a significant

fraction bound to goethite and hematite.

The mineralogy of the proximal metasediments is strongly influenced by the acidic sulfate-

rich solutions derived from the sulfide decay. Therefore, weathering products like kaolinite

and alunite are frequently observed.

8.2 Distribution of base metals in sulfide decay textures - colloform textures as a

proxy for the "invisible" base metal contents in goethite and hematite

Unequivocal textures of sulfide decay can be preserved in supergene ores under favourable


160

conditions (e.g. large grain sizes, euhedral crystals, and a relatively high pH especially at

the beginning of the oxidation process). These early precipitates are then retained

throughout the further development of the oxidation zone (NICKEL & DANIELS 1986) and

typical textures are known for most of the frequent sulfide minerals (BLANCHARD 1968;

NICKEL & DANIELS 1986). At Sanyati textures typical for the decay of chalcopyrite and

sphalerite were distinguished from textures typical for pyrrhotite and pyrite. Additionally,

textures with a varying degree of voids were examined. This group of residual textures can

be distinguished from a group of colloform textures that have been formed from

transported metal loads.

Most precipitation textures show a distinctive zonation characterised by light and dark

reflecting zones. In colloform textures these zonations are most pronounced. The light

reflecting zones are generally higher in Fe, but the Fe contents vary strongly and only

rarely reach the composition of pure hematite (i.e. 69.94 wt% Fe). The Fe contents in the

dark reflecting zones are also very variable and often below the Fe contents of pure

goethite (62.8 wt% Fe). A negative correlation between Fe and O is found in both cases.

Therefore the light and dark areas are best characterised as cryptocrystalline, partly

hydrated hematite- and goethite-rich bands.

Generally Cu and Zn are enriched in the dark reflecting zones. For As, Pb, and Al the

distributions suggest that As and Al are perferentially enriched in the dark reflecting zones,

while Pb is enriched in the light reflecting zones. However, the behaviour of these

elements is less systematic. Similar trends for Cu, Zn, Pb, and Al are reported by SCOTT

(1992). For As, SCOTT (1992) describes a random distribution in both goethite and

hematite. The distribution of trace elements (Ga, Ge, Se, Ag, Cd, and Sb) also points to an

enrichment of these elements in the dark reflecting, goethite-rich zones.

The parallel textures, which have been examined with respect to their varying degree of

filling with dark reflecting, goethite-rich material, show an increasing Fe content with

decreasing degree of filling. In contrast, the Cu and Zn contents are decreasing with

increasing degree of filling. This can be interpreted as a maturation trend. With increasing

dehydration goethite alters to hematite, which incorporates less base metals. The base

metals are in turn subsequently released into the solution with increasing degree of

maturity.

Mn and Co contents are low in goethite and hematite and the observations point to the

occurrence of manganese-oxyhydroxides that incorporate Co.


161

The textural and microchemical investigations show that the mobility of minor and trace

elements in the system was so high, that correlations are found on a sample scale rather

than on a mineral grain scale. Neither are Cu and Zn contents especially high in textures

typical for the base metal sulfides, nor are they depleted in textures typical for Fe sulfides.

Additionally, the trace elements Ga, Ge, Se, Ag, Cd, and Sb are controlled on a local scale

by solution chemistry and not by their textural setting.

These uniform metal distributions suggest a relative high element mobility leading to a

rather uniformly composed liquid phase on at least the dm3-scale. Therefore the colloform

textures precipitated from transported liquids mirror the average composition of goethites

and hematites best (Fig. 8.2). The observed deviations in sample 268 are due to the fact

that the colloform texture analysed is precipitated in a fracture and might have been fed

from a wider surrounding and/or reflect the information of different fluid pulses.

These findings suggest that the colloform textures are a powerful tool that might be used to

approximate the "invisible" base metal fractions retained in the goethite and hematite.

Fig. 8.2: Comparison of the mean Zn concentration in goethite and hematite in a section

and their mean in the corresponding colloform textures.

In summary, precipitation textures, which are mainly composed of cryptocrystalline, partly

hydrated hematite- and goethite-rich bands, incorporate metals in a systematic fashion. For

most metals (Cu, Zn, As, Al, Ge, Ga, Se, Ag, Cd, Sb) higher enrichment is found in the

dark reflecting goethite-rich zones. The light reflecting hematite-rich zones are enriched in

Pb. The degree of void filling of the textures shows a maturation trend: with increasing

Zn

0

1

2

3

4

294

294 c

oll21

2/2

212/2

coll

268

268 c

oll29

4

294 c

oll 7979

coll

212/2

212/2

coll

268

268 c

oll

mea

n co

nc. [

%]

light areas dark areas


162

maturation goethite alters to hematite and the base metals (e.g. Cu and Zn) are released.

A relative high element mobility is found on a sample scale. The average composition of

the ironoxide and ironoxyhydroxide precipitations is reflected best and easy to determine

in the colloform textures. These textures are easily found using optical microscopy.

Therefore, these textures might provide a powerful tool to gain information about the

"invisible" base metals fixed in limonites. However, the high element mobility prevents a

correlation of base metal concentrations in the characteristic boxwork textures to their

precursor sulfides.

8.3 Fixation of base metals by goethite and hematite

Enrichment of base metals in iron phases is important in various geological environments

and can control the trace metal budget of, e.g. soils, lake sediments, and oxidation zones.

The scavenging of metals to iron-rich rocks of the oxidation zone is used as a tool in

exploration for ore deposits (e.g. CHAO & THEOBALD 1976; SCOTT 1986). Adsorption of

copper to ironoxyhydroxide phases controls the release of copper during sulfide oxidation.

For example, ÖHLANDER et al. (2003) examined till which transported sulfide ore. They

showed that the finer fraction of the sulfides decayed and Cu, Co, Ni, and Zn were

redistributed and subsequently retained in newly formed ironoxyhydroxides.

In acid mine drainage systems Cu and Zn adsorb to the iron phases precipitating. Under

acidic conditions adsorption of Cu and Zn is enhanced on sulfated goethite relative to pure

goethite, ferrihydrite, and schwertmannite (SWEDLUND & WEBSTER 2001). While for

schwertmannite the high amounts of adsorbed Cu and Zn can be satisfactory modelled by

the formation of ternary complexes (FeOHMeSO4), the observed Cu and Zn adsorption

onto sulfated goethite remains difficult to explain (WEBSTER et al. 2000; WEBSTER et al.

1998). Adsorption experiments performed in natural and synthetic systems at different pH

reveal that 100 % of the initial Cu, Zn, Cd, and Pb present is adsorbed at high pH by both

natural and synthetic goethite. However, in low pH-environments (pH = 3-4) the

adsorption for these elements is below 20 % of the initial amount present, and the

adsorption to natural goethite is generally larger compared to synthetic goethite (WEBSTER

et al. 1998).

Adsorption and coprecipitation processes both play a role in metal scavenging. In theory,

the former is the two-dimensional accumulation of ions at the interface between a


163

preformed solid phase and an aqueous phase, and the latter is the simultaneous removal of

ions during the formation of metal precipitates, e.g. ironoxyhydroxides (LEE et al. 2002).

However, coprecipitation may include adsorption, the formation of clusters, occlusions,

solid solutions, or a combination of these processes (KARTHIKEYAN et al. 1997; MARTINEZ

& MCBRIDE 1998).

As shown in the previous chapters, the base metal contents of the iron phases present in the

Sanyati ores are elevated and bound to the ubiquitous ironoxyhydroxide goethite and the

ironoxide hematite, and to a minor extent the iron sulfates plumbojarosite/jarosite.

Laboratory experiments have demonstrated that certain base metals are showing

contrasting leaching behaviour and hence different element mobilities. This observation

can be attributed to the different relative amounts adsorbed onto mineral surfaces or

incorporated into the lattice via coprecipitation. In the following, these two fundamentally

different processes and their importance for the incorporation of base metals into goethite

and hematite during natural and technogene leaching are discussed in more detail.

8.3.1 Adsorption to goethite and hematite

The fundamental goal of all adsorption studies is the formulation of a mechanistic model

that describes the metal adsorption to the solid phase in question. If the mechanism that

leads to metal adsorption is well established, the adsorption process is amenable to

thermodynamic treatment and quantitative predictions of element mobilities as a function

of the physico-chemical conditions are possible (if sufficient thermodynamic data are

available).

Numerous studies investigated the adsorption behaviour of base metals onto

ironoxyhydroxides. However, the focus of most of these studies has been the behaviour of

goethite in sea- and freshwater environments (e.g. BALISTRIERI & MURRAY 1982; FORBES

et al. 1976; KOONER 1992) and less attention has been drawn to the reaction of

ironoxyhydroxides in sulfate-rich, acidic (< pH 2) systems (JUANG & WU 2002)

comparable to those on the heap leach pads.

In the supergene ores of Sanyati only a relatively small amount of Cu and Zn is adsorbed


164

to mineral surfaces. As was shown in extraction experiments, only 19 % of the Cu and 6 %

of the Zn belongs to the exchangeable fraction that can be attributed to adsorption on

mineral surfaces (DOLD 1999; CARDOSO FONSECA 1986).

Key parameters of adsorption are the charges of the mineral surfaces and the sorbent. The

goethite surface is formed by hydroxyl groups and is therefore generally amphoteric

(CORNELL & SCHWERTMANN 1996). This means that the surface charge decreases with

increasing pH. Therefore adsorption of cations to this surface is generally significant only

above pH 4, where Cu forms inner-sphere complexes on the goethite surface (PEACOCK &

SHERMAN 2004). PARKMAN et al. (1999) describe them as Jahn-Teller distorted octahedral

complexes. BONNISSEL-GISSINGER et al. (1998) found that especially above a pH of 4

ironoxyhydroxide species (i.e. Fe3+, FeOH2+, Fe(OH)2+) were found on the surface of

goethite, but they do not passivate it. These species play an important role in cation

adsorption during the decay of sulfides.

Fig. 8.3: Schematic representation of the distribution of positive, negative and neutral

surface hydroxyl groups on an ironoxide surface as a function of pH (from

CORNELL & SCHWERTMANN 1996).

In the case of base metals, the adsorption of Cu is the most pronounced and metal

adsorption decreases in the following order: Cu > Pb > Zn > Cd > Co > Ni > Mn

(CORNELL & SCHWERTMANN 1996). Cation adsorption is usually rapid while desorption is

rapid only for Pb. For Cu, Zn, Cd, Ni, and Co desorption is slower with a hysteresis effect

(CORNELL & SCHWERTMANN 1996).


165

Fig. 8.4: Effect of pH on the adsorption of selected metal cations on goethite and hematite

(from CORNELL & SCHWERTMANN 1996 and authors therein).

In acidic systems the surface charge of goethite and hematite at low pH (< 5) is positive

with FeOH2+-species dominating (Fig. 8.3; CORNELL & SCHWERTMANN 1996). Cation

adsorption thus decreases with decreasing pH and only minor amounts of negatively

charged metal complexes can be adsorbed below pH 4 (Fig. 8.4; CORNELL &

SCHWERTMANN 1996; ALI & DZOMBAK 1996). Concurrently anion adsorption increases,

which has important implications for the adsorption behaviour of positive charged

complexes, e.g. arsenates.

These observations suggest that in acidic systems, like the decay of sulfidic ore or acid

leaching of ore, adsorption of metals like Cu and Zn should be negligible. However, the

fraction of Cu and Zn bound exchangeable in the supergene ore of Sanyati is relatively

small, but significant and the total amounts of metals adsorbed are in disagreement with

the above described findings. Therefore, other factors are expected to strongly influence

the adsorption of metals onto ironoxyhydroxides at low pH.

In the presence of anionic ligands, metal uptake may be enhanced or inhibited via

processes like the formation of ternary complexes, site competition, the formation of

solution complexes, and the alteration of surface charges (BENJAMIN & LECKIE 1981,

1982). Trace metal adsorption on goethite is increased by the presence of sulfate ions. This

process leads, e.g., to an increased adsorption of Cu on goethite in seawater (BALISTRIERI

& MURRAY 1982).


166

Fig. 8.5: Fraction of surface sites of goethite occupied by ions in seawater (from

BALISTRIERI & MURRAY 1982). Sulfate adsorption increases significantly with

decreasing pH.

The increasing sulfate density on the goethite surface as a function of decreasing pH is

shown in Fig. 8.5. The increased sulfate density influences electrostatic effects, site

availability, and solution complex formation (BALISTRIERI & MURRAY 1982), which in turn

influences metal adsorption.

ALI & DZOMBAK (1996) investigated the sorption of Cu in the presence of sulfate at

relatively low sulfate concentrations (0.25 mM and 1 mM) and also found enhanced

adsorption of Cu at low pH. They concluded that the enhanced adsorption of Cu at low pH

is mainly due to the formation of the ternary complex ≡FeOHCuSO40 (modelled with the

Generalized Two Layer Model of DZOMBAK & MOREL 1990), whereas in a sulfate free

system the positively charged ≡FeOCu+ is the dominant complex (ALI & DZOMBAK 1996).

JUANG & WU (2002) report that Cu adsorption is enhanced by the presence of SO42-, while

in turn the adsorption of SO42- is reduced by the presence of Cu2+.

From the experimental work published the adsorbed Cu fraction is calculated from the

decrease of the Cu concentration of the solution. The amount of adsorbed Cu can be

calculated as:

q=(Co-Ce)V/W [mmol/kg] (8.1)

where Co is the initial concentration in the solution, Ce is the concentration at the end of the

experiment (assuming equilibrium at this stage), V is the solution volume [dm3] and W is

the weight of the dry goethite [kg] (e.g. JUANG & WU 2002).


167

In Table 8.1 the experimental conditions of representative experiments are compiled and

the amount of adsorbed Cu qCu is calculated. In sulfuric acid systems sulfate concentrations

are estimated from pH assuming complete dissociation.

Table 8.1: Comparison of the observed Cu adsorption to goethite during experiments in

different sulfate-bearing systems. Data are from JUANG & WU (2002), BALISTRIERI

& MURRAY (1982), ALI & DZOMBAK (1996), and this study.

Cu source pH initial Cu C0

[mmol/l] end Cu Ce [mmol/l]

initial SO42- C0

[mmol/l] goethite W

[kg] qCu

[mmol/kg] JUANG & WU (2002) Cu sulfate 2.5 52 - 52 0.00005 ~ 10 BALISTRIERI & MURRAY 1982 sea water 3 3.10E-04 2.95E-04 27 0.0005 0.03 ALI & DZOMBAK (1996) Cu nitrate 3.25 0.098 0.093 0.25 0.0016 3.06 this study malachite 2.36 1.44 0.75 2.18 0.01 68.14 this study Cu sulfate 2.14 1.61 0.68 4.68* 0.01 92.21 this study chrysocolla 2.22 0.97 0.97 3.01 0.01 0.00 * sulfate adds up from Cu mineral and sulfuric acid

From the exchangeable fraction determined in the extraction experiments (V12 - V14) an

adsorption of ~ 7.5 mmol Cu/kg goethite was calculated (cf Fig. 7.8). Considering the

simplifications made for the mass balance calculations, this finding is in very good

agreement with the 3.06 mmol Cu/kg goethite experimentally determined by ALI &

DZOMBAK (1996) under acidic, sulfate-rich conditions (Table 8.1).

Compared to the literature data the observed amount of adsorbed Cu in the adsorption

experiments with Cu minerals onto goethite (V8-V10) is one order of magnitude higher in

the malachite and Cu sulfate systems examined in this study (Table 8.1). Variations in the

surface area of the goethite used as starting material and the presence of other anions might

account for minor differences between the experimental results, but can not explain the

observed more than one order of magnitude higher Cu adsorption found in this study.

In addition to sulfate, H3SiO4- is usually present in natural solutions and might have

considerable influence on Cu adsorption by changing the goethite surface charge and the

Cu-complexation in the solution. However, up to date only few studies address to Cu


168

adsorption in silica-bearing systems, because Si-polymerisation complicates the data

interpretation. HINGSTON et al. (1967) found an adsorption maximum of ~ 3 µmol H3SiO4-

/m2 on goethite at pH ≈ pK1 (~9) for silicic acid and a slow decrease with sinking pH

(CORNELL & SCHWERTMANN 1996). HERBERT (1999) proposed a change of the surface

charge of goethite by condensation of silica at the goethite surface, allowing metal

adsorption (e.g. adsorption of Al3+): The positive surface charge of goethite under acidic

conditions promotes adsorption of aqueous anionic species like SO42- and H3SiO4

-

(LÖVGREN et al. 1990). The goethite surface may act as a catalyst initiating condensation

of silica polymers and subsequent silica precipitation (WILLIAMS et al. 1985; OHMARI &

MATIJEVIC 1992). Aluminium may substitute for Fe3+ in the goethite lattice (SCHULZE

1984) but adsorption of minor amounts of Al3+ also occurs (LÖVGREN et al. 1990).

Dissolved aluminium and silica can combine to form alumosilicate coatings (ILER 1973).

Aluminium and silica both inhibit the transformation of low crystalline ironoxyhydroxides

to goethite and the formation of goethite from solution (e.g. ZHAO et al. 1994; CORNELL &

SCHWERTMANN 1996). HERBERT (1999) proposed a model for the formation of

ironoxyhydroxides from solutions in which the Al(Si)/Fe concentrations of the solution

controls the compositional zonation. High Al(Si)/Fe ratios promote adsorption and

precipitation of silica, which inhibits the transformation to and neoformation of goethite. In

periods of low Al and Si concentrations in the solution goethite growth increases again.

The principal reactions that lead to the ironoxyhydroxide formation according to the model

by HERBERT (1999) are shown in Fig. 8.6.

Fig. 8.6: Reaction scheme for the condensation of silica polymers on an ironoxyhydroxide

surface and subsequent adsorption of Al3+ on the silica precipitate (from

HERBERT 1999).

This might be a reasonable mechanism also for other metals, e.g. Cu. However, direct

evidence for an ironoxyhydroxide formation according to this model in Sanyati is only

found in LPO (Fig. 6.22 and 6.23). It is likely that this process also takes place in the

unleached samples, but there it is superimposed by the adsorption signatures.


169

The goethite- and hematite-rich zones from Sanyati examined by EMPA show generally ~

1 wt % Si with a maximum value of 12 wt%. However, in the adsorption experiment with

chrysocolla Cu is not adsorbed to the goethite present. In this experiment Cu is presumably

stabilized in the solution by silica-complex formation (in contrast to the experiments with

silicate free systems).

The experiments in this study thus indicate that at strongly acidic conditions (i.e. at a pH =

2) in a sulfuric system Cu is adsorbed to the goethite surface as long as no further

complexant (e.g. silica gel) is present that stabilizes Cu in solution via complexation. At

low pH sulfate (and possibly silicate) induce a surface charge switch (SCS) of the

otherwise positively charged surface of goethite and enables cation adsorption via the

formation of ternary complexes. Possibly this mechanism allows some metal adsorption to

the ironoxide and -oxyhydroxide (i.e. hematite and goethite) surface and can be interpreted

as a first step of metal fixation. However, the overall observed magnitude of Cu removal

from the solution can not be attributed to this adsorption mechanism alone. As depicted in

Fig. 7.10 dissolution of some goethite (rise of Fe in the solution) can be observed during

the initial stage of the experiments. In the further course of the experiment a part of the

iron initially liberated is removed again from the solution, i.e. is re-precipitated. Therefore,

it is highly likely that a considerable fraction of Cu is coprecipitated with goethite from the

solution.

Coprecipitation of metals is well known and is discussed in more detail in Chapter 8.3.2. A

surface precipitation model by FARLEY et al. (1985) predicts that at low sorbate/sorbent

ratios, surface complexion dominates. Increasing surface loading of, e.g., Me2+ onto HFO

(nominal formula Fe(OH)3) leads to precipitation on the surface in the form of a solid

solution Fe(OH)3-Me(OH)2 until all surface sites have become saturated (ZHU 2002). This

would explain the excess amount of Cu removed from the solution and is another

indication that other processes beside adsorption control the base metal fixation.

In natural samples from acid mine drainage consisting of poorly ordered goethite and

quartz as main constituents, WEBSTER et al. (1998) report SO4 concentrations of 5 - 11

wt% and trace metal concentrations of 0.026 wt% Cu, 0.064 wt% Pb, 0.015 wt% As, and

0.034 wt% Zn. These metal concentrations could not be mobilized by HNO3 attack (pH 3.0

- 4.0 for 24 h) and are reported by WEBSTER et al. 1998 to have become incorporated into

the goethite structure. The sulfate concentrations in goethite and hematite of the Sanyati


170

ores are ~ 2 wt%.

8.3.2 Lattice incorporation of base metals in goethite and hematite

The major part of Cu and Zn is only mobilized from the iron phases when goethite or

hematite is dissolved, as could be shown in the extraction experiments. This observation

points to an incorporation of Cu and Zn into the crystal lattice of ironoxides and -

oxyhydroxides by substitution for Fe3+, a mechanism by which the metals are bound

relatively strong.

The degree of possible isomorphic substitution of the base metals is already discussed in

Chap. 6.1. The high degree of base metal substitution in the Sanyati goethites leads to

systematical changes in the unit cell (i.e. bo and vo) plotting along an array typical for Cu

and Zn-substituted goethites (Fig. 6.2).

8.3.3 Jarosites/plumbojarosite

In the oxidation zone of Sanyati no depletion of jarosite and plumbojarosite is found in the

oxidation zone. The base metal substitution of the plumbojarosites in the supergene ore

reaches up to 1.14 wt% for Cu and 0.4 wt% for Zn, which is below the maximal

substitution for both metals (JAMBOR & DUTRIZAC 1985; SCOTT 1987). Therefore these

phases are a significant scavenger only for Cu. However, in acidic systems, e.g. AMD,

plumbojarosite and argentojarosite often play an important role in limiting the Pb and Ag

mobility (HOCHELLA et al. 1999). Plumbojarosite is known to occur as a submicroscopic

phase in limonites as pointed out by HOCHELLA et al. (2002). Therefore plumbojarosite has

to be considered as a sink for Pb and to a lower extent Cu, respectively, in the Sanyati

supergene ores.

To summarize, in the Sanyati ores only a relatively small amount of Cu (19%) and Zn

(6%) is liberated from the exchangeable fraction in extraction experiments. This fraction

can be attributed to adsorption of Cu to goethite and hematite surfaces which are

negatively charged due to sulfate (and possibly silicate) loading.

Extraction experiments reveal that the major fraction of Cu and Zn is only mobilized when

goethite and hematite are dissolved. This observation is consitent with results from unit

cell calculations that demonstrate a high degree of Cu and Zn substitution into the crystal


171

lattice of the investigated goethites. However, coprecipitation with reprecipitating iron

phases is also shown to play an important role in element redistribution as revealed by the

experiments with synthetic goethite. Only a minor fraction of Cu and Zn is bound to

jarosites and plumbojarosites.

8.4 Consequences of the base metal retention on the extraction success of heap

leaching

For the economic profitability of a mine the information on the extraction recovery of the

desired metals from the ore is equally important to the ore grade and the amount of proven

and expected reserves. The recovery from the heap leaching in the Sanyati mine is ~ 20 –

30 % below expectations.

The prime source of metal values in the supergene ores in Sanyati are base metal phases, of

which the most important are chrysocolla and malachite. Beside these, other minerals, like

arsenates, are frequent and add to the release of Cu in the heap leaching process. These

main Cu-carriers are generally readily soluble (e.g. malachite with a solubility constant of

Ks = -33.8 in aqueous solution; NICKEL & DANIELS 1986). It was observed though that

malachite and the slightly slower soluble chrysocolla are often strongly intergrown, so that

Cu liberation from the two phases is probably controlled by chrysocolla dissolution.

However, this does not generally prevent mobilization when the ore is treated with highly

acidic solutions on the heap leach pad. The percolation experiments of LPO show that

leaching of base metals is only restricted for the previously extracted Cu. In contrast,

previously not extracted metals, e.g. Zn, are readily liberated during percolation

experiments from individual metal-bearing phases from the supergene ore, metals adsorbed

to mineral surfaces, and from efflorescences.

The most important mineral phases responsible for the observed metal retention are

hematite and goethite. A comparison of the mineralogical composition of ROM and LPO

shows that goethite and jarosite are more frequent in LPO, while hematite occurs more

often in ROM. This indicates a conversion and neo-formation of these phases during

leaching. Microprobe analysis revealed that the amount of metals retained in hematite and

goethite in LPO is in the range observed for the same phases in ROM, even after a

residence time of several years on the heap leach pad during which the ore has been


172

leached continuously. This demonstrates that significant amounts of base metals are fixed

in hematite and goethite and are not accessible to leaching.

Extraction experiments show that 65 and 85 % of Cu and Zn, respectively, are bound in the

goethite and hematite lattice. Only 19 and 6 % of Cu and Zn, respectively are adsorbed on

hematite and goethite surfaces and are thus belonging to the easily leachable fraction.

Adsorption experiments in synthetic systems carried out at conditions close to those during

leaching on the heap leach pad reveal that initial dissolution and subsequent reprecipitation

of hematite and goethite occurs during leaching. This dissolution – reprecipitation is

accompanied with the coprecipitation of large amounts of Cu, and most likely also Zn,

with hematite and goethite. These findings suggest that repeated cycles of dissolution and

reprecipitation - coprecipitation of base metals is responsible for the fixation of the bulk of

the “invisible” base metals into the crystal lattice of hematite and goethite. The textural and

microchemical evidence, especially from colloform textures in LPO, also points to a cyclic

dissolution - reprecipitation process as the most likely mechanism for the fixation of base

metals in hematite and goethite. On the heap leach pad unspecified precipitations (possibly

hematite and goethite?) on a macroscopic scale are reported by MAGOMBEZE & SANDVIK

(2002).

The metal values refixed in this fashion as well as those never mobilized from the ROM

are lost for recovery, because high iron contents in the solution are not wanted in the SX-

EW process and complete dissolution of the iron phases is therefore prevented by keeping

the pH of the solution used for heap leaching in the range ~1.5 - 2.


173

Fig. 8.7: Schematic sketch of the different types of primary and secondary sinks for metals

in the heap leach pad of the Sanyati mine.

The different sinks of metals in the circulation system of the heap leach pad are presented

schematically in Fig. 8.7. The conditions on the heap leach pad favour the additional

formation of plumbojarosite, a potentially important sink for Pb and base metals. However,

the modal abundances of plumbojarosites and jarosites are low and their base metal

contents are with max. 1.14 wt% Cu and 0.4 wt% Zn significantly below the saturation

limit (see Chapter 4.2.3.2). Hence, their contribution to the observed base metal retention is

limited, but is important for Pb.

The acid used in circulation is a temporary sink for e.g. Zn (mg/l), Mn (mg/l), and Co

(mg/l). Additionally, some potential metal values are precipitated in the efflorescences that

are found on the dry parts of the heap leach pad. However, these sinks are only temporary

as metals are easily mobilized from them.

A semi-quantitative mass-balance was performed for ROM and LPO based on the average

Cu and Fe contents in ROM and LPO (cf. Fig. 5.2). Because exact modal mineral

distributions are not available, it was estimated that the ironoxides and -oxyhydroxides are

dominated by goethite and hematite and are equidistributed. The Cu content fixed by

goethite and hematite was calculated using the microanalytical data of the ironoxide and -

oxyhydroxide textures.


174

Fig. 8.8: Semi-quantitative estimation of the distribution of Cu in ROM and LPO based on

the average composition of ROM and LPO (cf. Fig. 5.2), average contents of Cu

in goethite and hematite of the supergene ore (as determined by microprobe

analysis), and assuming that goethite and hematite are equidistributed in both

ore types.

The distribution of Cu in ROM shows clearly that a significant amount of Cu is bound in

the ironoxides and -oxyhydroxides (Fig. 8.8). This fraction has to be regarded as

dominantly unleachable. As shown by extraction experiments, only a minor fraction is

bound exchangeable.

In the LPO almost all leachable Cu from the base metal minerals is liberated and extracted

by the acid. The remaining Cu is dominantly fixed in the ironoxides and -oxyhydroxides.

These findings are in excellent agreement with the results from the extraction experiments

performed.

To conclude, base metals in LPO are fixed in temporary (e.g. efflorescences, leaching acid)

and permanent sinks (e.g. ore centres never leached, metals bound in goethite and hematite

lattices, and re/coprecipitated metals bound to iron oxides and -oxyhydroxides). From the

latter significant liberation is prevented by pH control (~1.5 - 2) of acid used for leaching.

A semiquantitative mass-balance shows that most of the leachable Cu is extracted from

LPO.

0

0.2

0.4

0.6

0.8

1

1.2

1.4

ROM LPO

[wt%

]

base mineral-bound

goethite-bound

hematite-bound

Cu

9. Conclusions and general perspective

175

9 Conclusions and general perspective

Supergene ores in Sanyati: geochemical and mineralogical characteristics

The mineralogical and geochemical investigations reveal that base metal carriers are very

heterogeneously distributed in ores of the oxidation zone in all open pits and in

mineralisations potentially to be mined in the future. This is partly due to the

heterogeneous distribution of the primary ore, but also due to the variable access of

meteoric water to the ore, because of the steep dipping banks of host rock and primary ore

lenses.

The maturity of the oxidation zone is very low. With the exception of the instable

precursor phases of goethite (e.g. ferrihydrite, schwertmannite) all maturity indicators

examined (Table 4.10) point to immature characteristics. Additional evidence for low

maturity is provided by an only rudimentarily (as ore lenses) developed supergene

enrichment zone. These observations suggest that the oxidation zone is unzoned.

Therefore, the separation of metal depleted zones is not needed and consequently the

mining strategy is relatively straightforward.

The most important Cu-carriers in the supergene ore are the base metal minerals malachite

and chrysocolla. Additionally, a significant amount of Cu is bound to numerous arsenates,

sulfates, and carbonates, amongst other phases (Table 4.6). It is common knowledge that

the ubiquitous limonites of supergene ores carry significant amounts of base metals.

However, only few studies report microanalytical data for these phases (e.g. SCOTT 1986,

1992, HERBERT 1999). The detailed microanalytical investigations carried out in this study

reveal that the base metal contents in goethite and hematite in Sanyati ores are significantly

higher compared to the available data from other oxidation zones (e.g. SCOTT 1992, 1986,

Chap. 6.1). Their average contents range from 1.0 wt% Cu and 1.2 wt% Zn in the

hematite-rich zones to 1.5 wt% Cu and 2.3 wt% Zn in the goethite-rich zones of the

sections examined.

Extraction experiments revealed that the major part of the base metals (65 % Cu and 85 %

Zn) in goethite and hematite is fixed in the crystal lattice. Only a minor part (19 %) of the

base metals retained by goethite and hematite is adsorbed. The absolute amount of Cu


176

adsorbed to goethite estimated from extraction experiments is in very good agreement with

the experimentally determined adsorbility of Cu in acidic, sulfate-rich systems (ALI &

DZOMBAK 1996). Adsorption experiments of Cu onto goethite in acidic, sulfate-rich

systems revealed, that under these conditions beside adsorption, coprecipitation plays a

significant role in scavenging Cu and presumably other base metals like Zn.

Microprobe analyses showed that the colloform textures in ROM (and also LPO) reflect

the limonite chemistry of the surrounding on the dm3 scale satisfyingly, and can serve as a

proxy for the limonite bound fraction of metals on this scale. The base metal contents of

colloform textures might be used for the estimation of the quality of the ore from the

different open pits because they yield information about the unleachable base metal

fraction.

Recommendations for an improved recovery in Sanyati

Investigations on the run-of-mine ore showed that the grain size spectrum applied onto the

heap leach pad is far from being ideal. Coarse blocks are leached insufficiently even after

many years of leaching and the fine fractions < 63 µm prevent fast acid entry into the heap

leach pad and clot up the access to the coarser, still unleached blocks. As a consequence

the surface of the heap leach pad is ploughed by heavy machines occasionally. The use of

heavy machinery for both initial emplacement of the ore on the heap leach pad (instead of

using conveyor belts) and ploughing is highly likely to cause considerable compaction at

lower levels. Therefore it is reasonable to subject the ore to crushing prior to leaching and

redesign the built-up of the heap leach pad (e.g. by using conveyor belts instead of dump

trucks) in order to enhance the permeability of the heap leach pad.

The metal contents of the fine fraction of ROM (< 63 µm) are not very high (~ 0.65 wt%

Cu and ~ 0.7 wt% Zn), but this fraction has a high liberation rate. It might be economic to

treat this fraction separately using an agitated leaching process. This would also resolve

their negative contribution to the permeability.

MAGOMBEZE & SANDVIK (2002) state that cuprite is the major Cu-oxide (together with

azurite, malachite, tenorite, chryoscolla, and dioptas). They therefore propose the addition

of Fe2+ to enhance the degradation of cuprite and the liberation of Cu. The findings of this

study, however, suggest that cuprite is present, but that modal abundances are insignificant


177

for the overall Cu-budget of the ore. Instead, chrysocolla and malachite have been found to

be the most important minerals for the Cu-budget of the supergene ore. Therefore, addition

of Fe2+ is not expected to have a positive effect on the Cu-recovery.

During the leaching process base metals are mobilized mainly from the base metal

minerals (see above), and some liberation is also expected from the limonite phases.

However, as shown in the percolation and adsorption experiments, these phases are only

partly leached and coprecipitate parts of the base metal load of the solution they are in

contact with during reprecipitation. This fraction of base metals is lost for the further

extraction process, since high Fe contents in the acid are not wanted in the SX-EW plant.

Reprecipitation on a macroscopic scale is indicated by precipitations at a depth of 6 - 8 m

of the heap leach pad reported by MAGOMBEZE & SANDVIK (2002). However, these authors

do not report a mineralogical characterisation of these precipitations. On a micro-scale

reprecipitation is indicated by additional colloform textures in the LPO that are neoformed

on the heap leach pad.

Likewise, these neoformed colloform textures can be used as an indicator for the limonite

chemistry. Geochemical evaluation of these neoformed colloform textures could help to

understand the precipitation processes in the heap leach pad. An improved understanding

of these processes provides useful information for controlling the pH conditions in an

attempt to minimize the Cu-loss due to re- and coprecipitation processes.

To summarize, in order to improve the unsatisfying Cu recovery of the Sanyati mine (if

wanted) it is proposed to investigate the effects of grain size distributions on base metal

leachability from the ores more systematically. This might help to find an optimal grain

size distribution for the heap leach pad. Additionally, the fine fraction should be

investigated with regard to the actual metal liberation rate and the possibility of separation

and agitated leaching. These investigations will have to include economic constraints,

which where beyond the scope of this study. Further investigation of the chemistry of the

colloform textures in the ROM can help to estimate the quality of unworked orebodies, and

in the LPO may be helpful to monitor and optimise the pH distribution on the heap leach

pad.


178

Colloform textures as proxies for "invisible" base metal content in supergene ores and

metal retention during heap leaching

The "invisible" and to a large extent immobile base metal contents of supergene ores can

influence the ore quality and might have important implications for the economy of a mine

operation. Knowledge of the distribution and quantities of these "invisible" base metal

contents is therefore of crucial importance and may determine the ultimate recovery from

supergene ores in HL/SX/EW processes.

The results of this case study suggest that microanalytical data for ironoxides and -

oxyhydroxides have the potential to be a powerful tool for the characterisation of

supergene ores. Detailed microchemical information for these minerals can potentially help

to classify the ore quality during exploration, characterize the ore quality during mining,

and support the control and optimisation of heap leaching processes by indicating the

quantities of metals retained in the ore during the leaching phase. Colloform textures are

especially suitable because they provide information on average metal contents. These

textures can be detected easily in reflected and transmission polarized microscopy.

Suggestions for further work

The results of this work suggest some avenues for future investigations. For example, the

general database describing the variations of major, minor, and trace elements in

ironoxides and -oxyhydroxides is limited and there are only a few studies available that

report microchemical data (i.e. EMPA and LA-ICP-MS) for these important rock-forming

minerals. Hence, there is a need for more high-quality compositional data for these phases.

These investigations should also comprise additional phases, such as lepidocrocite.

Combined TEM and AFM studies of experimental products might gain important insights

into the growth processes of goethite and the mechanism responsible for the

incorporation/occlusion of base metals on an atomistic scale. These studies should also

investigate the transformation mechanisms of ferrihydrite and schwertmannite to goethite

and its bearing on base metal incorporation.

In the Sanyati ores the colloform textures have proved to be useful proxies for the average

base metal contents in limonite precipitations. Therefore, it should be evaluated if this

approach is applicable to supergene ore deposits in general. A combination of this textural


179

and microanalytical approach with laboratory leaching experiments is especially

promising.

Future experimental work should systematically investigate base metal adsorption and

coprecipitation in acidic systems. These experiments should consider the influence of

additives like sulfate or silica with respect to ternary complex formation on surfaces. They

should also include the precursor phases of goethite (e.g. ferrihydrite and schwertmannite).

The zero-point-of-charge might prove as a simple tool to further characterise the mineral

surfaces. Last, but not the least, there is a need for an extensive set of solubility data of

metals in highly acidic, sulfate-rich solutions with a high load of total dissolved solids.

