development of an improved, fast throughput, test …€¦ · fernando perez perez de obanos bsc...

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DEVELOPMENT OF AN IMPROVED, FAST THROUGHPUT, TEST PROCEDURE, TO IDENTIFY THE METALLURGICAL EXTRACTIVE POTENTIAL OF GOSSANOUS AND SURFICIAL ORE DEPOSITS

CONTAINING REFRACTORY AND!REACTIVE!GOETHITE!

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Fernando Perez Perez de Obanos

BSc (Biochem); BSc (Kinetics Biology); Diploma of Forensic Science.

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Centre!for!Forensic!Science!

University!of!Western!Australia!!!

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This!thesis!is!presented!for!the!degree!of!

Master!of!Forensic!Science!of!the!University!of!Western!Australia!!

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2013

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ABSTRACT

Nickel (Ni) is one of the main elements used in stainless steels, coinage,

rechargeable batteries and magnets. It is an extremely important export

commodity for the Australian mining sector as the country is the fourth largest

producer of nickel internationally (90% in Western Australia). However due to

the continued use of the reserves of high-grade nickel sulphide ores, laterite

ores have now become an important source from which this element is

extracted.

The use of mineralized surficial material as a source of metals has increased in

recent years due to the diminishing availability of traditional primary

mineralization. Furthermore, because these ores occur at the surface, and

therefore have easy access for mining, they can be far more economic than

primary ores as a source of metals. This is true even though the actual

concentrations of the economic minerals are lower than in more traditional

primary ores. However, there is a significant problem associated with extraction

of metals from surficial deposits in that the relative refractory nature of goethite

(a secondary hydrated iron oxide mineral produced by weathering iron ores)

present can severely limit the extractive efficiency of both chemical and

bacteriological agents. Consequently, even relatively high grade deposits can

be extractively largely sterile. Therefore it would be economically extremely

beneficial to be able to characterise the degree of extractability, and therefore

economic value, of ores prior to establishing costly plant facilities on site and

only then find out that the ores being processed are refractory. At present the

test procedures involved in determining the extraction potential of these ores

includes processes that can last over six months and the expenditure of a

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significant amount of money with concomitant waste of time if the ores prove to

be largely economically un-extractable. Characterization of nickel laterite ores

suited to heap leach processing is therefore important in order to select those

that are the most reactive and provide the highest recovery of the metal.

Heap leaching, albeit an expensive and time consuming process, is the main

technique used to extract Ni from nickel laterites. Consequently, the research

detailed in this thesis is designed to develop a fast extractive chemical leaching

protocol to assess the reactiveness of nickel laterites to heap leaching regimes

and thereby overcome the significant expenditure and time requirements

associated with the traditional longer term field trials now used to determine

nickel extractability.

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ACKNOWLEDGEMENTS

First of all I will like to thank my two friends and supervisors Professor John Watling

and Sven Fjastad, without their help and patience I couldn’t have done it. Also I will

like to thank the Centre for Forensic Science, in particular Professors Ian Dadour and

Daniel Franklin for their help and advice. Secondly I will like to thank Dr H R Watling of

CSIRO Division of Process Science and Engineering for providing the ores and for

technical support throughout the project.

Thank you also all the team of TSW Analytical, all the students, Alex Martin, Natasha

Kreitals and Matthew Murphy and all the staff, Allen Thomas, Christopher May,

Cameron and Rachel Scadding, Anna Bradley and Jenna Valentin.

I will also like to thank the staff from Townsend Lodge for helping me since I arrived in

Australia, and helping me in every single thing, especially Raylene Hindle, Maeve

O’Sullivan and Nickolas Kerr.

Finally I will like to thank all my friends and family back home, specially my brother

Alejandro Perez, my mother and father Teresa and Carlos Perez. Thank you for being

always there and for keeping the faith, even in the most difficult circumstances, Thank

you for being such as good model to learn from and for always being there supporting

me.

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TABLE OF CONTENTS

Abstract 2

Acknowledgements 4

Table of contents 5

List of Figures 8

List of Tables 10

Chapter 1: Introduction 11

1.1 Background of to the project 11

1.2 Nickel market 15

1.3 Source of Nickel 17

1.3.1 Genesis of lateritic ores 19 1.3.2 Genesis of sulphide ores 20

1.4 Techniques for the extraction of nickel 20

1.5 Techniques for refining nickel 23

1.6 Pre-treatment of the ore 24

1.7 Lateritic ores 25 1.7.1 Hydrometallurgy of lateritic ores 25 1.8 Goethite 27 1.9 Summary of the problem 28 1.10 Aim of this research 29

Chapter 2: Experimental 31

2.1 Equipment 31

2.1.1 Inductive coupled plasma Atomic Emission Spectroscopy (ICP-AES) 32

2.1.2 Inductively Couple Plasma Mass Spectrometry (ICP-MS) 33

2.1.3 X-Ray Diffraction (XRD) 34

2.2 Background to the samples 35

2.3 Phase 1: Preliminarily Experiment using sample A and B 36

2.3.1 General 36

2.3.2 Total Dissolution 37

2.3.3 Inorganic Acid Leach-General 38

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2.3.3.1 Sulphuric Acid (H2SO4) Leach 38

2.3.3.2 Hydrochloric Acid (HCl) Leach 39

2.3.3.3 Aqua Regia Acid Leach 39

2.3.3.4 Nitric Acid (HNO3) Leach 39

2.3.4 Organic Leach 39

2.3.4.1 Oxalic Acid (H2C2O4) Leach 39

2.3.4.2 Citric Acid Leach (C6H8O7) Leach 40

2.3.4.3 Tartaric Acid Leach (H6C4O6 ) Leach 40

2.3.5 Mixed Acids 40

2.3.5.1 10% v/v Sulphuric (H2SO4) and 0.5M oxalic Acid (H2C2O4) Leach 40

2.3.5.2 10% v/v Sulphuric (H2SO4) and 0.5M citric acid (C6H8O7) Acid Leach 40

2.3.5.3 10% v/v Sulphuric (H2SO4) and 0.5M tartaric acid (H6C4O6 ) Acid Leach 41

2.3.7 Oxidation leach 41

2.3.7.1 Sulphuric Acid (H2SO4) Leach 41

2.3.7.2 Hydrochloric Acid (HCl) Leach 41


2.3.7.4 Nitric Acid (HNO3) Leach 42

2.4 XRD 42

2.5 Phase 2: validation samples. 42

2.6 Sample and Data Analysis 43

Chapter 3: Results and Discussion 44

3.1 Phase 1: Preliminarily Experiment using “Ore Sample A” and “Ore Sample B” 44

3.1.1 Total Dissolution 45

3.1.2 Inorganic Acid Leaching 45

3.1.2.1 Sulphuric Acid Leach 45

3.1.2.2 Hydrochloric Acid Leach 47


3.1.2.4 Nitric Acid Leach 50

3.1.2.5 Summary of Inorganic Leach Experiments 51

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3.1.3 Organic Acid Leaching 52

3.1.3.1. Oxalic Acid Leach 52

3.1.3.2 Citric Acid Leach 54

3.1.3.3 Tartaric Acid Leach 55

3.1.3.4 Summary Organic Acid based Leaching 58

3.1.4 Mixed Organic and Sulfuric Acid Leaching 61

3.1.4.1. Sulfuric Acid and Oxalic Acid Leach 61

3.1.4.2 Sulphuric acid and Citric Acid Leach 62

3.1.4.3 Sulphuric and Tartaric Acid Leach 63

3.1.1.4 Summary for the Mixed Acid Leaches 64

3.1.5 Single inorganic acid Leaching following pre-oxidation of ore materials 66

3.1.5.1 20% Sulfuric Acid Leach of Pre-oxidised Ore Samples 68

3.1.5.2 20% Hydrochloric Leach of Pre-oxidised Ore Samples 69

3.1.5.3 20% Aqua Regia Acid Leach of Pre-oxidised Ore Samples 70

3.1.5.4 20% Aqua Regia Acid Leach of Pre-oxidised Ore Samples 71

3.1.5.5 Comparative Leachability for all Study Analytes from “Ore Sample A”

and “Ore Sample B” Pre-oxidised Ore 72

3.1.6 Estimation of long term leachability of nickel based on short term

calibration against “standard samples” 74

Chapter 4: Conclusion & Future work 83

References 87

Appendix 91

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LIST OF FIGURES

Figure 1.1 - Use of Nickel (Bradley 2011) 12

Figure 1.2 - Australia nickel production (Hoatson et al 2006) 13

Figure 1.3 - Nickel and Copper prices since 1990-2011 (London Metal Exchange 2012) 16

Figure 1.4 - Nickel production 1950-2010 (Dalvi 2004) 18

Figure 1.5 - World nickel laterite deposits (Dalvi 2004) 19

Figure 1.6 - Genesis of a sulphide nickel deposit 20

Figure 1.7 - A schematic representation of a heap leaching process. (Oxley et al 2006) 23

Figure 1.8 - Laterite profiles from Western Australia and Indonesia (Dalvi et al 2004) 26

Figure 1.9 - Basic distance of Fe-Fe in goethite (Cornell & Schwertmann 2003.) 27

Figure 1.10 - Structure of goethite showing different projections (Cornell & Schwertmann,

2003) 28

Figure 2.1 - General representation of the experimental protocol used in this study 35

Figure 3.1 - Nickel extraction efficiency for “Ore Sample A” (1A) and “Ore Sample B”

(1B) using 20% v/v H2SO4 over a 240h period 46


(2B) using 20% v/v HCl over a 240h period. 49


(3B) using 20% v/v Aqua Regia over a 240h period. 49


(4B) using 20% v/v nitric acid over a 240h period 50

Figure 3.5 - Comparison of the nickel extraction efficiencies for all inorganic acid

based leach solutions 52

Figure 3.6 - Analyte leach curves for “Ore Sample A” and “Ore Sample B” using selected

molarity oxalic acid solutions 53


molarity citric acid solutions 56


molarity tartaric acid solutions 57

Figure 3.9 - Comparison of Leach Efficiencies of 1M Organic Acid Leach Solutions 60

Figure 3.10 - Extraction efficiency for oxalic acid/sulfuric acid mixed acid leaching of

“Ore Sample A” and “Ore Sample B” 62

Figure 3.11 - Extraction efficiency for sulfuric acid/citric acid mixed acid leaching of


Figure 3.12 - Extraction efficiency for sulfuric acid/tartaric acid mixed acid leaching of


Figure 3.13 - Extraction efficiencies for nickel from “Ore Sample A” and “Ore Sample B”

using three different mixed acid leach solutions 65

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Figure 3.14 - X-ray diffractograms of samples “Ore Sample A” (A and B) and “Ore Sample B”

(C and D) before and after heating at 1000oC for 2 hours. Green lines

correspond to Goethite and purple to Hematite. 67

Figure 3.15 - Variation in 20% v/v sulfuric acid based nickel extractability for ore samples “Ore Sample

A” and “Ore Sample B” following pre-oxidation of the ores at selected temperatures for a two hour

period. 68

Figure 3.16 - Variation in 20% v/v hydrochloric acid based nickel extractability for ore samples “Ore

Sample B” and “Ore Sample A” following pre-oxidation of the ores at selected temperatures for a two

hour 69

Figure 3.17 - Variation in 20% v/v Aqua Regia acid based nickel extractability for ore samples “Ore

Sample A” and “Ore Sample B” following pre-oxidation of the ores at selected temperatures for a two

hour period. 70

Figure 3.18 - Variation in 20% v/v nitric acid based nickel extractability for ore samples

“Ore Sample A” and “Ore Sample B” following pre-oxidation of the ores at selected temperatures for a

two hour period. 71

Figure 3.19 - Calibration leach curves for “Ore Samples C, D, F, G and H” in 20% sulfuric acid for

selected time periods – ICP-AES data. 78

Figure 3.20 - Calibration leach curves for “Ore Samples C, D, F, G and H” in 20% sulfuric acid for

selected time periods – ICP-MS data. 79

Figure 3.21 - Calibration leach curves for “Ore Samples C, D, F, G and H” and unknown “Ore sample

E”, in 20% sulfuric acid for selected time periods – ICP-MS data. 80

Figure 3.22 - Calibration leach curves for “Ore Samples C, D, F, G and H” and unknown “Ore sample

E”, in 20% sulfuric acid for selected time periods – ICP-MS data. 81

Figure 3.23 - Calibration curves produced for the relative relationship between the extractability of

nickel using both ICP-AES and ICP-MS and describing the relationship between long and short term

extractability of this metal from various refractory ores 82

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LIST OF TABLES

Table 1.1 - Presence of Nickel in different Minerals (Bolt et al 1967) 14

Table 2.1 - Analytical emission lines used in this study 32

Table 2.2 - List of isotopes used in the ICP-MS based analytical procedure 34

Table 2.3 - Description of samples used in this study together with their nickel extraction efficiencies as determined by CSIRO Division of Process Science and Engineering 36

Table 3.1 - Comparison of analytical data for Certified Reference Materials SARM1 and SARM2 obtained for this study. Results expressed as parts per million in the original material. 45

Table 3.2 - Data for Fe, Ni, Co and Mn concentrations in Samples “Ore Sample A” and “Ore Sample B” 45

Table 3.3 - Leach data for extraction efficiencies of all study analytes after a 240h leach period detailing all inorganic acid based leach solutions studied. 51

Table 3.4 - Summary of results for leaching “Ore Sample A” and “Ore Sample B” ores using selected organic acids under pre-determined molarity concentrations. 59

Table 3.5 - Summary of the extraction efficiencies of all study analytes from “Ore Sample A” and “Ore Sample B” using mixed acid leaches. 65

Table 3.6 - Leach data for all study analytes from samples “Ore Sample A” and “Ore Sample B” following pre-oxidation at selected temperatures for two hours. 73

Table 3.7 - Long Term Nickel Extractability of Analytes from the “Standard Samples” 75

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CHAPTER 1

INTRODUCTION

1.1 BACKGROUND TO THE PROJECT Metals have been a significant part of human culture for many thousands of years with

specific cultural periods being named after the metal predominant in society at the time.

The earliest evidence of the European “Copper Age” appears in Belevode, Serbia

(Radivojević et al 2010). Copper smelting took place in this area some 7000 BP and the

recovery of cast objects here challenges the concept of the origin of extractive metallurgy

as being in the Near East. Blending of metals (copper and tin) technology has been

recognized from the early Bronze Age (3800BP) to the start of the Iron Age (2750BP)

(Gimbutas 1965), However, only recently has the discovery of steel, and the development

of a commercial process for its manufacture, encouraged the use of a wide variety of

chemical elements in the manufacture of ferroalloys and an increase in the extraction and

beneficiation of elements such as nickel, cobalt, vanadium and chromium (Mudd 2010).

The extraction and beneficiation of one such element, nickel, from refractory ores is the

subject of this thesis.

Nickel was discovered in 1751 by Axel Fredrik Cronstedt, a Swedish metallurgist and

chemist, and is one of the main elements used in stainless steel production, coinage,

rechargeable batteries and magnets (Cunat 2004). Nickel is an extremely important export

commodity for the Australian mining sector as the country is the fourth largest producer of

nickel internationally (90% in Western Australia), (Mudd 2009). Nickel plays a very

important role in today’s society with approximately 61% of the nickel extracted used to

produce nickel steel, 9% for alloys such as German lead or Nichrome with the rest,

approximately 30%, being used in electrical components and batteries (Figure. 1.1). Such

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a widespread use can be explained in part by nickel’s characteristic slow oxidation rate at

ambient temperature, a characteristic that it also imparts to alloying metals where addition

of nickel to these mixtures adds both strength and corrosion resistance (Lee et al 2012).