10. References

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WEBSTER, J., LANE, V., HOWARTH, R., SWEDLUND, P. & SAUL, D. (2000): Factors

influencing the precipitation of sulphate-rich iron oxides, and their ability to absorb

trace metals. - Goldschmidt 2000, 3-8. Sept., J. of Conference Abstracts, 5, 2: 1073.

WEBSTER, J., SWEDLUND, P. & WEBSTER, K. (1998): Trace metal adsorption onto an acid

mine drainage iron(III)oxyhydroxy sulphate. - Environ. Sci. Techn., 32: 1361-1368.

WEDEPOHL, K.H. (1978): Handbook of Geochemistry. - Heidelberg (Springer Verlag).

WELLMER, F.-W. (2002): Leachable supergene base and precious metal deposits

worldwide - an overview. – Erzmetall, 55: 25-34.

WILLIAMS, L.A., PARKS, G.A., CRERAR D.A. (1985): Silica diagenesis. I. Solubility

controls. - J. Sed. Petr., 55: 301-311.

WILMSHURST, J.R. & FISCHER, N. (1982): Classification scheme for gossans. - In: SMITH,

R.E. (ed.): Geochemical exploration in deeply weathered terrains: 70-72; Wembley,

W.A. (CSIRO).

WILSON, M.J. (1970): A study of weathering in a soil derived from a biotite-hornblende

10. References

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rock: I. Weathering of biotite. - Clay Minerals, 8: 291-303.

WHITE, A.F. & YEE, A. (1985): Aqueous oxidation-reduction kinetics associated with

coupled electron-cation transfer from iron-containing silicates at 25°C. - Geochim.


WHITTON, D.G.A. & BROOKS, J.R.V. (1981): The Penguin Dictionary of Geology. - 520

p.; Aylesbury Buck (Viking press).

WRONKIEWICZ, D.J. & CONDIE, K.C. (1987): Geochemistry of archean shales from the

Witwatersrand supergroup, South Africa: Source-area weathering and provenance.-

Geochim. Cosmochim. Acta, 51: 2401-2416.

XIE, J. & DUNLOP, A.C. (1998): Dissolution rates of metals on Fe oxides: implications for

sampling ferrugenious materials with significant relict Fe oxides. - J. Geochem.

Expl., 61: 213-232.

YATES, D.M., JOYCE, K.J. & HEANEY, P.J. (1998): Complexation of copper with polymeric

silica in aqueous solution. - Appl. Geochem., 13: 235-241.

ZHAO, J., FENG, Z., HUGGINS, F.E. & HUFFMAN, G.P. (1994): Binary iron oxide catalysts

for direct coal liquefaction. - Energy and Fuels, 8: 38-43.

ZHU, C. (2002): Estimation of surface precipitation constants for sorption of divalent

metals onto hydrous ferric oxide and calcite. - Chem. Geol., 188: 23-32.

11 List of figures and tables

199

11 Lists of figures and tables

List of figures

page

Fig. 1.1: Sketch of the development of a supergene ore deposit (from POHL 1992). 11

Fig. 1.2: Copper production from 1800-2000 (from FALVEY 2003). 13

Fig. 1.3: a. Relative market share of the currently most important Cu producing companies, and b. relative share of extraction routes on the global Cu production (from AME 2004).

16

Fig. 1.4: Share and value of the global by-prodution from Cu mining relative to the global primary metal production in 1998 (from GRASSMANN 2003).

17

Fig. 1.5: SX-EW contribution to the mine production 1980-2001 (from WALLIS & CHLUMSKY 1999). 17

Fig. 1.6: Proven and probable reserves of supergene copper ores [in 1000 t of Cu] (without "Eastern Countries", DR Congo, and Zambia), which contribute 47.5* 106 t Cu, i.e. 11.5 % of the worlds total (from WELLMER 2002).

18

Fig. 1.7: General flow-sheet for the extraction of Cu by a heap leaching - solvent extraction - electrowinning process (from WELLMER 2002).

21

Fig. 2.1: Topographical map of the Sanyati region (from Zimbabwe Sheet SE-35-8 Copper Queen, ed. 3, published by the Surveyor-General, Zimbabwe 1977).

24

Fig. 2.2: Simplified geological map of Zimbabwe (after TRELOAR 1988) showing the Archean craton and the surrounding mobile belts. The Magondi Belt is positioned to the NW of the craton. D = Dett Inlier, SAN = Sanyati mine.

25

Fig. 2.3: Simplified geological map of the Magondi Belt (from TRELOAR 1988). The Sanyati mine (SAN) is located in the SW of the thrust belt close to the Copper Queen Dome (CQD).

26

Fig. 2.4: Simplified geological map of the Sanyati area (from NEWHAM 1986) with the position of the orebodies (termed cupriferous gossans).

28

Fig. 2.5: Morphological provinces of Zimbabwe (from LISTER 1987). 30

Fig. 2.6: Palaeomorphological map of NW Zimbabwe (from LISTER 1987). 31

Fig. 2.7: View to the southern wall of Copper-Queen open pit in the Sanyati deposit. The orebodies are deformed and exhibited to the surface in steep dipping folds. At the base sulfide ore is exposed.

32

Fig. 2.8: Map of the supergene ore deposits at Sanyati. The seven underlined orebodies are mined. Detail on the basis of the map: Sanyati Joined Venture: "Ore zone location and schematic plant layout."- Reunion Mining (Zimbabwe) Ltd.

36

Fig. 2.9: Flow-sheet of the HL-SX-EW plant at the Sanyati mine. 37

Fig. 2.10: a. Overview of the heap leach pad; b. Scetch map of the heap leach pad with trenches for profile sampling; c. Older trickle system for acid dispersion on the heap leach pad; d. New sprinkler system for acid dispersion on the heap leach pad.

38

Fig. 3.1: a. Profile in F-Body-S open pit (P-FBS) (base of picture = 2 m) b. Chrysocolla-rich ore lens in F-Body-N open pit (base of picture = 5 m).

40

Fig. 3.2: a. Run-of-mine ore dump north of Copper Queen open pit. b. Separated big ore blocs (Ø 1 m). 41

Fig. 3.3: Photograph and sketch of profile A2-4 taken on the heap leaching pad. (Grain size distribution: FG < 3 cm, MG 3 - 15 cm, CG > 15 cm).

43

Fig. 3.4: a. Experimental set-up with filters for acid percolation experiments (VR, V15); b. Set-up for the experiments performed in closed plexiglas vessels (V7). c. Close-up of the sample vessel.

47

Fig. 3.5: Ablation behaviour for 69Ga and 72Ge during analysis of NIST 610 and NIST 612 standard glasses and an goethite unknown. For 69Ga manual evaluation was necessary because of the fractionated signal at the beginning of ablation. The high background count rates for 72Ge leads to high detection limits and high analytical errors for concentrations close to the detection limit.

54


200

List of figures continued:

Fig. 4.1: Composition of weathered host rocks (phyllite and amphibolite) of the Sanyati in the ternary A-CN-K diagram of Nesbitt & Young (1984; 1989) a.) In weathered distal phyllites only the centres of the “core stones” (blue filled squares) and a phyllite from Copper Queen Dome (blue dot) are in the initial weathering stage. Most samples from the rim and matrix around the “core stones” are weathered along the “granite trend” (blue arrow) and have reached the illitic stability field. b.) In acidic altered proximal phyllites and amphibolites nearly all phyllite relicts (blue squares) have reached the kaolinitic weathering stage along the “granite trend” (blue arrows). The acid influenced alteration of the amphibolites represents the mafic weathering trend (red arrow) often up to the chloritic stage. (The advanced weathering trend follows the dashed arrow. Black dots represent ideal composition of minerals).

59

Fig. 4.2: Normalized element distributions of the "core stone" phyllite samples from Copper Queen open pit (CQ) and the surrounding matrix (according to WRONKIEWICZ & CONDIE 1987). The samples are shown with increasing degree of weathering from the "core stone" to the matrix. Concentrations are normalized to the upper continentral crust (TAYLOR & MCLENNAN 1985).

60

Fig. 4.3: Chemical index of alteration (CIA) vs acid neutralisation capacity of phyllites and amphibolites distal and proximal to the zone of sulfide decay based on alkali and alkali earth cations only (ANCA). For comparison, the black line marks the CIA of kaolinite, the blue line the lowest CIA for fresh basalts and the red line the lowest CIA for granites.

61

Fig. 4.4: Chemical index of alteration (CIA) vs acid neutralisation capacity of phyllites and amphibolites distal and proximal to the zone of sulfide decay based on alkali and alkali earth cations and including Al, Fe, and Mn as cations as well as S, and P as anions (ANCB). For comparison, the black line marks the CIA of kaolinite, the blue line the lowest CIA for fresh basalts and the red line the lowest CIA for granites. The red triangle highlights the samples with high ANCB and CIA plotted in Fig. 4.5.

63

Fig. 4.5: Correlation of ANCB and Al, Fe, and Mn in intensely altered samples (indicated with red triangle in Fig. 4.3).

64

Fig. 4.6: Polished section images of primary and secondary sulfide ores from the Copper Queen, F-Body-S, J-Lines, and J-Body open pits (microphotograph width: 400 µm, all taken with parallel nicols, except c. taken with crossed nicols).

70 - 71

Fig. 4.7: SEM photomicrographs of iron-rich supergene ore (sample 230b). a. & b.: dominantly hummockey surfaces, c.: flat surface in the center, d.: porous rough surfaces are preparation artefacts.

75

Fig. 4.8: Alteration sequence for oxidation of sulfides and products in the oxidation zone of Sanyati (restricted to the more frequent supergene phases).

79

Fig. 4.9: Metal enrichment factor in supergene ore samples calculated from medians of supergene ore normalised to proximal host rock. Elements are arranged in order of increasing enrichment.

80

Fig. 4.10: Cu vs Fe2O3 in supergene ore proximal and distal to the surface. 80

Fig. 4.11: Frequency distribution diagram of Cu concentrations in supergene ore to exclude outliers. Black: normal element distribution, white: excluded outliers presumably because of Cu mineralisations.

81

Fig. 4.12: Boxplots depicting mean, lower quartile, upper quartile, min. and max. of selected metals Al, As, Cu, Mn, Pb, Ti, and Zn (wt%) and Ag, Bi, Cd, Co, Sb, Se, and V (mg/kg) of samples with metal contents assumed to be limonite bound.

82

Fig. 4.13: Variation of selected metals bound to limonite phases (i.e. the metal fraction that is not easy removable during heap leaching) in the Sanyati open pits. Also given is the mean of all open pits examined.

83

Fig. 4.14: BSE image of a Pb-Mn-mineral, probably coronadite surrounded by goethite-rich precipitates in the oxidation zone sample (212-2, microphotograph width: 380 µm).

84

Fig. 4.15: Co and Cu correlation observed in coronadite. 85

Fig. 4.16: Plumbojarosite in the supergene ore surrounded by goethite-rich precipitates. Transmitted, linear polarized light. Microphotograph width: 1 mm.

85

Fig. 4.17: As, Cu, and Zn contents (wt%) of arsenates from Sanyati compared to their ideal composition. For abbreviations see Table 4.7. A wide range of adamite compositions is observed. The highest observed Co content in cobalto-adamite is 0.21 wt%.

86


201


Fig. 5.1: Grain size distribution of the fraction < 1 cm of run-of-mine ore (ROM sample 210) and leach pad ore (LPO samples profile A2-2) after wet sieving.

91

Fig. 5.2: Average Cu and Zn, and Fe2O3 contents of the supergene ores from the Sanyati mine: “Munyati Mining” denotes data reported by CHADWICK (1996; no data for Fe2O3 available); supergene "Ore from all open pits", ROM and LPO (profile A2-4) are based on own samples.

92

Fig. 5.3: Relative contents of selected metals estimated from supergene ore samples of all Sanyati open pits compared to the mean of samples from the run-of-mine ore dump.

93

Fig. 5.4: Cu, Zn, As, Co, and Pb contents in the grain size fractions of ROM (sample 210). 93

Fig. 5.5: Puddles (inhabited by algea and bacteria colonies) developed on the heap leaching pad because of insufficient acid drainage.

94

Fig. 5.6: Cu and Zn concentrations of the dry and wet sieved fractions of profile A2-2 (grain size in mm). 96

Fig. 5.7: Cu and Zn concentrations in LPO samples from profile A2-2 with regard to their position in the heap leach pad and their grain size.

98

Fig. 5.8: H2SO4 (pH ~ 1.5) is used to leach the Cu from the supergene ore. The acid is recycled after the solvent extraction of Cu and shows high concentrations of Zn and Co, as well as Mn, Al, Mg, and Ca. These elements which are, unlike Cu, not extracted do not show any change in concentration after the passage through the heap leach pad. Elevated Zn, Co, and Mn are of interest, for a potential economic use in the future.

99

Fig. 5.9: a. View on dry areas of heap leach pad. b. SEM picture of soluble sulfates. 100

Fig. 5.10: Relative metal compositions of soluble sulfates precipitated on the heap leach pad, the emergency pond, and the acid used for leaching prior to and after the passage through the heap leach pad.

103

Fig. 5.11: Concentration of Cu, Zn, and Co in the water-soluble fraction of samples of profile A2-4. Samples were washed with nine 1 l aliqouts of fresh, demineralized water and every second aliquot was analysed. After nine washings the water-soluble metal fraction of the leached ore was removed.

104

Fig. 5.12: Depths variation of Cu, Zn, and Co concentrations in the water-soluble fraction of samples in a vertical profile through the heap leach pad (profile A2-2).

104

Fig. 6.1: Structure of α-FeOOH. The double chains for the FeO3(OH)3 octahedra run parallel to the c-axis. (OH groups are indicated while the unmarked corners of the octahedra represent oxygen (from Klein & Hurlbut 1993).

107

Fig. 6.2: Deviation of the unit cell edge length of the b0 axis and the unit cell volume v0 of base metal containing goethites from Sanyati to those from pure goethite (expressed as relative deviation from pure goethite). Also depicted are the deviations observed for endmember compositions with the maximum observed substitutions of Al, Mn, Cu, and Zn for Fe (calculated from values given in Table 6.1).

109

Fig. 6.3: Chemical variation of all investigated goethite and hematite from boxwork textures (as determined by EMPA) and its bearing on the observed BSE-contrast.

112

Fig. 6.4: Photomicrographs of rectangular boxwork in supergene ore (sample 294Z1) a. reflected light microscopy (photomicrograph width: 600 µm). b. BSE image with analysis points and concentrations for Cu and Zn (photomicrograph width: 250 µm).

114

Fig. 6.5: Boxplots (median, lower and upper quartile) of the Fe, Cu, and Zn contents in light (l) and dark (d) zones in rectangular boxworks in sample 294Z1. Abbreviations A and B refer to analysed areas 294Z1-d-B1 and 294Z1-d-A3, respectively (cf. Table D2 in the Appendix D).

114

Fig. 6.6: Photomicrographs of triangular boxwork in supergene ore (LPO sample 294Z1). a. reflected light microscopy (photomicrograph width: 400 µm). b. BSE image of the same area with analysis points and concentrations for Cu and Zn (photomicrograph width: 125 µm).

115

Fig. 6.7: Boxplots (median, lower and upper quartile) of the Fe, Cu, and Zn contents in light (l) and dark (d) zones in triangular boxworks in samples 294Z1 and 4F2. Abbreviations C, D, and E refer to analysed areas 294Z1-d-A1, 4F2-d-A, and 4F2-d-B, respectively (cf. Table D2 in the Appendix D).

116

Fig. 6.8: Photomicrographs of euhedral grain boundary boxworks in supergene ore. a. BSE image with analysis points and concentrations for Cu and Zn (LPO sample 294Z1; photomicrograph width: 400 µm). b. BSE image with analysis points and concentrations for Cu and Zn (LPO sample 294Z1-d; photomicrograph width: 250 µm).

116


202


Fig. 6.9: Boxplots (median, lower and upper quartile, min. and max.) of the Fe, Cu, and Zn contents in light (l) and dark (d) zones in euhedral grain boundary boxworks in samples 294Z1 and 294R1. Abbreviations F, G, and H refer to analysed areas 294Z1-B, 294Z1-B, and 294R1-d-A2, respectively (cf. Table D2 in the Appendix D).

117

Fig. 6.10: Photomicrographs of trellis type boxwork in supergene ore (sample 79) a. reflected light microscopy (photomicrograph width: 2500 µm). b. BSE image with analysis points and concentrations for Cu and Zn (photomicrograph width: 500 µm).

118

Fig. 6.11: Boxplots (median, lower and upper quartile, min and max.) of the Fe, Cu, and Zn contents in trellis type boxworks in sample 79. Abbreviations I and J refer to analysed areas 79-d and 79-e, respectively. Abbreviation l refers to the light zones analysed (cf. Table D2 in the Appendix D).

118

Fig. 6.12: Photomicrographs of parallel closed and open boxworks in supergene ore: a. BSE image of parallel closed boxwork with analysis points and concentrations for Cu and Zn (sample 26-A3; photomicrograph width: 250 µm). b. BSE image of parallel open boxwork with analysis points and concentrations for Cu and Zn (sample 212-2-B-1 photomicrograph width: 400 µm).

119

Fig. 6.13: Boxplots (mean, lower and upper quartile, min. and max.) of the Fe, Cu, and Zn contents in light (l) and dark (d) zones in parallel open and closed boxworks in samples 26, 209c, and 212-2. Abbreviations K, L, M, N, and O refer to analysed areas 26-A3, 209-c-d-B, 209c-d-A, 212-2-d-B1, and 212-2-B1, respectively (cf. Table D2 in the Appendix D).

120

Fig. 6.14: Photomicrograph of cellular sponge boxwork in supergene ore (sample 268a-d): a. reflected light microscopy (photomicrograph width: 800 µm). b. BSE image with analysis points and concentrations for Cu and Zn (photomicrograph width: 250 µm).

121

Fig. 6.15: Boxplots (mean, lower and upper quartile) of the Fe, Cu, and Zn contents in light (l) and dark (d) zones in cellular sponge boxworks in samples 82 and 268a. Abbreviations P, Q, and R refer to analysed areas 82-d-A2, 268a-d-B, and 268a-B, respectively (cf. Table D2 in the Appendix D).

121

Fig. 6.16: Photomicrograph of open rib boxwork in supergene ore (sample 212/2-d-A2). BSE image with analysis points and concentrations for Cu and Zn (photomicrograph width: 500 µm).

122

Fig. 6.17: Boxplots (mean, lower and upper quartile) of the Fe, Cu, and Zn contents in light (l) zones in open rib boxworks in samples 212/2, 212/2-d, and 4F2-d. Abbreviations S, T, U, and V refer to analysed areas 212/2-A5, 212/2-d-A2, 212/2-d-B4, and 4F2-d-C, respectively (cf. Table D2 in the Appendix D).

122

Fig. 6.18: Cubic boxwork in supergene ore (sample 268a-d-C). BSE image with analysis points and concentrations for Cu and Zn (photomicrograph width: 500 µm).

123

Fig. 6.19: Boxplots (mean, lower and upper quartile, min. and max.) of the Fe, Cu, and Zn contents in light (l) and dark (d) zones in cubic boxworks in samples 268a and 72a. Abbreviations W, X, Y, and Z refer to analysed areas 268a-d-C, 268a-A2, 268a-d-A, and 72a, respectively (cf. Table D2 in the Appendix D).

123

Fig. 6.20: Microphotographs of colloform textures in supergene ores: a. BSE image with analysis points and concentrations for Cu and Zn (sample 294Z1-A; photomicrograph width: 150 µm) b. BSE image with analysis points and concentrations for Cu and Zn (sample 212-2-d-A1; photomicrograph width: 500 µm).

124

Fig. 6.21: Boxplots (median, lower and upper quartile, min and max.) of the Fe, Cu, Zn, and As contents in light (l) and dark (d) zones in colloform textures in samples 212-2, 294-Z1, 79, and 268d. Abbreviations 1, 2, 3, 4, 5 and 6 refer to analysed areas 212-2-d-A1, 212-2-A2, 294Z1-A, 294Z1, 79-d-B and 268d-d-A, respectively (cf. Table D2 in the Appendix D).

125

Fig. 6.22: Fe+S vs Si+Al in the goethite- and hematite-rich zones of LPO sample 294Z1 A (=LPO 3 coll) and B (LOP F euhedral).

126

Fig. 6.23: Si vs Al in the goethite- and hematite-rich zones of the LPO sample 294Z1 areas A and B and all other samples measured.

127

Fig. 6.24: Mean Cu and Zn contents in dark reflecting zones of different types of boxworks from both ROM and LPO. The solid line represents the gross average of all determinations.

128

Fig. 6.25: Mean Cu and Zn contents in dark reflecting zones of different types of boxworks from both ROM and LPO. The solid line represents the gross average of all determinations.

128


203


Fig. 6.26 a: Chemical variation (expressed as standard deviations) in light reflecting zones for boxwork types and cases in each section to compare the variabilty for each situation. The solid line represents the overall mean for all boxwork types and sections in ROM and LPO. The variability is always lower in the compilation of sections.

129

Fig. 6.26 b: Chemical variation (expressed as standard deviations) in dark reflecting zones for boxwork types and cases in each section to compare the variabilty for each situation. The solid line represents the overall mean for all boxwork types and sections in ROM and LPO. The variability is always lower in the compilation of sections.

130

Fig. 6.27: Average element concentrations (determined by EMPA) of goethites that are in direct proximity (i.e. in grain contact) to frequent base metal minerals. Abbreviations are: Mal = malachite, Ol = olivenite, Cl = clinochlore, Ad = adamite, and Hem = hemimorphite.

131

Fig. 6.28: Mean Zn concentrations in the light reflecting and dark reflecting zones of the sections of ROM (solid bars) and LPO block 294 (open bars).

133

Fig. 6.29: Cu and Zn in rim and center of the leach pad ore sample 294. 134

Fig. 6.30: Observed distributions of Ga, Ge, Se, Ag, Cd, and Sb in goethite- and hematite-rich zones as determined by LA-ICP-MS. Blue solid diamonds represent dark reflecting zones and orange solid diamonds represent light reflecting zones. Open symbols indicate samples with on average unusually high trace metal contents compared to the other samples (see text for explanation).

135

Fig. 6.31: Comparison of the chemical composition of Sanyati goethites from colloform textures with those precipitated from acid mine drainage (Herbert 1999). The boxplots depict the median, upper and lower quartile, min. and max. value.

136

Fig. 6.32: Comparison of EMPA analyses of goethite and hematite from base metal bearing mineralisations in turbidites (Tu), volcaniclastics (Vo) (Scott 1992), dolomitic shale (Scott 1986), and Sanyati ores (colloform textures). The boxplots depict the median, upper and lower quartile, min. and max. value.

137

Fig. 7.1: Comparison of the observed relative elemental enrichments and depletions in previously unleached ore (ROM = red) and previously leached ore (LPO =black) after leaching in percolation experiments (leached with H2SO4 at a pH of 1.5 recirculated 10 times).

142

Fig. 7.2: Relative fractions of the total Zn released as a function of the relative fractions of the total Cu released at each sampling interval during percolation experiments VR and V15, depicting the relative dissolution rates for these elements. See text for details.

143

Fig. 7.3: Comparison of the observed relative elemental enrichments and depletions in fine grained ROM (sample 210) and in an embedded coarse grained, massive ore block (sample 212/4) after leaching in percolation experiments (leached with H2SO4 at a pH of 1.5 recirculated 10 times).

144

Fig. 7.4: Concentration of Fe, Cu, Zn, As, Mn, and Co in the acid used (H2SO4 at a pH = 1.5) in the permanently stirred leaching experiments V1 and V6 as a function of leaching time.

146

Fig. 7.5: Compositional changes in the residual ore (V1) and three finest fractions sampled (F1, F2, and F3) relative to the starting composition resulting from leaching under idealized conditions.

147

Fig. 7.6: Comparison of relative dissolution rates during idealized leaching (V6): a. Relative fractions of the total Zn released as a function of the relative fractions of the total Cu released at each sampling interval. b. Relative fractions of the total Cu, Zn, and Co as a function of the relative fractions of the total Fe released at each sampling interval.

148

Fig. 7.7: Cu, Zn, and Fe concentrations vs time in the "ideal" leaching experiment V6. 148

Fig. 7.8: Relative fractions of metals and As mobilized from the goethite- and hematite-rich supergene ore sample 263 by the different extraction techniques used.

152

Fig. 7.9: Extraction of Fe, Cu, Zn, Al, and Si in goethite- and hematite-rich supergene ore with a.) NH4Ox-d performed in darkness at room temperature and b.) NH4Ox-80° performed in daylight at 80°C. The data are normalised to concentration at the end of the experiment. Systematic dissolution is only found in NH4Ox extraction in daylight at 80°C.

153

Fig. 7.10: Dissolution of malachite, chrysocolla, and copper sulfate (syn.) in the presence of fine grained goethite (syn.) at pH 1.5.

154

Fig. 8.1: Schematic sketch depicting the weathering and alteration situation of the proximal and distal rocks of the Sanyati region.

157


204


Fig. 8.2: Comparison of the mean Zn concentration in goethite and hematite in a section and their mean in the corresponding colloform textures.

161

Fig. 8.3: Schematic representation of the distribution of positive, negative and neutral surface hydroxyl groups on an ironoxide surface as a function of pH (from CORNELL & SCHWERTMANN 1996).

164

Fig. 8.4: Effect of pH on the adsorption of selected metal cations on goethite and hematite (from CORNELL & SCHWERTMANN 1996 and authors therein).

165

Fig. 8.5: Fraction of surface sites of goethite occupied by ions in seawater (from BALISTRIERI & MURRAY 1982). Sulfate adsorption increases significantly with decreasing pH.

166

Fig. 8.6: Reaction scheme for the condensation of silica polymers on an ironoxyhydroxide surface and subsequent adsorption of Al3+ on the silica precipitate (from HERBERT 1999).

168

Fig. 8.7: Schematic scetch of the different types of primary and secondary sinks for metals in the heap leach pad of the Sanyati mine.

173

Fig. 8.8: Semi-quantitative estimation of the distribution of Cu in ROM and LPO based on the average composition of ROM and LPO (cf. Fig. 5.2), average contents of Cu in goethite and hematite of the supergene ore (as determined by microprobe analysis), and assuming that goethite and hematite are equidistributed in both ore types.

174

List of tables page

Table 1.1: Raw materials used for leaching (from HABASHI 2003). 14

Table 1.2: Comparison of leaching methods and equipment (from HABASHI 2003). 15

Table 2.1: Stratigraphy of the Magondi Belt and related basement and cover units (TRELOAR 1988). 27

Table 2.2: Erosion cycles in Zimbabwe (LISTER 1987). 30

Table 2.3: Paragenesis of the primary ore of the Sanyati deposit (BAHNEMANN 1961; OBERTHÜR & KOCH 1999);*: ordered with decreasing abundance, **: alphabetical order.

33

Table 2.4: Frequently occurring Fe and nonferrous metal phases in the supergene ore of the Sanyati mine, according to VETTER et al. (1999).

35

Table 3.1: Locations and numbers of supergene ore samples. 41

Table 3.2: Locations and numbers of distal host rock samples. 42

Table 3.3: Overview of the laboratory experiments. 44

Table 3.4: Conditions and materials used for the partial extraction experiments. 48

Table 3.5: Instrument parameters for XRD measurements. 50

Table 3.6: Instrument parameters applied for microprobe measurements in Berlin (Cameca Camebax) and Bristol (JEOL JXA 8600).

52

Table 4.1: Mineral phases present in "core stones" and the weathered matrix of distal host rock.. 57

Table 4.2: Cu, Zn, and Pb contents of samples from weathering outcrops distal to the mineralisation as indicator for range of solutions derived from the ore lenses.

64

Table 4.3: Mean of minor and trace metals in proximal rocks (amphibolites and phyllites). 65

Table 4.4: Major minerals in the oxidation zone of the open pits (Copper-Queen, Copper-Queen-Beacon, F-Body-S, F-Body-N).

74

Table 4.5: Goethite/hematite ratio of supergene ore samples from the open pits (Copper-Queen, Copper-Queen-Beacon, F-Body-S, F-Body-N).

75


205

List of tables continued:

Table 4.6: List of the Sanyati supergene ore minerals compiled from Vetter et al. (1999), Frei & Germann (2001b) and this study.

77-78

Table 4.7: Trace metal contents observed in frequently occurring base metal phases in the supergene ores of Sanyati.

87

Table 4.8: Chemical composition of water samples from the bases of Copper-Queen and J-Body open pits. 87

Table 4.9: Minerals constituting the efflorescences found in the open pits of Sanyati. 88

Table 4.10: Indicators for maturity of the oxidation zone. 89

Table 5.1: Minerals present in run-of-mine ore (ROM) and leach pad ore (LPO). 95

Table 5.2: Fe, As, Cu, Zn, and Co concentrations of the fractions 100 - 2 mm (A) and <0.063 mm (B) of the wet sieved samples from profile A2-2. The higher metal recovery from the finer fraction can be estimated from the max./min. ratio.

97

Table 5.3: Mineralogical composition of "efflorescences" sampled on the heap leach pad and emergency pond..

101

Table 5.4: Mineral names and formulas of the "efflorescences" detected on the Sanyati heap leach pads and emergency pond (classification by STRUNZ & NICKEL 2001).

101

Table 5.5: Chemical composition of three representative soluble sulfate efflorescence samples from Sanyati. 102

Table 6.1: Cations substituting for Fe3+ in goethite - maximum substitution and unit cell parameters (compiled from CORNELL & SCHWERTMANN 1996; SHANNON 1976; SMYTH & BISH 1988).

108

Table 6.2: Characteristic sulfide decay textures (see images Fig. 6.4 to Fig. 6.20) observed in Sanyati ROM and LPO. (For abbreviations see Appendix A, Table A1).

111

Table 6.3: Mean of the standard deviations of the data sorted by boxwork types and examined sections. 133

Table 6.4: Element concentrations in bulk supergene ore samples determined by XRF. 135

Table 7.1: Metal production rates calculated from leaching experiment V6. 149

Table 7.2: Rate equations used for modeling the dissolution kinetics of experiment V6 (from BROWN et al. 1980; GIOVANOLI & BRÜTSCH 1975, cited in CORNELL & SCHWERTMANN 1996; and HABASHI 1969).

149

Table 7.3: k and R2 for rate equation modeling of experiment V6. 150

Table 7.4: Experimental conditions and materials used for extraction experiments on sample 263. 151

Table 8.1: Comparison of the observed Cu adsorption to goethite during experiments in different sulfate-bearing systems. Data are from JUANG & WU (2002), BALISTRIERI & MURRAY (1982), ALI & DZOMBAK (1996), and this study.

167

12. Acknowledgements

206

12 Acknowledgements

This study was initiated and performed at the Department of Economic Geology and

Applied Geochemistry of the Technical University of Berlin (TUB) under the critical as

well as tolerant supervision of Prof. Dr. Klaus Germann. I would like to thank him for

scientific support and encouragement throughout the entire project. Many thanks are due to

the official referees Prof. Dr.-Ing. Dr. h. c. mult. F.-W. Wellmer and Prof. Dr. G. Franz.

The project was hooked on a research project of the section of Economic Geology of the

BGR (Federal Institute of Geosciences and Natural Resources) in Hannover. I am thankful

for financial support within the framework "BGR-Hochschulvergabeprojekte" for the field

campaign and some analytical work as well as a lot of scientific and logistic information

from Dr. Thomas Oberthür and Dipl.-Min. Ulrich Vetter. At the Sanyati mine my special

thanks go to the mine manager, R. Mushangwe, and lots of helpful miners who supported

our field campaign in every posssible way. In the geochemical lab of the TUB Dr. Günther

Matheis has to be thanked for giving assistance to find individual ways for "out of range"-

samples. Thanks to the many helpful hands in the TU labs (Iris Pieper in the experimental

lab., Dr. Heinz Holl, Cordelia Lange (XRD), Lothar Domin (XRF), Wolfgang Becker

(ICP-OES and AAS), François Galbert (EMPA), Silke Stöwer and Cornelia von

Engelhardt in the thin section lab). Some EMPA and the LA-ICP-MS measurements were

carried out at the Large Scale Geochemical Facility supported by the European Community

- Access to Research Infrastructure action of Improving Human Potential Programme,

contract number HPRI-CT-1999-00008 awarded to Prof. B. J. Wood (University of

Bristol). My special thanks go to Stuart Kearns and Dr. Bruce Paterson for introducing me

to work at the SEM, EMPA and LA-ICP-MS as well as to Dr. John Dalton for the perfect

organisation of my research visits at the University of Bristol. Thanks are also due to Dr.

Matthias Koch-Moeck for frequent computer-related advice, to Rosemarie Geffe for

helping with some graphic design, and to Dr. Dirk Frei for critical reading of the

manuscript.

Beside all the scientific encouragement mentioned above my special thanks go to my

family and parents, who gave so much support when needed.

13. Curriculum vitæ

207

13 Curriculum vitæ

I was born in 1969 in Munich, Germany, as the child of Dr. Hans Joachim Frei (surgeon)

and his wife Rosemarie Frei (medical assistent) and spent most of my youth and school

time in southern Germany, interrupted by a three year excursion to Johannesburg, Republic

of South Africa. In 1989 I finished school with the "Abitur" in Rastatt, Germany.

In October 1989 I took up my studies in chemistry and mineralogy at the University of

Tübingen. In 1993, after having finished my BSc in both subjects, I changed to the

University of Göttingen for master courses in geology. With main focus on geochemistry

and a master thesis in environmental geochemistry, supervised by Prof. Dr. H. Heinrichs in

the Department of Geochemistry, I finished my studies in October 1998.

After a halfyear project contract in an environmental and geotechnical engineering and

consulting company in Neuss, Germany, I took up a five years position as a research,

administration, and teaching assistant at the Department of Economic Geology and

Applied Geochemistry, Institute of Applied Geosciences, Technical University of Berlin.

This position offered me the possibility to initiate and carry out the research on which this

PhD thesis is founded.

Appendix Index

208

Appendix Index

Appendix A:

Table A1: List of abbreviations Fig. A1: Quality control of XRF methods Table A2: Detection limits of ICP-OES, F-AAS and G-AAS

Appendix B: Sample lists and results of XRF analysis

Sample lists

Table B1: Sample list of the supergene ore and proximal host rock Table B2: Sample list of distal host rocks Table B3: Sample list of secondary sulfide ore and primary sulfide ore Table B4: Sample list of LPO from profiles A2-2 and A2-4 XRF results

Table B5: XRF results of ores and proximal host rocks (profiles and single samples) Table B6: Selected sample set from Table B5 showing normal distibution in histograms (see Chap. 4.2.3.2) Table B7: XRF results of distal host rocks (weathering profiles and single samples) Table B8: XRF results of ROM: 2 samples < 3 cm, sieved sample and 10 lump ores Table B9: XRF results of LPO from profiles A2-2 and A2-4 XRD results

Table B10: Position of the (111)-peak in selected samples

Appendix C: Results of the laboratory experiments

Table C1: Position of samples used in the experiments in the profiles on the heap leach pad

Table C2: XRF data of ore samples used in the laboratory experiments and filtrates Table C3: Composition of the water-soluble fraction of the samples from profile A2-2 and A2-4 (V0) Table C4: Composition of the H2SO4-soluble fraction leached from Run-of-mine ore (VR) Table C5: Composition of the H2SO4-soluble fraction leached from leach pad ore (V15) Table C6: Composition of the H2SO4-soluble fraction leached from leach pad ore (V1, V6, V7) Table C7: Composition of the partial extraction experiments with NH4Acetate (V12),

NH4Oxalate dark (V13), NH4Oxalate (V14) Table C8: Composition of sorption experiments of malachite (V8), chrysocolla (V9)

and synthetic Cu-sulfate (V10) in an acidic sulfate-rich environment

Appendix Index

209

Appendix D: Results of EMPA and Laser-ICP-MS

Table D1: EMPA results of goethite- and hematite-rich zones in ROM and LPO Table D2: List of abbreviations used in the boxplots in Chapter 6 Table D3: EMPA results of primary sulfides and minerals in the secondary sulfide

ores Table D4: EMPA results of some chlorites in the supergene ore Table D5: EMPA results of coronadite Table D6: EMPA results of plumbojarosite Table D7: EMPA results of individual base metal minerals and goethite in their

direct vicinity Table D8: LA-ICP-MS results of goethite- and hematite-rich zones in selected

sections of ROM and LPO

Appendix E: Results of the chemical composition of the leaching acid

Table E1: Chemical composition of the leaching acid prior and after the passage through the heap leach pad

Appendix A

210

Appendix A Table A1: Abbreviations ad: Adamite AMD: Acid mine drainage asp: Arsenopyrite cc: Chalcocite chr: Chrysocolla CJ: Copper-Joker orebody CK: Copper-King orebody clin: Clinoclase coll. py: Colloform pyrite cp: Chalcopyrite Co ad: Cobalto-adamite CQ: Copper-Queen open pit CQB: Copper-Queen-Beacon open pit CQD: Copper Queen Dome (intrusion) CQSW: Copper-Queen-South-West open pit cub: Cubanite cup: Cuprite cv: Covellite dg: Digenite EMPA: Electron microprobe analyses F-AAS: Flame atomic absorption spectrometry FBN: F-Body-North open pit FBS: F-Body-South open pit G-AAS: Graphite furnace atomic absorption spectrometry goe: Goethite hem: Hemimorphite ICP-OES: Inductively coupled plasma - optical emissions spectrometry JB: J-Body open pit JLN: J-Lines-North open pit LA-ICP-MS: Laser ablation - inductively coupled plasma - mass spectrometry Low Cu ad: Low Cu-adamite LPO: Leach pad ore mal: Malachite Me: metal ion mio: Million mol: Mol

Appendix A

211

Table A1 cont.: Abbreviations nat. cu: Native copper n.d.: not detected ol: Olivenite P: Pressure P-CQ: Profile in Copper-Queen open pit P-CQB: Profile in Copper-Queen-Beacon open pit P-FBN: Profile in F-Body-North open pit P-FBS: Profile in F-Body-South open pit po: Pyrrhotite py: Pyrite ROM: Run-of-mine ore SAN: Sanyati mine sp: Sphalerite std. dev.: Standard deviation T: Temperature ten: Tenorite XRD: X-ray defraction XRF: X-ray fluorescence

Appendix A

212

Fig. A1: Quality control by comparison of XRF data measured with the instruments Philips PW 1404 at TU-Berlin and Philips PW 1480/2400 BGR (Hannover).