Figure 1.1. Use of Nickel (Bradley 2011)

Nickel has five isotopes, 58Ni (68.3%), 60Ni (26.1%), 61Ni (1.13%), 62Ni (3.59%) and 64Ni

(0.91%), and is found in combination with elements such as sulfur, arsenic and iron, producing

naturally occurring minerals such as pentlandite (Ni,Fe)9S8, nickeline (NiAs) and millerite

(NiS), the main nickel minerals are detailed in Table 1.1. Nickel is predominantly extracted

from two types of ore deposits, magmatic sulphide deposits where it is extracted from

pentlandite, and secondly from laterites where garnierite (Ni, Mg)3Si2O5(OH)4 and limonite (Fe,

Ni)O(OH) are the main ores (Dalvi et al 2004).

Due to their higher extraction efficiency and the lower cost of extraction, the majority of the

worldwide production of Ni is derived from processing sulphide rich deposits (Figure 1.2).

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However, this kind of deposit is becoming scarce and therefore alternative nickel deposits,

such as laterites, have to be considered as alternative sources even though they are generally

lower in nickel and processing costs are higher (McDonald et al 2008).

Figure 1.2. Australia nickel production (Hoatson et al 2006)

Nickel is one of the main elements used in stainless steels, coinage, rechargeable batteries

and magnets (Pearce 1987). Is an extremely important export commodity for the Australian

mining sector as the country is the fourth largest producer of nickel internationally (90% in

Western Australia) (Watling et al 2011). However due to the continued use of the reserves of

high-grade nickel sulphide ores, laterite ores have now become an important source from

which this element is extracted (Chander.1982).

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Table 1.1. Presence of Nickel in different Minerals (Bolt et al 1967)

Formula % of Nickel

Sulfides

Pentlandite (Ni,Fe)9S8 34.2 More common mined

Millerite NiS 64.7 Minor constituents

Heazlewoodite Ni3S2 73.3 Highest nickel content

Linnacite series (Fe,Co,Ni)3S4 Variable Minor constituent

Polydymite Ni3S4 57.9

Violarite Ni2FeS4 38.9

Siegenite (Co,Ni)3S4 28.9

Arsenides Typical from Nickel-bearing

cobalt ores

Niccolite NiAs 43.9

Maucherite Ni11As8 51.9

Rammelsbergite NiAs2 28.1

Gersdorffite NiAsS 35.4

Antimonide

Breithauptite NiSb 32.5

Arsenate

Annabergite Ni3As2O8·8H2O 29.4

Silicate and Oxide

Garnierite (Ni,Mg)6Si4O10(OH)8 Variable (up to

47%)

Most abundant nickel

carrier in silicate ores

Nickeliferous Low concentration of Ni

Laterite deposits

Limonite (Fe,Ni)O(OH)·nH2O Low

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The use of mineralized surficial material as a source of metals has increased in recent

years due to the diminishing availability of traditional primary mineralization. Furthermore,

because these ores occur at the surface, and therefore have easy access for mining, they

can be far more economic than primary ores as a source of metals (Tang & Valix 2006).

This is true even though the actual concentrations of the economic minerals in these ores

are lower than in more traditional primary ores. However, there is a significant problem

associated with extraction of metals from surficial deposits in that the relative refractory

nature of goethite (a secondary hydrated iron oxide mineral produced by weathering iron

ores) present can severely limit the extractive efficiency of both chemical and

bacteriological agents (Tzeferis 1994). Consequently, even relatively high grade deposits

can be extractively and economically largely sterile.

Because of this aspect it would be economically extremely beneficial to be able to

characterise the degree of extractability, and therefore economic value, of ores prior to

establishing costly plant facilities on site and only then find out that the ore being

processed is refractory. At present the test procedures involved in determining the

extraction potential of these ores include processes that can last over six months and the

expenditure of a significant amount of money. The concomitant waste of time and money if

the ores prove to be largely economically un-extractable (Chander 1982; Norgate &

Jahanshahi 2010), can be considerable. Characterization of nickel laterite ores suited to

heap leach processing is therefore important in order to select those that are the most

reactive and provide the highest recovery of the metal.

1.2 NICKEL MARKET.

Subsequence to its discovery in the eighteenth century, nickel production and use has

increased enormously. However, nickel production and consumption seem to reflect periods

of prosperity and recession and the trends in its production seem to closely mirror those of the

steel industry (Figure. 1.3). Nickel was first extracted in Germany (1823) followed by America

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(1830) and Sweden (1838). In 1902, Charles Schwab opened “the International Nickel

Company of Canada (INCO), a company which played a major role world-wide in nickel

production over the following 50 years (Bolt et al, 1967. p.26).

Figure 1.3 Nickel and Copper prices since 1990-2011 (London Metal Exchange 2012)

The First World War was a period of high nickel production where most of the nickel was

consumed for military purposes. In 1929, due to the economic crisis in Europe and America,

production decreased and by 1934, 81% of the nickel worldwide was produced in Canada with

the remaining production from New Caledonia (Pacific) where nickel was mainly mined from

silicate ore material.

During the Second World War, nickel was produced and refined in a number of countries

including Germany and especially Russia which expanded both its mining and refining

capability, primarily for military use, until 1945. After the Second World War, between 1945

and 1970, there was a significant research push to establish different uses for nickel and by

1950, Russia was responsible for about 20% of the world’s production. Nickel production over

the next 25 years was influenced primarily by the Korean War (1950-1953). During this period,

new high strength alloys were created primarily for the military and aerospace industries. In

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addition, new ways of mining were also developed and went hand in hand with modernisation

of smelting and refining processes (Bolt et al 1967. P.74). During the 1960’s Australia and

Indonesia started exploring for nickel deposits and by the end of 1970 several countries such

as Greece, Peru and Venezuela where also producing nickel (Jaques et al 2005). This

expansion in nickel exploration and production generated a total world production in 1970

which was double the production in 1960. This huge increase in nickel production was the

main cause of a significant drop in the price of nickel in 1976 and nickel production essentially

stopped from around this period .

In 1985, twenty three countries were producing nickel with Australia producing approximately

10% of nickel worldwide. From 2001 to 2007 the price of nickel increased twelvefold. This was

due to three main factors: Chinas open market and the in demand for steel, the appearance of

multinational auto industries around the world and the increase of the Indian technology sector

during the latter half of the decade. Nowadays, the price of nickel is stable at approximately

$20,000 AUD/Tonne (London Metal Exchange 2012), mainly due to the fact that new mines

are being opened and new and better metallurgical technology has been developed to

increase the production of this element (Mudd G.M, 2007).

1.3 SOURCE OF NICKEL.

Nickel in nature is associated with two main sources, basic or mafic rocks that usually contain

a high concentration of nickel within iron and magnesium rich minerals, and acid or silicic

rocks where the nickel is largely present as sulfides. The concentration of nickel in these types

of igneous rocks depends on the relationship between iron/magnesium and silicon/aluminium

(Dean 2006.).

Nickel is economically distributed in two types of ores, primary or silicate rich ores where the

nickel has been introduced during petrogenesis as a sulfide mineral, (Reference 14), and

secondary or lateritic ores, where the nickel is present as an oxide or hydrated oxide, nearly

always associated with iron, following weathering of the primary ore minerals. To date,

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extraction of nickel from the primary sulfides is the main economic route for the production of

nickel (Nickel 2008).

Figure 1.4 Nickel production 1950-2010 (Dalvi 2004)

Pyrrhotite (FeS to Fe7S8) is an extremely important mineral, often associated with sulphide

nickel ores, and is usually more abundant in its own right than any primary nickel. In

pyrrhotite, a small amount of iron can be substituted by nickel. This mineral is important

because it is magnetic and therefore is amenable to geophysical prospecting to determine the

position of a nickel rich ore body (Bolt et al 1967).

It is also is important to realise that nickel rich deposits can also be found on the ocean floor

associated with manganese nodules. The two most important ores, sulphides and laterites

need to be processed in totally different ways from one another in order to recover as much

nickel as possible

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1.3.1 Genesis of Lateritic Nickel ores

This type of ore is generated by a process called lateritic sub-aerial oxidative weathering. The

basis of this process is that a rock, exposed to the atmosphere, is gradually affected by

different environmental factors such as heat, water and air (oxygen) that generate mechanical

and chemical erosion. The original rock is broken down into oxide based minerals which can,

depending on their solubility, be sequentially taken into solution and either transported away

from the primary deposit, if they are soluble, or remain in situ and are concentrated. The result

of this oxidation and transportation process is the production of lateritic deposits rich in both

nickel and iron oxides (Howard-White 1963).

The iron oxides precipitate as ferric hydroxide and eventually lose water to create hematite

(Fe2O3) and goethite (FeO(OH)). During the course of thousands to millions of years, erosion

continues affecting this iron-nickel rich layer resulting in build up of significantly thick iron oxide

rich layers which are also rich in nickel. Due to the immense size of the deposits and their

worldwide occurrence, this type of ore, is considered as the future source of nickel (Figure.

1.5).

Figure 1.5. World nickel laterite deposits (Dalvi 2004).

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1.3.2 Genesis of Sulfide Nickel ores.

The genesis of nickel sulfide ores is controlled by the solubility of nickel sulphide in the liquid

produced when mafic and ultramafic rocks melt. These types of ores are hosted in rocks that

are usually high in iron and magnesium and relatively low in silicon (e.g., gabbro, peridotite)

(Bames & Lightfoor 2005). Nickel is a chalcophile element (like sulphur) and consequently,

when a gabbroic or peridotitic magma forms, sulfides of iron, copper and nickel form an

immiscible association which, being heavier than the parent rock-melt, drop to the base of the

melt carrying with them elements such as Co, Pd, Pt, Au, and Rh (mainly group VIII). Here

there elements become more and more concentrated in this sulfide rich oxide liquid, forming

layers which are essentially the ore zones (Melekestseca et al 2013).

Figure 1.6. Genesis of a sulphide nickel deposit.

1.4 Techniques for the extraction of Nickel.

Today, the majority of nickel is extracted from sulfide ores. These ores, while expensive to

extract, are usually high grade. Because of this aspect, extraction costs per tonne of nickel are

less than those associated with the production of nickel from lateritic ores (Kaya & Topkaya,

2011). Usually nickel is extracted from the ore by roasting and different reduction processes

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that generate a nickel purity of approximately 75%, this purity, for some stainless-steel

producing companies is directly suitable for steel production. The type of process, and

techniques uses for the extraction of nickel, depends on the characteristics of the ore.

Sulphide ores

Sulphides ores are crushed and milled to a fine powder prior to flotation and concentration of

the nickel rich ore minerals. The concentrate is roasted to convert the sulfide minerals into

oxides. Most of these types of ores are pre-concentrated using froth flotation process followed

by a pyrometallurgical extraction. However, hydrometallurgical processes followed by the

Sherritt-Gordon process (Plasket & Romanchuk, 1978) can also be used. Depending on the

purity of the final metal product required, the process can undergo further refining stages to

obtain even pure metal.

Lateritic Ores.

A) Pyrometallurgical process: Lateritic ores that are extracted using this technique

usually require pre-drying, calcining/reduction and electrical furnace smelting, with

nickel being recovered from this process as a ferro-nickel matte. However this type

of extraction is quite difficult because the nickel-bearing minerals (eg: garnierites

and saprolitic deposits) may often be finely disseminated throughout the ore body.

In order to optimize the amount of electric energy used for smelting, it is important

to first reduce the moisture (which can be up to 40%) and which is lattice-bound to

the iron-hydroxide minerals. Another problem that must be taken into consideration

is the chemical and metallurgy diversity of laterite ores, as the SiO2/MgO ratio and

iron content can influence the production of slag (and therefore if not controlled,

adversely affect the production costs). The dehydration and feedstock control

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processes are both energy and time intensive, however the recovery of nickel can

be increased to nearly 95 % and cobalt to 50% if these processes are controlled

correctly. Nowadays, there is almost twice as much hydrometallurgical processing

as pyrometallurgical treatment. In the pyrometallurgical process, the ore is roasted

in the presence of carbon, sulfur is converted to sulfur dioxide and released as a

gas which is trapped and eventually reacted with water to from sulfuric acid. The

final product is a metal based concentrate.

B) Hydrometallurgical Process: Usually saprolitic laterites have a higher concentration

of nickel in the upper completely weathered zones (Agatzini-Leonardou et all,

2009). However, the presence of magnesium in high concentration, increases the

acid consumption, and therefore, for this technique, limonitic laterites are preferred.

Different hydrometallurgical processes are used depending on the nature of the

ore:

a. Caron process: Mixture of pyro and hydrometallurgy that involve roasting

followed by an ammonia/ammonium carbonate leaching stage which dissolves

the cobalt and nickel as ammine complexes (Lower overall Ni & Co extraction)

(Graaf, 1980).

b. Atmospheric-pressure acid leaching process (AI) (Buyukakinci & Topkaya

2009). This process sometimes involves the use of sulphuric acid leaching but

can involve solely ammoniacal leaching that precipitates the nickel as an

hydroxide.

c. High-pressure acid leaching (HPAL): Used in some Australian projects such as

Cawse or Murrin Murrin. This is one of the best extraction techniques for these

kinds of ore, providing an extraction efficiency for Ni & Co of approximately

90%. (Kaya & Topkaya 2011)

! 23!

d. Heap leaching (HL) (Figure 1.7): This process involves the use of limonitic and

saprolitic ores. This method involves the use of large amounts of acid over a

long period of time, however, if applied correctly can be an economically viable

process. In using this technique the grain size of the ore plays and important

role, therefore crushing the ore into different grain sizes is cruzial in order to

facilitate acid attack and minimize channelling (Ghorbani et al 2011). However,

limonitic ores, containing refractory goethite, sometimes do not leach well,

representing a loss of money to the company undertaking beneficiation

(Dhawan et al 2013).

Figure 1.7 A schematic representation of a heap leaching process. (Oxley et al 2006).

1.5 Techniques for refining of Nickel.

Once the nickel is extracted, depending on the purpose for which it is ultimately destined, the

concentrate can undergo different refining process that will generate a high purity nickel

product. Some of these methods involve:

A) Electrometallurgy: This process involves mainly electrowinning, electro refining and

fused salt electrolysis. Electrowinning recovers materials from aqueous solution, and is usually

one of the last stages of the hydrometallurgical process. The process uses a cathode of the

metal of interest. In fused salt electrolysis, the metal is dissolved into a salt that acts as an

! 24!

electrolyte and the value metal is collected on the cathode of the cell, (usually using high

temperature). Electro-refining is a process involving dissolution of an impure metallic anode

and the producion of a high purity cathode (Evans 2003).

B) Vapometallurgy: At a temperature between 50oC and 100oC, Ni can be combined with

four molecules of CO to create nickel tetracarbonyl (Frank et al. 2011). Nickel tetracarbonyl is

volatile and can be distilled from the ore at approximately 50oC and then either condensed

and collected or dissociated at temperatures between 250-600oC back to metallic nickel with

release of carbon monoxide. The iron and cobalt carbonyls tend to decompose before

evaporating and as such are left behind in the ore to be either processed at another stage or

simply discarded.

Ni+4CO = Ni(Co)4 (50oC)

Ni(Co)4 = Ni + 4CO (250oC)

In choosing a method for beneficiation, several characteristics need to be considered

including the expense of the chemicals, the facility in which the process needs to take place,

transport of the ore and type of ore. Therefore, each case can be similar or totally different

and must be costed on its own merit (Mishra 2001).