Appendix A

213

Fig. A1 contiunued: Quality control by comparison of XRF data measured with the instruments Philips PW 1404 at TU-Berlin and Philips PW 1480/2400 BGR (Hannover).

Appendix A

214

Table A2: Detection limits of ICP-OES, F-AAS, G-AAS

ICP-OES F-AAS G-AAS [mg/l] [mg/l] [mg/l] Al 0.20 0.10 As 0.01 Ca 0.01 0.10 Cd 0.05 0.05 0.001 Co 0.10 0.05 0.01 Cu 0.05 0.05 0.005 Fe 0.10 0.05 0.005 Mg 0.05 0.005 Mn 0.02 0.05 0.005 Na 0.10 Pb 0.05 0.10 0.01 Si 0.10 Zn 0.05 0.05 0.001

Appendix B

215

Table B1: Sample list of the supergene ore and proximal host rock. Note: Single samples are listed in chronological order, while profile samples are listed in the order taken in the profile e.g. from east (E) to west (W)

Des

crip

tion

whi

te, f

ine

grai

ned,

fria

ble,

por

ous,

with

bro

wn

impr

egna

tions

of i

rono

xyhy

drox

ide/

oxid

e an

d in

ters

pers

ed w

ith

light

blu

e ch

alca

nthi

te

whi

te, p

artly

red,

fine

gra

ined

, cla

yey,

with

red

irono

xyhy

drox

ide/

oxid

e im

preg

natio

ns, p

artly

den

se re

d ar

eas

light

gre

y, fi

ne g

rain

ed

dark

blu

e-gr

een,

fine

gra

ined

, cla

yey,

few

whi

te, c

laye

y ar

eas,

frac

ture

s with

iron

oxyh

ydro

xide

/oxi

de c

oatin

gs

red,

ear

thy

oliv

e gr

een,

coa

rse

grai

ned,

par

tly fr

actu

res w

ith ir

onox

yhyd

roxi

de/o

xide

coa

tings

da

rk g

reen

, mid

dle

grai

ned,

rich

in b

iotit

e, fr

actu

res w

ith ir

onox

yhyd

roxi

de/o

xide

coa

tings

dark

blu

e-gr

een,

few

whi

te a

reas

, coa

rse

grai

ned,

cla

yey

as st

rong

ly w

eath

ered

whi

te, p

artly

red,

fine

gra

ined

, cla

yey,

impr

egna

ted

by ir

onox

yhyd

roxi

de/o

xide

brow

n-bl

ack,

com

pact

whi

te, p

artly

red,

fine

gra

ined

, cla

yey,

par

ts w

ith re

d iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

whi

te, p

artly

turq

uois

e, fi

ne g

rain

ed, c

laye

y, w

ith re

d iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

dark

gre

en, f

ine

grai

ned,

com

pact

brow

n, p

artly

den

se, p

artly

ear

thy

brow

n, c

ompa

ct, b

oxw

ork-

stru

ktur

e, w

ith su

lphi

de/a

rsen

ide

relic

ts

whi

te, f

ine

grai

ned,

cla

yey

but s

ilici

fied,

few

iron

oxyh

ydro

xide

/oxi

de im

preg

natio

ns

red,

par

tly b

row

n, fr

iabl

e, p

orou

s, re

d an

d br

own

area

s 1:1

brow

n, fr

iabl

e, p

orou

s, fe

w b

lack

frac

ture

coa

tings

dark

red

partl

y bl

ack,

har

d, c

ompa

ct

blac

k, p

artly

com

pact

, par

tly c

ellu

lar s

pong

e

blac

k, p

artly

com

pact

, par

tly c

ellu

lar s

pong

e w

hite

, cla

yey,

fria

ble,

stro

ngly

wea

ther

ed

yello

w to

ligh

t bro

wn,

sand

y ap

pear

ence

, with

iron

oxyh

ydro

xide

/oxi

de im

preg

natio

ns

blac

k, p

artly

com

pact

, par

tly c

ellu

lar s

pong

e ye

llow

- lig

ht b

row

n, c

ellu

lar s

pong

e

Rock

type

phyl

lite

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed c

hlor

ite sc

hist

supe

rgen

e or

e

wea

ther

ed a

mph

ibol

ite

wea

ther

ed a

mph

ibol

ite

wea

ther

ed a

mph

ibol

ite

wea

ther

ed p

hylli

te

supe

rgen

e or

e

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

chlo

rite

schi

st

supe

rgen

e or

e

supe

rgen

e or

e

wea

ther

ed p

hylli

te

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

supe

rgen

e or

e

supe

rgen

e or

e

Loca

tion/

Out

crop

Cop

per-

Que

en

Cop

per-

Que

en p

rofil

e (E

)

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e (W

)

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en p

rofil

e

Cop

per-

Que

en-B

eaco

n

Cop

per-

Que

en-B

eaco

n

Cop

per-

Que

en-B

eaco

n

Cop

per-

Que

en-B

eaco

n

Cop

per-

Que

en-B

eaco

n pr

ofile

(S)

Cop

per-

Que

en-B

eaco

n pr

ofile

Cop

per-

Que

en-B

eaco

n pr

ofile

Cop

per-

Que

en-B

eaco

n pr

ofile

Cop

per-

Que

en-B

eaco

n pr

ofile

Cop

per-

Que

en-B

eaco

n pr

ofile

(N)

Sam

ple

nr.

255n

265

264

208

263

207

262

206

205a

205b

261

204

203

268

a-d

209

a-c

153

155a

-1

155a

-2

155c

13

14

18

20

21a

22

Appendix B

216

Table B1 continued: Sample list of the supergene ore and proximal host rock.

Des

crip

tion

purp

le, p

artly

bro

wn,

par

tly b

recc

ious

mat

eria

l sur

roun

ded

by ir

onox

yhyd

roxi

de/o

xide

, har

d, d

ense

, with

stro

ng

irono

xyhy

drox

ide/

oxid

e im

preg

natio

ns

purp

le, p

artly

bro

wn,

fine

gra

ined

, stro

ng ir

onox

ide

impr

egna

tions

red,

par

tly b

row

n, fi

ne g

rain

ed, h

ard,

den

se

red,

par

tly b

row

n, fi

ne g

rain

ed, h

ard,

den

se, h

emim

orph

ite fr

actu

re c

oatin

g

red,

par

tly b

row

n, fi

ne g

rain

ed, h

ard,

den

se, w

hite

Cu-

free

Ada

mite

frac

ture

coa

ting

bro

wn,

fria

ble,

par

tly c

ellu

lar s

pong

e, re

licts

of a

n am

phib

olite

, litt

le b

lue-

gree

n Cu

min

eral

isat

ion

red,

mid

dle

grai

ned,

har

d, c

ompa

ct, i

rono

xide

impr

egna

tions

, rel

icts

of a

n am

phib

olite

dar

kbro

wn

partl

y br

own,

fria

ble,

par

tly c

ellu

lar s

pong

e, re

licts

of a

n am

phib

olite

bro

wn,

fria

ble,

par

tly c

ellu

lar s

pong

e, re

licts

of a

n am

phib

olite

whi

te, p

artly

turq

uois

e, m

iddl

e gr

aine

d, c

laye

y, st

rong

ly w

eath

ered

grey

gree

n to

red,

with

stro

ng ir

onox

yhyd

roxi

de/o

xide

impr

egna

tions

and

coa

tings

on

frac

ture

s, re

licts

of a

n

amph

ibol

ite

brow

n to

red,

with

stro

ng ir

onox

yhyd

roxi

de/o

xide

impr

egna

tions

and

coa

tings

on

frac

ture

s, re

licts

of a

n am

phib

olite

whi

te, p

artly

red,

mid

dle

grai

ned,

cla

yey,

inte

nsiv

e iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

brow

n, fi

ne g

rain

ed, d

ense

, iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

light

bro

wn,

fine

gra

ined

, stro

ngly

wea

ther

ed, s

andy

app

eara

nce,

with

iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

gree

n, st

rong

ly w

eath

ered

, with

coa

rse

grai

ned

biot

ite a

nd a

mph

ibol

e

light

gre

en, f

riabl

e, st

rong

ly w

eath

ered

blac

k, m

iddl

e gr

aine

d, ri

ch in

bio

tite,

with

iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

grey

-gre

en, f

ine

grai

ned,

few

frac

ture

s with

iro

noxy

hydr

oxid

e/ox

ide

coat

ings

turq

uois

e, h

ard,

com

pact

, with

few

iron

oxyh

ydro

xide

/oxi

de im

preg

natio

ns

red,

mid

dle

grai

ned,

fria

ble,

cel

lula

r spo

nge

, stro

ng ir

onox

yhyd

roxi

de/o

xide

impr

egna

tions

dark

brow

n-bl

ack,

har

d, c

ompa

ct

yello

w to

ligh

tbro

wn,

fine

gra

ined

, par

tly c

ompa

ct p

artly

cel

lula

r spo

nge,

stro

ngly

wea

ther

ed, s

trong

iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

light

brow

n, fi

ne g

rain

ed, p

artly

com

pact

, stro

ngly

wea

ther

ed, s

trong

iron

oxyh

ydro

xide

/oxi

de im

preg

natio

ns

whi

te p

artly

turq

uois

e, fi

ne g

rain

ed, c

laye

y, st

rong

ly w

eath

ered

gray

gree

n, fi

ne g

rain

ed, c

ompa

ct

whi

te, p

artly

red,

mid

dle

grai

ned,

cla

yey,

stro

ngly

wea

ther

ed

whi

te, f

ine

grai

ned,

com

pact

dark

bro

wn,

par

tly re

d, p

artly

den

se, p

artly

cel

lula

r spo

nge

light

bro

wn,

ear

thy,

relic

t am

phib

olite

, gre

en C

u-m

iner

alis

atio

n

Roc

k ty

pe

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

supe

rgen

e or

e

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

supe

rgen

e or

e

wea

ther

ed. a

mph

ibol

ite

supe

rgen

e or

e

supe

rgen

e or

e

wea

ther

ed a

mph

ibol

ite

supe

rgen

e or

e

supe

rgen

e or

e

wea

ther

ed a

mph

ibol

ite

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed a

mph

ibol

ite

wea

ther

ed a

mph

ibol

ite

wea

ther

ed p

hylli

te

rich

ore

wea

ther

ed a

mph

ibol

ite

supe

rgen

e or

e

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed a

mph

ibol

ite

wea

ther

ed p

hylli

te

supe

rgen

e or

e

supe

rgen

e or

e

Loca

tion/

Out

crop

F-Bo

dy-N

(dum

p)

F-Bo

dy-N

F-Bo

dy-N

F-Bo

dy-N

F-Bo

dy-N

F-Bo

dy-N

F-Bo

dy-N

F-Bo

dy-N

(len

s)

F-Bo

dy-N

(len

s)

F-Bo

dy-N

pro

file

(SE)

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

F-Bo

dy-N

pro

file

(NW

)

F-Bo

dy-S

F-Bo

dy-S

F-Bo

dy-S

F-Bo

dy-S

pro

file

(NW

) F-

Body

-S p

rofil

e

F-Bo

dy-S

pro

file

F-Bo

dy-S

pro

file

F-Bo

dy-S

pro

file

F-Bo

dy-S

pro

file

F-Bo

dy-S

pro

file

F-Bo

dy-S

pro

file

(SE)

Sam

ple

nr.

39

45

55a

55b

55c

58

147

149-

a

149-

b

53

52

51

50

49

48

47

145

144

46

37

65

33

276/

1 27

6/2

275

274

273

272

271/

1

271/

2

Appendix B

217


Des

crip

tion

blac

k, m

iddl

e gr

aine

d, ri

ch in

bio

tite,

with

iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

grey

-gre

en, f

ine

grai

ned,

few

frac

ture

s with

iro

noxy

hydr

oxid

e/ox

ide

coat

ings

turq

uois

e, h

ard,

com

pact

, with

few

iron

oxyh

ydro

xide

/oxi

de im

preg

natio

ns

red,

mid

dle

grai

ned,

fria

ble,

cel

lula

r spo

nge

, stro

ng ir

onox

yhyd

roxi

de/o

xide

impr

egna

tions

dark

brow

n-bl

ack,

har

d, c

ompa

ct

yello

w to

ligh

tbro

wn,

fine

gra

ined

, par

tly c

ompa

ct p

artly

cel

lula

r spo

nge,

stro

ngly

wea

ther

ed, s

trong

iro

noxy

hydr

oxid

e/ox

ide

impr

egna

tions

light

brow

n, fi

ne g

rain

ed, p

artly

com

pact

, stro

ngly

wea

ther

ed, s

trong

iron

oxyh

ydro

xide

/oxi

de im

preg

natio

ns

whi

te p

artly

turq

uois

e, fi

ne g

rain

ed, c

laye

y, st

rong

ly w

eath

ered

gray

gree

n, fi

ne g

rain

ed, c

ompa

ct

whi

te, p

artly

red,

mid

dle

grai

ned,

cla

yey,

stro

ngly

wea

ther

ed

whi

te, f

ine

grai

ned,

com

pact

dark

bro

wn,

par

tly re

d, p

artly

den

se, p

artly

cel

lula

r spo

nge

light

bro

wn,

ear

thy,

relic

t am

phib

olite

, gre

en C

u-m

iner

alis

atio

n

red,

par

tly b

row

n, fr

iabl

e, p

artly

har

d, c

ompa

ct, p

artly

cel

lula

r spo

nge,

chr

ysoc

olla

-impr

egna

tions

red,

par

tly b

row

n, fr

iabl

e, p

artly

har

d, c

ompa

ct, p

artly

cel

lula

r spo

nge,

chr

ysoc

olla

-impr

egna

tions

red,

par

tly b

row

n, fr

iabl

e, p

artly

har

d, c

ompa

ct, p

artly

cel

lula

r spo

nge,

chr

ysoc

olla

-impr

egna

tions

dark

brow

n, h

ard,

com

pact

red

to b

row

n, e

arth

y, so

ft, p

orou

s

light

to d

ark

brow

n, h

ard,

com

pact

, rel

ict a

mph

ibol

ite

dark

bro

wn,

par

tly re

d, h

ard,

par

tly c

ompa

ct, p

artly

cel

lula

r spo

nge

yel

low

, fin

e gr

aine

d, h

ard,

par

tly c

ompa

ct, r

elic

t am

phib

olite

, con

tain

s sul

phid

es/a

rsen

ides

brow

n, e

arth

y, re

lict a

mph

ibol

ite

brow

n, e

arth

y

bro

wn,

ear

thy,

relic

t am

phib

olite

blac

kbro

wn,

har

d, c

ompa

ct, r

elic

t am

phib

olite

blac

kbro

wn,

ear

thy

dark

brow

n to

ligh

tbro

wn,

har

d, c

ompa

ct, r

elic

t am

phib

olite

red

partl

y br

own,

har

d, p

artly

com

pact

, par

tly e

arth

y, p

artly

bre

ccio

us p

hylli

tes i

ncor

pora

ted

brow

n, p

artly

red,

har

d, c

ellu

lar s

pong

e

red,

har

d, c

ellu

lar s

pong

e

red,

har

d, p

artly

den

se, p

artly

ear

thy,

par

tly c

ellu

lar s

pong

e

Rock

type

wea

ther

ed a

mph

ibol

ite

wea

ther

ed p

hylli

te

rich

ore

wea

ther

ed a

mph

ibol

ite

supe

rgen

e or

e

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed a

mph

ibol

ite

wea

ther

ed p

hylli

te

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

wea

ther

ed a

mph

ibol

ite

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

Loca

tion/

Out

crop

F-B

ody-

N p

rofil

e

F-B

ody-

N p

rofil

e (N

W)

F-B

ody-

S

F-B

ody-

S

F-B

ody-

S

F-B

ody-

S pr

ofile

(NW

) F-

Bod

y-S

prof

ile

F-B

ody-

S pr

ofile

F-B

ody-

S pr

ofile

F-B

ody-

S pr

ofile

F-B

ody-

S pr

ofile

F-B

ody-

S pr

ofile

F-B

ody-

S pr

ofile

(SE)

J-Li

nes-

N

J-Li

nes-

N

J-Li

nes-

N

J-Li

ne-N

J-Li

ne-N

J-B

ody

J-B

ody

J-B

ody

(dum

p)

J-B

ody

J-B

ody

J-B

ody

J-B

ody

J-B

ody

J-B

ody

Cop

per-

Joke

r

Cop

per-

Joke

r

Cop

per-

Joke

r

Cop

per-

Joke

r

Sam

ple

nr.

144

46

37

65

33

276/

1 27

6/2

275

274

273

272

271/

1

271/

2

230a

230b

230c

235a

235b

24

26

141

255-

1

255-

2

255-

3

256

257

258

156

158-

a

158-

b

160

Appendix B

218


Des

crip

tion

redb

row

n w

ith li

ghtb

row

n sp

ecs,

fine

grai

ned,

ear

thy,

por

ous

red,

fine

gra

ined

, ear

thy

redb

row

n fr

ame

of a

box

wor

k te

xtur

e w

ith li

ghtb

row

n fil

ling

of c

aviti

es

dark

brow

n-bl

ack

partl

y lig

htbr

own,

par

tly e

arth

y, p

artly

com

pact

, por

ous,

band

ed

brow

n pa

rtly

light

brow

n, h

ard,

com

pact

, rel

ict o

f am

phib

ole,

gre

en m

iner

alis

atio

n on

frac

ture

s (ol

iven

ite)

red

partl

y da

rkbr

own,

har

d bo

xwor

k te

xtur

e w

ith e

arth

y ca

vity

filli

ngs,

partl

y ea

rthy

partl

y co

mpa

ct

light

brow

n to

dar

kbro

wn,

har

d, c

ompa

ct

red

to d

arkb

row

n, h

ard,

por

ous

dark

brow

n pa

rtly

blac

k, h

ard,

com

pact

, rel

ict a

mph

ibol

ite

red

partl

y br

own,

har

d, p

artly

com

pact

, par

tly c

ellu

lar s

pong

e

red,

ear

thy,

por

ous

brow

n, h

ard,

par

tly c

ompa

ct p

artly

cel

lula

r spo

nge

dark

brow

n, h

ard,

por

ous

dark

brow

n, h

ard,

por

ous

Roc

k ty

pe

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

wea

ther

ed p

hylli

te

wea

ther

ed a

mph

ibol

ite

supe

rgen

e or

e

wea

ther

ed p

hylli

te

supe

rgen

e or

e

wea

ther

ed a

mph

ibol

ite

wea

ther

ed a

mph

ibol

ite

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

supe

rgen

e or

e

Loca

tion/

Out

crop

Cop

per-

Que

en-S

W

Cop

per-

Que

en-S

W

Cop

per-

Que

en-S

W

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Cop

per-

Kin

g

Sam

ple

nr.

1F

3F

4F

75

76

77

78

79

80

81

82

69

72

73

Appendix B

219

Table B2: Sample list of distal host rocks.

Des

crip

tion

sapr

olith

, red

dish

, fin

e gr

aine

d, fr

iabl

e

redd

ish,

fine

gra

ined

, ear

thy

redd

ish,

fine

gra

ined

, ear

thy

light

bro

wn,

fine

gra

ined

, ear

thy

grey

, fin

e gr

aine

d, c

ompa

ct, "

core

-sto

ne" c

ente

r

light

gre

y, fi

ne g

rain

ed, f

riabl

e, ri

m o

f "co

re-s

tone

"

light

gre

y, fi

ne g

rain

ed, f

riabl

e, su

rrou

ndin

g of

"co

re-s

tone

" >2

mm

light

gre

y, fi

ne g

rain

ed, f

riabl

e, su

rrou

ndin

g of

"co

re-s

tone

" <2

mm

grey

, fin

e gr

aine

d, c

ompa

ct, "

core

-sto

ne" c

ente

r

grey

, fin

e gr

aine

d, fr

iabl

e, ri

m o

f "co

re-s

tone

"

redd

ish,

fine

gra

ined

, ear

thy,

surr

ound

ing

of "

core

-sto

ne" >

2 m

m

grey

, fin

e gr

aine

d, e

arth

y, su

rrou

ndin

g of

"co

re-s

tone

" <2

mm

grey

, cla

yey,

fria

ble,

moi

st

grey

, fin

e gr

aine

d, c

ompa

ct

grey

, fin

e gr

aine

d, fr

iabl

e, ri

m o

f "co

re-s

tone

"

whi

te c

rust

filli

ng a

frac

ture

red,

fine

gra

ined

, ear

thy

redd

ish,

fine

gra

ined

, fria

ble,

sapr

olith

light

gre

y, fi

ne g

rain

ed, f

riabl

e

light

gre

y, fi

ne g

rain

ed, e

arth

y

gree

n-gr

ey, f

ine

grai

ned,

com

pact

, fol

iate

d

gree

n-gr

ey, f

ine

grai

ned,

com

pact

, fol

iate

d

whi

te-b

eige

, fin

e gr

aine

d, c

ompa

ct

whi

te-b

eige

, fin

e gr

aine

d, c

ompa

ct

whi

te-g

rey,

mid

dle

grai

ned,

with

kao

liniz

ed fe

ldsp

ars,

garn

et b

earin

g

brow

m, c

lay-

fine

sand

brow

n, fi

ne g

rain

ed, e

arth

y

Roc

k ty

pe

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

unco

nsol

idat

ed se

dim

ent

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

wea

ther

ed p

hylli

te

dolo

mite

dolo

mite

wea

ther

ed p

hylli

te

unco

nsol

idat

ed se

dim

ent

unco

nsol

idat

ed se

dim

ent

Loca

tion/

Out

crop

Cop

per-

Joke

r wea

ther

ing

prof

ile

Cop

per-

Joke

r wea

ther

ing

prof

ile

Cop

per-

Joke

r wea

ther

ing

prof

ile

Cop

per-

Joke

r wea

ther

ing

prof

ile (t

erm

ite h

ill)

"Cor

e-St

one"

-out

crop

Cop

per-

Que

en

"Cor

e-St

one"

-out

crop

Cop

per-

Que

en

"Cor

e-St

one"

-out

crop

Cop

per-

Que

en

"Cor

e-St

one"

-out

crop

Cop

per-

Que

en

"Cor

e-St

one"

-out

crop

J-Bo

dy

"Cor

e-St

one"

-out

crop

J-Bo

dy

"Cor

e-St

one"

-out

crop

J-Bo

dy

"Cor

e-St

one"

-out

crop

J-Bo

dy

J-B

ody

open

pit

J-B

ody

open

pit

J-B

ody

open

pit

J-B

ody

open

pit

J-B

ody

open

pit

J-B

ody

open

pit

J-B

ody

open

pit

J-B

ody

open

pit

Dirt

road

out

crop

Dirt

road

out

crop

Cop

per-

Que

en o

pen

pit

Cop

per-

Que

en o

pen

pit

Cop

per-

Que

en o

pen

pit

Val

ley

wes

t of C

oppe

r-Q

ueen

ope

n pi

t

East

of C

oppe

r Que

en D

omes

(ter

mite

hill

)

Sam

ple

nr.

161a

161b

162

159

181a

181b

181d

181e

246a

246b

247

248

249

250

251a

251b

252

253

254b

254a

168

167a

201

195

5 169

281

Appendix B

220

Table B2 continued: Sample list of distal host rocks.

Des

crip

tion

brow

n, fi

ne g

rain

ed, e

arth

y

grey

, fin

e gr

aine

d

dark

gre

y, fi

ne g

rain

ed, c

ompa

ct

grey

, fin

e gr

aine

d, fo

liate

d

grey

, fin

e gr

aine

d, fo

liate

d

grey

, unc

onso

lidat

ed m

ater

ial <

2 m

m

grey

, unc

onso

lidat

ed m

ater

ial <

1 m

m

grey

, fin

e gr

aine

d, fo

liate

d

grey

, unc

onso

lidat

ed m

ater

ial <

2 m

m

grey

, unc

onso

lidat

ed m

ater

ial <

1 m

m

light

pin

k, fi

ne g

rain

ed, c

ompa

ct

light

pin

k, fi

ne g

rain

ed, c

ompa

ct

dark

gre

y-gr

een,

med

ium

gra

ined

Rock

type

unco

nsol

idat

ed se

dim

ent

quar

tz-p

hylli

te

quar

tzite

phyl

lite

phyl

lite

unco

nsol

idat

ed m

ater

ial

unco

nsol

idat

ed m

ater

ial

phyl

lite

unco

nsol

idat

ed m

ater

ial

unco

nsol

idat

ed m

ater

ial

gran

ite

gran

ite

amph

ibol

ite

Loca

tion/

Out

crop

East

of C

oppe

r Que

en D

ome

(term

ite h

ill)

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Copp

er Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Cop

per Q

ueen

Dom

e

Sam

ple

nr.

281

277c

d1

277c

d2

280c

d

282c

d1

282c

d2<2

282c

d2<1

283c

d1

283c

d2<2

283c

d2<1

279c

d1

279c

d2

278c

d

Appendix B

221

Table B3: Sample list of secondary sulfide ore and primary sulfide ore.

Des

crip

tion

dark

grey

, tur

quoi

se, m

assi

ve, l

ayer

ed, d

ense

ore

chal

copy

rite,

ars

enop

yrite

, gal

ena,

cov

ellit

e, b

roch

antit

e

dark

grey

, tur

quoi

se-g

reen

, mas

sive

, den

se o

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Appendix B

222

Table B4: Sample list of LPO from profiles A2-2 and A2-4.

sample nr. profile hight relative to surface sieving grain size fraction m N=wet, T=dry mm

97 A2-2 -1 N >297 A2-2 -1 N >0.7197 A2-2 -1 N >0.2597 A2-2 -1 N >0.06397 A2-2 -1 N <0.06397 A2-2 -1 T >297 A2-2 -1 T >0.7197 A2-2 -1 T >0.2597 A2-2 -1 T >0.06397 A2-2 -1 T <0.063



113< 1cm A2-4 -1.5 N < 10 114< 1cm A2-4 -1 N < 10

115< 1cm A2-4 -0.5 N < 10

116< 1cm A2-4 0 N < 10

113> 1cm A2-4 -1.5 N > 10114> 1cm A2-4 -1 N > 10

115> 1cm A2-4 -0.5 N > 10

116> 1cm A2-4 0 N > 10

Appendix B

223

Table B5: XRF results of ores and proximal host rocks (profiles and single samples).

Grey marking highlights samples with >30% Fe2O3

Sample nr. Location SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 SO3

% % % % % % % % % %

255n CQ 22.6 9.51 3.17 n.d. 4.78 n.d. 0.54 0.28 0.10 15.26

203 P CQ 25.5 19.36 20.82 19.12 0.15 <0.1 <0.05 0.25 0.03 <0.05204 P CQ 47.2 24.45 0.47 <0.3 <0.1 <0.1 0.30 0.31 0.01 0.68205a P CQ 37.9 36.34 6.62 <0.3 0.38 <0.1 1.27 0.52 0.03 3.58205b P CQ n.d. 12.69 62.37 <0.9 <0.3 <0.3 <0.15 0.12 0.05 0.71206 P CQ 34.2 28.06 9.03 4.68 <0.1 <0.1 0.25 0.43 0.06 0.48207 P CQ 37.3 7.66 22.16 5.09 4.09 <0.1 0.14 0.17 0.06 0.24208 P CQ 28.2 22.84 27.99 21.63 <0.1 <0.1 <0.05 0.24 0.03 0.13209a P CQ n.d. 0.78 30.12 8.01 4.35 <0.3 <0.15 <0.09 0.23 <0.15209b P CQ n.d. <0.6 40.59 11.01 <0.3 <0.3 <0.15 <0.09 0.26 0.51209c P CQ n.d. 0.63 56.58 <0.9 <0.3 <0.3 0.18 <0.09 0.04 1.48261 P CQ 45.4 33.85 8.57 0.39 0.27 <0.1 1.07 0.64 0.11 0.58262 P CQ 28.0 14.97 17.77 8.35 0.18 <0.1 1.96 0.37 0.14 0.06263 P CQ n.d. 2.64 61.89 <0.9 <0.3 <0.3 <0.15 0.45 0.07 0.32264 P CQ 54.6 18.42 3.74 1.10 <0.1 0.19 5.14 0.36 0.04 <0.05265 P CQ 47.3 39.33 3.11 <0.3 <0.1 <0.1 0.11 0.52 0.06 0.63268a P CQ n.d. 1.77 46.95 2.28 <0.3 <0.3 <0.15 <0.09 0.44 0.19268b P CQ n.d. 1.59 31.65 10.02 5.13 <0.3 <0.15 <0.09 0.13 0.15268c P-CQ n.d. <0.6 68.52 1.50 <0.3 <0.3 <0.15 <0.09 0.07 0.54268d P-CQ n.d. <0.6 71.01 <0.9 <0.3 <0.3 <0.15 <0.09 0.09 0.27153 CQB 65.3 12.05 0.44 <0.3 <0.1 <0.1 0.25 0.38 0.05 <0.05155a/1 CQB n.d. 3.81 60.33 <0.9 <0.3 <0.3 <0.15 0.21 0.28 0.41155a/2 CQB 10.9 6.86 59.79 <0.3 0.15 <0.1 0.05 0.27 0.38 0.59155c CQB n.d. 1.92 66.48 <0.9 <0.3 <0.3 <0.15 <0.09 0.16 0.4613 P-CQB n.d. <0.6 60.15 <0.9 <0.3 <0.3 <0.15 <0.09 0.11 0.3814 P-CQB n.d. 2.25 49.80 <0.9 <0.3 <0.3 <0.15 <0.09 0.08 0.4518 P-CQB 72.2 7.27 0.89 <0.3 0.12 <0.1 <0.05 0.07 <0.01 <0.0520 P-CQB 52.2 6.86 5.82 <0.3 0.23 <0.1 0.12 0.13 0.05 0.9821a P-CQB n.d. 2.34 33.39 <0.9 <0.3 <0.3 0.33 0.12 0.08 1.5322 P-CQB 28.9 6.86 5.38 <0.3 0.12 <0.1 0.12 0.04 0.02 1.8939 FBN n.d. 11.97 18.30 <0.9 <0.3 <0.3 <0.15 0.18 0.03 0.4545 FBN n.d. 7.65 15.66 <0.9 <0.3 <0.3 0.21 0.18 0.07 <0.1555a FBN 15.1 6.86 68.35 0.72 0.65 <0.1 0.09 0.02 0.18 0.7855b FBN 38.0 6.86 22.26 3.37 1.79 0.54 0.05 <0.03 0.17 0.1055c FBN 35.2 6.86 20.86 2.92 1.26 <0.1 <0.05 <0.03 0.13 0.0458 FBN 11.8 0.76 51.42 3.61 0.15 <0.1 0.02 0.03 0.33 0.29147 FBN n.d. 5.49 26.13 13.32 2.64 <0.3 <0.15 0.12 0.09 0.18149-a FBN 28.1 6.86 54.45 6.90 1.19 <0.1 0.02 0.07 0.22 0.12149-b FBN n.d. 2.31 30.12 5.16 1.53 <0.3 <0.15 0.09 0.20 <0.1546 P-FBN 36.4 10.09 15.33 7.84 <0.1 <0.1 0.17 0.16 0.06 0.0947 P-FBN 45.9 9.21 5.54 4.44 <0.1 <0.1 0.32 0.13 0.03 0.4548 P-FBN 40.7 1.92 7.11 <0.3 <0.1 <0.1 0.11 0.14 0.05 0.4849 P-FBN 69.8 1.31 8.10 <0.3 <0.1 <0.1 0.08 0.04 0.09 0.1450 P-FBN 32.5 17.35 32.58 0.41 0.20 <0.1 <0.05 0.18 0.19 0.3551 P-FBN 26.1 5.68 32.00 7.73 2.57 <0.1 0.05 0.07 0.14 0.1052 P-FBN 31.1 7.40 23.70 8.93 4.50 <0.1 <0.05 0.13 0.10 0.0553 P-FBN 41.1 8.70 17.31 0.93 0.38 <0.1 <0.05 0.09 0.28 0.20144 P-FBN 27.9 17.33 28.03 8.38 <0.1 <0.1 1.33 0.57 0.12 0.08145 P-FBN 37.5 8.17 10.65 14.30 5.57 <0.1 0.06 0.15 0.02 <0.0537 P-FBS 63.3 0.70 0.74 1.13 <0.1< <0.1 <0.05 0.14 0.02 <0.0565 P-FBS 20.1 2.78 11.79 0.62 <0.1 <0.1 0.16 0.10 0.34 0.7733 P-FBS n.d. 1.08 49.89 2.76 <0.3 <0.3 <0.15 <0.09 0.19 1.07

Appendix B

224

Table B5 continued: XRF results of ores and proximal host rocks (profiles and single samples).

Sample nr. Location F Ag As Ba Bi Br Cd Cl Co Cu

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

255n CQ n.d. n.d. n.d. n.d. n.d. n.d. n.d. n.d. n.d. 188550

203 P CQ <800 4 19 <30 17 <4 6 <300 48 198204 P CQ <800 3 2877 411 <7 <4 <5 <300 <10 5040205a P CQ <800 4 3594 454 32 5 10 <300 10 1614205b P CQ <2400 <8 1590 <90 83 <12 <15 <900 207 10790206 P CQ <800 <3 120 97 81 <4 <5 <300 18 1701207 P CQ <800 <3 59 38 24 <4 <5 <300 29 42840208 P CQ <800 <3 184 <30 53 5 <5 <300 35 1089209a P CQ <2400 21 3642 <90 115 <12 73 <900 33 1518209b P CQ <2400 37 1992 <90 121 15 29 <900 237 10250209c P CQ <2400 54 23170 <90 469 22 517 <900 45 13810261 P CQ <800 6 159 171 366 <4 5 <300 <10 2508262 P CQ <800 3 100 236 19 <4 9 <300 37 55730263 P CQ <2400 <8 <30 <90 298 <12 <15 <900 <30 5754264 P CQ <800 7 2220 489 <7 <4 <5 <300 <10 1352265 P CQ <800 16 89 113 18 <4 <5 <300 47 2544268a P CQ <2400 19 336 <90 110 14 93 <900 240 9800268b P CQ <2400 <3 1128 <90 43 <12 27 <900 33 5600268c P-CQ <2400 <3 648 <90 79 <12 28 <900 39 2286268d P-CQ <2400 8 2319 <90 71 <12 41 <900 72 3663153 CQB <800 <3 229 194 <7 <4 <5 <300 <10 254155a/1 CQB <2400 <8 6120 345 67 <12 31 <900 108 5760155a/2 CQB 5392 33 7284 460 195 11 47 <300 209 6630155c CQB <2400 28 4410 339 67 <12 16 <900 75 408313 P-CQB <2400 10 3351 <90 59 <12 16 <900 <30 365114 P-CQB <2400 38 3696 96 54 <12 <15 <900 <30 417918 P-CQB <800 <3 164 <30 <7 <4 <5 <300 <10 24720 P-CQB 1325 18 7290 101 154 <4 28 <300 6 164421a P-CQB <2400 100 17940 339 235 17 323 <900 <30 464422 P-CQB 2145 35 44150 196 161 11 41 <300 6 96239 FBN <2400 11 4623 <90 <21 <12 <15 <900 <30 255945 FBN <2400 <8 2940 <90 <21 <12 <15 <900 <30 542755a FBN 3675 <3 6441 <30 199 5 39 <300 55 1166055b FBN <800 15 7557 <30 149 <4 80 <300 143 654355c FBN 860 5 37650 <30 54 9 58 <300 393 1809058 FBN 5220 <3 22680 <30 144 8 68 <300 21 23170147 FBN <2400 16 3393 114 <21 <12 54 <900 102 7000149-a FBN 6789 <3 7254 <30 135 <4 75 <300 62 2106149-b FBN <2400 <8 4134 <90 37 <12 35 <900 51 107146 P-FBN <800 6 472 61 36 <4 6 <300 199 125147 P-FBN <800 <3 154 37 <7 <4 <5 <300 12 222948 P-FBN <800 <3 2535 31 <7 <4 10 <300 <10 233749 P-FBN <800 <3 2385 <30 15 <4 9 <300 <10 111850 P-FBN <800 <3 2330 <30 78 <4 6 <300 18 456051 P-FBN <800 32 83 60 47 <4 15 <300 520 924052 P-FBN <800 <3 2534 <30 21 <4 18 <300 58 585053 P-FBN <800 <3 209 55 59 <4 10 <300 16 133130144 P-FBN <800 <3 609 155 41 <4 7 <300 14 1068145 P-FBN <800 <3 11 49 74 <4 <5 <300 108 3137 P-FBS <800 <3 275 <30 10 0.7< <5 <300 35 8057765 P-FBS <800 21 40730 38 185 8 49 <300 35 4113333 P-FBS 9903 41 8058 183 127 <12 52 <900 453 8610

Appendix B

225


Sample nr. Location Cr Cs Ga Mn Mo Ni Pb Rb Sb Se Sn Sr

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

255n CQ n.d. n.d. n.d. n.d. n.d. n.d. 1240 n.d. n.d. n.d. n.d. n.d.