1.6 Pre-treatment of the ore The initial stage of ore processing is mechanical breakdown and separation of the ore from

the unmineralized material (gangue). The amount of comminution depends on the grain size

of the mineralized material, with the smaller grain sized ore minerals requiring finer

comminution to facilitate separation from gangue. Firstly, the ore is broken into medium to fine

grained particles, using a jaw or gyratory crusher. This process facilitates liberation of the ore

from large sized gangue material. Following this, the ore is ground using a rod mill and/or ball

mill which is usually attached to a water spray to avoid the problem of dust. The final stage

! 25!

often involves gravity flotation or magnetic separation. These processes take advantage of the

relative ore/gangue friability differences. Concentration for the ore minerals can be achieved

using, for example, magnets or slurry separation using variations in the magnetic potential of

the ore minerals with respect to the gangue and also differences in the density and particle

shape between these materials (Bolt et al, 1967).

Oxide ores are more expensive to process than sulfide ores, not only because they are

essentially low-grade but also because techniques such as flotation or magnetic separation

cannot be used to concentrate them due to the physical and chemical nature of the ore

material. In addition to this, naturally occurring iron oxides usually contain high levels of water

which require removal before processing and therefore increases energy costs.

Compensating factors at the mine site (e.g. easy mining and transport) are factors which,

together with the extremely extensive nature of the deposits, make the processing of this ore

type more attractive by employing economies of scale (Kaya & Topkaya 2011).

1.7 Lateritic Ores

There are two types of oxide ores, one is limonitic, and usually found close to the surface, and

the other is a silicate type found at depth. The limonitic type ore consists of a wide variety of

ferric oxide minerals (Figure. 1.8), particularly limonite, where the nickel has replaced iron in

some lattice sites ((Fe,Ni)O(OH)*nH2O). In the silicate type ore, nickel is associated with

silicate minerals, particularly serpentine (Mg6Si4O10(OH)8) (Basile et al 2010).

1.7.1 Hydrometallurgy of lateritic ores.

Hydrometallurgical extraction for nickel from laterite ores may involve roasting the ore in the

presence of sulphate and chloride ions and subsequently reducing the nickel product directly

! 26!

to a metal, or leaching the ore either at room or elevated temperatures and pressures using

sulphuric, hydrochloric and nitric acids.

Figure 1.8. Laterite profiles from Western Australia and Indonesia (Dalvi et al 2004).

Pyrometallurgical processing of laterite ore is expensive and not usually considered. Instead,

direct leaching (hydrometallurgical extraction) of the concentrate or the ore itself, is the

preferred option. The leaching can be achieved using mineral acids or other suitable solvents

or may involve bacterial leaching (Castro et al 2000).

The main ore containing mineral in nickel laterites is goethite (Chang et al. 2010) and it is the

degree to which this mineral sequesters the nickel into its lattice that is the main problem

when it comes to beneficiation for nickel rich laterites. To understand the problem it is first

necessary to consider the mineralogical and chemical characteristics of goethite.

! 27!

1.8 Goethite

Iron oxides are created by a process of aerobic weathering of the primary iron minerals

contained in magmatic rocks. There are sixteen different hydroxy-oxides of iron common in

nature and their relative abundance in a lateritic nickel ore is dependent on the conditions

under which they were created during the weathering process (Cornell, M. Schwertmann, U.

2003). Goethite (α-FeOOH) (Figure.1.9) is one of the predominant iron oxides found in lateritic

ores. Its structure is based on a hexagonal close packing (hcp) cell (diaspore symmetry) and

as such is very stable at ambient temperature. Goethite is a non-stoichiomentric compound

with a variable crystalline structure (Figure. 1.10) and chemical characteristics (Cornell &

Schwertmann, 2003). Because of the variability of the crystal structure and the fact that

substitution of elements such as manganese, aluminium and nickel can occur for iron in this

lattice, goethite can apparently be passivated to leaching.

.

Figure 1.9 Basic distance of Fe-Fe in goethite (Cornell & Schwertmann 2003.).

Consequently nickel extraction efficiencies can range from a few percent upwards without any

apparent change in the appearance of the ore. Consequently, it is of fundamental importance

to establish a fast test to determine the percentages of refractory and non-refractory goethite

present in an ore to be able to indicate the profitability for final nickel extraction. This

information should ideally be made available to plant metallurgist before a significant expense

is incurred in setting up leaching situations on site for essentially refractory ores.

! 28!

Figure 1.10. Structure of goethite showing different projections (Cornell &

Schwertmann, 2003)

1.9 Summary of the problem

The use of mineralized lateritic surficial material as a source of metals has increased in recent

years due to the diminishing availability of traditional primary sulfide based mineralization.

Furthermore, because these ores occur at the surface and are therefore easily accessible,

they can be far more economic than primary ores as a source of metals. This is true even

though the actual concentrations of the economic minerals in laterites are lower than in the

more traditional primary sulfide ores.

However, there is a significant problem associated with extraction of metals from surficial

oxide deposits in that the relative refractory nature of the primary nickel rich mineral, goethite,

a secondary hydrated iron oxide mineral produced by weathering iron ores, can severely limit

! 29!

the extractive efficiency of both chemical and bacteriological agents. Consequently, even

relatively high grade deposits can be rendered extractively uneconomic depending on the

percentages of refractory goethite present.

With this obviously significant problem in mind, it would be economically extremely beneficial

to be able to characterise the degree of extractability, and therefore economic value, of ores

prior to establishing costly plant facilities on site and only then find out that the ore being

processed is refractory. At present the test procedures involved in determining the extraction

potential of these ores includes processes that can last over six months and the expenditure

of a significant amount of money; with concomitant waste of time if the ores prove to be

largely economically un-extractable. Characterization of nickel laterite ores, suited to heap

leach processing, is therefore important in order to select those that are the most reactive and

provide the highest recovery of the metal.

Consequently, the research detailed in this thesis is designed to try and develop a fast

extractive chemical leaching protocol to assess the reactiveness of nickel laterites for heap

leaching regimes and thereby overcome the significant expenditure and time requirements

associated with the traditional longer-term field trials currently used to determine nickel

extractability.

1.10 Aim of this research

The purpose for the current research is to develop a fast, economic and easy to perform test

to determine the extractability of nickel from nickel laterites which will identify and quantify the

amount of both refractory and non refractory goethite. Initially, different chemical leaching

agents (weak and strong acids, complexation compounds and chelation compounds) were

trialled on selected ores that are known to be refractive and non-refractive with respect to

nickel extraction. These ores were supplied by CSIRO Division of Process Science and

Engineering. Staff from this Division of CSIRO jointly supervised the project. The extraction

chemicals were trialled individually and in combination and at different pH levels where

! 30!

possible. In addition, the crystal structure of the goethite will also be investigated to establish if

crystalline structure affects the reactivity of the goethite. This detail of information will possibly

facilitate:

1 Understanding how nickel is bound in refractory and non-refractory goethite.

2 Determination of the optimum extraction method and leach conditions necessary to

separate the nickel from the goethite.

3 Determining how calcination affects the leachability of goethite.

4 Developing a mathematical model, that can be used on material that has been leached

between 48 and 250 hours, to establish the ultimate extraction efficiencies for nickel from

different ores that contain goethite.

It must be noted that this project is not designed to develop a new leaching technique for ores

that are rich in refractory goethite. It is designed simply to develop a fast predictive technique

that will establish the potential of ores to be leached by existing processes. Consequently, the

chemical mixtures that are used are extremely diverse as far the purpose of this study; they

do not have to be economic. Data from this study will be applied to produce a predictive model

of ore extractability. If successful, any developed technique could be used by mining

companies to quickly establish the extent to which an ore is chemically leachable and thereby

assess the economic potential of these surficial deposits in a far more efficient, cost effective

and timely manner.

! 31!

CHAPTER 2

EXPERIMENTAL

2.1 Equipment

Three main analytical techniques were use during this research, Inductively Couple Plasma

Mass Spectrometry (ICP-MS), Inductive coupled plasma Atomic Emission Spectroscopy (ICP-

AES) and X-Ray Diffraction (XRD).

All chemically related analytical data were compared to standard leach graphs prepared

following field trials undertaken by staff from CSIRO Division of Process Science and

Engineering. Inductively Coupled Plasma Mass Spectrometry (ICP-MS) and Inductively

Coupled Plasma Atomic Emission Spectroscopy (ICP-AES) were used to determine the

elemental concentration of all study analytes in the leach experiments. Both ICP-MS and

ICP-AES have been used in previous mineral characterization studies to determine

elemental composition and are accepted analytical techniques (Kisakurek, B. et al 2004).

These techniques have variable levels of sensitivity, depending on the specific emission

lines and isotopes determined and can therefore be used to produce optimum, high

accuracy, analytical data for the study analytes. A drift solution was used for both ICP-AES

and ICP-MS analyses to ensure inter-run comparability of data and Rh and Ir (2ngmL-1 in

solution) were also used as internal standards for ICP-MS analyses to correct for both drift

and matrix effects in the samples and also to ensure inter-run comparability of data

(Saitoh, K. et al 2002). X-ray diffractometry (PANalytical Empyrean XRD) was used to

determine the mineralogy of the samples and help interpret any relationship that may exist

between hematite and goethite.

! 32!

2.1.1 Inductively Coupled Plasma Atomic Emission Spectroscopy (ICP- AES)

Inductively Coupled Plasma Atomic Emission Spectroscopy (ICP-AES) is an analytical

technique that uses an inductively coupled argon plasma to generate excited atoms and ions

which emit characteristic electromagnetic radiation specific for an individual analyte. The

technique quantifies the emission intensity of both ion and atom lines and relates these values

to a calibration curve for each analyte to quantify the relevant analyte concentration.

ICP-AES was used in this study to quantify the concentrations of Na, K, Al, Mg, Ca, and Fe as

these analytes are present at relatively high concentrations in samples and therefore

quantification does not require the high sensitivity of ICP-MS to ensure accuracy. Analytical

emission lines used in this study are detailed in Table 2.1.

Table 2.1 Analytical emission lines used in this study, wavelengths are given in

nanometres

Na 589.5 Na 818.3 Mg 279.5 Mg 285.2 Al 167.0 Al 396.1 Si 251.6 P 177.4

P 178.2 S 180.7 K 766.4 Ca 422.6 Ti 334.9 Cr 283.5 Cr 359.3 Mn 257.6

Mn 279.4 Fe 239.5 Fe 259.9 Co 230.7 Co 237.8 Ni 221.6 Ni 341.4 Cu 324.7

Zn 206.2 Zn 213.8 Sr 407.7 Ba 455.4 Pb 220.3 !! !!

Although 19 elements were determined only data for Fe are reported in this thesis. Initially, as

it was not possible to be specific as to the analytes that would assist in the interpretation of

data, as many as possible were determined using both ICP-AES and ICP-MS techniques.

Following analysis and preliminary data interpretation, a refined list of analytes was developed

and data for the unreported analytes were used for cross validation of ICP-MS data to confirm

reproducibility and accuracy of the final data sets rather than as interpretive analytes in their

own right.

! 33!

A Thermo Scientific ICAP 6000 series ICP spectrophotometer was used throughout this

thesis. Analytical solutions needed to be diluted in order to contain <2% TDS. Following

dilution, quantization is achieved with reference to data for calibration standards (0-100µgmL-

1). These standards were supplied by AccuTrace (2986 Scott Blvd Santa Clara, CA 95054,

United States) and were diluted to the appropriate range using 18MegΩ deionised water.

Redistilled nitric acid (quartz sub-boiling still made by Walhofstrasse 14, 25474 Ellerbek

Germany) was added to the final solution to give a final acid strength of 5% v/v before sample

dilution to the appropriate final volume. Analytical samples were diluted to an appropriate final

volume in the same manner.

2.1.2 Inductively Coupled Plasma Mass Spectrometry (ICP-MS)

ICP-MS was used to determine the concentration of 46 analytes. The list of isotopes and their

mass to charge ratios used (m/v) is given in Table 2.2. Although ICP-MS is an extremely

sensitive analytical technique, it is apparent that the study analytes are present at relatively

high concentration. Consequently, in this study, the technique was not being used to

determine low concentrations but more specifically to overcome spectral overlap problems,

and resulting data inaccuracy, that could occur when using ICP-AES. Nevertheless, while the

analytes eventually selected for this study were at relatively high concentrations, many of the

analytes determined were at very low levels and did require a technique with extremely high

sensitivity to be used for their accurate quantitation. As with data produced using the ICP-

AES, not all the analytes determined were directly used in data interpretation in this study. In

some cases, as with ICP-AES data, some analytes, not directly reported in the text, were used

for cross calibration and validation between the two techniques but in addition, as at the

beginning of the study the list of relevant analytes could not be established without

experimentation, it was decided to record all analytical data and eliminate data retrospectively,

that were unnecessary to the interpretation of extraction efficiency.

! 34!

The ICP-MS instrument used was an Agilent 7500 Cs Inductively Coupled Plasma Mass

Spectrometer (Agilent Technologies, Inc., Santa Clara, CA, USA). The Standards used for the

calibration were obtained from AccuTrace (2986 Scott Blvd Santa Clara, CA 95054, United

States). The instrument calibration standards (1, 2, 5 and 10 ngmL-1) were prepared by mixing

the relevant standards (taking account of insolubility criteria) and diluting them to the

appropriate concentration using 18MegΩ deionised water. The final solutions were made to

volume to contain 2% v/v quartz redistilled nitric acid. Before analysis, samples were made up

to final volume with a TDS of <500 µgmL-1 by dilution using 2% v/v quartz redistilled nitric acid,

Rh and Ir were then added to these solutions, as internal standards, to give a final

concentration of each of 2ngmL-1.

Table 2.2 List of isotopes and their m/z values used in the ICP-MS based analytical procedure

7Li 23Na 24Mg 25Mg 27Al 39K 42Ca 44Ca 45Sc 48Ti

49Ti 51V 52Cr 53Cr 55Mn 57Fe 59Co 60Ni 63Cu 64Zn

65Cu 66Zn 71Ga 72Ge 85Rb 88Sr 89Y 90Zr 91Zr 93Nb

96Mo 98Mo 118Sn 120Sn 121Sb 133Cs 138Ba 139La 140Ce 141Pr

146Nd 147Sm 153Eu 157Gd 159Tb 163Dy 165Ho 166Er 169Tm 172Yb

175Lu 178Hf 181Ta 206Pb 207Pb 208Pb 232Th 238U

Rhodium and iridium were added as internal standards to facilitate normalization of data in the

event of drift during the analytical run. Samples were diluted to the appropriate concentrations

using 2% v/v quartz redistilled nitric acid in 18MegΩ quartz redistilled water.

2.1.3 XRD

A PANalytical Empyream XRD was used to determine the mineralogical composition of the

study samples. The system operated at 10mA and 40 kV using a copper Kα source.

Sample preparation, run time and angular movement (10o 2θ to 80o 2θ) per min from start

! 35!

and finish were also pre-determined and a typical set of instrumental parameters used

throughout this study is detailed in Appendix Table 1.

2.2 Background to the samples

Laterite samples were provided by the CSIRO Division of Process Science and

Engineering. All information regarding grain size, element composition and nickel

extraction capability were also provided by CSIRO. Due to confidentiality issues sample

location cannot be provided in this thesis and samples are referred to by an alpha numeric

code only in the text. However samples with the same alpha character are from the same

deposit and the numeric character reflects a different geographic location within that

deposit.