203 P CQ 53 <8 42 1223 17 40 37 <10 12 <3 39 <10204 P CQ 46 <8 8 <150 <5 <12 417 <10 <7 <3 <10 97205a P CQ 135 <8 15 <150 <5 <12 8161 94 10 45 <10 194205b P CQ 108 <24 <9 897 <15 <36 1197 <30 54 98 <30 <30206 P CQ 59 <8 22 804 <5 <12 806 22 74 21 17 53207 P CQ <30 <8 <3 1433 <5 18 488 <10 20 14 38 31208 P CQ 58 <8 26 1107 <5 15 102 <10 77 16 37 <10209a P CQ <90 <24 <9 2817 <15 <36 21729 <30 275 53 256 <30209b P CQ <90 <24 <9 25923 <15 <36 19344 <30 64 55 158 <30209c P CQ <90 <24 <9 <450 <15 <36 19917 <30 250 77 <30 <30261 P CQ 123 <8 37 <150 <5 <12 3451 79 575 18 <10 143262 P CQ 47 28 <3 1098 <5 <12 542 122 19 10 <10 45263 P CQ <90 <24 <9 <450 <15 <36 174 <30 74 86 <30 <30264 P CQ 49 15 18 212 <5 <12 416 132 7 <3 19 68265 P CQ 174 <8 116 <150 95 <12 586 14 28 17 25 42268a P CQ <90 <24 <9 61455 <15 42 29040 33 31 75 <30 <30268b P CQ <90 <24 <9 5493 <15 <36 8262 <30 173 22 <30 <30268c P-CQ <90 <24 <9 2991 <15 <36 8160 <30 503 34 <30 <30268d P-CQ <90 <24 <9 999 <15 60 6408 <30 274 19 <30 <30153 CQB 83 <8 12 <150 <5 <12 1861 10 38 4 <10 <10155a/1 CQB <90 <24 <9 1230 <15 <36 6513 <30 61 26 <30 84155a/2 CQB <30 <8 25 2505 <5 <12 10152 17 112 35 <10 137155c CQB <90 <24 <9 1437 <15 <36 3618 <30 80 18 <30 <3013 P-CQB <90 <24 <9 <450 <15 <36 4905 <30 345 17 <30 <3014 P-CQB <90 <24 <9 2058 <15 <36 6177 <30 230 18 <30 <3018 P-CQB <30 <8 7 <150 <5 <12 358 <10 26 <3 71 <1020 P-CQB <30 <8 <3 <150 <5 <12 15348 <10 524 31 <10 <1021a P-CQB <90 <24 <9 <450 <15 <36 23505 <30 378 68 <30 <3022 P-CQB <30 <8 <3 <150 <5 <12 119650 12 691 190 <10 <1039 FBN <90 <24 <9 771 <15 <36 12186 <30 <21 31 <30 <3045 FBN <90 <24 15 n.d. <15 <36 273 <30 <21 <9 <30 <3055a FBN 17 <8 <3 2547 <5 28 13512 15 34 32 <10 <1055b FBN <30 <8 <3 9040 <5 27 19815 15 19 45 <10 <1055c FBN <30 <8 <3 13823 <5 59 2177 <10 281 11 89 1358 FBN 15 <8 <3 1190 <5 33 14890 <10 165 30 <10 <10147 FBN <90 <24 <9 8694 <15 <36 6777 <30 <21 23 <30 <30149-a FBN 15 <8 12 1112 <5 46 3955 <10 65 20 <10 <10149-b FBN <90 <24 <9 1065 <15 <36 2319 <30 30 <9 <30 <3046 P-FBN <30 10 9 2078 <5 20 261 33 <7 <3 <10 <1047 P-FBN <30 <8 11 329 <5 <12 19 13 <7 <3 <10 <1048 P-FBN <30 <8 4 <150 <5 <12 353 <10 26 <3 15 <1049 P-FBN <30 <8 <3 <150 <5 <12 384 12 42 <3 <10 <1050 P-FBN <30 <8 <3 2868 <5 <12 7447 <10 23 24 511 <1051 P-FBN <30 <8 <3 21179 <5 32 736 20 9 8 32 <1052 P-FBN <30 <8 4 3624 <5 18 903 <10 11 5 116 1353 P-FBN <30 <8 <3 1046 <5 <12 6630 <10 16 25 1493 11144 P-FBN 168 44 20 3242 <5 78 6938 262 10 19 <10 <10145 P-FBN <30 <8 16 2426 <5 20 319 <10 <7 <3 <10 <1037 P-FBS <30 <8 <3 <150 <5 <12 1546 <10 6 4 27 <1065 P-FBS <30 <8 <3 129 <5 7 31990 9 202 75 <10 2133 P-FBS <90 <24 <11 27280 <15 <36 16270 <30 <21 44 <30 <30

Appendix B

226


Sample nr. Location Th Tl U V W Zn Zr sum l.o.i sum+l.o.i.

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg % % %

255n CQ n.d. n.d. n.d. n.d. n.d. 540 n.d. - - -

203 P CQ 16 <3 8 53 <15 1521 144 85.44 10.0 95.42204 P CQ <10 5 <5 18 <15 165 54 74.17 10.3 84.44205a P CQ 17 10 <5 94 <15 333 98 87.95 19.7 107.65205b P CQ <30 <9 <15 159 <45 1386 <120 - 12.6 -206 P CQ <10 <3 7 28 <15 612 92 77.43 15.2 92.6207 P CQ <10 <3 6 28 <15 2765 41 80.06 9.3 89.33208 P CQ <10 <3 6 42 <15 2948 103 118.85 8.9 127.72209a P CQ <30 <9 <15 <30 <45 29553 <120 - 12.0 -209b P CQ 36 <9 <15 <30 <45 60246 <120 - 5.4 -209c P CQ <30 27 <15 <30 <45 16473 <120 - 12.1 -261 P CQ 32 <3 13 69 <15 1405 139 91.73 16.1 107.84262 P CQ <10 7 <5 29 <15 1649 80 75.64 10.5 86.1263 P CQ <30 <9 <15 51 <45 744 132 - 6.2 -264 P CQ 12 7 <5 37 <15 618 74 84.08 2.3 86.4265 P CQ 18 7 6 60 44 426 179 91.39 17.1 108.52268a P CQ 105 <9 <15 39 <45 48210 <120 - 11.9 -268b P CQ <30 <9 <15 33 <45 11550 <120 - 4.8 -268c P-CQ <30 <9 <15 84 <45 12330 <120 - 9.0 -268d P-CQ <30 <9 <15 <30 <45 27141 <120 - 11.5 -153 CQB <10 <3 <5 28 <15 819 89 78.94 11.5 90.45155a/1 CQB <30 <9 <15 <30 <45 9171 <120 - 11.5 -155a/2 CQB <10 18 10 17 <15 17204 160 - 14.0 -155c CQB <30 <9 <15 <30 <45 6456 <120 - 7.2 -13 P-CQB <30 <9 <15 57 <45 3714 <120 - 6.3 -14 P-CQB <30 <9 <15 84 <45 5448 <120 - 6.8 -18 P-CQB <10 <3 <5 <10 <15 2132 <40 81.07 7.0 88.0620 P-CQB <10 11 <5 17 <15 1891 <40 63.56 4.1 67.6621a P-CQB <30 24 <15 <30 <45 6279 <120 - 9.1 -22 P-CQB 17 29 <5 <10 <15 650 <40 41.51 6.9 48.3739 FBN <30 <9 <15 <30 <45 1887 <120 - 7.8 -45 FBN <30 <9 <15 36 <45 141 <120 - 4.9 -55a FBN <10 <3 <5 72 <15 27495 19 94.52 12.0 106.5255b FBN 13 8 8 50 116 71976 19 76.97 6.0 82.9655c FBN <10 32 <5 18 59 69117 <40 72.29 6.2 78.4558 FBN <10 32 8 129 <15 35640 22 79.39 - -147 FBN 33 <9 <15 <30 <45 2724 <120 - 6.2 -149-a FBN <10 10 <5 74 36 38388 27 100.24 7.1 107.37149-b FBN <30 <9 <15 75 <45 19251 <120 - 10.5 -46 P-FBN <10 <3 <5 28 <15 4448 50 70.94 5.0 75.8947 P-FBN <10 <3 <5 <10 <15 1111 <40 66.27 3.1 69.3948 P-FBN <10 <3 <5 <10 <15 341 <40 50.9 2.7 53.6449 P-FBN <10 <3 <5 <10 <15 590 <40 80.1 4.0 84.0850 P-FBN <10 <3 <5 35 <15 7589 57 86.07 13.9 10051 P-FBN 20 <3 <5 23 <15 20426 <40 87.23 9.6 96.852 P-FBN <10 <3 <5 60 <15 9791 <40 78.05 5.8 83.8753 P-FBN <10 15 <5 51 <15 8727 <40 80.48 19.8 100.3144 P-FBN 22 <3 <5 104 <15 17226 161 86.52 7.9 94.37145 P-FBN 13 <3 <5 15 <15 2030 66 76.85 6.1 82.9337 P-FBS <10 9 2.5< <10 <15 5891 24 69.54 3.5 73.0465 P-FBS <10 53 0.0< <10 <15 20940 29 43.58 7.2 50.7533 P-FBS <30 <9 <15 <30 <45 7473 <120 - 4.9 -

Appendix B

227


Sample nr. Location SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 SO3

% % % % % % % % % %

271/1 P-FBS n.d. <0.6 68.70 <0.9 <0.3 <0.3 <0.15 <0.09 0.47 0.68

271/2 P-FBS 50.6 1.02 32.01 0.40 <0.1 <0.1 0.04 0.06 0.41 0.30272 P-FBS 81.4 n.d. 11.26 1.63 1.48 n.d. n.d. n.d. 0.13 n.d.273 P-FBS 46.9 12.56 49.00 2.08 0.13 <0.1 <0.05 0.16 0.07 0.16274 P-FBS 28.7 21.69 38.54 15.12 <0.1 <0.1 <0.05 0.35 0.03 0.12275 P-FBS 44.5 35.64 1.27 <0.3 <0.1 <0.1 0.04 0.26 0.07 0.57276/1 P-FBS 4.1 4.04 15.24 0.31 <0.1 <0.1 0.04 0.08 0.07 0.14276/2 P-FBS 5.6 5.02 20.15 0.31 <0.1 <0.1 0.04 0.06 0.07 0.14230a JLN 6.9 0.93 59.89 n.d. <0.1 n.d. n.d. 0.05 0.23 0.24230b JLN 5.8 0.58 62.94 n.d. 0.39 n.d. n.d. 0.07 0.19 0.21230c JLN 19.1 0.82 31.34 n.d. 1.99 n.d. n.d. n.d. 0.17 0.15235a JLN n.d. 1.17 56.64 <0.9 <0.3 <0.3 <0.15 <0.09 0.23 1.06235b JLN n.d. 1.29 66.36 <0.9 <0.3 <0.3 <0.15 <0.09 0.28 1.3224 JB n.d. <0.6 50.79 3.21 <0.3 <0.3 <0.15 <0.09 0.17 1.2926 JB n.d. <0.6 69.15 1.41 <0.3 <0.3 <0.15 <0.09 0.14 0.87255/1 JB n.d. 2.70 33.09 9.87 2.70 <0.3 <0.15 <0.09 0.33 0.46255/2 JB n.d. 1.71 29.55 5.85 2.64 <0.3 <0.15 <0.09 0.49 0.16255/3 JB n.d. 1.20 60.18 2.28 0.72 <0.3 <0.15 <0.09 0.37 1.04257 JB n.d. 2.07 35.31 6.75 <0.3 <0.3 <0.15 <0.09 0.64 <0.15256 JB n.d. <0.6 58.32 4.71 <0.3 <0.3 <0.15 <0.09 0.29 0.54258 JB n.d. 2.01 23.76 <0.9 0.42 <0.3 <0.15 <0.09 0.18 0.66141 JB 42.7 0.19 41.39 0.82 1.47 <0.1 0.03 <0.03 0.08 0.18156 CJ n.d. 2.52 36.96 <0.9 <0.3 <0.3 <0.15 0.27 0.22 0.71158-a CJ n.d. 1.20 68.25 <0.9 <0.3 <0.3 <0.15 0.24 0.11 0.38158-b CJ n.d. 0.69 66.90 <0.9 <0.3 <0.3 <0.15 0.21 0.10 0.33160 CJ n.d. 1.89 50.64 <0.9 <0.3 <0.3 0.18 <0.09 0.32 2.281F CQSW n.d. 1.50 49.95 <0.9 <0.3 <0.3 <0.15 <0.09 0.17 9.143F CQSW n.d. 3.87 68.64 <0.9 <0.3 0.33 <0.15 <0.09 0.17 2.634F CQSW n.d. 0.93 73.38 1.83 <0.3 <0.3 <0.15 <0.09 0.27 0.6575 CK n.d. <0.6 14.25 2.94 1.77 1.02 <0.15 <0.09 0.27 <0.0976 CK n.d. <0.6 27.63 2.55 <0.3 <0.3 <0.15 <0.09 0.22 0.3477 CK n.d. <0.6 51.81 <0.9 <0.3 <0.3 <0.15 <0.09 0.19 0.5678 CK n.d. 3.39 19.02 <0.9 <0.3 <0.3 <0.15 0.21 0.03 0.2579 CK n.d. <0.6 34.44 1.71 <0.3 <0.3 <0.15 <0.09 0.09 1.6780 CK n.d. 2.34 44.55 2.31 <0.3 0.63 <0.15 0.15 0.12 0.5581 CK n.d. 1.44 41.94 2.43 <0.3 0.33 <0.15 <0.09 0.87 0.2482 CK n.d. <0.6 46.56 1.20 <0.3 <0.3 <0.15 <0.09 0.20 <0.0969 CK n.d. 1.50 38.13 7.23 <0.3 0.33 <0.15 <0.09 0.53 <0.0972 CK n.d. 0.81 46.65 1.74 1.20 <0.3 <0.15 <0.09 1.15 0.7973 CK n.d. 0.99 30.85 3.54 <0.3 <0.3 <0.15 <0.09 0.31 0.34

Grey marking highlights samples with >30% Fe2O3

Appendix B

228


Sample nr. Location F Ag As Ba Bi Br Cd Cl Co Cu

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

271/1 P-FBS <2400 <8 20060 <90 <21 <12 78 <900 <30 34910

271/2 P-FBS <800 10 17000 113 68 8 87 <300 58 33200272 P-FBS n.d. n.d. n.d. n.d. n.d. n.d. n.d. n.d. n.d. 5730273 P-FBS <800 9 3590 25 571 6 42 <300 340 1400274 P-FBS 2839 <3 474 <30 356 <4 <5 <300 131 841275 P-FBS <800 17 8538 <30 497 8 7 <300 188 6130276/1 P-FBS <800 9 170200 <30 1183 28 5 <300 170 8460276/2 P-FBS 1040 4 151500 <30 1281 34 11 <300 143 4480230a JLN n.d. n.d. 5981 n.d. n.d. n.d. n.d. n.d. n.d. 47337230b JLN n.d. n.d. 4833 n.d. n.d. n.d. n.d. n.d. n.d. 41487230c JLN n.d. n.d. 2647 n.d. n.d. n.d. n.d. n.d. n.d. 154044235a JLN 6474 9 6339 129 97 <12 19 <900 303 44610235b JLN 10419 <8 8055 249 107 <12 34 <900 480 3360024 JB 4227 <8 3057 <90 119 <12 <15 <900 <30 611026 JB <2400 10 2934 <90 165 <12 <15 <900 234 14330255/1 JB <2400 40 11067 942 61 <12 151 <900 588 11240255/2 JB 6129 <8 11805 <90 102 <12 137 <900 >600 18000255/3 JB 2766 <8 273500 <90 71 15 242 <900 75 18480257 JB <2400 14 6018 828 77 <12 42 <900 1530 9710256 JB <2400 19 4968 312 55 13 68 <900 36 1200258 JB <2400 <8 2217 225 33 <12 <15 <900 120 1602141 JB 917 124 60 165 87 <4 29 <300 55 5073156 CJ <2400 <8 18530 <90 98 <12 <15 <900 <30 6190158-a CJ 6990 <8 4176 <90 64 <12 <15 <900 <30 3936158-b CJ 5463 <8 4296 <90 86 13 <15 <900 39 4137160 CJ <2400 14 7734 123 156 15 54 <900 48 66401F CQSW 13557 14 738 <90 258 <12 30 <900 <10 82403F CQSW 12384 145 603 <90 135 <12 <15 <900 <10 106304F CQSW 12492 <8 486 <90 86 <12 <15 <900 33 628075 CK <2400 <8 420 <90 63 <12 66 <900 387 863076 CK <2400 <8 1068 <90 107 <12 <15 <900 <30 7717077 CK 7701 18 11349 <90 115 <12 76 <900 66 4878078 CK 6999 50 8205 <90 29 <12 142 <900 36 707079 CK 8472 54 19520 <90 138 13 139 <900 <30 793080 CK 6528 <8 2757 <90 161 <12 <15 <900 39 645081 CK 11976 28 7569 <90 97 <12 113 <900 <30 515182 CK 11778 <8 41270 153 113 18 381 <900 2640 1984069 CK 5277 13 1644 501 32 <12 34 <900 72 528372 CK 8115 <8 7026 642 57 <12 94 <900 48 662073 CK 6321 <8 1137 132 71 <12 17 <900 147 4503

Appendix B

229


Sample nr. Location Cr Cs Ga Mn Mo Ni Pb Rb Sb Se Sn Sr


271/1 P-FBS <90 <24 <9 <450 <15 <36 2090 <30 206 67 282 <30

271/2 P-FBS <30 <8 <3 255 8 4 2864 <10 58 21 97 23272 P-FBS n.d. n.d. n.d. 7530 n.d. n.d. 3110 n.d. n.d. n.d. n.d. n.d.273 P-FBS 25 <8 87 1732 45 <12 144 <10 104 11 56 26274 P-FBS 84 <8 68 3324 <5 31 87 <10 <7 9 15 12275 P-FBS 106 8 114 <150 <5 <12 4965 <10 22 108 <10 157276/1 P-FBS <30 <8 35 207 34 <12 1066 <10 56 19 <10 44276/2 P-FBS 15 8 65 <150 47 <12 1018 <10 60 33 <10 49230a JLN n.d. n.d. n.d. 2232 n.d. n.d. 4757 n.d. n.d. 443 n.d. n.d.230b JLN n.d. n.d. n.d. n.d. n.d. n.d. 5120 n.d. n.d. 715 n.d. n.d.230c JLN n.d. n.d. n.d. n.d. n.d. n.d. 4487 n.d. n.d. 462 n.d. n.d.235a JLN <90 <24 <9 3792 <15 <36 6174 <30 23 33 <30 <30235b JLN <90 <24 <10 6453 <15 <36 8373 <30 40 44 130 <3024 JB <90 <24 <9 21264 <15 <36 19059 <30 <21 51 <30 <3026 JB <90 <24 <9 9348 <15 <36 14145 <30 77 50 828 <30255/1 JB <90 <24 <9 32763 <15 <36 11595 <30 541 29 <30 <30255/2 JB 141 <24 <9 22269 <15 63 24183 <30 >1200 59 104 <30255/3 JB <90 <24 <9 5877 39 <36 4257 <30 254 16 <30 <30257 JB <90 <24 <9 137950 61 39 4446 33 437 14 <30 111256 JB <90 <24 <9 25860 <15 <36 6492 <30 31 19 31 <30258 JB <90 <24 <9 12876 20 <36 4308 <30 23 17 <30 <30141 JB <30 <8 <3 6575 <5 <12 10835 10 8 34 <10 53156 CJ <90 <24 <9 n.d. <15 <36 9735 <30 506 28 343 36158-a CJ <90 <24 <9 783 <15 <36 3018 <30 <21 19 <30 <30158-b CJ <90 <24 <9 1146 <15 <36 3045 <30 <21 19 <30 <30160 CJ <90 <24 <9 1938 <15 <36 28761 <30 148 134 652 <301F CQSW <90 <24 <12 1995 <15 <36 63960 33 45 189 4739 <303F CQSW <90 <24 <13 1161 <15 <36 19710 <30 <21 70 <30 <304F CQSW <90 <24 <14 5316 <15 <36 7332 <30 <21 31 118 <3075 CK <90 <24 <9 31479 <15 51 22401 <30 77 49 <30 <3076 CK <90 <24 <9 8508 <15 <36 7155 <30 51 23 <30 <3077 CK <90 <24 <9 18354 <15 <36 19458 <30 51 59 <30 <3078 CK <90 <24 <9 891 <15 <36 1530 <30 132 <9 <30 <3079 CK <90 <24 <9 11895 <15 <36 30972 <30 268 80 664 <3080 CK <90 <24 <9 6177 <15 <36 8040 <30 32 26 <30 <3081 CK <90 <24 <9 9912 <15 <36 18042 <30 37 44 <30 <3082 CK <90 <24 <9 8952 <15 <36 2931 <30 886 15 <30 <3069 CK <90 <24 <9 72366 <15 51 2892 <30 <21 <9 94 4572 CK <90 <24 <9 21267 <15 <36 6741 <30 56 20 159 <3073 CK <90 <24 <9 34686 <15 <36 14199 <30 35 33 151 <30

Appendix B

230


Sample nr. Location Th Tl U V W Zn Zr sum l.o.i sum+l.o.i.

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg % % %

271/1 P-FBS <30 <9 <15 <30 <45 6610 <120 - 8.2 -

271/2 P-FBS <10 23 <5 55 <15 7576 <40 90.88 6.5 97.35272 P-FBS n.d. n.d. n.d. n.d. n.d. 3260 n.d. - 1.9 -273 P-FBS <10 11 8 15 <15 3807 46 112.41 7.2 119.57274 P-FBS 13 <3 10 58 <15 1625 137 105.52 8.4 113.96275 P-FBS <10 31 <5 13 <15 709 77 84.87 18.9 103.76276/1 P-FBS <10 256 <5 10 <15 1353 <40 36.89 16.4 53.33276/2 P-FBS <10 244 <5 16 <15 1127 <40 43.55 21.0 64.55230a JLN n.d. n.d. n.d. n.d. n.d. 8950 n.d. - 13.2 -230b JLN n.d. n.d. n.d. n.d. n.d. 7781 n.d. - 12.8 -230c JLN n.d. n.d. n.d. n.d. n.d. 8537 n.d. - 16.0 -235a JLN <30 <9 <15 <30 <45 4395 <120 - 13.9 -235b JLN <30 <9 <15 <30 <45 4107 <120 - 12.0 -24 JB <30 9 <15 45 <45 1749 <120 - 6.7 -26 JB <30 <9 <15 33 <45 11754 <120 - 11.9 -255/1 JB 53 16 17 273 <45 18771 <120 - 11.2 -255/2 JB 41 21 15 306 <45 15894 <120 - - -255/3 JB <30 22 27 366 <45 16290 <120 - - -257 JB 175 <9 18 117 <45 31371 <120 - 15.8 -256 JB <30 <9 <15 93 <45 4410 <120 - 1.2 -258 JB <30 <9 21 162 558 2364 <120 - 3.9 -141 JB <10 <3 <15 <10 <15 19711 <40 91.25 12.6 103.81156 CJ <30 17 <15 <30 <45 642 <120 - 9.0 -158-a CJ <30 <9 <15 69 <45 9783 <120 - 8.5 -158-b CJ <30 <9 <15 60 <45 9900 <120 - 8.8 -160 CJ <30 <9 <15 33 <45 2241 <120 - 12.8 -1F CQSW 47 <9 <15 <30 <45 1596 <120 - 15.2 -3F CQSW <30 <9 <15 69 <45 5388 <120 - 12.0 -4F CQSW <30 <9 <15 48 <45 8439 <120 - 9.5 -75 CK 57 <9 <15 <30 195 134250 <123 - 13.3 -76 CK <30 <9 <15 36 <45 5784 <124 - 7.6 -77 CK <30 16 <15 <30 <45 9537 <125 - 10.1 -78 CK <30 11 <15 <30 <45 5442 <126 - 5.9 -79 CK <30 28 <15 <30 <45 13383 <127 - 8.4 -80 CK <30 <9 <15 33 <45 19629 <128 - 5.6 -81 CK <30 <9 <15 42 <45 21210 <129 - 6.8 -82 CK <30 65 <15 <30 <45 77920 <130 - 8.3 -69 CK 89 <9 <15 42 <45 16812 <120 - 2.6 -72 CK <30 <9 <15 <30 <45 10164 <121 - 13.5 -73 CK 47 <9 <15 <30 <45 4452 <122 - 3.5 -

Appendix B

231

Table B6: Selected sample set from Table B5 showing normal distibution in histograms (see Chap. 4.2.3.2).

Sample open pit Al As Co Cu Mn Pb Sb Se Zn sum mol% mol% mol% mol% mol% mol% mol% mol% mol% mol% 212/5 ROM 5.573 0.706 0.030 1.348 0.818 1.942 0.010 0.011 0.861 11.298160 CJ 4.531 1.304 0.010 1.320 0.450 1.746 0.016 0.022 0.437 9.836210 T >0.25 ROM 18.335 0.932 0.055 1.134 2.328 1.378 0.011 0.011 2.619 26.804210 N >0.25 ROM 17.007 0.982 0.012 1.044 1.471 1.359 0.014 0.011 1.247 23.146212/2 ROM 2.210 1.204 0.057 2.156 3.886 1.275 0.010 0.011 2.027 12.837F3 CQSW 8.858 0.103 <0.001 2.097 0.270 1.238 <0.001 0.011 1.044 13.62124 JB <0.001 0.520 <0.001 1.216 4.721 1.164 <0.001 0.008 0.341 7.970210 T >0.71 ROM 14.823 0.898 0.046 1.001 2.135 1.153 0.010 0.009 2.155 22.23181 CK 3.490 1.277 <0.001 1.027 2.258 1.103 0.004 0.007 3.988 13.153210 N >0.71 ROM 11.883 0.954 0.018 1.009 1.697 1.067 0.011 0.010 1.304 17.95233 FBS 2.641 1.358 0.098 1.705 5.977 0.995 <0.001 0.007 1.442 14.22426 JB <0.001 0.499 0.051 2.806 2.132 0.866 0.008 0.008 2.250 8.620210 T >2 ROM 9.726 0.798 0.035 1.153 2.095 0.811 0.006 0.007 1.924 16.554212/1 ROM 0.574 0.189 0.012 1.028 3.779 0.801 0.009 0.007 0.859 7.260210 N >2 ROM 11.413 0.588 0.058 0.809 1.984 0.729 0.008 0.006 1.584 17.179212/9 ROM 1.776 0.430 0.018 0.707 0.195 0.720 0.005 n.d 1.410 5.261255/1 CQ 6.350 1.856 0.128 2.214 7.093 0.711 0.057 0.005 3.545 21.959141 JB 0.475 0.010 0.012 1.012 1.509 0.665 0.001 0.005 3.716 7.405155a/2 CQB 14.696 1.229 0.045 1.318 0.580 0.623 0.012 0.006 3.259 21.769268b CQ 3.840 0.192 0.007 1.116 1.264 0.508 0.018 0.004 2.212 9.160268c CQ <0.001 0.111 0.008 0.458 0.692 0.502 0.053 0.005 2.358 4.18780 CK 5.550 0.469 0.008 1.283 1.419 0.494 0.003 0.004 3.701 12.93350 FBN 30.348 0.397 0.004 0.910 0.664 0.458 0.002 0.004 1.464 34.251F4 CQSW 2.282 0.083 0.007 1.249 1.224 0.451 <0.001 0.005 1.626 6.92772 CK 1.994 1.186 0.010 1.316 4.722 0.415 0.006 0.003 1.951 11.604155a/1 CJ 8.732 1.035 0.023 1.147 0.286 0.401 0.006 0.004 1.764 13.399257 FBS <0.001 0.842 0.008 0.241 5.684 0.400 0.003 0.003 0.856 8.036F1 CQSW 3.630 0.126 <0.001 1.633 0.463 0.394 0.005 0.031 0.312 6.59214 CQB 5.348 0.628 <0.001 0.835 0.477 0.380 0.024 0.003 1.056 8.751212/7 ROM 0.599 0.351 0.010 0.856 3.622 0.314 0.003 0.006 1.015 6.77713 CQB <0.001 0.569 <0.001 0.730 <0.001 0.302 0.036 0.003 0.722 2.363230c JLN 2.010 0.450 <0.001 n.d. <0.001 0.276 <0.001 0.075 1.644 4.456255/3 JB 2.925 n.d. 0.016 3.589 1.351 0.262 0.027 0.003 3.091 11.265155c CQB 4.600 0.748 0.016 0.816 0.334 0.223 0.008 0.003 1.248 7.996158-b CJ 1.703 0.729 0.008 0.827 0.266 0.188 <0.001 0.003 1.902 5.626158-a CJ 2.925 0.709 <0.001 0.787 0.182 0.186 <0.001 0.003 1.880 6.67269 CK 3.630 0.280 0.016 1.053 14.430 0.178 <0.001 <0.001 3.187 22.775212/10 ROM 0.871 0.581 0.025 0.407 1.982 0.168 0.004 n.d 1.778 5.817212/8 ROM 0.822 0.274 0.007 0.395 2.262 0.153 0.002 0.006 1.081 5.002149-b FBN 5.483 0.701 0.011 0.215 0.248 0.143 0.003 <0.001 3.633 10.437212/4 ROM 0.970 0.525 0.056 1.008 0.286 0.117 0.004 0.004 1.617 4.587205b CQ 24.167 0.271 0.045 2.127 0.209 0.074 0.006 0.016 0.271 27.18551 FBN 12.483 0.014 0.113 1.827 4.703 0.045 0.001 0.001 3.846 23.035263 CQ 6.218 <0.001 <0.001 1.146 <0.001 0.011 0.008 0.014 0.145 7.541273 FBS 23.978 0.610 0.074 0.281 0.402 0.009 0.011 0.002 0.740 26.107274 FBS 35.262 0.081 0.028 0.169 0.769 0.005 <0.001 0.001 0.317 36.633211/1 ROM 14.696 0.777 0.022 0.913 1.439 n.d. 0.010 n.d 1.777 19.634max 35.262 1.856 0.128 3.589 14.430 1.942 0.057 0.075 3.988 36.633min <0.001 <0.001 <0.001 0.169 <0.001 0.005 <0.001 <0.001 0.145 2.363mean 7.222 0.621 0.026 1.162 2.016 0.596 0.009 0.008 1.770 13.379

Appendix B

232

Table B7: XRF results of distal host rocks (weathering profiles and single samples).

Sample nr. SiO2 Al2O3

Al2O3/

SiO2 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 SO3 F

% % % % % % % % % % mg/kg

Copper Jokerweathering profile 161a 44.3 27.81 0.63 10.41 2.48 <0.1 0.13 7.38 0.76 0.03 <0.05 1065161b 44.0 23.89 0.54 8.89 2.15 <0.1 0.10 5.42 0.67 0.05 0.02 1008162 42.8 22.25 0.52 7.69 2.05 <0.1 <0.1 3.64 0.76 0.02 <0.05 887159 45.8 18.86 0.41 8.12 2.31 0.78 <0.1 2.67 0.62 0.06 0.02 1158"Core-stone" outcrop Copper Queen 181a 59.4 13.8 0.23 3.54 1.63 2.64 0.75 2.86 0.34 0.13 0.02 1008181b 49.8 19.51 0.39 2.99 1.17 <0.1 0.12 4.11 0.37 0.10 <0.05 <800181d 46.9 20.63 0.44 3.46 1.10 <0.1 0.10 3.80 0.38 0.13 <0.05 661181e 54.9 19.66 0.36 3.11 <0.3 <0.1 0.12 3.74 0.37 0.09 <0.05 773"Core-stone" outcrop J-Body 246a 51.9 13.87 0.27 6.09 2.89 2.35 0.88 2.82 0.57 0.09 <0.05 1016246b 48.6 16.61 0.34 6.34 2.76 0.99 0.50 3.91 0.62 0.09 <0.05 1076247 40.0 6.86 0.17 6.85 1.88 0.75 <0.1 3.23 0.62 0.02 <0.05 831248 47.7 18.99 0.40 6.65 2.55 0.52 0.12 3.43 0.58 0.01 <0.05 922J-Body open pit 249 40.8 16.95 0.42 9.19 3.50 0.34 2.96 1.62 0.53 0.03 0.28 <800250 40.5 17.81 0.44 7.22 3.15 0.97 1.11 4.05 0.5 0.08 <0.05 1042251a 45.0 24.28 0.54 9.32 3.54 <0.1 0.20 5.12 0.57 0.03 0.07 1178252 39.3 6.86 0.17 8.88 2.28 0.17 0.18 5.66 0.67 0.03 0.41 881253 53.9 19.04 0.35 3.67 2.81 <0.1 0.16 5.11 0.4 <0.01 <0.05 1361254b 34.4 6.86 0.20 7.44 5.91 0.29 0.31 4.00 0.47 0.02 3.97 <800254a 43.4 6.86 0.16 9.32 3.74 0.62 0.32 5.88 0.55 0.02 0.38 845Dirt road outcrop 168 50.6 18.49 0.37 6.50 2.92 <0.1 0.13 3.82 0.54 0.03 <0.05 1301167a 45.8 19.02 0.42 7.52 3.84 <0.1 <0.1 3.11 0.61 0.02 <0.05 1182Copper Queen open pit 201 0.4 <0.2 0.50 5.02 13.77 24.68 <0.1 <0.05 <0.03 0.06 0.04 <800195 2.5 0.96 0.38 8.87 14.39 25.71 <0.1 <0.05 0.05 0.03 0.11 <8005 48.3 22.74 0.47 7.36 1.4 2.5 1.69 2.65 0.48 0.03 <0.05 <800West of Copper Queen open pit 169 47.5 19.14 0.40 7.69 2.39 0.19 0.19 3.69 0.68 0.06 0.02 1548Copper Queen Dome 281 42.8 13.07 0.31 4.2 1.82 5.95 0.55 2.13 0.51 0.08 0.06 <800277cd1 86.0 7.03 0.08 1.18 0.44 0.79 0.45 0.5 0.15 <0.05 <0.05 <800277cd2 89.3 3.38 0.04 0.53 <0.1 0.98 0.21 0.61 0.07 <0.05 <0.05 <800280cd 58.7 15.38 0.26 5.88 5.18 1.26 0.7 0.21 0.5 <0.05 <0.05 <800282cd1 53.4 13.04 0.24 14.39 4.4 8.58 2.3 0.33 1.25 0.138 <0.05 <800282cd2<2 57.1 13.36 0.23 11.41 2.56 3.48 1.43 1.88 0.98 <0.05 <0.05 <800282cd2<1 52.8 13.9 0.26 11.65 2.67 3.12 1.44 1.78 1.04 0.054 <0.05 <800283cd1 56.7 16.89 0.30 5.9 5.24 1.63 1.14 0.28 0.54 0.073 <0.05 <800283cd2<2 55.4 16.24 0.29 6.93 3.26 0.97 <0.2 1 0.66 <0.05 <0.05 <800283cd2<1 52.3 16.76 0.32 7.35 3.05 1.13 0.25 1.23 0.74 0.064 <0.05 <800279cd1 74.2 13.03 0.18 0.82 n.d. 0.12 2.99 5.01 0.07 <0.05 <0.05 <800279cd2 76.2 12.34 0.16 0.47 n.d. 0.13 4.33 3.31 0.04 <0.05 <0.05 <800278cd 45.6 14.05 0.31 16.32 6.63 10.19 1.89 0.55 1.49 0.15 <0.05 <800

Appendix B

233

Table B7 continued: XRF results of distal host rocks (weathering profiles and single samples).