A total of eight samples were provide to test and develop a method that could ultimately

facilitate the development of a suitable predictive protocol to determine the relative

refractory and non refractory nature of the goethite present. All the samples contain

refractory and non refractory goethite in different quantities and therefore exhibit different

extraction efficiencies. A list of the different samples with their nickel extraction efficiencies

as provide by CSIRO is given in Table 2.3.

Figure 2.1. General representation of the experimental protocol used in this study

! 36!

Table 2.3. Description of samples used in this study together with their nickel extraction efficiencies

as determined by CSIRO Division of Process Science and Engineering

Sample Description Percentage long term nickel extraction achievable

Sample H Deposit A Area 1 4%

Sample B Deposit A Area 2 8%

Sample C Deposit B Area 1 50%

Sample D Deposit C Area 1 25%

Sample E Deposit D Area 1 40%

Sample F Deposit D Area 2 40%

Sample A Deposit D Area 3 66%

Sample G Deposit E Area 1 70%

2.3 Phase 1 Preliminarily Experiment using Sample A and B.

2.3.1 General

Initial experiments to determine nickel extractability focussed on two ores, sample A and B.

These ores represented the two extremes of extractability in percentage terms, with sample A

having 69% goethite and 66% nickel extraction efficiency and sample B having 50% goethite

and 8% nickel extraction efficiency. It was considered that using ores with such disparate

nickel extraction efficiencies would facilitate the development of a preliminary model to

determine the extraction efficiency of the other ores and develop a model into which all ores

could then be incorporated or referred.

Consequently, for this preliminary stage only samples A and B were used. Initially,

samples were analysed following total digestion, to determine if there would be any

significant difference between CSIRO results and study analyses. Samples were dissolved

using aqua regia and a mixture of perchloric, nitric and hydrofluoric acids. As the samples

! 37!

have been provided already crushed to a fine powder the leach was conducted without the

additional sample modifications that are outlined previously (Li J et al 2009).

Once the total elemental concentration of the samples had been determined, three basic

generic leaching experiments, organic, inorganic and mixed organic/inorganic leaching, were

conducted both samples over a period of 240h.

In addition, these samples were leached in H2SO4 following calcination at pre-determined

temperatures (250ºC, 400ºC, 500ºC, 600ºC, 750ºC and 1000ºC). An added advantage of

using these two samples was that they were provided by CSIRO in a relatively large quantity

(approximately 200g each) making it possible to undertake a series of different leach

experiments without completely running out of sample. It must be stressed that it is extremely

difficult to obtain these ores in even relatively small quantities so a mass of this amount is

usually very difficult to obtain. Consequently they were ideal for the preliminary leach studies

as they represent this extreme range of extractability and differences between results within

the leach regimes used would be easily identifiable. XRD was also undertaken on both

samples before and after calcination at the predetermined temperatures selected.

2.3.2 Total Dissolution

To ensure accuracy of data, a total digestion was also undertaken of two South Africa

Reference Materials, SARM 1 (granite) and SARM 2 (syenite). These Certified Reference

Materials (CRM’s) were dissolved in exactly the same manner as the study samples and

comparison of the data obtained from their analysis was used to check and determine the

accuracy of the instrumental analysis protocols used in the current study. Although

accuracy of leach protocols is not really possible to determine, the determination of

analytical accuracy of the method does give a significant amount of confidence in data

used to construct the final data set. Approximately Two hundred and fifty milligrams of

each sample were accurately weighed to 3 decimal places and dissolved inside 50mL

! 38!

Teflon beakers using 20mL of a 4:1 mixture of redistilled nitric:perchloric acids and 10mL

of hydrofluoric acid. The samples were heated and left to reflux at 150oC overnight by

placing a Teflon watch glass on top of each beaker. After this, the watch glasses were

removed and the samples evaporated to incipient dryness after which 5 mL of 18 MegΩ

water were added to solubilise the residues. Finally the samples were made up to 50mL

final volume using 18MegΩ water and the final mass of solution determined by weighing to

three decimal places. Final dilution to the required TDS content, and to achieve optimum

analytical range for the analytes, was undertaken by performing a 25 fold dilution (ICP-

AES sample) and 50 fold dilution (ICP-MS sample). Data produced were used as a

reference for the potential amount of nickel that could theoretically be leached and were

compared with data from leach tests undertaken separately on each sample.

2.3.3 Inorganic Acid Leach – General

Leach tests were undertaken on samples A and B (particle size less than 75µm), using

specific acid leach solutions. Approximately 0.25g of each sample was placed into 55mL

polypropylene screw topped tubs, to which 50mL of the appropriate leach solution were

added. Duplicate extraction trials were undertaken on each sample. The tubs were sealed and

transferred to a bottle roller for a total period of 240 hours, during which time the tubes were

rolled at a constant 30rpm.

2.3.3.1 Sulphuric Acid (H2SO4) Leach

This procedure was based on the reaction of 20% v/v sulphuric acid with samples over a

period of 240 hours. The samples were rolled at 30rpm and subsamples removed at pre-

specified times throughout the run for analysis. Data for these analyses were used to produce

leach curves based on time of association of the leach solution with the sample.

! 39!

2.3.3.2 Hydrochloric Acid (HCL) Leach

Extraction protocols have already been described in section 2.3.3.1. The only modification

was that in this leach procedure 20% v/v hydrochloric acid (HCl) was used as the leach

solution.

2.3.3.3 Aqua Regia Acid Leach


was that in this leach procedure a solution of Aqua Regia (1:3 HNO3 : HCl) was used as the

leach solution.

2.3.3.4 Nitric Acid (HNO3) Leach


was that in this leach procedure 20% v/v nitric (HNO3) acid was used as leach solution.

2.3.4 Organic Leach-General

Similar to section 2.3.3, an organic leach was undertaken using three different organic acids,

oxalic (H2C2O4), citric (C6H8O7) and tartaric (H6C4O6 ). The concentrations trialled for these

leach experiments were 0.2, 0.5 and 1M as outlined in the literature (McKenzie, D. et al 1987).

Samples and their duplicates were placed on a bottle roller, rotating at 30rpm, and leached for

a total of 240 hours. Representative leach solutions were taken from each of these samples at

24h intervals throughout the experiment.

2.3.4.1 Oxalic Acid (H2C2O4) Leach

Similar to the extraction protocol that has already been described in section 2.3.3.1, This

procedure was based on the reaction of 0.2, 0.5 and 1M oxalic (H2C2O4) acid with samples

over a period of 240 hours. The samples were rolled at 30rpm and subsamples removed at

pre-specified times throughout the run for analysis. Data for these analyses were used to

produce leach curves based on the time of association of the leach solution with the sample.

! 40!

2.3.4.2 Citric Acid (C6H8O7) Leach


was that in this leach procedure citric acid (C6H8O7) was used as the leach solution.

2.3.4.3 Tartaric Acid (H6C4O6 ) Leach


was that in this leach procedure tartaric acid (H6C4O6 ) was used as the leach solution.

2.3.5 Mixed Acids

Similar to the protocol describe in section 2.3.3 and 2.3.4, samples were leached with a

mixture of 10% sulphuric and 0.5M organic acid solutions (oxalic, citric and tartaric). The

leaches were undertaken for a total time of 264h, and samples were taken at time intervals

of 24h throughout the run.

2.3.5.1 10% v/v Sulphuric (H2SO4) and 0.5M oxalic Acid (H2C2O4) Leach .

Similar to the extraction protocol that has already been described in section 2.3.4.1; this

procedure was based on the reaction between samples and 10% v/v sulphuric (H2SO4) mixed

with 0.5M oxalic acid (H2C2O4) over a period of 264 hours. The samples were rolled at 30rpm

and subsamples removed at pre-specified times throughout the run for analysis. Data for

these analyses were used to produce leach curves based on time of association of the leach

solution with the sample.