Sample nr. Ag As Ba Bi Br Cd Cl Co Cu Cr Cs Ga


Copper Jokerweathering profile 161a 3 589 1697 <7 <4 <5 <300 26 2191 254 63 32161b <3 609 1188 <7 <4 <5 <300 25 2384 187 48 25162 <3 477 746 <7 <4 5 <300 20 1880 165 49 19159 <3 23 587 <7 4 <5 <300 30 80 166 <8 18"Core-stone" outcrop Copper Queen 181a <3 153 478 <7 <4 <5 <300 74 762 55 <8 15181b <3 116 859 <7 <4 <5 <300 <10 1211 47 13 18181d <3 288 822 <7 <4 <5 <300 11 605 43 12 16181e <3 202 786 <7 <4 <5 <300 <10 527 57 11 14"Core-stone" outcrop J-Body 246a <3 10 687 <7 <4 <5 <300 71 35 135 <8 17246b 3 30 1109 <7 <4 <5 <300 51 43 158 14 21247 <3 39 704 <7 <4 <5 <300 26 174 149 17 17248 <3 22 816 <7 <4 <5 <300 22 47 148 17 14J-Body open pit 249 <3 119 728 <7 <4 <5 <300 33 5098 190 56 16250 <3 <10 668 <7 <4 <5 <300 37 11 166 30 21251a <3 62 1225 <7 <4 <5 <300 28 19 216 13 31252 <3 125 1262 <7 <4 <5 <300 27 78 226 31 35253 <3 13 1319 <7 <4 <5 <300 10 62 55 11 19254b <3 92 1130 <7 <4 <5 <300 77 494 166 16 22254a <3 14 1657 <7 <4 <5 <300 27 303 254 16 33Dirt road outcrop 168 <3 99 690 <7 <4 <5 <300 15 268 156 9 16167a <3 31 471 <7 <4 <5 <300 14 89 196 6 18Copper Queen open pit 201 <3 <10 <30 <7 <4 <5 <300 <10 83 <30 <8 <3195 <3 <10 <30 <7 <4 <5 <300 <10 82 <30 <8 175 30 1927 664 83 <4 <5 <300 128 2782 76 29 24West of Copper Queen open pit 169 <3 91 755 <7 <4 <5 <300 29 334 158 13 19Copper Queen Dome 281 <3 <10 555 <7 7 <5 <300 17 33 89 <8 14277cd1 <3 <10 191 <10 <40 <5 <300 n.d. 24 32 <8 10277cd2 <3 <10 351 <10 <40 <5 <300 n.d. <20 <20 <8 <8280cd <3 <10 <50 13 102 <5 <300 20 <20 138 <8 22282cd1 <3 <10 82 26 <40 <5 <300 46 88 <20 <8 18282cd2<2 <3 <10 354 17 48 <5 <300 57 77 35 <8 21282cd2<1 <3 <10 331 16 42 <5 <300 49 89 30 <8 16283cd1 <3 <10 62 <10 50 <5 <300 25 <20 143 <8 25283cd2<2 <3 <10 225 <10 89 <5 <300 28 39 152 <8 25283cd2<1 <3 <10 319 <10 92 <5 <300 31 46 157 <8 25279cd1 <3 <10 384 <10 <40 <5 <300 <20 <20 <20 <8 15279cd2 <3 <10 185 13 <40 <5 <300 <20 31 <20 <8 24278cd <3 <10 70 19 43 <5 <300 51 <20 332 <8 20

Appendix B

234


Sample nr. Mn Mo Ni Pb Rb Sb Se Sn Sr Th Tl U V

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

Copper Jokerweathering profile 161a 1138 <5 104 74 426 27 <3 11 21 17 7 <5 130161b 1337 <5 83 123 298 26 <3 <10 22 17 4 <5 106162 1118 <5 80 63 253 <7 <3 <10 11 15 <3 <5 86159 1184 <5 65 44 119 <7 <3 <10 33 14 <3 <5 104"Core-stone" outcrop Copper Queen 181a 2209 <5 47 90 213 <7 <3 <10 156 16 <3 <5 32181b 591 <5 105 50 239 <7 <3 <10 92 11 <3 <5 36181d 1746 <5 49 153 358 <7 <3 <10 137 20 <3 <5 40181e 1287 <5 45 126 355 <7 <3 <10 103 14 <3 5 32"Core-stone" outcrop J-Body 246a 1005 <5 52 13 122 <7 <3 <10 95 13 <3 <5 100246b 1001 <5 57 16 176 <7 <3 <10 60 12 <3 <5 104247 938 <5 52 57 168 <7 <3 <10 29 13 <3 <5 103248 903 <5 55 24 173 <7 <3 <10 26 11 <3 <5 93J-Body open pit 249 1413 <5 86 48 172 <7 <3 66 54 14 <3 <5 72250 1071 <5 66 31 198 <7 <3 <10 80 13 <3 <5 96251a 1319 <5 90 27 172 <7 <3 <10 21 22 <3 <5 148252 1021 <5 70 51 242 <7 <3 <10 27 16 <3 6 137253 710 <5 25 35 224 <7 <3 13 16 16 <3 <5 40254b 1597 <5 71 178 178 <7 <3 <10 30 14 <3 <5 106254a 1252 <5 95 38 214 <7 <3 <10 45 17 <3 <5 161Dirt road outcrop 168 717 <5 79 24 128 <7 <3 <10 17 <10 <3 <5 119167a 627 <5 65 22 154 <7 <3 <10 46 10 <3 <5 119Copper Queen open pit 201 5858 <5 <12 215 <10 11 <3 <10 285 <10 <3 <5 <5195 4753 <5 <12 787 <10 <7 <3 <10 271 <10 <3 <5 145 2388 <5 70 673 209 62 <3 14 530 20 8 6 57West of Copper Queen open pit 169 1310 <5 73 187 191 <7 <3 <10 32 14 <3 <5 98Copper Queen Dome 281 823 <5 39 35 97 <7 <3 <10 233 10.5 <3 <5 75277cd1 <150 <7 <20 <20 <30 <15 <6 <10 89 <20 <3 <9 25277cd2 256 <7 <20 <20 <30 <15 <6 <10 29 <20 <3 <9 11280cd 542 <7 97 <20 <30 <15 <6 <10 98 <20 <3 <9 76282cd1 1642 <7 26 <20 31 38 <6 <10 126 <20 <3 <9 350282cd2<2 1681 <7 41 24 130 29 <6 <10 131 <20 <3 <9 263282cd2<1 1387 <7 42 26 143 39 <6 <10 138 <20 <3 <9 264283cd1 488 <7 121 <20 <30 <15 <6 <10 134 <20 <3 <9 73283cd2<2 666 <7 105 24 62 15 <6 <10 82 <20 <3 <9 155283cd2<1 782 <7 102 27 77 20 6 <10 85 <20 <3 <9 172279cd1 <150 <7 <20 37 226 <15 <6 <10 75 38 <3 <9 <10279cd2 <150 <7 <20 37 142 <15 <6 <10 39 22 <3 21 13278cd 2107 <7 84 22 58 45 <6 <10 159 <20 <3 <9 346

Appendix B

235


Sample nr. W Zn Zr sum l.o.i. sum+l.o.i

mg/kg mg/kg mg/kg % % %

Copper Jokerweathering profile 161a 15 4926 158 94.6 6.3 100.9161b <15 3736 129 86.3 5.5 91.9162 <15 2386 183 80.2 5.6 85.8159 <15 255 159 79.7 8.3 88.0"Core-stone" outcrop Copper Queen 181a 586 888 85 85.8 2.0 87.8181b <15 1609 83 78.7 3.6 82.4181d <15 2335 89 77.3 4.3 81.7181e <15 2346 84 83.9 4.0 87.9"Core-stone" outcrop J-Body 246a 389 73 151 81.9 2.3 84.1246b 255 78 161 80.9 3.5 84.4247 <15 225 155 72.5 5.6 78.1248 <15 315 160 80.9 5.6 86.6J-Body open pit

249 <15 8150 120 77.8 5.8 83.6250 99 276 109 75.8 2.7 78.4251a <15 1429 139 88.7 - -252 <15 1913 142 82.6 7.3 89.9253 <15 765 89 85.6 3.3 88.8254b <15 2469 105 76.0 13.0 89.1254a <15 2031 146 91.4 5.9 97.3Dirt road outcrop 168 <15 888 138 83.6 4.8 88.3167a <15 409 161 80.4 7.7 88.1Copper Queen open pit 201 18 1385 <40 44.8 44.4 89.2195 <15 1406 <40 53.4 41.6 95.05 <15 2266 125 88.4 4.8 93.2West of Copper Queen open pit 169 <15 491 129 82.1 6.5 88.6Copper Queen dome 281 <15 69 191 71.4 - 71.4277cd1 <10 <20 67 98.1 1.7 99.9277cd2 <10 <20 <50 96.5 1.5 98.0280cd 17 99 131 96.2 8.5 104.7282cd1 24 109 111 99.0 1.1 100.1282cd2<2 21 86 108 99.5 2.7 102.2282cd2<1 28 92 127 97.5 0.8 98.3283cd1 19 98 152 95.2 7.0 102.2283cd2<2 21 82 126 95.9 1.3 97.2283cd2<1 22 82 126 96.3 1.3 97.6279cd1 <10 <20 68 96.6 0.8 97.4279cd2 35 <20 <50 96.8 0.4 97.2278cd 27 140 106 98.1 1.0 99.1

Appendix B

236

Table B8: XRF results of ROM: 2 samples < 3 cm, sieved sample and 10 lump ores.

Location/ Outcrop Sample type SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 SO3

ROM % % % % % % % % % %

210/1 <3 cm 24.9 6.85 27.56 3.61 0.87 0.10 0.84 0.18 0.14 2.62

210 N >2 sieved sample n.b. 5.13 35.31 3.63 0.90 <0.3 0.84 0.18 0.25 1.71 210 N >0.71 sieved sample n.b. 5.37 42.72 4.26 0.69 <0.3 0.60 0.18 0.26 1.88 210 N >0.25 sieved sample n.b. 8.16 39.96 3.93 0.60 <0.3 1.29 0.27 0.28 2.13 210 N >0.063 sieved sample n.b. 8.97 28.56 4.11 0.63 <0.3 1.47 0.42 0.23 1.67 210 N <0.063 sieved sample n.b. 12.51 28.62 4.44 0.57 <0.3 1.65 0.33 0.29 3.42

210 T >2 sieved sample n.b. 4.29 39.12 3.75 0.60 <0.3 0.54 0.15 0.20 2.77 210 T >0.71 sieved sample n.b. 6.93 34.83 4.17 0.66 <0.3 0.87 0.21 0.25 3.85 210 T >0.25 sieved sample n.b. 8.94 32.85 3.99 0.54 <0.3 1.26 0.27 0.27 5.22 210 T >0.063 sieved sample n.b. 8.19 26.25 3.75 0.54 <0.3 1.14 0.33 0.25 5.57 210 T <0.063 sieved sample n.b. 8.82 26.88 3.63 0.54 <0.3 1.35 0.27 0.24 5.44

211/1 < 3cm 26.5 6.86 31.37 4.63 0.68 <0.1 0.80 0.16 0.15 1.62

212/1 lump ores 16.0 0.23 68.12 3.88 0.14 <0.01 0.02 0.01 0.15 0.08 212/2 lump ores 3.2 0.9 75.72 0.08 0.02 <0.01 0.01 0.07 0.23 <0.01 212/3 lump ores 18.8 13.4 28.31 12.57 0.04 <0.01 0.01 0.39 0.21 0.04 212/4 lump ores 1.7 0.39 87.02 0.07 0.02 <0.01 0.02 0.05 0.07 0.09 212/5 lump ores 3.3 2.35 80.26 1.08 0.02 <0.01 0.02 0.11 0.27 0.02 212/6 lump ores 28.5 0.19 39.17 1.26 0.05 <0.01 0.01 0.01 3.27 <0.01 212/7 lump ores 27.0 0.24 59.03 4.42 0.27 <0.01 0.02 0.03 0.07 0.03 212/8 lump ores 19.8 0.33 64.97 5.9 0.21 <0.01 0.02 0.01 0.1 0.04 212/9 lump ores 2.3 0.72 84.45 0.19 0.01 <0.01 0.02 0.01 0.11 0.09 212/10 lump ores 13.1 0.35 74.88 2.32 0.06 <0.01 0.02 0.02 0.3 0.03

F Ag As Ba Bi Br Cd Cl Co Cu

ROM Sample type mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

210/1 < 3 cm <800 12 3998 309 126 <4 37 <299 191 4372

210 N >2 sieved sample <2400 21 3462 285 80 <12 <15 <900 267 4047

210 N >0.71 sieved sample 4503 12 5637 213 143 <12 32 <900 81 5061

210 N >0.25 sieved sample 3309 17 5802 372 166 14 34 <900 57 5238

210 N >0.063 sieved sample <2400 15 4227 456 110 <12 25 <900 60 4089

210 N <0.063 sieved sample <2400 38 5910 612 151 15 21 <900 102 6540

210 T >2 sieved sample <2400 18 4707 195 97 <12 23 <900 162 5790

210 T >0.71 sieved sample <2400 13 5301 324 123 <12 32 <900 213 5019

210 T >0.25 sieved sample <2400 16 5508 429 158 <12 34 <900 252 5694

210 T >0.063 sieved sample <2400 29 4614 414 121 13 37 <900 279 5106

210 T <0.063 sieved sample <2400 33 5454 549 143 <12 40 <900 342 5910

211/1 < 3cm 1571 12 4583 315 126 <4 31 <300 102 4575

212/1 lump ores <0.05 <3 1109 6 17 <4 9 90 55 5158

212/2 lump ores <0.05 21 7133 37 36 5 45 90 263 10939

212/3 lump ores <0.05 10 37383 76 59 11 27 130 906 58243

212/4 lump ores <0.05 <3 3087 8 12 14 16 60 259 5055

212/5 lump ores <0.05 4 4158 11 56 <4 11 110 138 6784

212/6 lump ores <0.05 20 3215 33 35 <4 81 40 710 61919

212/7 lump ores <0.05 8 2061 16 3 7 19 60 48 4284

212/8 lump ores <0.05 <3 1606 10 <3 8 7 40 31 1967

212/9 lump ores <0.05 <3 2526 8 21 14 28 80 84 3535

212/10 lump ores <0.05 6 3420 22 39 9 102 40 117 2030

Grey marking highlights samples with >30% Fe2O3. T= dry sieving N=wet sieving

Appendix B

237

Table B8 continued: XRF results of ROM: 2 samples < 3 cm, sieved sample and 10 lump ores.

Location/ Outcrop Sample type Cr Cs Ga Mn Mo Ni Pb Rb Sb Se Sn

ROM mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

210/1 <3 cm 42 <8 <3 7983 <5 31 14641> 66 81 48> <10

210 N >2 sieved sample <90 25 <9 8685 <15 <36 11889 45 77 38 <30 210 N >0.71 sieved sample <90 <24 <9 7407 <15 <36 17448 36 102 60 117 210 N >0.25 sieved sample <90 <24 <9 6405 <15 <36 22290 90 128 68 409 210 N >0.063 sieved sample <90 <24 <9 4731 <15 <36 15828 93 93 47 555 210 N <0.063 sieved sample <90 <24 <9 4401 <15 <36 25683 78 96 71 <30

210 T >2 sieved sample <90 <24 <9 9183 <15 <36 13227 33 59 43 <30 210 T >0.71 sieved sample <90 <24 <9 9360 <15 <36 18879 45 94 57 <30 210 T >0.25 sieved sample <90 <24 <9 10230 <15 <36 22620 75 107 66 161 210 T >0.063 sieved sample <90 <24 <9 10296 <15 <36 19260 75 92 54 196 210 T <0.063 sieved sample <90 <24 <9 9543 <15 <36 23472 66 93 65 <30

211/1 < 3cm 51 <8 <3 6265 <5 20 13762> 56 96 45> 76

212/1 lump ore 35 <5 4 16855 <2 <3 13076 6 83 46 314 212/2 lump ore 30 <5 10 17351 95 <3 20908 <2 98 67 295 212/3 lump ore 45 <5 12 6576 3 19 840 <2 30 9 20 212/4 lump ore 8 <5 31 1232 <2 11 1892 4 36 24 342 212/5 lump ore 55 <5 41 3540 49 <3 32054 <2 91 66 404 212/6 lump ore <3 <5 8 14369 <2 3 1409 8 47 15 15 212/7 lump ore 3 <5 42 16127 <2 <3 5098 3 26 40 107 212/8 lump ore 18 <5 10 9930 <2 <3 2483 <2 23 40 78 212/9 lump ore 5 <5 5 837 8 <3 11737 <2 46 112> 1204 212/10 lump ore 7 <5 21 8675 13 <3 2723 6 42 48> 73

Sr Th Tl U V W Zn Zr sum l.o.i sum+l.o.i.

ROM Sample type mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg % %

210/1 total sample 15 14 7 <5 46 <15 11599 57 - - -

210 N >2 sieved sample <30 <30 <9 <15 48 <45 8220 <120 - - -

210 N >0.71 sieved sample <30 <30 <9 <15 54 <45 6747 <120 - - -

210 N >0.25 sieved sample <30 <30 15 <15 69 <45 6447 <120 - - -

210 N >0.063 sieved sample <30 <30 10 <15 69 <45 5367 <120 - - -

210 N <0.063 sieved sample <30 <30 14 <15 90 <45 7044 <120 - - -

210 T >2 sieved sample <30 <30 10 <15 63 <45 10017 <120 - - -

210 T >0.71 sieved sample <30 <30 10 <15 78 <45 11250 <120 - - -

210 T >0.25 sieved sample <30 <30 16 <15 75 <45 13737 <120 - - -

210 T >0.063 sieved sample <30 <30 11 <15 63 <45 13776 <120 - - -

210 T <0.063 sieved sample <30 <30 16 <15 51 <45 15474 <120 - - -

211/1 < 3cm <10 <10 8 <5 53 <15 9237 54 - - -

212/1 lump ore 5 122 <3 9 149 <5 4426 <3 90.82 6.33 97.15

212/2 lump ore <2 200 10 3 106 <5 10564 <3 82.49 11.51 94

212/3 lump ore <2 29 33 14 43 <5 15890 77 74.67 10.86 85.53

212/4 lump ore <2 36 <3 16 35 <5 8394 3 89.62 8.03 97.65

212/5 lump ore <2 308 4 15 41 <5 4436 <3 87.87 6.63 94.5

212/6 lump ore <2 28 11 13 13 <5 44297 <3 74.33 11.69 86.02

212/7 lump ore <2 47 6 <3 46 <5 5238 <3 93.23 4.70 97.93

212/8 lump ore <2 33 4 5 31 <5 5583 <3 92.70 5.78 98.48

212/9 lump ore <2 112 <3 13 16 <5 7306 <3 88.02 8.96 96.98

212/10 lump ore 4 46 4 12 36 <5 9245 <3 92.20 5.57 97.77

Appendix B

238

Table B9: XRF results of LPO from profiles A2-2 and A2-4.

N/T KG SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 SO3

A2-2-0 mm % % % % % % % % % % 97 N >2 39.4 11.46 26.40 3.90 0.16 <0.1 0.13 0.19 0.10 0.62 97 N >0.71 37.7 13.47 23.82 3.33 0.16 <0.1 0.17 0.26 0.12 0.71 97 N >0.25 46.7 16.92 18.66 3.30 0.13 <0.1 0.23 0.59 0.12 1.06 97 N >0.063 50.4 10.53 8.31 3.42 <0.1 <0.1 0.10 0.38 0.06 0.72 97 N <0.063 48.2 12.39 9.39 4.21 <0.1 <0.1 0.07 0.21 0.06 0.38 97 T >2 33.8 15.44 16.02 3.32 <0.1 <0.1 0.12 0.30 0.07 1.76 97 T >0.71 32.4 17.86 15.01 3.41 <0.1 <0.1 0.12 0.35 0.07 2.01 97 T >0.25 36.9 17.10 14.99 3.76 0.11 <0.1 0.14 0.38 0.08 1.98 97 T >0.063 42.3 12.85 10.78 3.73 0.14 <0.1 0.10 0.25 0.07 1.45 97 T <0.063 38.3 14.95 16.76 5.06 0.19 <0.1 0.14 0.30 0.08 1.61

A2-2-100 95 N >2 33.5 5.01 34.83 10.35 1.19 <0.1 0.11 0.10 0.15 1.27 95 N >0.71 28.3 5.88 34.83 8.22 1.09 <0.1 0.20 0.11 0.15 1.27 95 N >0.25 30.1 7.14 31.74 7.23 0.84 <0.1 0.55 0.22 0.18 1.67 95 N >0.063 41.3 7.59 21.90 7.59 1.38 <0.1 0.54 0.23 0.13 1.05 95 N <0.063 44.9 6.90 12.45 3.17 0.58 <0.1 0.47 0.14 0.15 0.67 95 T >2 30.5 6.55 27.63 5.68 1.07 <0.1 0.38 0.14 0.13 2.18 95 T >0.71 28.8 7.29 25.45 5.19 0.83 <0.1 0.41 0.18 0.15 2.78 95 T >0.25 30.8 7.45 24.09 5.16 0.84 0.12 0.60 0.22 0.17 2.80 95 T >0.063 35.8 6.53 20.75 5.19 1.17 0.11 0.56 0.20 0.15 2.33 95 T <0.063 32.6 7.94 22.37 5.25 1.16 <0.1 0.72 0.18 0.16 2.32

A2-2-200 91 N >2 33.7 8.79 28.26 5.49 0.88 <0.1 0.20 0.17 0.13 1.33 91 N >0.71 32.6 9.96 30.51 5.37 0.52 <0.1 0.29 0.21 0.14 1.12 91 N >0.25 32.2 10.08 28.71 4.56 0.46 <0.1 0.48 0.33 0.15 1.59 91 N >0.063 45.2 9.99 17.55 4.95 0.61 <0.1 0.46 0.33 0.10 0.82 91 N <0.063 46.8 9.90 12.78 3.78 0.27 <0.1 0.45 0.20 0.18 0.41 91 T >2 26.2 10.49 20.78 3.71 0.40 <0.1 0.22 0.23 0.09 1.92 91 T >0.71 23.3 10.27 19.43 3.46 0.34 <0.1 0.38 0.23 0.09 2.47 91 T >0.25 27.4 9.78 20.08 3.66 0.40 <0.1 0.54 0.26 0.11 2.35 91 T >0.063 33.7 8.70 16.24 3.77 0.53 0.12 0.52 0.24 0.10 1.88 91 T <0.063 36.1 11.26 19.52 4.61 0.60 <0.1 0.76 0.27 0.11 2.05

* no further dilution because of lack of material. N=wet T=dry

sample nr. SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 SO3

A2-4 % % % % % % % % % % 113< 1cm N < 1cm n.d. 7.86 26.28 2.76 0.63 <0.3 0.51 0.24 0.22 1.24 114< 1cm N < 1cm n.d. 7.41 28.65 1.29 <0.3 <0.3 0.78 0.30 0.19 4.88 115< 1cm N < 1cm n.d. 5.67 25.41 1.23 <0.3 <0.3 0.63 0.27 0.20 3.50 116< 1cm N < 1cm n.d. 5.13 28.59 1.05 <0.3 <0.3 0.54 0.30 0.23 4.24 113> 1cm N > 1cm n.d. 2.94 29.52 4.59 3.63 <0.3 <0.15 <0.09 0.51 1.04 114> 1cm N > 1cm n.d. 9.39 34.17 1.50 n.d. <0.3 1.44 0.30 0.26 1.62 115> 1cm N > 1cm n.d. 6.51 31.20 3.18 0.54 <0.3 0.36 0.27 0.41 1.92 116> 1cm N > 1cm n.d. 4.53 33.48 2.07 0.60 <0.3 0.63 0.21 0.36 1.72

Appendix B

239

Table B9 continued: XRF results of LPO from profiles A2-2 and A2-4.

N/T KG F Ag As Ba Bi Br Cd Cl Co A2-2-0 mm mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

97 N >2 2248 10 12309 42 101 7 112 <300 48 97 N >0.71 1933 12 12240 40 127 6 82 <300 42 97 N >0.25 1911 7 10347 53 123 5 54 <300 31 97 N >0.063 1345 5 3090 <30 48 <4 15 <300 10 97 N <0.063 <800 4 2520 <30 37 <4 7 <300 10 97 T >2 972 9 >8391 31 80 5 55 <300 85 97 T >0.71 <800 8 >6833 30 84 <4 45 <300 92 97 T >0.25 <800 10 >6788 43 89 <4 37 <300 98 97 T >0.063 990 9 >4261 33 59 <4 22 <300 65 97 T <0.063 1289 13 >6740 51 100 <4 34 <300 89

A2-2-100 95 N >2 3753 9 12960 65 218 8 67 <300 105 95 N >0.71 3980 6 11988 79 207 <4 81 <300 60 95 N >0.25 3707 10 11355 143 217 <4 51 <300 51 95 N >0.063 3140 8 7257 149 139 <4 29 <300 33 95 N <0.063 <800 9 4348 219 >81 <4 11 <300 13 95 T >2 2506 5 >8748 167 147 7 40 <300 102 95 T >0.71 2488 16 >9731 173 183 5 50 <300 112 95 T >0.25 2709 17 >9969 230 197 6 41 <300 117 95 T >0.063 2723 19 >8465 222 163 6 38 <300 94 95 T <0.063 2417 22 >9646 375 182 5 35 <300 79

A2-2-200 91 N >2 2409 11 12027 63 133 9 58 <300 78 91 N >0.71 2962 13 13473 92 194 6 62 <300 54 91 N >0.25 2898 10 13185 123 188 5 48 <300 45 91 N >0.063 2298 9 7011 118 105 5 30 <300 21 91 N <0.063 <800 8 4540 183 64 4 14 <300 17 91 T >2 <800 9 >8824 80 107 5 48 <300 105 91 T >0.71 <800 14 >9205 155 130 <4 46 <300 115 91 T >0.25 1195 15 >9360 206 139 5 43 <300 104 91 T >0.063 1368 14 >7135 184 104 5 29 <300 81 91 T <0.063 1862 15 >8411 353 132 8 33 <300 90

sample nr. F Ag As Ba Bi Br Cd Cl Co A2-4 mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

113< 1cm N < 1cm 4761 16 10983 294 79 <12 31 <900 177 114< 1cm N < 1cm <2400 37 13200 180 159 <12 41 <900 111 115< 1cm N < 1cm <2400 31 13410 171 205 13 24 <900 84 116< 1cm N < 1cm <2400 34 11570 192 223 16 25 <900 117 113> 1cm N > 1cm <2400 20 10290 93 107 12 56 <900 255 114> 1cm N > 1cm <2400 31 15310 261 104 14 31 <900 153 115> 1cm N > 1cm <2400 14 7590 93 168 <12 18 <900 144 116> 1cm N > 1cm <2400 48 9318 183 158 <12 28 <900 378

Appendix B

240


N/T KG Cu Cr Cs Ga Mn Mo Ni Pb Rb A2-2-0 mm mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

97 N >2 6008> <30 <8 10 1101 12 <12 3903 <10 97 N >0.71 5716> <30 <8 6 1254 13 <12 5907 12 97 N >0.25 4749 <30 <8 14 855 13 <12 5760 19 97 N >0.063 1302 <30 <8 11 471 <5 <12 2274 <10 97 N <0.063 1052 <30 <8 18 623 6 <12 1517 <10 97 T >2 3438 31 <8 18 2269 7 <12 3918 <10 97 T >0.71 3026 37 <8 22 2542 7 <12 3867 <10 97 T >0.25 2997 39 <8 21 2611 7 <12 4405 <10 97 T >0.063 1826 <30 <8 13 1808 <5 <12 3051 <10 97 T <0.063 2795 36 <8 18 2503 <5 <12 4984 12

A2-2-100 95 N >2 3888 <30 <8 <3 4149 <5 <12 20748 18 95 N >0.71 3975 <30 <8 <3 2760 <5 <12 21408 22 95 N >0.25 3669 <30 8 <3 2592 <5 24 22440 65 95 N >0.063 2229 <30 8 <3 2193 <5 12 13881 55 95 N <0.063 1304 <30 <8 <3 1062 <5 <12 7507 22 95 T >2 3106 <30 <8 <3 4200 <5 <12 14411> 27 95 T >0.71 3105 <30 <8 <3 3828 <5 19 17438> 40 95 T >0.25 3014 <30 <8 <3 3729 <5 19 19150> 59 95 T >0.063 2499 30 <8 <3 3251 <5 20 16141> 52 95 T <0.063 2779 39 <8 <3 3345 <5 17 18345> 47

A2-2-200 91 N >2 5650 <30 <8 <3 5451 <5 <12 12810 16 91 N >0.71 4521 31 <8 <3 3255 <5 <12 11970 26 91 N >0.25 4362 33 <8 <3 2553 <5 <12 13938 47 91 N >0.063 2289 33 <8 3 1599 <5 <12 8064 41 91 N <0.063 1482 <30 <8 8 1049 <5 <12 4824 21 91 T >2 3241 <30 <30 6 3918 <5 <12 7314> 19 91 T >0.71 3003 <30 <30 5 3458 <5 <12 9016> 29 91 T >0.25 2944 37 <8 <3 3439 <5 <12 10483> 44 91 T >0.063 2241 32 <8 <3 2738 <5 15 8281> 39 91 T <0.063 2604 45 <8 4 3124 <5 15 10197> 46

sample nr. Cu Cr Cs Ga Mn Mo Ni Pb Rb A2-4 mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

113< 1cm N < 1cm 3969 102 <24 <9 7617 <15 <36 8613 33 114< 1cm N < 1cm 4881 <90 <24 <9 2154 <15 <36 9414 36 115< 1cm N < 1cm 5478 <90 <24 <9 1947 <15 <36 10032 33 116< 1cm N < 1cm 4710 <90 <24 <9 3084 <15 <36 11679 <30 113> 1cm N > 1cm 6200 <90 <24 <9 11694 <15 <36 10224 <30 114> 1cm N > 1cm 6030 <90 <24 12 4995 <15 <36 4380 57 115> 1cm N > 1cm 3819 <90 <24 <9 6627 <15 <36 5487 <30 116> 1cm N > 1cm 3890 <90 <24 <9 13527 <15 <36 6312 30

Appendix B

241


N/T KG Sr Th Tl U V W Zn Zr sum l.o.i sum+l.o.i. A2-2-0 mm mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg % % %

97 N >2 <10 <10 16 14 70 <15 5061 51 82.3 9.39 91.7 97 N >0.71 14 <10 18 11 95 <15 4548 58 78.6 9.95 88.5 97 N >0.25 15 10 17 17 109 <15 3642 78 84.5 7.57 92.1 97 N >0.063 <10 <10 5 8 43 <15 1257 46 73.0 4.54 77.6 97 N <0.063 <10 <10 6 5 48 <15 1097 59 75.7 - - 97 T >2 <10 <10 15 10 76 <15 6837 75 73.5 12 85.5 97 T >0.71 11 11 11 10 85 <15 7151 88 73.8 12.3 86.1 97 T >0.25 12 14 13 12 86 <15 7316 88 78.0 12.2 90.2 97 T >0.063 <10 <10 6 7 56 <15 4608 62 73.4 8.3 81.7 97 T <0.063 <10 <10 11 10 83 <15 6594 89 80.0 11.9 91.9

A2-2-100 95 N >2 <10 12 16 <5 23 <15 10239 39 82.0 9.25 91.3 95 N >0.71 <10 15 16 <5 79 <15 8655 39 76.5 10.3 86.8 95 N >0.25 <10 <10 19 <5 38 <15 8268 49 76.7 10.6 87.3 95 N >0.063 <10 11 12 <5 33 <15 5214 45 79.7 7.47 87.2 95 N <0.063 <10 <10 6 <5 29 <15 2820 <40 71.2 - - 95 T >2 <10 <10 15 <5 34 <15 11438 48 78.8 12.1 91 95 T >0.71 <10 14 15 <5 38 <15 11655 57 76.1 13.9 90 95 T >0.25 <10 20 16 <5 38 <15 11825 60 77.3 13.8 91.1 95 T >0.063 <10 13 14 5 35 <15 9467 57 77.1 11.7 88.8 95 T <0.063 <10 12 15 <5 43 <15 9440 64 77.5 - -

A2-2-200 91 N >2 <10 <10 18 <5 45 <15 7056 51 78.3 10 88.3 91 N >0.71 <10 <10 21 6 53 <15 6111 53 82.2 10.4 92.6 91 N >0.25 <10 <10 20 9 61 <15 6054 59 76.7 10.7 87.4 91 N >0.063 <10 <10 13 6 44 <15 3375 52 78.6 6.98 85.6 91 N <0.063 <10 <10 8 <5 35 <15 2405 54 76.3 - - 91 T >2 <10 <10 13 7 48 <15 8633 63 67.3 13.7 81 91 T >0.71 <10 <10 15 <5 48 <15 9188 66 63.5 14.3 77.8 91 T >0.25 <10 10 16 6 52 <15 9287 67 68.4 14.3 82.7 91 T >0.063 <10 10 13 <5 41 <15 7041 61 68.8 11.1 79.9 91 T <0.063 <10 12 13 <5 47 <15 7923 79 78.9 - -

sample nr. Sr Th Tl U V W Zn Zr sum l.o.i sum+l.o.i. A2-4 mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg % % % 113< 1cm N < 1cm <30 <30 14 <15 60 <45 7575 <120 75.6 - - 114< 1cm N < 1cm 33 <30 23 <15 60 <45 4977 <120 72.6 - - 115< 1cm N < 1cm <30 <30 23 <15 39 <45 4251 <120 67.1 - - 116< 1cm N < 1cm <30 <30 18 <15 57 <45 5253 <120 69.0 - - 113> 1cm N > 1cm <30 <30 <9 <15 48 <45 17028 <120 85.4 - - 114> 1cm N > 1cm 36 <30 19 <15 93 <45 3825 <120 85.7 - - 115> 1cm N > 1cm 36 <30 <9 <15 72 <45 5556 <120 82.6 - - 116> 1cm N > 1cm <30 <30 11 <15 <30 <45 8655 <120 82.3 - -

Appendix B

242

Table B10: Position of the (111)-peak in selected goethite-rich samples

Sample nr. Open pit (111)-peak

268d CQ 2.45

263 CQ 2.45

209c CQ 2.45

209b CQ 2.45

209a CQ 2.45

205b CQ 2.43

13 CQB 2.44

14 CQB 2.44

51 FBN 2.45

52 FBN 2.45

276 FBS 2.44

Appendix C

243

Appendix C: Results of the laboratory leaching experiments

Table C1: Position of samples used in the experiments in the profiles on the heap leaching pad.

sample nr. location / profile on the

leach pad position in the profile

[m]

97 A2-2 0

95 A2-2 -1

91 A2-2 -2

116/2 A2-4 0

116/4 A2-4 0

115 A2-4 -0.5

113 A2-4 -1.5

114 A2-4 -1

Matrix 1/M1 A2-3 +1

211FK storage heap random

210FK storage heap random

212/4MK storage heap random

263 Copper Queen open pit

Appendix C

244

Table C2: XRF data of ore samples used in the laboratory experiments and filtrates.

Sample nr. SiO2 Al2O3 Fe2O3 MgO CaO Na2O K2O TiO2 P2O5 SO3

% % % % % % % % % % Run-of-mine ore before leaching experiment 210/1 24.9 6.85 27.56 3.61 0.87 0.1 0.84 0.18 0.14 2.62 211/1 26.5 6.86 31.37 4.63 0.68 <0.1 0.8 0.16 0.15 1.62 212/4 1.7 0.39 87.02 0.07 0.02 <0.01 0.02 0.05 0.07 0.09 Run-of-mine ore after leaching experiment 210g - 6.12 43.59 6.30 1.56 <0.3 0.90 0.21 0.26 1.50 211g - 8.31 35.49 3.81 0.84 <0.3 1.05 0.24 0.23 1.75 212/4g - 0.63 83.34 <0.9 <0.3 <0.3 <0.15 <0.09 0.08 1.04 A2-4 before leaching experiment 113 < 1cm - 7.86 26.28 2.76 0.63 <0.3 0.51 0.24 0.22 1.24 114 < 1cm - 7.41 28.65 1.29 <0.3 <0.3 0.78 0.3 0.19 4.88 115 < 1cm - 5.67 25.41 1.23 <0.3 <0.3 0.63 0.27 0.20 3.50 116 < 1cm - 5.13 28.59 1.05 <0.3 <0.3 0.54 0.3 0.23 4.24 A2-4 after leaching experiment 113g 25.4 7.64 27.88 2.86 0.52 <0.1 0.39 0.22 0.16 1.46 114g 18.1 8.29 28.96 1.37 0.11 <0.1 0.66 0.25 0.13 1.88 115g - 7.62 26.76 1.32 <0.3 <0.3 <0.15 0.36 0.24 1.88 116/2g - 6.18 26.22 1.38 <0.3 <0.3 <0.15 0.30 0.22 1.62 116/4g - 6.93 28.62 1.11 <0.3 <0.3 <0.15 0.33 0.27 2.08 Extraction experiment 263 - 2.64 61.89 <0.9 <0.3 <0.3 <0.15 0.45 0.07 0.32 26 - <0.6 69.15 1.41 <0.3 <0.3 <0.15 <0.09 0.14 0.87 Leach pad ore before leaching experiment MM1 - 7.8 28.83 2.61 0.39 <0.3 0.84 0.3 0.198 2.10 MM1 - 8.19 31.41 2.58 0.36 <0.3 0.87 0.27 0.189 2.25 MM1/V1 5.94 28.02 2.52 0.51 <0.3 0.75 0.3 0.144 1.62 MM1/V1 - 5.79 27.96 2.34 0.48 <0.3 0.72 0.3 0.153 1.69 MM1/V6/F1 - 14.2 38.9 4 <1.0 <1.0 1.4 0.4 0.34 3.84 MM1/V6/F3 - 11.8 38 3.3 <1.0 <1.0 1.2 0.4 0.31 4.13 MM1/V6/F5 - 8.5 43.2 <3.0 <1.0 <1.0 0.9 0.5 0.31 3.64

MM1 Matrix1 after H2O and H2SO4 washing MM1a Matrix1 after experiment V1 V6/Fx filter residue after sampling

Chrysocolla/malachite C1(cry/mal) 17.0 2.12 0.44 0.28 0.1 4.39 - 0.01 - -

Appendix C

245

Table C2 continued: XRF data of ore samples used in the laboratory experiments and filtrates. Sample nr. F Ag As Ba Bi Br Cd Cl Co Cu mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg Run-of-mine ore before leaching experiment 210/1 654 12 3998 309 126 <4 37 n.d. 191 4372 211/1 1571 12 4583 315 126 <4 31 <300 102 4575 212/4 <0.05 <3 3087 8 12 14 16 60 259 5055 Run-of-mine ore after leaching experiment 210g <2400 14 3804 405 91 <12 33 <900 72 3081 211g <2400 22 4356 396 121 <12 25 <900 114 4785 212/4g <2400 <8 2244 <90 90 <12 <15 <900 150 4446 A2-4 before leaching experiment 113 < 1cm 4761 16 10983 294 79 <12 31 <900 177 3969 114 < 1cm <2400 37 13200 180 159 <12 41 <900 111 4881 115 < 1cm <2400 31 13410 171 205 13 24 <900 84 5478 116 < 1cm <2400 34 11570 192 223 16 25 <900 117 4710 A2-4 after leaching experiment 113g <800 15 12980 229 105 8 33 <300 141 4080 114g <800 32 13700 169 206 9 26 <300 42 4830 115g <2400 37 12738 204 263 <12 28 <900 51 5142 116/2g <2400 26 10485 189 254 <12 <15 <900 45 3999 116/4g <2400 35 11472 225 269 <12 22 <900 42 4296 Extraction experiment 263 <2400 <8 <30 <90 298 <12 <15 <900 <30 5754 26 <2400 10 2934 <90 165 <12 <15 <900 234 14330 Leach pad ore before leaching experiment MM1 2775 17 8949 348 144 <12 23 <900 72 3270 MM1 4329 16 10092 405 162 <12 25 <900 78 3741 MM1/V1 9042 9 8268 288 142 <12 16 <900 60 2934 MM1/V1 7470 <7.5 8172 282 140 <12 22 <900 54 2940 MM1/V6/F1 12740 <25 11140 600 150 <40 <50 n.d. 120 4440 MM1/V6/F3 <8000 30 11660 580 181 <40 <50 n.d. <100 4500 MM1/V6/F5 8800 34 12860 460 179 <40 <50 n.d. <100 4640 MM1 Matrix1 after H2O and H2SO4 washing MM1a Matrix1 after experiment V1 V6/Fx filter residue after sampling

Chrysocolla/malachite C1(cry/mal) - - 102 1081 100 - - - - 452072

Appendix C

246

Table C2 continued: XRF data of ore samples used in the laboratory experiments and filtrates.

Sample nr. Cr Cs Ga Mn Mo Ni Pb Rb Sb Se

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg Run-of-mine ore before leaching experiment 210/1 42 n.d. n.d. 7983 n.d. 31 14641> 66 81 48> 211/1 51 n.d. n.d. 6265 3.5< 20 13762> 56 96 45> 212/4 8 <5 31 1232 <2 11 1892 4 36 24 Run-of-mine ore after leaching experiment 210g <90 <24 <9 4287 <15 <36 10632 42 83 35 211g <90 <24 <9 5316 <15 <36 15783 48 176 47 212/4g <90 <24 27 840 <15 <36 2469 <30 26 29 A2-4 before leaching experiment 113 < 1cm 102 <24 <9 7617 <15 <36 8613 33 66 30 114 < 1cm <90 <24 <9 2154 <15 <36 9414 36 146 65 115 < 1cm <90 <24 <9 1947 <15 <36 10032 33 167 70 116 < 1cm <90 <24 <9 3084 <15 <36 11679 <30 150 75 A2-4 after leaching experiment 113g 63 <8 <3 8108 <5 23 10902 40 54 37 114g 57 <8 4 991 <5 <12 10510 47 176 66 115g <90 <24 <9 669 <15 <36 11253 42 152 76 116/2g <90 <24 <9 1521 <15 <36 11238 <30 146 63 116/4g <90 <24 <9 1323 <15 <36 11751 <30 168 69 Extraction experiment 263 <90 <24 <9 <450 <15 <36 174 <30 74 86 26 <90 <24 <9 9348 <15 <36 14145 <30 77 50 Leach pad ore before leaching experiment MM1 <90 33 <9 4071 <15 <36 12309 45 117 43 MM1 <90 <24 <9 4017 <15 <36 13236 51 121 48 MM1/V1 <90 <24 <9 4149 <15 <32 12129 39 101 44 MM1/V1 <90 <24 <9 4002 <15 <36 11958 36 101 43 MM1/V6/F1 n.d. <80 n.d. 4680 <50 <120 14980 <100 110 39 MM1/V6/F3 n.d. <80 n.d. 4870 n.d. <120 17980 <100 143 48 MM1/V6/F5 n.d. 84 n.d. 4490 n.d. <120 19460 <100 127 61 MM1 Matrix1 after H2O and H2SO4 washing MM1a Matrix1 after experiment V1 V6/Fx filter residue after sampling chrysocolla/malachite C1(cry/mal) - - - - 44 - 2308 2 - -

Appendix C

247

Table C2 continued: XRF data of ore samples used in the laboratory experiments and filtrates.

Sample nr. Sn Sr Th Tl U V W Zn Zr sum l.o.i sum+l.o.i.

mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg % % % Run-of-mine ore before leaching experiment 210/1 <10 15 14 7 <5 46 <15 11599 57 72.08 - - 211/1 76 <10 <10 8 <5 53 <15 9237 54 76.91 - - 212/4 342 <2 36 <3 16 35 <5 8394 3 90 8 98 Run-of-mine ore after leaching experiment 210g 922 <30 <30 <9 <15 93 <45 5448 <120 62 - - 211g 238 <30 <30 <9 <15 78 <45 6081 <120 58.8 - - 212/4g 278 <30 <30 <9 <15 57 <45 7365 <120 61.85 - - A2-4 before leaching experiment 113 < 1cm 73 <30 <30 14 <15 60 <45 7575 <120 85 - - 114 < 1cm 118 33 <30 23 <15 60 <45 4977 <120 86 - - 115 < 1cm 82 <30 <30 23 <15 39 <45 4251 <120 84 - - 116 < 1cm 227 <30 <30 18 <15 57 <45 5253 <120 82 - - A2-4 after leaching experiment 113g 22 <10 14 15 <5 58 <15 7902 57 67.19 - - 114g 100 25 <10 19 <5 48 <15 2095 64 84.5 - - 115g 50 <30 <30 21 <15 51 <45 1644 <120 60.21 - - 116/2g 149 <30 <30 16 <15 42 <45 2043 <120 57.22 - - 116/4g 131 <30 <30 19 <15 48 <45 2214 <120 61.56 - - Extraction experiment 263 <30 <30 <30 <9 <15 51 <45 744 132 - 6 - 26 828 <30 <30 <9 <15 33 <45 11754 <120 - 12 - Leach pad ore before leaching experiment MM1 <30 <30 <30 13 <15 57 <45 4593 <120 63.13 - - MM1 <30 <30 <30 17 <15 60 <45 5247 <120 57.86 - - MM1/V1 208 <30 <30 11 <15 45 <45 4038 <120 61.26 - - MM1/V1 220 <30 <30 <9 <15 60 <45 4074 <120 61.59 - - MM1/V6/F1 n.d. <100 <100 <30 <50 <100 <150 5830 <400 58.33 - - MM1/V6/F3 n.d. n.d. <100 <30 <50 120 <150 5890 <400 61.01 - - MM1/V6/F5 n.d. <100 <100 <30 n.d. <100 <150 6020 <400 59.89 - - MM1 Matrix1 after H2O and H2SO4 washing MM1a Matrix1 after experiment V1 V6/Fx filter residue after sampling Chrysocolla/malachite C1(cry/mal) - 28 - - 6 5 - 3846 - 67.45 - -

Appendix C

248

Table C3: Composition of the water-soluble fraction of the samples from profile A2-2 and A2-4 (V0).

Ore sample Solute sample take at l nr. Co Cu Zn Cd Pb

mg/l mg/l mg/l mg/l mg/l

A2-2-0 5 2.6 4.8 133 0.27 0.16

A2-2-100 5 2.4 3.2 123 0.27 1.26

A2-2-100 10 0.31 0.4 19 <0.05 1.84

A2-2-100 15 <0.1 <0.05 0.53 <0.05 <0.1

Σ 2.71 3.6 143 0.54 3.25

A2-2-200 5 4.3 8.3 270 0.46 <0.1

A2-2-200 10 0.39 0.7 230 0.05 <0.1

A2-2-200 15 0.18 0.3 12 <0.05 0.06

A2-2-200 20 <0.1 0.1 2.7 <0.05 0.22

A2-2-200 25 <0.1 0.07 1.48 <0.05 0.13

Σ 4.87 9.47 516 0.51 0.95

Ore sample Solute sample take at l nr. Co Cu Zn Cd Pb

mg/l mg/l mg/l mg/l mg/l

A2-4-50 1 0.06 0.74 4.22 - - A2-4-50 3 0.04 0.37 2.78 - - A2-4-50 5 0.02 0.29 2.15 - - A2-4-50 7 <0.10 0.1 0.7 - - A2-4-50 9 <0.10 0.12 0.84 - <0.10

A2-4-0 1 0.25 1.65 14.23 - -

A2-4-0 3 0.12 0.7 6.56 - -

A2-4-0 5 0.06 0.51 4.31 - -

A2-4-0 7 0.01 0.25 1.45 - -

A2-4-0 9 0.02 0.27 1.65 - 0.13

A2-4-0 1 0.13 1.12 8.53 - -

A2-4-0 3 0.1 0.6 5.8 - -

A2-4-0 5 0.05 0.48 3.82 - -

A2-4-0 7 0.02 0.22 1.33 - -

A2-4-0 9 0.01 0.24 1.5 - <0.10

Appendix C

249

Table C4: Composition of the H2SO4-soluble fraction leached from Run-of-mine ore (VR).

Ore sample litre pH Leitf. Mg Al Mn Co Cu Zn Pb

mg/l mg/l mg/l mg/l mg/l mg/l mg/l

211 FK 1 1.5 - 89.6 40 420.6 20.2 114 961.6 1.2

211 FK 2 1.74 14.6 114.4 57 530.2 24.6 163.2 1241 1.7

211 FK 3 - - 122.6 74.5 549.6 25.3 196.8 1327 2

211 FK 4 1.7 13.4 124.6 84 556.7 25.2 217.8 1334 2.0

211 FK 5 1.72 12.8 130 98 569.8 25.9 250.2 1401 2.0

211 FK 6 1.76 12.5 135.2 117.5 600.2 27 284.6 1445 2.2

211 FK 7 1.78 12.2 147.2 132.8 671.8 28 317.8 1568 2.1

211 FK 8 - - 147 144.2 636.5 27.7 327 1534 2.1

211 FK 9 - 12.2 155 151 644.1 26.7 340.2 1638 2.1

211 FK 10 2.04 12.1 155.6 157.5 638.7 28.1 359.2 1621 2.1

210 FK 1 1.5 - 168 76.5 822.9 33.2 137.4 1633 1.1

210 FK 2 1.91 15.1 272 143 1263.5 45.0 232.6 2610 1.7

210 FK 3 - - 287.2 165.8 1331.1 45.8 256.2 2900 1.8

210 FK 4 1.88 14.5 282.8 178.2 1318.95 45.7 273 2676 1.9

210 FK 5 1.85 14.5 283 186 1295.6 45. 9 287.8 2940 2.0

210 FK 6 1.86 14.2 293 206 1343.7 46.1 317.6 2960 2.1

210 FK 7 1.92 14.3 292.6 210 1322.4 46.0 318 2970 2.1

210 FK 8 - - 292.4 216.8 1343.3 46.2 326 3000 2.1

210 FK 9 - 14.3 294.8 225.8 1339.9 46.6 333.8 3100 2.1

210 FK 10 2.09 14.2 300.8 230.4 1358.7 46.6 351.2 3100 2.2

212/4 MK 1 1.5 - 1 3.4 2 0.1 1.6 4.6 0.1

212/4 MK 2 1.49 15.9 1.0 3.4 3 0.1 3.1 8.2 0.2

212/4 MK 3 - - 1.2 3.4 3.6 0.2 4.1 10.7 0.2

212/4 MK 4 1.56 16.5 1.5 3.8 4.8 0.2 4.9 14.1 0.3

212/4 MK 5 1.54 16.7 1.7 5 5.4 0.3 5.7 16.1 0.3

212/4 MK 6 1.57 17 2.0 4.4 6.2 0.3 6.8 19.7 0.4

212/4 MK 7 - - 2.1 4.6 6.6 0.4 8.1 22.3 0.5

212/4 MK 8 - 17.5 2.1 4.9 6.8 0.3 8.3 22.1 0.5

212/4 MK 9 - - 2 4.6 6.8 0.3 8.1 21.6 0.5

212/4 MK 10 1.71 17.3 2 4.6 6.6 0.3 8.6 22.2 0.6

Appendix C

250

Table C5: Composition of the H2SO4-soluble fraction leached from leach pad ore (V15).

Ore sample

Solute sample take at l

nr. pH cond. Mg Al Si Mn Co Cu Zn Pb

mS/cm mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l

113 0.2 1.61 23.4 - - - - 2.5 89.7 189 2.5

113 0.4 1.7 20.2 - - - - 1.6 72.6 127 2.3

113 0.6 1.69 18.7 - - - - 1.2 59.8 96 1.9

113 1 - - 1.3 19.3 3.06 5.9 0.8 20.0 38 1.5

113 2 - - 6.4 34.7 11.2 15.3 1.0 54.0 112 1.8

113 3 - - 9.3 60.0 17.1 16.6 1.2 71.0 135 1.8

113 4 - - 10.7 62.6 21.4 16.9 1.3 77.0 122 1.8

113 5 - - 14.6 68.5 25.6 17.4 1.3 82.0 146 1.9

113 6 - - 16.5 71.3 27.4 17.9 1.3 87.0 151 1.9

113 7 - - 16.7 74.5 30.1 18.2 1.4 88.0 157 1.9

113 8 - - 20.3 86.1 36.6 19.0 1.6 105 167 1.9

113 9 - - 23.6 94.8 40.5 19.5 1.6 110 168 1.9

113 10 1.7 14 25.7 97.0 47.4 20.8 1.8 120 188 2.0

114 0.2 1.59 20.4 - - - - 0.7 9.3 37.1 0.1

114 0.4 1.68 19.6 - - - - 0.5 6.0 21.1 0.1

114 0.6 1.59 20.4 - - - - 0.23 4.2 12.4 <0.1

114 1 - - 4.3 29.4 2.6 3.6 0.2 4.2 14.3 <0.1

114 2 - - 5.1 31.7 9.0 3.8 0.3 5.7 15.6 <0.1

114 3 - - 6.5 55.7 14.7 4.1 0.2 7.4 17.2 <0.1

114 4 - - 8.0 58.1 19.3 4.3 0.3 8.7 17.7 <0.1

114 5 - - 9.4 64.6 21.3 4.3 0.3 9.3 17.8 <0.1

114 6 - - 10.6 68.9 24.2 4.4 0.3 10.3 18.1 <0.1

114 7 - - 10.6 70.2 25. 5 4.4 0.3 10.6 18.6 <0.1

114 8 - - 13.9 70.5 31.5 4.7 0.3 11.9 19.2 <0.1

114 9 - - 15.5 72.6 35.8 4.9 0.3 12.3 19.7 <0.1

114 10 - - 18 78.6 40.4 5.1 0.3 12.9 20.4 <0.1

115 1 1.5 9.3 0.6 29.6 <1.0 1.4 0.1 3.7 5.9 <0.1

115 2 1.79 - 0.4 37.3 5.5 1.9 0.1 6.6 8.5 <0.1

115 3 1.79 10.9 0.8 38.9 7. 7 2.1 0.1 7.4 9.5 <0.1

115 4 - - 1.3 40.5 11.2 2.3 0.1 8.3 10.4 <0.1

115 5 1.79 12.4 2.1 46.0 15.2 2.5 0.1 9.2 11.9 <0.1

115 6 1.85 12.4 2.8 49.4 18.5 2.8 0.1 9.8 13.8 <0.1

115 7 - - 5.3 73.4 35.1 4.5 0.1 15.5 21.1 0.1

115 8 1.62 16.2 14.2 94.9 37.3 6.6 0.2 22.1 29.1 0.2

115 9 - - 13.9 100.2 39.4 6.6 0.2 22.2 29.3 0.2

115 10 1.64 15.9 14.8 103.0 41.4 6.9 0.2 23.1 29.9 0.2

Appendix C

251

Table C5 contiuned: Composition of the H2SO4-soluble fraction leached from leach pad ore (V15).

Ore sample

Solute sample take at l

nr. pH cond. Mg Al Si Mn Co Cu Zn Pb


116/2 1 1.50 9.5 0.2 26.8 2.9 3.1 0.1 5.4 6.9 2.2

116/2 2 1.71 - 0.4 31.8 9.0 5 0.2 7.6 10.0 2.5

116/2 3 1.75 11.1 20.5 35.2 12.9 5.7 0.2 8.2 11.3 2.6

116/2 4 - - 1.9 37.3 15.8 6.6 0.2 8.6 12.2 2.5

116/2 5 1.75 12.0 2.4 43.3 23.3 8 0.3 9.6 14.7 2.6

116/2 6 1.72 12.2 2.5 45.8 25.0 8.3 0.3 9.8 15.2 2.5

116/2 7 - - 5.6 67.3 445 13.5 0.4 14.2 25.0 3.7

116/2 8 1.59 16.2 14.6 96.6 46.6 21.9 0.6 19.0 33.4 2.6

116/2 9 - - 14.2 92.1 48.2 21.6 0.6 18.4 33.2 2.7

116/2 10 1.62 15.9 14.8 95.7 50.9 22.7 0.6 18.7 33.6 2.6

116/4 1 1.52 11.1 0.5 38.6 6.4 4.1 0.1 6.7 9.0 3.2

116/4 2 1.75 - 1.7 43.1 14.9 5.6 0.2 8.1 11.1 2.5

116/4 3 1.68 13.2 1.6 44.7 20.3 6.7 0.2 8.9 13.1 2.5

116/4 4 - - 1.8 44.7 20.3 6.6 0.2 8.6 12.5 2.3

116/4 5 1.81 13.8 3.9 65.2 36.1 9.4 0.3 10.0 17.7 3.8

116/4 6 1.81 13.7 3.1 50.2 28.1 8.2 0.2 9.4 15.2 2.1

116/4 7 - - 6.9 79.2 53.4 14.0 0.4 13.6 25.7 3.1

116/4 8 1.67 18.7 18.2 108.7 55.5 22.7 0.6 18.0 33.9 2.2

116/4 9 - - 17.1 107.5 55.6 22.1 0.6 17.4 33.4 2.5

116/4 10 1.61 18.3 19.3 121.9 68.2 24.7 0.7 19.3 35.8 -

Appendix C

252

Table C6: Composition of the H2SO4-soluble fraction leached from leach pad ore (V1, V6, V7).

V1

Ore sample sampling time Mg Al Si Ca Mn Fe Co Cu Zn As

h mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l

Matrix 1 0.25 1.3 3.6 1.9 0.6 1.6 7.4 0.1 1.3 2.1 4.2

Matrix 1 0.5 2.2 7 1.8 2.9 2.9 18.2 0.1 2.3 3.6 8.3

Matrix 1 1 2.2 7.3 2.2 4.2 3.1 20.2 0.2 2.5 3.8 8.9

Matrix 1 2 2.3 8.3 2.4 2.6 3.5 23.8 0.1 2.7 4.1 10.0

Matrix 1 4 2.4 8.9 3.2 1.7 3.8 27.5 0.2 2.9 4.4 9.9

Matrix 1 9 3.5 13.1 7.1 3.5 6.0 39.0 0.2 4.4 6.0 11.1

Matrix 1 20.5 4.2 15.5 10.2 1.5 8.6 48.8 0.3 5.5 7.3 10.5

Matrix 1 42.5 6.8 21.7 16.3 2.4 13.8 71.6 0.3 8.0 10.1 8.5

V6

Ore sample sampling time ph cond. Al Si Mn Fe Co Cu Zn As

h mS/cm mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l

Matrix 1 0.083 1.5 16.7 2.8 0.1 2.1 9.3 <0.3 2.3 2.7 6.9

Matrix 1 0.17 - - 2.9 0.7 2.3 11.9 <0.3 2.6 3 9.2

Matrix 1 0.25 - 16.1 3 0.4 2.3 12.4 <0.3 2.2 2.8 8.6

Matrix 1 0.33 - 3 0.5 2.3 12.3 <0.3 2.2 2.9 9.1

Matrix 1 0.5 1.6 16.1 3.7 1.5 2.8 17.2 <0.3 2.5 3.2 11.7

Matrix 1 1 1.5 15.9 3.6 1.6 2.9 18.1 <0.3 2.2 3.2 12.4

Matrix 1 3 1.5 15.8 3.8 2.0 2.8 20.7 <0.3 2.5 3.4 12.9

Matrix 1 5 1.6 15.6 5.2 5.5 4.0 31.9 <0.3 3.8 4.3 16.4

Matrix 1 16 1.5 15.2 6.7 9.5 5.2 43.9 <0.3 4.7 5.4 17.7

Matrix 1 20 1.5 15 - - - - - - - -

Matrix 1 44 1.6 14.5 12.1 20.3 9.4 69.5 <0.3 7.1 7.9 17.1

Matrix 1 68 1.5 14 16.4 27.7 12.9 87.0 0.4 8.2 9.4 15.7

Matrix 1 92 1.6 14.6 19.9 32.9 15.8 99.7 0.9 8.7 10.3 14.6

Matrix 1 116 1.5 13.1 38.9 38.2 34.7 140.6 4.1 11.1 13.3 13.6

Matrix 1 140 1.6 12.6 59.9 43.0 54.0 178.5 9.1 11.9 15.7 12.8

Matrix 1 164 1.6 11.8 87.7 46.0 77.5 229.7 20 13.3 19 12.5

V7

Ore sample ph cond. Al Si Mn Fe Co Cu Zn As


Matrix 1 1.55 10.5 43.3 44.8 14.8 95.4 0.8 13 18. 13.4

Matrix 1 1.67 10.5 46.8 58.4 15.1 93.6 0.8 12.4 18 12.2

Matrix 1 1.75 9.7 63.4 79.4 20.5 115.3 1.1 14.7 21.5 9.0

Matrix 1 1.71 10 70.7 85.9 22.2 121.1 1.3 14.9 22.2 8.7

Matrix 1 1.69 9.7 74.6 89.4 23.2 128.6 1.1 15.1 22.6 8.3

Matrix 1 2.06 10.5 75.4 88.8 24 136.1 1.2 13.7 21.1 10.5

Matrix 1 1.71 10.3 77.3 92.6 25.3 138.2 1.3 14 21.8 9.7

Matrix 1 1.72 10.2 83.2 96.7 26.4 146.8 1.3 14.1 22.3 8.5

Matrix 1 1.43 9.9 86.7 98.8 27.5 147.9 1.3 14.3 22.5 8.8

Matrix 1 1.71 9.9 88.4 101.7 28 153.9 1.3 14.5 22.8 8.9

Appendix C

253

Table C7: Composition of the partial extraction experiments with NH4acetate (V12), NH4oxalate dark (V13), NH4oxalate 80°C(V14).

V 12

Sampling time pH cond. Al Si Mn Fe Co Cu Zn As


0.083 4.55 - n.d. 0.4 n.d. 0.2 n.d. 1.7 0.3 2.8

0.25 4.53 55.7 n.d. 0.2 n.d. 0.2 n.d. 1.8 0.1 5.5

0.5 4.51 55.7 n.d. 0.5 n.d. 0.6 n.d. 2.2 0.3 2.8

0.75 4.50 55.2 n.d. 0.4 n.d. 0.6 n.d. 2.3 0.2 2.8

1 4.49 55.0 n.d. 0.5 n.d. 0.5 n.d. 2.3 0.2 2.8

2 4.47 55.7 n.d. 0.8 n.d. 0.8 n.d. 2.8 0.7 3

4 4.46 55.6 n.d. 0.9 n.d. 1.0 n.d. 3.0 0.3 3

8 4.53 55.6 n.d. 1 n.d. 1.2 n.d. 3.2 0.2 2.9

V 13



0.083 3.08 32.2 0.12 0.7 n.d. 1.6 0.1 1.9 0.2 0.3

0.25 3.06 31.9 0.2 0.8 n.d. 2.73 n.d. 2.328 0.21 0.2

0.5 3.06 32.5 0.3 2.01 n.d. 2.8 0.1 2.4 0.2 0.2

0.75 3.06 31.9 0.3 0.9 0.09 2.6 0.1 2.3 0.1 0.1

1 3.05 31.9 0.3 1 n.d. 5.0 n.d. 2.5 0.1 0.1

2 3.06 32.0 0.4 1.4 n.d. 6 n.d. 2.7 0.2 0.1

4 3.10 31.9 0.6 3.6 n.d. 7.0 n.d. 2.9 0.4 0.1

8 3.06 31.7 0.4 2.1 n.d. 8.6 n.d. 3 0.3 0.1

V 14



0.083 2.95 34.6 0.4 1.3 n.d. 9.9 0.1 2.7 0.3 0.1

0.25 2.94 35.0 0.5 1.0 n.d. 17.7 0.1 2.9 0.2 0.2

0.5 2.91 35.9 0.7 1.5 n.d. 33.3 0.1 3.4 0.3 0.1

0.75 2.92 36.3 1.1 2.0 n.d. 92.4 n.d. 4.4 0.2 0.2

1 2.96 35.7 1.8 3 0.1 184 n.d. 5.7 3.8 0.3

2 3.00 35.6 4.1 5.9 0.1 490 0.3 9.3 0.8 0.4

4 3.20 36.2 7.8 11.2 0.1 1135 0.1 11 1.6 0.7

8 3.19 36.8 8.5 13.1 0.2 1320 0.2 11.2 2.2 0.8

Appendix C

254

Table C8: Composition of sorption experiments of malachite (V8), chrysocolla (V9) and synthetic Cu-sulphate (V10) in an acidic environment.

Malachite + H2SO4 Chrysocolla + H2SO4

V8 V9

Sampling time pH cond. Cu Fe Sampling time pH cond. Cu Fe

h mS/cm mg/l mg/l h mS/cm mg/l mg/l

0.08 - 2.6 91.2 8.4 0.08 2.22 2.9 47.6 10.6

0.16 2.36 2.6 92.7 10.2 0.16 2.21 2.8 55 11.4

0.25 2.31 2.6 94.2 10.8 0.25 2.21 2.8 60.7 14

0.33 2.27 2.6 85.9 11.4 0.33 2.20 2.7 61.8 14.6

0.5 - 2.7 81.3 16.0 0.5 2.23 2.7 61.6 16.2

1 2.33 2.7 76.8 16.2 1 2.33 2.7 63.3 19.2

2 2.33 2.6 77.9 21.4 2 2.35 2.6 62.6 23.8

3 2.35 2.6 71.4 25.6 3 2.27 2.6 61.1 25.8

5 2.39 2.6 66.9 29.0 5 2.38 2.6 58 26.8

10 2.32 2.6 63.8 33.0 10 2.35 2.6 56.5 32.2

25 2.33 2.6 61.6 36.8 25 2.26 2.6 58.3 37

50 2.23 2.6 55.9 33.0 50 2.23 2.6 58 31.4

75 2.22 2.7 52.8 32.8 75 2.21 2.7 57.8 30

100 2.23 2.8 51.3 33.2 100 2.23 2.7 58.7 29

150 2.23 2.8 47.9 31.2 150 2.26 2.7 61.1 26.8

200 2.24 2.8 61.8 23.6 Cu sulfate + H2SO4

V10

Sampling time pH cond. Cu Fe

h mS/cm mg/l mg/l

0.08 2.14 3.54 102 14.4

0.16 2.19 3.47 101 16.4

0.25 2.29 3.46 96.1 21.6

0.33 2.53 3.45 90.1 24.2

0.5 2.43 3.38 86.6 28.2

1 2.20 3.24 77.5 34.4

2 2.09 3.18 77.1 41.6

3 2.08 3.14 59.3 48.8

5 2.17 3.11 79.0 62.4

10 2.16 2.98 67.6 56.2

25 2.16 2.9 60.4 55.0

50 2.18 3.05 55.9 51.0

75 2.19 3.09 54.0 50.4

100 2.19 3.12 50.2 48.2

150 2.17 3.18 49.4 44.0

200 2.15 3.22 43.3 38.6

Appendix D

255

Table D1: EMPA results of goethite- and hematite-rich zones in ROM and LPO.

Label col. O Mg Al Si S Ca K Mn Fe Co Cu Zn As Hg Pb Σ wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%rectangular boxwork 294Z1-d-b1-1 1 30 n.d. 0.21 1.06 0.25 0.01 n.d. 0.01 61.8 0.02 0.88 0.86 0.34 n.d. 2.00 97.6294Z1-d-b1-3 1 30 0.02 0.49 3.39 0.18 0.08 n.d. 0.09 57.2 0.06 1.08 1.14 0.81 n.d. 2.20 97.2294Z1-d-b1-4 1 29 n.d. 1.00 3.11 0.22 0.05 n.d. 0.05 57.4 0.06 1.08 1.85 0.55 n.d. 1.07 95.0294Z1-d-b1-7 1 31 n.d. 0.33 0.95 0.29 0.02 n.d. 0.01 61.4 0.05 0.53 1.07 0.16 n.d. 1.80 97.1rectangular boxwork 294Z1-d-a3-12 1 31 0.02 0.30 1.79 0.22 0.02 n.d. n.d. 58.1 0.06 0.99 0.91 0.29 n.d. 1.55 95.4294Z1-d-a3-13 1 31 n.d. 0.25 1.38 0.23 0.02 n.d. 0.01 59.3 0.04 0.62 0.81 0.31 n.d. 1.47 95.0triangular boxwork 294Z1-d-a1-8 1 31 n.d. 0.33 1.04 0.29 0.01 n.d. 0.01 58.1 0.04 0.83 1.12 0.17 n.d. 2.02 94.6294Z1-d-a1-16 1 31 n.d. 0.32 1.32 0.24 0.03 n.d. 0.03 58.9 0.06 0.66 0.80 0.17 n.d. 1.80 95.0euhedral boxwork 294Z1B_A 1 30 n.d. 0.59 1.86 0.19 n.d. n.d. n.d. 60.6 0.06 0.98 0.91 n.d. n.d. 1.79 96.8294Z1B_I 1 30 n.d. n.d. 0.04 0.30 n.d. n.d. n.d. 59.9 0.04 0.90 0.11 n.d. n.d. 2.46 93.9294Z1B_I2 1 31 n.d. 0.01 0.04 0.33 n.d. n.d. n.d. 60.3 0.06 0.84 0.01 n.d. n.d. 2.56 95.0294Z1B_I3 1 31 n.d. 0.19 0.90 0.30 n.d. n.d. n.d. 59.1 0.05 0.78 1.25 n.d. n.d. 2.52 96.0294Z1B_I4 1 32 n.d. 0.37 0.91 0.33 n.d. n.d. n.d. 59.6 0.05 0.93 1.24 n.d. n.d. 2.29 97.3294Z1B_Q 1 32 n.d. n.d. 0.95 0.32 n.d. n.d. n.d. 60.1 0.04 0.83 1.39 n.d. n.d. 2.34 97.5294Z1B_Q1 1 32 n.d. 0.28 0.98 0.35 n.d. n.d. n.d. 60.3 0.08 0.86 1.21 n.d. n.d. 2.46 97.7294Z1B_T 1 32 n.d. 0.28 0.91 0.33 n.d. n.d. n.d. 55.8 0.05 0.75 1.34 n.d. n.d. 2.38 93.7294Z1B_W 1 32 n.d. 1.75 3.82 0.13 n.d. n.d. n.d. 44.3 0.07 2.35 2.65 n.d. n.d. 0.89 87.7294Z1B_X 1 32 n.d. 0.60 2.15 0.20 n.d. n.d. n.d. 53.7 0.06 0.81 0.96 n.d. n.d. 1.81 92.1294Z1B_Y=0 1 31 n.d. 0.44 2.03 0.22 n.d. n.d. n.d. 50.8 0.07 0.62 0.72 n.d. n.d. 1.48 87.0294Z1B_Z 1 31 n.d. 1.03 2.50 0.18 n.d. 0.01 n.d. 50.3 0.07 1.06 2.01 n.d. n.d. 1.70 89.5euhedral boxwork 294Z1-d-a4-18 1 31 n.d. 0.24 0.60 0.28 0.01 n.d. 0.01 59.7 0.08 0.73 0.90 0.19 n.d. 1.99 94.9294Z1-d-a4-19 1 31 n.d. 0.24 1.43 0.20 0.04 n.d. 0.05 59.0 0.05 0.67 0.70 0.29 n.d. 1.43 95.2euhedral boxwork 294R1-d-a2-8 1 32 0.01 0.84 1.92 0.25 0.02 n.d. 0.04 56.9 0.03 0.97 2.15 0.21 n.d. 0.99 96.6294R1-d-a2-10 1 32 n.d. 0.62 1.80 0.22 0.02 n.d. 0.03 55.8 0.05 1.07 1.99 0.22 n.d. 1.16 95.1294R1-d-a2-12 1 29 0.01 1.03 2.17 0.22 0.00 n.d. 0.02 54.8 0.03 1.13 3.31 0.31 n.d. 0.95 92.9294R1-d-a2-14 1 30 0.01 0.07 0.62 0.30 0.03 n.d. n.d. 61.7 0.04 0.77 0.70 0.14 n.d. 1.52 96.3colloform textures 294Z1A_B 1 31 n.d. 0.39 1.54 0.39 n.d. 0.04 n.d. 58.3 0.04 1.31 1.05 n.d. n.d. 3.20 97.5294Z1A_D 1 31 n.d. 0.27 1.85 0.32 n.d. n.d. n.d. 62.2 0.05 0.40 0.79 n.d. n.d. 1.48 98.4294Z1A_F 1 31 n.d. 0.43 1.58 0.23 n.d. n.d. n.d. 59.0 0.05 1.20 1.11 n.d. n.d. 2.47 96.7294Z1A_S 1 31 n.d. 0.63 1.56 0.20 n.d. 0.01 n.d. 57.1 0.04 1.17 1.13 n.d. n.d. 2.38 94.8294Z1A_J 1 31 n.d. 0.33 2.30 0.21 n.d. 0.01 n.d. 59.9 0.03 0.55 0.86 n.d. n.d. 1.40 96.5294Z1A_H 1 30 n.d. 0.33 2.35 0.17 n.d. n.d. n.d. 59.1 0.05 0.66 1.00 n.d. n.d. 2.04 95.6294Z1A_N 1 31 n.d. 0.22 1.70 0.28 n.d. n.d. n.d. 60.9 0.07 0.40 0.87 n.d. n.d. 1.48 96.6294Z1A_M 1 33 n.d. 2.45 4.24 0.09 n.d. n.d. n.d. 54.5 0.06 0.79 0.73 n.d. n.d. 1.92 97.6294Z1A_L 1 30 n.d. 0.54 2.31 0.16 n.d. n.d. n.d. 59.2 0.05 0.86 1.10 n.d. n.d. 2.19 96.8mean 31 0.01 0.49 1.53 0.25 0.03 0.01 0.03 57.5 0.1 0.9 1.2 0.3 n.d. 1.8 94.7std dev 0.9 0.01 0.39 0.93 0.06 0.02 0.00 0.02 3.97 0.01 0.33 0.72 0.18 n.d. 0.51 2.80mean coll 31 n.d. 0.62 2.16 0.23 n.d. 0.02 n.d. 58.9 0.0 0.8 1.0 n.d. n.d. 2.1 96.7rectangular boxwork 294Z1-d-b1-2 3 31 0.02 0.53 3.59 0.20 0.10 n.d. 0.13 57.5 0.02 1.06 1.11 0.80 n.d. 1.53 97.2294Z1-d-b1-5 3 32 n.d. 0.60 3.65 0.20 0.11 n.d. 0.11 56.1 0.09 1.13 1.24 0.67 n.d. 1.73 97.3294Z1-d-b1-6 3 32 n.d. 0.53 3.50 0.20 0.10 n.d. 0.11 56.8 0.07 1.13 1.15 0.70 n.d. 1.78 97.8

Appendix D

256

Table D1 continued: EMPA results of goethite- and hematite-rich zones in ROM and LPO.

Label col. O Mg Al Si S Ca K Mn Fe Co Cu Zn As Hg Pb Σ wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%rectangular boxwork 294Z1-d-a3-14 3 35 0.01 0.59 3.06 0.20 0.02 n.d. 0.05 48.5 0.04 2.50 2.70 0.74 n.d. 1.34 95.1294Z1-d-a3-15 3 35 n.d. 0.61 2.92 0.20 0.03 n.d. 0.07 49.7 0.06 2.10 2.48 0.59 n.d. 1.24 95.4triangular boxwork 294Z1-d-a1-9 3 32 0.01 0.32 2.86 0.18 0.03 n.d. 0.08 55.9 0.06 1.16 1.05 0.30 n.d. 2.17 96.5294Z1-d-a1-10 3 34 0.01 0.65 3.03 0.19 0.01 n.d. 0.06 51.6 0.07 1.84 1.44 0.42 n.d. 1.54 95.2294Z1-d-a1-11 3 32 n.d. 0.29 2.95 0.14 0.05 n.d. 0.08 56.7 0.04 1.18 0.92 0.28 n.d. 2.15 96.8euhedral boxwork 294Z1B_B 3 33 n.d. 1.72 3.82 0.09 n.d. n.d. n.d. 47.9 0.02 2.40 2.90 n.d. n.d. 0.92 93.2294Z1B_R 3 30 n.d. 1.79 3.53 0.26 n.d. 0.02 n.d. 45.3 0.06 2.36 2.70 n.d. n.d. 0.87 86.7294Z1B_S 3 32 n.d. 1.80 3.57 0.14 n.d. n.d. n.d. 46.5 0.04 2.37 2.76 n.d. n.d. 0.81 89.5294Z1B_U 3 30 n.d. 1.93 4.17 0.14 n.d. 0.01 n.d. 45.2 0.05 2.37 2.79 n.d. n.d. 0.78 87.6294Z1B_U2 3 34 n.d. 1.65 3.85 0.11 n.d. 0.01 n.d. 45.5 0.04 2.18 2.73 n.d. n.d. 0.87 90.8294Z1B_V 3 31 n.d. 1.94 4.14 0.17 n.d. 0.01 n.d. 44.2 0.03 2.33 3.07 n.d. n.d. 0.76 87.6294Z1B_P 3 33 n.d. 1.45 3.68 0.10 n.d. n.d. n.d. 45.3 0.02 1.98 2.96 n.d. n.d. 0.86 89.7294Z1B_Y 3 34 n.d. 1.61 3.75 0.09 n.d. n.d. n.d. 45.4 0.05 2.07 2.79 n.d. n.d. 0.83 90.6294Z1B_Z1 3 33 n.d. 1.59 3.83 0.12 n.d. n.d. n.d. 44.4 0.03 2.07 2.96 n.d. n.d. 0.86 88.4euhedral boxwork 294Z1-d-a4-17 3 33 0.02 0.45 2.76 0.27 0.01 n.d. 0.11 54.9 0.04 1.37 0.95 0.46 n.d. 1.65 96.0294Z1-d-a4-20 3 34 n.d. 0.46 3.54 0.07 0.03 n.d. 0.03 52.6 0.04 2.64 2.22 0.30 n.d. 1.41 97.2euhedral boxwork 294R1-d-a2-9 3 34 0.01 0.51 2.81 0.28 0.06 n.d. 0.10 54.9 0.06 1.48 1.21 0.23 n.d. 1.26 96.4294R1-d-a2-11 3 33 0.02 0.58 2.72 0.31 0.04 n.d. 0.15 56.2 0.06 1.25 0.96 0.27 n.d. 1.21 96.6294R1-d-a2-13 3 33 0.04 0.55 2.71 0.29 0.04 n.d. 0.14 55.5 0.05 1.37 1.02 0.26 n.d. 1.24 96.7colloform textures 294Z1A_C 3 34 n.d. 2.80 2.02 0.28 n.d. n.d. n.d. 51.0 0.04 1.60 2.27 n.d. n.d. 1.49 95.2294Z1A_E 3 32 n.d. 1.47 2.89 0.15 n.d. n.d. n.d. 53.9 0.06 1.52 1.43 n.d. n.d. 1.65 94.2294Z1A_G 3 34 n.d. 2.06 3.49 0.13 n.d. n.d. n.d. 47.8 0.02 2.86 2.23 n.d. n.d. 1.26 93.7294Z1A_R 3 32 n.d. 1.36 3.02 0.11 n.d. 0.01 n.d. 54.6 0.06 1.46 1.36 n.d. n.d. 1.91 95.0294Z1A_I 3 34 n.d. 1.86 2.85 0.15 n.d. n.d. n.d. 51.5 0.05 1.73 2.66 n.d. n.d. 1.78 96.7294Z1A_I2 3 32 n.d. 1.52 2.24 0.22 n.d. n.d. n.d. 54.3 0.07 1.41 3.13 n.d. n.d. 1.83 96.9294Z1A_K 3 37 n.d. 4.08 2.07 0.30 n.d. n.d. n.d. 46.2 0.05 1.58 3.40 n.d. n.d. 1.09 95.5mean 37 0.02 1.0 3.4 0.2 0.0 n.d. 0.1 50.8 0.0 1.8 2.0 0.46 n.d. 1.3 93.5std dev 1.5 0.01 0.61 0.46 0.07 0.03 0.00 0.03 4.89 0.02 0.53 0.84 0.20 n.d. 0.43 3.76mean coll 33 n.d. 2.2 2.7 0.2 n.d. n.d. n.d. 51.3 0.1 1.7 2.4 n.d. n.d. 1.6 95.3triangular boxwork 4F2-d-A-8 1 31 n.d. 0.06 1.13 0.19 0.01 n.d. 0.07 55.3 0.09 0.78 0.33 0.24 n.d. 1.86 91.54F2-d-A-11 1 31 0.02 0.03 1.14 0.24 0.04 n.d. 0.10 58.4 0.00 0.88 0.23 0.41 n.d. 1.86 94.14F2-d-A-14 1 31 n.d. 0.04 1.05 0.21 0.02 n.d. 0.08 58.7 0.06 1.01 0.45 0.20 n.d. 2.14 94.9triangular boxwork 4F2-d-B-15 1 30 0.01 0.07 1.13 0.20 0.03 n.d. 0.06 61.3 0.07 0.74 0.37 0.14 n.d. 1.35 95.54F2-d-B-19 1 30 n.d. 0.04 1.02 0.21 0.06 n.d. 0.08 60.5 0.09 0.36 0.33 0.16 n.d. 0.87 94.04F2-d-B-21 1 30 n.d. 0.04 0.99 0.23 0.01 n.d. 0.09 58.8 0.06 0.86 0.32 0.27 n.d. 1.73 93.4open rib texture 4F2-d-C-22 1 29 0.02 0.04 0.69 0.18 0.04 n.d. 0.07 60.9 0.07 0.31 0.54 0.09 n.d. 0.67 92.54F2-d-C-23 1 30 0.02 0.05 0.76 0.16 0.03 n.d. 0.14 63.6 0.07 0.26 0.55 0.07 n.d. 0.59 95.84F2-d-C-26 1 29 n.d. 0.04 0.67 0.19 0.03 n.d. 0.13 61.4 0.08 0.24 0.67 0.16 n.d. 0.71 93.6mean 30 n.d. 0.05 0.95 0.20 0.03 n.d. 0.09 59.9 0.07 0.60 0.42 0.19 n.d. 1.31 93.9std dev 0.8 0.00 0.01 0.18 0.02 0.02 n.d. 0.03 2.2 0.02 0.29 0.13 0.10 n.d. 0.57 1.3

Appendix D

257


Label col. O Mg Al Si S Ca K Mn Fe Co Cu Zn As Hg Pb Σ wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%triangular boxwork 4F2-d-A-9 3 33 0.02 0.11 0.74 0.23 0.01 n.d. 0.12 51.4 0.06 0.28 1.29 0.08 n.d. 0.28 87.84F2-d-A-10 3 31 n.d. 0.14 0.95 0.31 0.02 n.d. 0.08 55.1 0.03 0.56 0.87 0.04 n.d. 0.99 90.54F2-d-A-12 3 34 0.01 0.05 0.59 0.27 0.02 n.d. 0.01 55.0 0.08 0.20 1.83 0.03 n.d. 0.27 91.84F2-d-A-13 3 36 n.d. 0.22 0.89 0.26 0.03 n.d. 0.00 57.6 0.12 0.49 1.06 0.02 n.d. 0.44 96.8triangular boxwork 4F2-d-B-16 3 36 n.d. 0.12 1.08 0.27 0.04 n.d. 0.04 54.2 0.03 0.57 1.88 0.10 n.d. 0.17 94.14F2-d-B-20 3 35 0.02 0.49 1.18 0.33 0.01 n.d. 0.02 50.9 0.04 1.35 0.97 0.12 n.d. 0.10 90.1open rib texture 4F2-d-C-24 3 31 n.d. 0.05 0.64 0.22 0.01 n.d. 0.02 62.2 0.06 0.24 1.07 0.05 n.d. 0.56 96.54F2-d-C-25 3 30 0.01 0.05 0.68 0.24 0.03 n.d. 0.06 57.6 0.11 0.26 0.66 0.15 n.d. 0.74 90.24F2-d-C-27 3 31 0.03 0.05 0.48 0.24 0.02 n.d. 0.05 59.7 0.03 0.19 0.91 0.08 n.d. 0.76 93.0mean 33 0.02 0.14 0.80 0.26 0.02 n.d. 0.04 56.0 0.06 0.46 1.17 0.07 n.d. 0.48 92.3std dev 2.1 0.01 0.13 0.22 0.03 0.01 n.d. 0.04 3.48 0.03 0.34 0.40 0.04 n.d. 0.29 2.86colloform textures 79-d-B86 1 31 n.d. 0.08 1.18 0.16 0.03 n.d. 0.07 52.2 0.03 1.59 2.22 1.77 n.d. 0.20 90.179-d-B89 1 33 n.d. 0.06 1.22 0.19 0.05 n.d. 0.03 53.1 0.05 1.71 2.32 1.78 n.d. 0.47 93.7colloform textures 79b_12 1 30 n.d. n.d. 0.31 0.21 n.d. n.d. n.d. 60.1 0.04 0.84 1.70 1.83 n.d. 0.74 95.179b_13 1 30 n.d. 0.01 0.31 0.21 n.d. n.d. n.d. 59.8 0.03 0.84 1.66 1.74 n.d. 0.69 95.3mean 31 n.d. 0.05 0.76 0.19 0.04 n.d. 0.05 56.3 0.04 1.24 1.98 1.78 n.d. 0.5 93.6std dev 1 n.d. 0.03 0.45 0.02 0.01 n.d. 0.02 3.67 0.01 0.41 0.30 0.03 n.d. 0.2 2.1 trellis-work texture 79-d-C81 3 33 n.d. 0.06 0.91 0.18 0.04 n.d. n.d. 55.3 0.07 1.67 2.13 1.88 n.d. 0.20 95.379-d-C82 3 32 n.d. 0.05 0.94 0.20 0.05 n.d. 0.01 55.7 0.05 1.71 2.30 2.06 n.d. 0.27 94.779-d-C83 3 33 n.d. 0.12 0.87 0.20 0.06 n.d. 0.05 54.9 0.06 1.78 2.00 1.98 n.d. 0.19 94.579-d-C84 3 31 n.d. 0.08 0.89 0.19 0.05 n.d. 0.02 54.9 0.03 1.60 2.14 2.14 n.d. 0.10 92.879-d-C85 3 29 n.d. 0.07 0.88 0.20 0.03 n.d. 0.04 54.6 0.07 1.62 2.12 2.17 n.d. 0.35 90.8trellis-work texture 79e_3 3 23 n.d. 0.06 0.73 0.12 n.d. n.d. n.d. 46.8 0.04 1.30 1.70 0.92 n.d. 0.23 74.879e_4 3 21 n.d. 0.04 0.74 0.11 n.d. n.d. n.d. 43.4 0.03 1.21 1.49 0.94 n.d. 0.22 69.079e_5 3 24 n.d. 0.05 0.74 0.14 n.d. n.d. n.d. 49.7 0.06 1.47 1.67 1.12 n.d. 0.23 79.579e_6 3 25 n.d. 0.06 0.75 0.12 n.d. n.d. n.d. 47.7 0.03 1.50 1.92 1.33 n.d. 0.19 78.8colloform textures 79-d-B90 3 23 n.d. 0.11 1.16 0.13 0.06 n.d. 0.02 51.3 0.02 1.76 2.68 2.11 n.d. 0.37 82.779-d-B91 3 21 n.d. 0.13 1.24 0.13 0.04 n.d. 0.03 48.8 0.05 1.38 3.61 2.97 n.d. 0.67 80.0colloform textures 79b_2 3 34 n.d. 0.18 0.24 0.21 n.d. n.d. n.d. 50.4 0.04 1.35 2.91 2.62 n.d. 0.71 92.479b_3 3 34 n.d. 0.45 0.25 0.09 n.d. n.d. n.d. 50.4 0.03 1.24 2.68 3.06 n.d. 1.02 92.879b_4 3 34 n.d. 0.58 0.25 0.18 n.d. n.d. n.d. 49.9 0.03 1.20 2.88 2.62 n.d. 0.90 92.179b_5 3 32 n.d. 0.50 0.32 0.08 n.d. n.d. n.d. 51.8 0.04 1.41 2.58 2.63 n.d. 0.73 92.279b_6 3 34 n.d. 0.70 0.26 0.25 n.d. n.d. n.d. 49.8 0.04 1.21 2.84 2.65 n.d. 0.75 92.279b_7 3 34 n.d. 0.53 0.29 0.14 n.d. n.d. n.d. 50.7 0.03 1.38 2.65 2.81 n.d. 0.75 93.579b_8 3 34 n.d. 0.31 0.23 0.12 n.d. n.d. n.d. 50.0 0.06 1.31 2.52 2.94 n.d. 1.11 92.5mean 28 n.d. 0.07 0.83 0.16 0.05 n.d. 0.03 51.4 0.05 1.54 1.94 1.62 n.d. 0.22 85.6stabwn 4 n.d. 0.02 0.08 0.04 0.01 n.d. 0.01 4.3 0.02 0.18 0.25 0.50 n.d. 0.06 9.5 mean coll 31 n.d. 0.39 0.47 0.15 0.05 n.d. 0.02 50.3 0.04 1.36 2.82 2.71 n.d. 0.78 n.d.

Appendix D

258


Label col. O Mg Al Si S Ca K Mn Fe Co Cu Zn As Hg Pb Σ wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%cellular sponge 82-d-A2-30 1 29 n.d. 0.08 1.02 0.07 0.14 n.d. 0.08 56.1 0.07 2.47 3.08 1.87 n.d. 0.38 94.082-d-A-32 1 27 n.d. 0.13 1.00 0.02 0.03 n.d. 0.03 54.5 0.05 2.15 3.45 2.57 n.d. 0.20 90.682-d-A-36 1 26 n.d. 0.18 1.06 0.01 0.05 n.d. 0.03 50.5 0.03 2.13 2.61 1.94 n.d. 0.27 84.682-d-A-38 1 28 n.d. 0.11 0.96 0.01 0.07 n.d. 0.04 53.4 0.08 2.32 3.28 2.42 n.d. 0.33 90.9mean 27 n.d. 0.13 1.01 0.03 0.07 n.d. 0.05 53.6 0.06 2.27 3.10 2.20 n.d. 0.29 90.0std dev 1 n.d. 0.04 0.04 0.03 0.04 n.d. 0.02 2.0 0.02 0.14 0.32 0.30 n.d. 0.07 3.43cellular sponge 82-d-A-33 3 22 n.d. 0.22 0.87 0.02 0.02 n.d. 0.02 49.0 0.06 2.43 4.38 2.49 n.d. 0.46 82.482-d-A-34 3 28 n.d. 0.13 0.94 0.01 0.04 n.d. 0.05 52.5 0.02 2.29 3.49 2.57 n.d. 0.47 90.982-d-A-37 3 25 n.d. 0.18 0.97 0.01 0.06 n.d. 0.10 49.9 0.03 2.59 3.74 2.92 n.d. 0.42 85.6mean 25 n.d. 0.18 0.93 0.01 0.04 n.d. 0.06 50.5 0.04 2.4 3.9 2.7 n.d. 0.5 86.3std dev 2 n.d. 0.04 0.04 0.01 0.02 n.d. 0.03 1.5 0.02 0.1 0.4 0.2 n.d. n.d. 3.5 cellular sponge 268a_B_a 1 33 n.d. n.d. 6.54 n.d. n.d. n.d. 0.13 44.9 0.01 1.61 3.71 n.d. 0.09 5.50 95.5268a_B_f 1 33 n.d. n.d. 5.36 n.d. n.d. 0.01 0.08 47.1 0.07 1.58 3.33 n.d. 0.09 5.31 96.3268a_B_c 1 38 n.d. 0.03 3.37 0.09 n.d. n.d. 0.20 48.3 0.07 1.45 5.25 n.d. 0.08 1.30 97.6268a_B_d 1 37 n.d. n.d. 4.90 0.02 n.d. n.d. 0.09 48.0 0.04 1.82 4.76 n.d. 0.05 1.46 98.0cellular sponge 268a-d-B-77 1 32 0.02 n.d. 4.50 0.04 0.16 n.d. 0.04 44.8 0.02 1.52 3.21 n.d. n.d. 3.54 90.1268a-d-B-81 1 33 0.02 n.d. 4.78 0.01 0.20 n.d. n.d. 45.4 0.03 1.57 3.29 0.02 n.d. 3.32 90.9cubic boxwork 268a_A2_a 1 33 n.d. n.d. 2.07 0.06 n.d. n.d. 0.05 57.3 0.09 0.48 2.11 n.d. 0.03 1.22 96.8268a_A2_b 1 34 n.d. n.d. 2.16 0.05 n.d. n.d. 0.02 56.9 0.03 0.48 2.12 n.d. 0.01 1.21 97.3268a_A2_g 1 34 n.d. n.d. 1.87 0.10 n.d. n.d. n.d. 58.6 0.04 0.51 1.86 n.d. 0.00 0.68 97.5268a_A2_h 1 34 n.d. 0.01 2.06 0.11 n.d. n.d. n.d. 57.0 0.05 0.99 2.37 n.d. 0.08 0.79 97.4cubic boxwork 268a-d-A-66 1 30 n.d. n.d. 1.17 0.05 0.03 n.d. 0.03 56.4 0.06 0.54 1.89 0.28 n.d. 1.81 92.4colloform textures 268d-d-A-1 1 33 n.d. 0.05 2.35 0.04 0.04 n.d. 0.04 51.9 0.03 0.39 2.96 0.28 n.d. 0.91 92.1268d-d-A-4 1 32 n.d. 0.06 2.16 0.06 0.01 n.d. 0.03 50.7 0.03 0.51 3.11 0.31 n.d. 1.05 89.9268d-d-A-5 1 32 n.d. 0.07 2.14 0.07 n.d. n.d. 0.02 50.6 0.02 0.62 2.65 0.41 n.d. 1.06 89.2mean 34 0.02 0.05 3.24 0.06 0.09 n.d. 0.07 51.3 0.04 1.01 3.04 0.26 0.05 2.08 94.4std dev 2 n.d. 0.02 1.54 0.03 0.07 n.d. 0.05 4.7 0.02 0.51 0.95 0.12 n.d. 1.54 3.1 mean coll 32 n.d. 0.06 2.22 0.05 0.02 n.d. 0.03 51.1 0.03 0.51 2.91 0.33 n.d. 1.01 90.4cellular sponge 268a_B_b 3 33 n.d. n.d. 12.0 n.d. n.d. 0.06 0.07 41.9 0.07 1.27 3.40 n.d. n.d. 3.02 95.2268a_B_e 3 34 n.d. n.d. 9.48 0.02 n.d. 0.05 0.04 42.8 0.07 1.12 3.34 n.d. n.d. 3.35 94.7cellular sponge 268a-d-B-76 3 32 0.01 n.d. 3.68 0.06 0.07 n.d. 0.08 46.2 0.06 1.38 3.43 0.11 n.d. 0.89 87.9268a-d-B-78 3 36 0.01 0.03 8.45 0.00 0.41 n.d. 0.09 40.6 0.03 1.10 2.87 n.d. n.d. 3.31 92.6268a-d-B-79 3 34 0.03 0.02 8.43 0.01 0.41 n.d. 0.05 40.4 0.05 1.14 3.00 0.01 n.d. 3.07 90.3268a-d-B-80 3 34 0.01 0.01 3.71 0.03 0.08 n.d. 0.08 47.4 0.03 1.47 3.84 0.10 n.d. 0.81 91.5cubic boxwork 268a_A2_c 3 38 n.d. n.d. 3.42 0.08 n.d. n.d. 0.23 49.8 0.06 1.30 4.59 n.d. 0.08 0.57 97.7268a_A2_d 3 38 n.d. 0.01 3.00 0.03 n.d. n.d. 0.23 48.8 0.05 1.07 4.04 n.d. 0.00 0.63 95.1268a_A2_e 3 38 n.d. n.d. 4.39 0.07 n.d. 0.01 0.06 48.6 0.03 1.76 4.47 n.d. 0.12 0.64 97.6268a_A2_f 3 38 n.d. n.d. 2.52 0.35 n.d. n.d. 0.05 49.6 0.02 1.32 4.70 n.d. 0.17 0.54 97.3268a_A2_j 3 37 n.d. n.d. 4.60 0.09 n.d. n.d. 0.02 47.4 0.03 3.41 4.14 n.d. 0.06 0.40 97.5268a_A2_k 3 37 n.d. n.d. 5.05 0.05 n.d. n.d. 0.03 47.0 0.05 3.65 4.50 n.d. 0.00 0.60 97.5

Appendix D

259


Label col. O Mg Al Si S Ca K Mn Fe Co Cu Zn As Hg Pb Σ wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%cubic boxwork 268a-d-C-69 3 33 n.d. 0.01 2.35 0.06 0.07 n.d. 0.17 46.3 0.03 2.70 0.17 0.32 n.d. 1.12 85.7268a-d-C-70 3 34 n.d. n.d. 4.33 0.06 0.09 n.d. 0.19 43.3 0.04 2.35 3.87 0.25 n.d. 1.28 89.6268a-d-C-71 3 30 n.d. n.d. 2.16 0.08 0.06 n.d. 0.02 56.1 0.06 0.90 2.37 0.26 n.d. 1.75 93.8268a-d-C-72 3 28 0.03 0.01 1.70 0.08 0.06 n.d. 0.11 55.4 0.05 0.87 1.83 0.29 n.d. 2.16 91.0cubic boxwork 268a-d-A-67 3 34 n.d. n.d. 2.40 0.10 0.04 n.d. 0.04 46.1 0.01 1.67 3.79 0.23 n.d. 0.98 89.7colloform textures 268d-d-A-2 3 26 0.03 0.17 1.92 0.06 n.d. n.d. 0.02 54.2 0.05 0.61 2.12 0.30 n.d. 0.52 85.9268d-d-A-3 3 25 0.01 0.28 2.21 0.11 0.02 n.d. 0.02 51.4 0.06 0.68 1.73 0.26 n.d. 0.74 82.5268d-d-A-6 3 26 0.01 0.32 2.30 0.17 0.02 n.d. 0.05 55.0 0.03 0.85 1.48 0.21 n.d. 0.70 87.1268d-d-A-7 3 26 n.d. 0.36 2.50 0.16 0.01 n.d. 0.00 56.2 0.03 0.99 1.35 0.30 n.d. 0.56 88.2direct surrounding of hematite 268a_C_b 3 37 n.d. n.d. 2.80 0.06 n.d. n.d. 1.54 48.9 0.05 1.02 4.23 n.d. n.d. 0.81 96.2268a_C_c 3 35 n.d. 0.03 3.93 0.08 n.d. n.d. 0.12 51.3 0.02 1.89 3.76 n.d. n.d. 1.40 97.7268a_C_e 3 31 n.d. 0.06 3.38 0.06 n.d. 0.01 0.06 41.6 0.06 1.30 4.21 n.d. 0.05 0.99 83.1268a_C_f 3 35 n.d. 0.04 4.46 0.08 n.d. n.d. 0.09 50.8 0.06 1.79 4.46 n.d. n.d. 1.46 98.7268a_C_d 3 26 n.d. 0.09 4.22 0.09 n.d. n.d. 0.14 48.7 0.06 1.49 3.90 n.d. 0.04 0.90 85.5268a_C_g 3 24 n.d. 0.11 5.49 0.10 n.d. 0.02 0.01 43.7 0.03 1.34 4.86 n.d. n.d. 0.84 80.9268a_C_h 3 24 n.d. 0.11 5.10 0.13 n.d. 0.01 0.06 38.2 0.02 1.24 5.16 n.d. 0.01 0.62 74.5direct surrounding of hematite 268a_A1_e 3 34 n.d. 0.01 4.75 0.03 n.d. n.d. 0.25 50.4 0.03 2.00 5.09 n.d. 0.08 0.60 97.0268a_A1_f 3 34 n.d. 0.01 4.67 0.05 n.d. n.d. 0.53 49.3 0.01 2.18 5.09 n.d. 0.17 0.62 96.6268a_A1_g 3 36 n.d. n.d. 3.33 0.12 n.d. n.d. 1.05 50.4 0.05 1.06 4.75 n.d. n.d. 0.78 97.7268a_A1_h 3 38 n.d. 0.01 3.39 0.08 n.d. 0.01 2.51 47.5 0.06 1.13 4.34 n.d. n.d. 0.47 97.1mean 35 0.02 0.01 4.80 0.07 0.14 0.04 0.09 46.9 0.04 1.67 3.43 0.20 0.07 1.48 93.2std dev 2 0.01 0.01 2.88 0.08 0.14 0.02 0.07 4.3 0.02 0.82 1.12 0.10 0.06 1.05 3.7 mean coll 30 0.02 0.12 3.63 0.09 0.02 0.01 0.43 49.2 0.04 1.30 3.77 0.27 0.07 0.80 89.9parallel closed to open boxwork 26A_3_3 1 32 n.d. 0.06 1.20 0.36 0.02 n.d. 0.06 58.9 0.06 1.51 0.83 0.16 n.d. 1.66 96.626A_3_4 1 32 n.d. 0.07 1.16 0.34 n.d. n.d. 0.03 58.0 0.05 1.56 0.88 0.21 n.d. 1.62 95.526A_3_8 1 32 n.d. 0.08 1.11 0.35 0.01 n.d. 0.05 55.4 0.04 1.53 0.83 0.18 n.d. 1.49 93.226A_3_9 1 32 n.d. 0.05 1.17 0.36 0.03 n.d. 0.01 56.8 0.06 1.67 0.63 0.20 n.d. 1.74 94.326A_3_15 1 32 n.d. 0.05 1.18 0.37 n.d. n.d. 0.03 56.9 0.06 1.51 0.86 0.08 n.d. 1.65 94.4mean 32 n.d. 0.06 1.16 0.35 0.02 n.d. 0.04 57.2 0.05 1.56 0.81 0.16 n.d. 1.63 94.8stabwn 0.2 n.d. 0.01 0.03 0.01 0.01 n.d. 0.02 1.17 0.01 0.06 0.09 0.05 n.d. 0.08 1.18parallel closed to open boxwork 26A_3_1 3 33 n.d. 0.06 0.90 0.21 0.01 n.d. 0.04 53.1 0.10 1.09 2.60 0.23 n.d. 0.63 92.926A_3_2 3 34 0.01 0.04 1.12 0.29 0.00 n.d. 0.04 51.9 0.08 1.06 2.79 0.19 n.d. 0.47 91.926A_3_5 3 35 0.01 0.08 1.18 0.30 0.01 n.d. 0.07 52.8 0.07 1.13 3.01 0.19 n.d. 0.55 94.726A_3_6 3 35 n.d. 0.05 1.15 0.29 0.01 n.d. 0.02 53.6 0.06 0.98 3.14 0.16 n.d. 0.45 94.726A_3_7 3 35 0.01 0.04 1.03 0.31 0.00 n.d. 0.01 54.9 0.07 0.94 2.96 0.11 n.d. 0.53 95.3mean 35 n.d. 0.05 1.08 0.28 0.01 n.d. 0.04 53.3 0.07 1.04 2.90 0.18 n.d. 0.5 93.9stabwn 0.5 n.d. 0.01 0.10 0.03 n.d. n.d. 0.02 0.98 0.01 0.07 0.19 0.04 n.d. 0.1 1.3 parallel closed to open boxwork 209-c-d-B-55 1 28 n.d. 0.02 0.43 0.25 0.08 n.d. 0.03 55.9 0.04 1.90 1.64 2.33 n.d. 0.40 91.1209-c-d-B-60 1 27 n.d. 0.05 0.35 0.15 0.07 n.d. 0.03 58.9 0.05 1.36 1.73 1.89 n.d. 0.44 92.4

Appendix D

260


Label col. O Mg Al Si S Ca K Mn Fe Co Cu Zn As Hg Pb Σ wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%parallel closed to open boxwork 209-c-d-A-49 1 28 n.d. n.d. 0.30 0.14 0.08 n.d. 0.04 59.7 0.05 1.26 1.71 1.66 n.d. 0.37 93.3209-c-d-A-52 1 28 n.d. 0.01 0.32 0.18 0.08 n.d. 0.04 59.3 0.07 1.34 1.63 1.85 n.d. 0.28 93.0209-c-d-A-53 1 28 n.d. 0.02 0.32 0.14 0.09 n.d. 0.01 59.7 0.06 1.39 1.61 1.64 n.d. 0.42 93.6mean 28 n.d. 0.02 0.34 0.17 0.08 n.d. 0.03 58.7 0.06 1.45 1.67 1.87 n.d. 0.4 92.7std dev 0.3 n.d. 0.02 0.04 0.04 0.01 n.d. 0.01 1.44 0.01 0.23 0.05 0.25 n.d. 0.1 0.9 parallel closed to open boxwork 209-c-d-B-56 3 32 n.d. 0.12 0.44 0.33 0.05 n.d. n.d. 51.8 0.07 2.69 1.33 1.56 n.d. 0.21 90.1209-c-d-B-57 3 28 n.d. 0.05 0.36 0.22 0.08 n.d. 0.06 53.1 0.06 1.72 1.88 2.34 n.d. 0.36 87.9209-c-d-B-58 3 33 n.d. 0.11 0.47 0.35 0.07 n.d. 0.01 53.1 0.03 2.89 1.49 1.76 n.d. 0.20 93.4209-c-d-B-59 3 32 n.d. 0.09 0.41 0.26 0.09 n.d. n.d. 50.6 0.07 2.77 1.23 1.57 n.d. 0.23 89.3212_2B1-b 1 32 n.d. 0.02 0.05 0.32 n.d. 0.02 0.17 62.2 0.05 0.82 n.d. n.d. 0.16 0.49 95.9212_2B1-e 1 32 n.d. 0.02 0.05 0.29 n.d. 0.00 0.09 61.1 0.06 0.67 0.11 n.d. 0.04 0.61 94.5212_2B1-f 1 32 n.d. 0.03 0.09 0.46 n.d. 0.00 0.16 60.6 0.03 0.88 0.07 n.d. 0.07 0.40 93.9212_2B1-h 1 32 n.d. 0.01 0.06 0.30 n.d. 0.01 0.09 60.8 0.03 0.78 n.d. n.d. n.d. 0.57 93.8open rib texture 212_2A5_a 1 32 n.d. 0.19 0.66 0.38 n.d. n.d. n.d. 60.6 0.08 0.70 0.36 n.d. n.d. 0.73 95.7212_2A5_b 1 32 n.d. 0.26 0.82 0.47 n.d. 0.01 n.d. 58.9 0.04 0.77 0.42 n.d. n.d. 0.67 94.0212_2A5_c 1 32 n.d. 0.23 0.68 0.40 n.d. n.d. n.d. 60.1 0.08 0.76 0.44 n.d. n.d. 0.80 95.5212_2A5_d 1 32 n.d. 0.25 0.74 0.44 n.d. n.d. n.d. 59.9 0.04 0.81 0.47 n.d. n.d. 0.75 95.4212_2A5_e 1 33 n.d. 0.25 0.85 0.40 n.d. n.d. n.d. 58.9 0.07 0.77 0.54 n.d. n.d. 0.68 94.9open rib texture 212/2-d-a2-18 1 33 n.d. n.d. n.d. 0.30 0.01 n.d. 0.09 60.7 0.06 0.60 n.d. 0.68 n.d. 0.55 94.6212/2-d-a2-19 1 32 n.d. n.d. n.d. 0.29 0.03 n.d. 0.09 60.2 0.06 0.70 n.d. 0.79 n.d. 0.51 94.7212/2-d-a2-20 1 33 n.d. n.d. n.d. 0.30 0.02 n.d. 0.10 61.2 0.07 0.72 n.d. 0.73 n.d. 0.28 95.7open rib texture 212/2-d-b4-1 1 31 n.d. 0.19 0.82 0.27 0.02 n.d. 0.01 62.8 0.04 0.64 0.44 0.44 n.d. 0.54 97.4212/2-d-b4-3 1 32 n.d. 0.17 0.82 0.28 0.02 n.d. 0.03 62.8 0.04 0.56 0.48 0.48 n.d. 0.32 97.9212/2-d-b4-5 1 32 0.01 0.17 0.74 0.28 0.01 n.d. 0.02 63.8 0.07 0.59 0.42 0.40 n.d. 0.61 99.0212/2-d-b4-2 1 31 0.03 0.17 1.72 0.20 0.05 n.d. 0.09 61.3 0.06 0.76 0.50 0.55 n.d. 0.62 97.4colloform textures 212_2A2_a 1 32 n.d. 0.09 0.77 0.33 n.d. 0.01 n.d. 62.2 0.04 0.75 0.28 n.d. n.d. 0.81 97.0212_2A2_e 1 34 n.d. 0.18 2.74 0.27 n.d. 0.01 n.d. 57.2 0.08 0.76 0.44 n.d. n.d. 0.80 96.0212_2A3_a 1 31 n.d. 0.09 0.72 0.33 n.d. n.d. n.d. 61.4 0.06 0.71 0.39 n.d. n.d. 0.69 95.4212_2A3_b 1 32 n.d. 0.13 1.05 0.29 n.d. 0.02 n.d. 61.0 0.09 0.68 0.32 n.d. n.d. 0.87 95.8212_2A3_d 1 32 n.d. 0.07 1.14 0.26 n.d. 0.01 n.d. 60.2 0.04 0.55 0.31 n.d. n.d. 0.79 94.9colloform textures 212/2-d-a1-14 1 31 n.d. n.d. n.d. 0.25 n.d. n.d. 0.11 60.8 0.05 0.54 n.d. n.d. n.d. 0.40 92.8212/2-d-a1-16 1 30 n.d. n.d. 0.01 0.30 n.d. n.d. 0.08 63.1 0.04 0.63 n.d. n.d. n.d. 0.46 94.4mean 32 n.d. 0.15 0.62 0.34 0.02 0.01 0.1 61.0 0.05 0.72 0.39 0.58 0.09 0.57 95.6std dev 0.4 0.01 0.09 0.45 0.08 0.01 0.01 0.05 1.33 0.02 0.09 0.15 0.14 0.05 0.14 1.49mean coll 31 n.d. 0.11 1.07 0.29 n.d. n.d. 0.09 60.8 0.06 0.66 0.35 n.d. n.d. 0.69 95.2parallel closed to open boxwork 212/2-d-b1-7 3 32 n.d. 0.18 1.26 0.23 0.02 n.d. 0.09 61.3 0.10 0.88 0.32 0.59 n.d. 0.66 97.7212/2-d-b1-9 3 32 0.02 0.16 1.20 0.23 0.03 n.d. 0.03 61.6 0.04 0.88 0.31 0.52 n.d. 0.44 97.6212/2-d-b1-12 3 31 n.d. n.d. 0.04 0.21 0.02 n.d. 0.09 59.5 0.06 0.83 0.03 n.d. n.d. 0.48 92.7

Appendix D

261


Label col. O Mg Al Si S Ca K Mn Fe Co Cu Zn As Hg Pb Σ wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%

parallel closed to open boxwork 212_2B1-a 3 30 n.d. 0.04 0.04 0.38 n.d. 0.02 0.08 63.0 0.07 0.74 0.09 n.d. 0.15 0.55 95.5212_2B1-c 3 32 n.d. 0.02 0.05 0.31 n.d. 0.03 0.13 61.1 0.05 0.65 0.02 n.d. n.d. 0.54 94.6212_2B1-d 3 32 n.d. 0.04 0.08 0.27 n.d. 0.01 0.05 61.8 0.10 0.72 0.03 n.d. 0.03 0.66 95.6212_2B1-g 3 32 n.d. 0.02 0.06 0.32 n.d. 0.01 0.08 60.9 0.05 0.66 0.00 n.d. 0.03 0.58 94.3colloform textures 212_2A2_b 3 38 n.d. 1.09 1.43 0.54 n.d. n.d. n.d. 50.4 0.04 1.46 1.15 n.d. n.d. 0.27 94.8212_2A2_c 3 37 n.d. 1.04 1.42 0.56 n.d. n.d. n.d. 50.5 0.03 1.43 1.12 n.d. n.d. 0.25 93.8212_2A2_d 3 35 n.d. 1.11 1.26 0.47 n.d. n.d. n.d. 49.4 0.02 1.56 1.06 n.d. n.d. 0.23 90.4212_2A3_c 3 38 n.d. 1.03 1.48 0.48 n.d. 0.01 n.d. 49.7 0.04 1.57 1.08 n.d. n.d. 0.29 93.5colloform textures 212/2-d-a1-13 3 35 n.d. n.d. 0.01 0.36 0.01 n.d. 0.08 52.3 0.05 0.92 0.08 n.d. n.d. 0.16 88.9212/2-d-a1-15 3 35 n.d. n.d. n.d. 0.37 0.02 n.d. 0.10 53.0 0.04 1.11 n.d. n.d. n.d. 0.31 90.0212/2-d-a1-17 3 35 n.d. n.d. n.d. 0.43 0.01 n.d. 0.12 53.4 0.04 1.20 n.d. n.d. n.d. 0.31 90.3mean 32 n.d. 0.47 0.69 0.37 0.02 0.02 0.09 56.3 0.05 1.04 0.44 0.56 0.07 0.41 93.5std dev 0.6 n.d. 0.49 0.65 0.11 0.01 0.01 0.03 5.19 0.02 0.33 0.48 0.03 0.06 0.16 2.7

mean coll 36 n.d. 1.07 1.12 0.46 0.01 0.01 0.10 51.3 0.04 1.32 0.90 n.d. n.d. 0.26 91.7

Appendix D

262

Table D2: List of abbreviations used in the boxplots in Chapter 6.

Analysis nr. Area analyzed Abbreviation in boxplots Chapter 6

rectangular boxwork 294Z1-d B1 A 294Z1-d A3 B triangular boxwork 294Z1-d A1 C 4F2-d A D 4F2-d B E euhedral grain boundary 294Z1 B F 294Z1-d A4 G 294R1-d A2 H trellis 79 D I 79 E J closed-open 26 A3 K 209c-d B L 209c-d A M 212-2-d B1 N 212-2 B1 O cellular sponge 82-d A2 P 268a-d B Q 268a B R open rib 212/2 A5 S 212/2-d A2 T 212/2-d B4 U 4F2-d C V cubic 268a-d C W 268a A2 X 268a-d A Y 72a A Z colloform 212-2-d A1 1 212-2 A2 2 294Z1 A 3 79-d B 4 79 B 5 268d-d A 6

Appendix D

263

Table D3: EMPA results of primary sulfides and minerals in the secondary sulfide ore.

Primary sulfides Mineral Sample Si Fe Mn Zn Cu Co As Pb S Total

wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%

pyrrhotite 166 <0.01 45.53 <0.01 <0.01 3.37 0.13 <0.01 - 53.24 102.3pyrrhotite 166 <0.01 44.11 0.02 0.02 6.36 0.05 <0.01 - 51.66 102.2sphalerite 137a <0.01 7.25 0.01 48.90 0.11 0.04 <0.01 - 32.78 89.07chalcopyrite 137a <0.01 27.91 <0.01 <0.01 28.79 0.02 0.03 - 34.30 90.99chalcopyrite 137a <0.01 27.64 <0.01 <0.01 27.65 0.02 <0.01 - 32.06 87.34chalcopyrite 34 <0.01 27.71 <0.01 <0.01 29.35 0.03 0.41 - 33.91 91.38galena 137a <0.01 0.03 <0.01 <0.01 0.08 0.03 - 69.84 13.73 83.57arsenopyrite 137a <0.01 28.04 0.03 <0.01 0.04 3.74 94.59 - 18.73 145.1arsenopyrite 34 <0.01 30.62 <0.01 0.05 1.35 0.68 83.32 - 20.86 136.8arsenopyrite 34 <0.01 31.33 0.04 0.03 0.24 0.71 84.45 - 21.77 138.5

Secondary sulfides, oxides, and sulfates in the secondary sulfide ore lenses

bornite 166 0.06 10.15 <0.01 <0.01 58.9 0.08 0.01 - 27.01 96.1bornite 166 <0.01 7.63 0.01 <0.01 63.7 0.05 <0.01 - 26.49 97.9covellite 34 <0.01 0.54 <0.01 <0.01 65.3 0.01 0.09 - 21.53 87.4covellite 34 <0.01 0.32 <0.01 <0.01 66.2 <0.01 0.1 - 22.34 88.8brochanite 166 0.38 0.32 <0.01 <0.01 57.6 <0.01 0.12 - 5.64 63.9brochanite 166 0.03 0.4 <0.01 <0.01 66.8 <0.01 0.09 - 7.54 74.7brochanite 166 0.17 0.04 0.02 <0.01 70.4 0.01 0.04 - 7.93 78.6brochanite 34 <0.01 0.33 <0.01 1.91 59.2 0.06 0.54 - 6.29 68.3brochanite 34 <0.01 0.1 0.02 1.42 54.8 0.05 0.44 - 5.98 62.8brochanite 34 0.03 0.06 <0.01 1.58 53.7 0.02 0.13 - 6.86 62.2brochanite 34 <0.01 0.86 0.01 2.07 54.6 0.03 1.31 - 6.5 65.3cuprite 166 <0.01 0.25 <0.01 <0.01 83.6 <0.01 0.26 - 0.1 83.9cuprite 166 0.01 0.02 <0.01 <0.01 85.7 <0.01 <0.010 - 0.019 85.7

Table D4: EMPA results of some chlorites in the supergene ore.

MgO Al2O3 SiO2 SO3 CaO MnO FeO CoO CuO ZnO CdO PbO Total

wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%

chlorite 19.7 <0.01 56.2 0.02 0.21 1.79 20.2 0.04 0.0065 0.15 <0.01 <0.01 98.28chlorite 16.2 0.27 55.8 0.02 10.9 0.32 10.5 0.01 0.0002 0.09 <0.01 <0.01 94.1chlorite 16.2 0.09 54.1 <0.01 0.55 1.08 23.5 0.04 0.0002 0.1 <0.01 0.05 95.6chlorite 16.4 0.02 54.5 <0.01 0.43 1.18 23.7 0.03 0.0002 0.15 <0.01 0.01 96.5chlorite 15.6 0.05 54.6 0.02 0.49 0.98 23.8 0.04 0.0319 0.25 <0.01 <0.01 95.8chlorite 16.2 0.05 54.2 0.07 0.51 1.04 23.7 0.01 0.0163 0.18 <0.01 0.04 96.1

mean 16.70 0.08 54.92 0.02 2.18 1.07 20.90 0.03 0.01 0.15 0.00 0.02 96.06

Appendix D

264

Table D5: EMPA results of coronadite.

Mineral Sample O Al Mn Fe Co Cu Zn As Pb Total wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% coronatite 212/2 27 <0.01 34.53 2.96 0.62 2.23 0.10 <0.01 28.62 96.3 coronatite 212/2 28 <0.01 35.39 3.19 0.65 1.97 0.12 <0.01 27.64 96.7 coronatite 212/2 28 <0.01 39.94 0.04 0.24 4.42 0.08 <0.01 25.50 98.2 coronatite 212/2 27 0.02 36.40 2.43 0.38 2.11 0.05 <0.01 27.99 96.6 coronatite 212/2 27 0.03 39.18 <0.01 0.37 4.69 0.05 <0.01 25.80 97.9 coronatite 212/2 29 0.01 35.93 4.80 0.33 2.46 0.08 <0.01 22.23 94.7 coronatite 212/2 26 0.02 39.33 0.31 0.14 4.53 0.04 <0.01 25.78 96.6 coronatite 212/2 28 0.01 39.90 0.01 0.20 4.29 0.02 <0.01 24.23 96.4 mean 27 0.01 37.57 1.72 0.37 3.34 0.07 <0.01 25.97 96.7

Table D6: EMPA results of plumbojarosite.

Mineral Sample O Al Si S Mn Fe Co Cu Zn As Pb Total

wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt%

pb-jarosite 212/2 33 0.05 0.06 9.95 <0.01 23.69 <0.01 1.14 0.40 <0.01 19.07 87.4 pb-jarosite 212/2 33 0.04 0.01 9.77 <0.01 24.24 0.03 0.95 0.02 <0.01 18.58 86.3 212-2 mean 33 0.04 0.03 9.86 <0.01 23.96 0.02 1.05 0.21 <0.01 18.82 86.8

pb-jarosite F1a 37 0.01 0.04 11.26 0.02 26.72 0.01 0.22 0.08 0.03 19.38 94.6 pb-jarosite F1a 38 0.01 0.01 10.77 <0.01 26.38 0.04 0.22 0.06 0.04 18.69 94.4

mean all 35 0.03 0.03 10.32 <0.01 25.00 0.02 0.72 0.16 0.01 18.91 89.9

Appendix D

265

Table D7: EMPA results of individual base metal minerals and goethite in their direct vicinity (For abbreviations see Appendix A, Table A1).

Analysis nr. Min.eral MgO Al2O3 SiO2 SO3 MnO Fe2O3 CoO CuO ZnO As2O5 PbO Total wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% Oliv-1-1 ol n.d. 0.12 n.d. n.d. n.d. n.d. n.d. 53.54 0.39 40.06 n.d. 94.1 Oliv-1-2 ol n.d. n.d. n.d. n.d. n.d. 0.22 n.d. 53.7 0.41 39.97 n.d. 94.3 Oliv-1-3 ol n.d. n.d. 0.13 n.d. n.d. n.d. n.d. 58.05 0.051 33.84 n.d. 92.1 Oliv-1-4 ol n.d. n.d. 0.18 n.d. n.d. 0.03 n.d. 57.65 0.25 34.26 n.d. 92.4 Oliv-1-5 ol n.d. n.d. n.d. n.d. n.d. 0.07 n.d. 53.97 0.23 39.62 n.d. 93.9 mean ol n.d. n.d. 0.16 n.d. n.d. 0.11 n.d. 55.38 0.27 37.55 n.d. 93.4 ideal ol 56.21 40.61 Oliv-2-1 geo n.d. 3.97 0.22 0.32 n.d. 63.01 n.d. 4.94 0.31 6.89 0.1 79.8 Oliv-2-2 geo n.d. 1.44 0.76 0.17 n.d. 64.65 0.15 1.754 0.9 4.94 n.d. 74.8 mean geo n.d. 2.71 0.49 0.25 n.d. 63.83 0.15 3.347 0.60 5.91 n.d. 77.26HEMV-1-1 hem n.d. n.d. 24.9 n.d. n.d. n.d. n.d. 1E-04 57.15 n.d. n.d. 82.1 HEMV-1-2 hem n.d. n.d. 25.26 n.d. n.d. n.d. n.d. 0.012 63.81 n.d. n.d. 89.1 HEMV-1-3 hem n.d. n.d. 25.98 n.d. n.d. n.d. n.d. 1E-04 62.75 n.d. n.d. 88.7 mean hem n.d. n.d. 25.38 n.d. n.d. n.d. n.d. 0.004 61.24 n.d. n.d. 86.62ideal hem 24.94 67.58 HEMV-1-4 geo n.d. n.d. 4.23 n.d. n.d. 62.72 n.d. 0.471 3.84 2.19 1.37 74.8 HEMV-1-5 geo n.d. n.d. 5.18 n.d. n.d. 62.37 n.d. 0.537 4.34 2.22 1.84 76.5 HEMV-1-6 geo n.d. n.d. 0.18 13.41 n.d. 23.69 n.d. 0.054 6.94 2.57 34.74 81.6 mean geo n.d. n.d. 3.1 13.41 n.d. 49.59 n.d. 0.354 5.04 2.33 12.65 77.63CHRV-1-1 chr n.d. n.d. 38.76 n.d. n.d. n.d. n.d. 51.37 0.31 n.d. n.d. 90.4 CHRV-1-2 chr n.d. n.d. 36.21 n.d. n.d. n.d. n.d. 52.87 0.46 n.d. n.d. 89.5 CHRV-1-3 chr n.d. n.d. 39.85 n.d. n.d. n.d. n.d. 51.72 0.26 n.d. n.d. 91.8 CHRV-1-4 chr n.d. n.d. 39.89 n.d. n.d. n.d. n.d. 41.95 4.85 n.d. 0.56 87.3 CHRV-1-6 chr n.d. n.d. 45.64 n.d. n.d. n.d. n.d. 48.21 3.08 n.d. 0.56 97.5 CHRV-1-7 chr n.d. n.d. 45.75 n.d. n.d. n.d. n.d. 47.98 3.43 n.d. 0.47 97.6 CHRV-1-8 chr n.d. n.d. 39.16 n.d. n.d. n.d. n.d. 50.61 0.33 n.d. n.d. 90.1 mean chr n.d. n.d. 40.75 n.d. n.d. n.d. n.d. 49.24 1.82 n.d. 0.53 92.04ideal chr 2.77 52.24 30.26 MALV-1-1 mal n.d. n.d. n.d. n.d. n.d. n.d. n.d. 69.63 0.22 n.d. n.d. 69.8 MALV-1-2 mal n.d. n.d. 0.11 n.d. n.d. 0.06 n.d. 69.1 0.32 n.d. 0.095 69.7 MALV-1-3 mal n.d. n.d. 0.1 n.d. n.d. 0.47 n.d. 69.27 0.36 n.d. 0.13 70.3 mean mal n.d. n.d. 0.10 n.d. n.d. 0.26 n.d. 69.33 0.3 n.d. 0.11 69.95ideal mal 71.95 MALV-1-4 geo n.d. 1 14.8 0.16 n.d. 50.38 n.d. 10.43 0.16 0.15 0.47 77.6 MALV-1-5 geo n.d. 0.99 16.15 0.13 n.d. 47.18 n.d. 11.35 0.2 0.1 0.53 76.6 MALV-1-6 geo n.d. 0.86 11.92 0.19 n.d. 61.78 n.d. 6.44 0.09 0.15 0.69 82.1 mean geo n.d. 0.95 14.29 0.16 n.d. 53.11 n.d. 9.41 0.15 0.13 0.56 78.76AD2V-1-1 low Cu ad n.d. n.d. 0.13 n.d. n.d. n.d. n.d. 0.94 48.03 36.95 n.d. 86.1 AD2V-1-2 low Cu ad n.d. n.d. n.d. n.d. n.d. n.d. n.d. 0.07 51.33 38.96 n.d. 90.4 AD2V-1-3 low Cu ad n.d. n.d. n.d. n.d. n.d. n.d. n.d. 0.85 49.95 38.81 n.d. 89.6 AD2V-2-1 low Cu ad n.d. n.d. 0.01 n.d. n.d. n.d. n.d. 0.75 48.45 37.97 n.d. 87.2 AD2V-2-2 low Cu ad n.d. n.d. n.d. n.d. n.d. n.d. n.d. 0.21 48.87 38.15 n.d. 87.2 mean low Cu ad n.d. n.d. n.d. n.d. n.d. n.d. n.d. 0.57 49.32 38.17 n.d. 88.09

Appendix D

266

Table D7 continued: EMPA results of individual base metal minerals and goethite in their direct vicinity.

Analysis nr. Mineral MgO Al2O3 SiO2 SO3 MnO Fe2O3 CoO CuO ZnO As2O5 PbO Total wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% wt% AD2V-1-5 ad n.d. n.d. 0.45 n.d. n.d. n.d. n.d. 43.96 9.63 37.99 n.d. 92 AD2V-1-6 ad n.d. n.d. 0.39 n.d. n.d. n.d. n.d. 43.84 9.26 37.89 n.d. 91.4 mean ad n.d. n.d. 0.42 n.d. n.d. n.d. n.d. 43.9 9.44 37.94 n.d. 91.7 ideal ad 56.78 40.08 AD1V_A_1 ad n.d. n.d. 0.2 n.d. n.d. n.d. n.d. 23.74 28.73 38.23 0.77 91.7 AD1V_A_2 ad n.d. n.d. 0.29 n.d. n.d. n.d. n.d. 22.24 29.65 38.64 0.72 91.5 AD1V_A_3 ad n.d. n.d. 0.75 n.d. n.d. n.d. n.d. 27.04 25.52 37.88 0.93 92.1 AD1V_A_5 ad n.d. n.d. 0.2 n.d. n.d. n.d. n.d. 24.94 27.01 38.51 0.79 91.4 AD1V-A2-1 ad n.d. n.d. 0.09 n.d. n.d. n.d. n.d. 22.91 29.89 38.57 0.39 91.9 AD1V-A2-2 ad n.d. n.d. 0.32 n.d. n.d. n.d. n.d. 23.1 28.53 38.21 0.83 91 mean ad n.d. n.d. 0.31 n.d. n.d. n.d. n.d. 23.99 28.22 38.34 0.74 91.6 AD1V-B2-1 geo n.d. n.d. 2.86 n.d. 6.75 34.45 n.d. 4.13 12.42 3.85 0.95 65.4 AD1V-B2-2 geo n.d. n.d. 3.13 n.d. 4.79 35.45 n.d. 3.87 10.19 3.34 0.76 61.5 AD1V-B2-2 geo n.d. n.d. 3.68 n.d. 7.12 38.17 n.d. 4.67 13.47 4.46 0.88 72.4 AD1V-B2-3a geo n.d. n.d. 4.67 n.d. 0.85 59.04 n.d. 3.84 5.8 3.99 0.90 79.1 AD1V-B2-4 geo n.d. n.d. 3.74 n.d. 0.79 60.39 n.d. 3.47 4.70 4.04 0.88 78 AD1V-B2-5b geo n.d. n.d. 2.94 n.d. 7.52 39.78 n.d. 4.59 12.1 4.83 1.18 72.9 AD1V-B2-6 geo n.d. n.d. 3.39 n.d. 11.44 38.86 0.47 4.45 5.86 3.05 7.55 75.1 AD1V-B2-9 geo n.d. n.d. n.d. n.d. 2.55 53.12 n.d. 3.61 n.d. n.d. 0.64 59.9 AD1V-B2-9c geo n.d. n.d. 3.65 n.d. 9.3 30.05 n.d. 5.13 15.96 4.81 1 69.9 AD1V-B2-10 geo n.d. n.d. 2.69 n.d. 1.17 53.9 n.d. 2.97 4.27 4.23 0.62 69.8 AD1V-B2-11 geo n.d. n.d. 3.62 n.d. 0.89 60.39 n.d. 3.28 4.63 3.86 0.94 77.6 mean geo n.d. n.d. 3.44 n.d. 4.83 45.78 n.d. 4.00 8.94 4.04 1.48 71.1 AD1V-B-1 ad n.d. n.d. 0.15 n.d. n.d. n.d. n.d. 17.4 34.47 39.01 0.25 91.3 AD1V-B-2 ad n.d. n.d. 0.12 n.d. n.d. n.d. n.d. 27.7 24.34 39.88 0.26 92.2 AD1V-B-3 geo n.d. n.d. 2.91 0.13 2.69 52.39 n.d. 3.82 5.06 4.8 1.71 73.5 AD1V-B-4 geo n.d. 4.236 44.77 n.d. 2.23 22.49 n.d. 0.12 0.32 n.d. n.d. 74.2 AD1V-B-5 geo n.d. 0.15 49.19 n.d. 2.67 25.02 n.d. 0.29 0.45 n.d. 0.22 78 AD1V-B-6 geo n.d. n.d. 4.67 n.d. 1.01 58.54 n.d. 3.38 5.19 3.84 1.13 77.8 AD1V-B-8 geo n.d. n.d. 2.14 n.d. 15.52 27.28 0.49 4.21 5.14 2.86 10.48 68.1 AD1V-B-9 geo n.d. n.d. 3.93 0.1 2.44 52.63 n.d. 3.76 5.42 4.64 1.54 74.5 AD1V-B-10 geo n.d. n.d. 2.32 n.d. 1.82 47.17 n.d. 3.47 4.48 5.14 0.77 65.2 AD1V-B-13 geo n.d. n.d. 3.61 0.21 3.58 44.96 n.d. 3.87 5.67 5.17 2.99 70.1 mean geo n.d. 2.193 14.19 n.d. 3.99 41.31 0.49 2.87 3.96 4.41 2.69 72.65 AD3V-B-1 Co ad n.d. n.d. 0.02 n.d. n.d. n.d. 0.18 0.31 48.23 39.13 n.d. 87.9 AD3V-B-2 Co ad n.d. n.d. n.d. n.d. n.d. n.d. 0.24 0.13 51.28 41.87 n.d. 93.5 AD3V-B-3 Co ad n.d. n.d. n.d. n.d. n.d. n.d. 0.39 0.25 50.32 41.36 n.d. 92.3 AD3V-B-4 Co ad n.d. n.d. n.d. n.d. n.d. n.d. 0.10 1.03 47.72 38.38 n.d. 87.2 AD3V-B-5 Co ad n.d. n.d. n.d. n.d. n.d. n.d. n.d. 9.11 41.57 38.83 n.d. 89.5 AD3V-B-6 Co ad n.d. n.d. n.d. n.d. n.d. n.d. 0.28 1.85 47.57 39.21 n.d. 88.9 mean Co ad n.d. n.d. n.d. n.d. n.d. n.d. 0.24 2.11 47.78 39.8 n.d. 89.89 CL2V-B-1 clin n.d. n.d. 0.19 n.d. n.d. n.d. n.d. 59.73 0.3 28.82 n.d. 89 CL2V-B-2b clin n.d. n.d. 0.30 n.d. n.d. n.d. n.d. 61.16 0.16 30.2 n.d. 91.8 CL2V-A-1 clin n.d. 0.09 0.29 n.d. n.d. n.d. n.d. 59.95 0.25 29.52 n.d. 90.1 CL2V-A-2 clin n.d. 0.026 0.29 n.d. n.d. n.d. n.d. 60.86 0.23 30.46 n.d. 91.9 mean clin n.d. 0.056 0.27 n.d. n.d. n.d. n.d. 60.42 0.23 29.75 n.d. 90.71 ideal clin 62.7 30.2 CL2V-B-3 geo n.d. 0.39 0.40 0.48 n.d. 60.16 n.d. 9.46 0.65 6.48 0.25 78.3 CL2V-B-5 geo n.d. 0.30 0.18 0.51 n.d. 59.95 0.12 7.75 0.46 6.4 0.13 75.8 CL2V-B-7 geo n.d. 0.39 0.57 0.64 n.d. 61.46 0.07 7.49 0.48 5.19 0.27 76.6 mean geo n.d. 0.36 0.38 0.54 n.d. 60.53 n.d. 8.23 0.53 6.02 0.22 76.88

Appendix D

267

Table D8: LA-ICP-MS results of iron oxides and oxyhydroxides in selected sections of ROM and LPO.

Analysis nr. Color Tex. Si Co Cu Zn Ga Ge Se Ag Cd Sn Sb Au Bi

wt% mg/kg wt% wt% mg/kgmg/kgmg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

212/2d-A1-1 dark coll 2.5 55 0.87 1.27 25 11 111 2 33 159 107 n.d. n.d. 212/2d-A1-4 dark 2.6 72 1.08 1.34 28 14 172 2 39 5 171 n.d. 1 212/2d-A1-7 dark 4.06 61 1.09 1.37 16 22 201 2 58 1894 182 n.d. 1 212/2d-A1-9 dark 6.11 56 1.22 2.23 18 30 211 7 43 23 308 n.d. 2 212/2d-A1-10 dark 3.98 76 1.20 1.34 20 28 220 2 56 2 242 n.d. 2 212/2d-A1-2 light coll 2.15 33 0.56 0.82 10 15 144 1 36 3 106 n.d. 1 212/2d-A1-3 light 2.28 39 0.59 0.6 8 23 145 1 31 1 211 n.d. 1 212/2d-A1-6 light 2.51 27 0.59 0.27 5 21 101 2 10 1 255 n.d. 2 212/2d-A1-8 light 2.85 44 0.59 0.56 6 25 156 1 26 1 213 n.d. 2 212/2d-B1-1 light 2.29 29 0.62 0.69 5 15 608 2 29 1 83 n.d. n.d. 212/2d-B1-2 light 2.18 31 0.63 0.69 5 14 566 2 27 3 77 n.d. n.d. 212/2d-B1-3 light 2.53 37 0.63 0.70 5 16 705 2 27 1 90 n.d. 1 212/2d-B1-4 light 2.45 35 0.62 0.65 4 15 781 2 27 2 83 n.d. n.d. 212/2d-B1-5 dark 3.51 54 0.87 0.87 6 18 1232 2 32 2 70 n.d. 1 212/2d-B1-6 dark 3.06 53 0.81 0.67 5 13 1077 2 25 1 44 n.d. n.d. 212/2d-B1-7 dark 3.14 51 0.85 0.65 4 15 1217 2 20 2 47 n.d. n.d. 212/2d-B1-8 dark 3.41 62 0.83 0.78 4 17 1237 2 29 4 70 n.d. 1 212/2d-A2-1 light 2.48 45 0.61 0.75 5 15 784 2 34 2 119 n.d. 1 212/2d-A2-2 light 2.65 46 0.59 0.67 4 16 658 2 24 2 104 n.d. 1 212/2d-A2-3 light 3.47 54 0.63 0.79 5 31 764 3 26 4 115 n.d. 1 212/2d-A2-4 light 3.63 53 0.62 0.75 5 26 781 2 26 3 103 n.d. 1 212/2d-A2-5 light 3.09 54 0.76 0.83 5 22 1153 2 34 3 132 n.d. 1 212/2d-A2-6 light 3.31 63 0.76 1.07 6 25 906 8 44 10 146 n.d. 4 212/2d-A2-7 light 3.22 45 0.65 0.87 5 25 708 3 26 10 119 n.d. 1 212/2d-A2-8 light 3.05 50 0.65 0.73 5 22 776 2 26 5 112 n.d. 1 212/2d-A2-9 light 2.63 47 0.62 0.73 5 18 833 2 31 4 105 n.d. 1 212/2d-B4-1 light 1.73 17 0.29 0.97 2 7 129 1 8 4 42 n.d. n.d. 212/2d-B4-2 light 2.33 32 0.62 0.77 5 11 330 2 22 11 83 n.d. n.d. 212/2d-B4-3 light 2.38 33 0.54 0.77 6 15 448 3 21 5 56 n.d. n.d. 268d-A-1 light coll 3.1 51 0.42 3.01 n.d. 3 8 5 43 n.d. 748 n.d. n.d. 268d-A-2 light 3.48 41 0.41 3.42 n.d. 3 20 5 52 n.d. 388 n.d. n.d. 268d-A-3 light 4.04 33 0.41 3.3 1 4 29 5 37 n.d. 195 n.d. n.d. 268d-A-4 light 2.92 20 0.43 2.4 n.d. 3 21 4 38 0 111 n.d. n.d. 268d-A-7 light 4.89 35 0.42 4.63 1 4 40 4 46 n.d. 185 n.d. n.d. 268d-A-5 dark coll 7.43 63 1.04 6.07 1 6 58 12 77 n.d. 215 n.d. n.d. 268d-A-6 dark 9.03 70 1.10 8.07 1 8 71 11 81 n.d. 244 n.d. n.d. 268d-A-8 dark 8.32 59 1.06 6.98 1 5 59 12 91 n.d. 332 n.d. n.d. 268d-A-9 dark 8.89 78 1.04 9.22 2 7 85 13 113 n.d. 543 n.d. n.d. 268d-A-10 dark 8.24 76 1.04 8.46 n.d. 9 95 14 125 1 751 n.d. n.d. 294R1d-A2-1 dark 7.48 46 1.27 3.14 2 5 66 3 45 3 13 n.d. 3 294R1d-A2-1a dark 6.43 44 1.25 2.8 1 3 82 3 55 2 12 n.d. 3 294R1d-A2-2 dark 4.57 48 1.3 2.08 1 5 151 2 38 2 4 n.d. 1 294R1d-A2-3 dark 5.55 39 1.25 2.66 2 4 64 3 48 2 10 n.d. 2 294R1d-A2-8 dark 5.40 25 1.26 4.39 1 3 109 2 34 1 8 n.d. 4 294R1d-A2-4 light 2.51 10 0.83 2.63 2 3 169 2 54 6 48 n.d. 53 294R1d-A2-5 light 4.11 12 0.82 4.64 3 7 244 3 70 2 5 n.d. 4 294R1d-A2-6 light 3.43 10 0.82 4.23 3 9 177 2 59 2 10 n.d. 1 294R1d-A2-7 light 3.08 7 0.82 3.1 1 5 151 1 46 2 2 c n.d. 294R1d-A1-1 light 2.61 10 0.56 2.42 2 5 127 1 34 2 5 n.d. 2

Appendix D

268

Table D8 contiuned: LA-ICP-MS results of iron oxides and oxyhydroxides in selected sections of ROM and LPO.


wt% mg/kg wt% wt% mg/kg mg/kgmg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

294R1d-A1-2 light 2.81 11 0.6 2.07 2 4 109 1 31 3 19 n.d. 8 294R1d-A1-3 light 2.73 8 0.63 2.91 2 6 197 1 48 2 6 n.d. 1 294R1d-A1-4 light 4.94 15 0.59 2.68 2 3 93 1 24 29 14 n.d. 6 294R1d-A1-5 dark 9.85 52 1.48 4.41 8 7 143 4 56 6 15 n.d. 13 294R1d-A1-6 dark 8.78 54 1.52 4.11 10 4 106 3 48 2 13 n.d. 12 294R1d-A1-7 dark mix 11.32 60 1.53 4.62 10 7 127 4 57 6 15 n.d. 22 294R1d-A1-8 dark mix 9.14 50 1.47 5.98 10 7 97 4 59 2 11 n.d. 24 294R1d-B2-1 dark 4.74 28 1.46 4.05 10 3 87 4 40 1 7 n.d. 39 294R1d-B2-2 dark 6.26 34 1.50 5.03 8 9 101 5 58 1 8 n.d. 57 294R1d-B2-3 dark 6.03 60 1.89 7.11 34 7 226 8 97 18 53 n.d. 82 294R1d-B2-4 dark 5.8 52 1.90 6.36 36 5 302 7 108 3 35 n.d. 23 294R1d-B2-5 light 1.61 4 0.53 1.66 1 3 56 1 37 4 9 n.d. 19 294R1d-B2-6 light 2.23 7 0.53 2.10 3 4 75 1 31 7 16 n.d. 38 294Z1d-B1-1 light 2.57 10 0.93 1.7 24 5 71 3 72 n.d. 52 n.d. 6 294Z1d-B1-2 light 4.36 13 0.95 2.61 10 13 37 4 55 10 294 n.d. 98 294Z1d-B1-3 light 3.13 13 0.94 2.02 22 10 61 4 71 1 61 n.d. 11 294Z1d-B1-4 dark 15.02 48 1.19 11.13 12 18 78 9 101 127 98 n.d. 36 294Z1d-B1-5 dark 6.7 34 1.06 5.22 8 7 22 5 60 7 26 n.d. 8 294Z1d-B1-6 dark 7.8 38 1.23 5.48 8 9 28 6 73 13 31 n.d. 13 294Z1d-A3-1 light 3.27 10 0.67 1.65 6 9 61 3 38 6 168 n.d. 45 294Z1d-A3-2 light 3.42 8 0.67 1.82 5 8 96 3 45 3 132 n.d. 18 294Z1d-A3-3 light 2.42 4 0.67 1.77 4 6 51 2 45 9 123 n.d. 60 294Z1d-A3-4 dark 8.75 40 2.10 4.3 2 13 309 8 105 1 17 n.d. 6 294Z1d-A3-5 dark 11.86 60 2.28 4.4 4 25 356 9 130 1 21 n.d. 13 294Z1d-A3-6 dark 10.74 54 2.39 5.82 3 17 302 14 120 6 21 n.d. 9 268ad-A-1 dark 9.96 36 1.75 7.64 1 3 8 5 90 2 11 n.d. 1 268ad-A-2 dark 8.99 26 1.73 6.92 1 3 7 4 87 n.d. 4 n.d. n.d. 268ad-A-4 dark 8.36 37 1.54 6.86 1 4 22 4 70 1 15 n.d. 1 268ad-A-3 light 4.36 57 0.58 4.59 2 6 16 2 31 1 26 n.d. 2 268ad-B-1 dark 20.95 15 1.39 6.84 1 5 7 10 130 n.d. 1 n.d. 0 268ad-B-2 dark 12.84 31 1.38 7.65 1 4 15 6 122 1 8 n.d. 2 209cd-A-1 light 0.68 26 1.22 1.56 1 3 101 6 249 n.d. 14 n.d. n.d. 209cd-A-3 light 0.74 44 1.26 2.44 2 5 86 21 342 n.d. 17 1 21 209cd-A-2 dark 1.42 52 2.91 3.26 3 5 210 12 460 1 27 2 1 209cd-A-4 dark 1.26 41 2.72 2.76 4 6 200 27 413 4 37 85 4 4F2d-A-1 light 2.32 2 0.7 1.54 4 4 51 1 16 3 5 n.d. n.d. 4F2d-A-4 light 2.37 1 0.76 0.92 3 4 59 1 24 1 4 n.d. n.d. 4F2d-A-2 dark 1.60 1 0.52 0.89 3 3 80 1 14 1 3 n.d. n.d. 4F2d-A-3 dark 1.44 2 0.5 0.82 2 2 71 1 7 n.d. 2 n.d. n.d. 4F2d-B-1 dark coll 1.45 1 0.33 1.42 3 1 8 1 5 n.d. 6 n.d. n.d. 4F2d-B-2 dark coll 3.25 3 0.40 2.78 3 n.d. 20 1 8 1 4 n.d. n.d. 4F2d-B-6 dark coll 2.77 n.d. 0.39 2.65 4 n.d. n.d. n.d. 3 1 5 n.d. n.d. 4F2d-B-7 dark coll 2.42 n.d. 0.39 1.75 4 n.d. 25 n.d. 3 n.d. 2 n.d. n.d. 4F2d-B-3 light 2.86 n.d. 0.77 0.48 3 8 95 1 22 2 3 n.d. n.d. 4F2d-B-4 light 2.80 n.d. 0.8 0.65 4 n.d. 67 1 14 4 8 n.d. n.d. 4F2d-B-5 light 2.60 3 0.73 0.46 3 5 68 1 18 1 3 n.d. n.d. 4F2d-B-8 light 1.21 1 0.38 0.39 2 4 36 1 6 1 3 n.d. n.d. 4F2d-C-1 dark 1.64 n.d. 0.25 0.99 5 4 19 1 4 18 6 n.d. 5 4F2d-C-2 dark 2.04 n.d. 0.24 0.95 5 n.d. 45 1 4 12 6 n.d. 6 4F2d-C-3 light 1.98 3 0.26 0.74 4 11 n.d. 1 3 3 7 n.d. 7

Appendix D

269

Table D8 contiuned: LA-ICP-MS results of iron oxides and oxyhydroxides in selected sections of ROM and LPO.


wt% mg/kg wt% wt% mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg mg/kg

4F2d-C-4 light 2.80 4 0.33 0.98 4 7 n.d. 1 9 8 n.d. 13 4F2d-C-5 light 2.08 3 0.27 0.67 3 6 38 1 4 2 6 n.d. 17 79d-B-1 light coll 2.25 13 1.55 4.99 9 6 104 5 421 31 651 n.d. 15 79d-B-2 light coll 2.45 12 1.61 4.12 9 7 150 4 495 26 635 n.d. 15 79d-B-3 light coll 3.38 66 1.66 7.92 9 9 110 46 370 42 573 n.d. 12 79d-B-5 light coll 2.65 15 1.50 4.22 9 7 112 8 417 15 482 n.d. 11 79d-B-4 dark coll 2.71 14 1.73 4.70 10 7 107 3 374 32 650 n.d. 12 79d-C-1 light 2.02 6 1.61 3.71 12 8 49 3 348 352 339 n.d. 1 79d-C-2 light 2.64 6 1.64 3.93 11 31 72 4 246 637 308 n.d. 5 79d-C-3 light 1.87 7 1.49 3.65 11 10 44 2 336 232 327 n.d. 1

Appendix E

Table E1: Chemical composition of the leaching acid prior and after the passage through the LPO.

Cu Zn Co Pb Cd As Fe Na Si Ca Mg Al Mn

mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l mg/l acid at sprinkler 110 18200 360 11.3 4 7.8 32 6.7 65 485 2240 2550 6405 acid at heap drainage 400 17500 370 12.2 4.1 8.2 26 7.5 66 506 2272 2700 6460

Composition, formation, and leaching behaviour of ... · 3.2.4 Chemical analysis (XRF, EMPA,...

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