2.3.5.2 10% v/v Sulphuric (H2SO4) and 0.5M citric acid (C6H8O7) Acid Leach


was that instead of sulfuric and oxalic acids, sulfuric and citric acids were used as the leach

solution.

! 41!

2.3.5.3 10% v/v Sulphuric (H2SO4) and 0.5M tartaric acid (H6C4O6 ) Acid Leach


was that instead of sulfuric and oxalic acids, sulfuric and tartaric acids were used as the leach

solution.

2.3.7 Oxidation leach

To further understand the crystalline structure of the laterite, and aid in the development of

a mathematical model to determine their extraction efficiency, solid samples of the laterites

were also analysed using X-ray powder diffraction (XRD). These data were used as a

baseline for calcination studies to determine the effect of increased temperature on the

nickel extraction efficiency of the samples. To investigate the effect of calcination on nickel

extraction, samples A and B were heated in a furnace at 250 oC, 400 oC, 500 oC, 600 oC,

750 oC and 1000oC for two hours. Each sample was subjected to XRD analyses to identify

the mineralogical changes associated with calcination. These samples were then leached

inside 100mL plastic containers placed on a bottle roller operating at 30rpm, for 240h. The

leach solutions used in this study consisted of 20% v/v H2SO4, 20% v/v HNO3 and 20 % v/v

HCl.

2.3.7.1 Sulphuric Acid (H2SO4) Leach

This procedure was based on the reaction of 20% v/v sulphuric acid with samples over a

period of 240 hours. The samples were rolled at 30rpm and subsamples removed at pre-

specified times throughout the run for analysis. Data for these analyses were used to produce

leach curves based on time of association of the leach solution with the sample.

2.3.7.2 Hydrochloric Acid (HCl) Leach


was that in this leach procedure 20% v/v hydrochloric acid was used as the leach solution.

! 42!



was that in this leach procedure Aqua Regia was used as the leach solution.

2.3.7.4 Nitric Acid (HNO3) Leach


was that in this leach procedure 20% v/v nitric acid was used as the leach solution.

2.4 XRD

To determine if changes in the mineralogy of the samples were associated with each stage

of the calcination experiment, and thereby understand the effect of temperature on the

samples and any variation in the leach characteristics of these materials, all sub-samples

of A and B, generated throughout the calcination experiment, were subjected to x-ray

diffraction analysis.

2.5 Phase 2: Validation samples.

In addition to the two initial samples on which the majority of research time was expended, the

CSIRO provided an additional six samples from deposits around western Australia that had

different nickel extraction characteristics. The initial concept was that these samples could be

used to validate any novel analytical protocols that were developed during the research

undertaken in this thesis. However, in addition to this potential use it was considered that

these samples could possibly be used to produce a direct calibration of long term nickel

extraction efficiency (data from field trials) with respect to time dependent nickel extraction

efficiency, under lab-based experimental conditions. In this way it was hoped that a

‘calibration graph’ may be able to be produced which would mathematically relate the long

term nickel extraction efficiency of these samples (field trials) to nickel extraction efficiency

over a much shorter time (lab-based experimental procedure). It was envisioned that these, or

! 43!

equivalent samples, could then be used in lab based experiments when new ores were

investigated to quickly give an indication of what would be the long term nickel extraction

efficiency of the new ore being trialled under field conditions. This approach would save a

significant amount of time in determining the refractivity of the goethite present in this ore.

Of the samples provided by the CSIRO, only six samples were suitable for use in developing

the relevant extraction efficiency graphs as only these had sufficient mass of material for the

experiments required (>20g) (Table 3).

For this series of experiments both the total concentration of the study elements and their

extraction efficiency in sulphuric acid had to be undertaken. Total nickel concentrations in

samples C (50%), D (25%), E (40%), F (40%), G (70%), H (4%) were determined by

dissolving 0.25 g of each sample using the same procedure as detailed in section 2.3.3. For

the leach procedure, 0.25 gram of each sample was leached under the same protocol as in

section 2.3.3.1.

2.6 Samples and Data Analysis

Experiments were undertaken in duplicate and as data reproducibility gave results that

different by less than +/-1% relative, average data were used for comparison purposes

throughout this thesis. The main method of data comparison and interpretation was graphical

plotting and visual comparison of the leach curves. This approach, while it may not be

considered as sophisticated, is the industry standard method for this type of data

interpretation, and is entirely adequate when interpreting leach characteristics of ores. This

relatively simple approach also means that data cannot be over-interpreted which can be a

danger especially when considering the variable nature of ores within and between deposits

and trying to develop a generic method associated with their extraction potential.

! 44!

Chapter 3

Results and Discussion

3.1 Phase 1: Preliminarily Experiment using “Ore Sample A” and “Ore Sample B”

3.1.1. Total Dissolution

Analytical data obtained for concentrations of Fe, Mn, Co and Ni in SARM1 and SARM2 were

compared with the Certified Values (Table 3.1) to provide an indication of the accuracy of the

analytical technique and results for the nickel extraction efficiency experiments. The data were

also used to validate the accuracy of total concentrations of analytes in the original samples.

While it is realised that it is not possible to have Certified Reference Materials for partial leach

experiments, and that the data detailed in Table 3.1 represent data produced for total

dissolution, this exercise did serve to validate the methodological accuracy and thus

confirmed the accuracy of the standards used for partial leach data calibration. Data obtained

for nickel in the Reference Materials vary slightly from certified values. However, this is largely

due to the fact that in these samples this analyte is at an extremely low level (near the

detection limit for the technique), a situation that does not occur into the lateritic samples used

in the study and no Reference Materials were available for use to establish the accuracy of

the nickel in iron ore. The reason for the choice of these four analytes is that the relationship

between nickel and iron, and manganese and cobalt has shown a direct correlation between

these two binary sets of metals (Hallberg et al. 2011; Yongue-Fouateu et al. 2006).

Following instrumental calibration, “Ore Sample A” and “Ore Sample B”” were analysed for

the four study elements and results detailed in Table 3.2.

! 45!

Table 3.1 Comparison of analytical data for Certified Reference Materials SARM1 and

SARM2 obtained for this study. Results expressed as parts per million in the original material.

CRM Type Type Mn ppm Fe ppm Co ppm Ni ppm

SARM 1 Certified Value Granite 150 14100 0.3 6.2

SARM 1 Study Value Granite 155 14300 0.4 8

SARM 2 Certified Value Syenite 79.9 10200 3 8.6

SARM 2 Study Value Syenite 77.5 10100 3 7

Table 3.2. data for Fe, Ni, Co and Mn concentrations in Samples “Ore Sample A” and “Ore Sample B”

Ore Sample

Percentage extraction efficiency in field leach long term experiments

Iron concentration %

Nickel concentration %

Cobalt Concentration (ppm)

Manganese concentration (ppm)

A 69% 45.6 2.43 496 3760

B 8% 33.3 0.944 692 2110

It is apparent from data in Table 3.2 that there is no direct correlation between nickel

concentration and the extraction efficiency of the ores, with nickel being only approximately

two and a half times more concentrated in “Ore Sample A” than in “Ore Sample B” while

the nickel extraction efficiency for “Ore Sample A” is over eight times higher than for “Ore

Sample B”.

3.1.2 Inorganic Acid Leaching

Results for all experimental data are detailed in Appendix Table 2.

3.1.2.1 Sulphuric Acid Leach

The 240hour extraction profiles of the four analyte elements, using 20% v/v sulfuric acid, are

detailed in Figure 3.1A and B. It is obvious from these figures that the relative extraction

efficiency of the analytes from the two different samples is completely dissimilar. Sulphuric

acid attacks the sample, and to a certain extent dissolves the iron, which is the major

constituent of goethite, leaching available nickel from the goethite matrix. The amount of

nickel leached is directly dependent of the degree of decomposition of the host matrix, which

! 46!

in turn is directly related to the amount of refractory goethite present in the sample and the

amount of nickel associated with this material.

!The final nickel extraction efficiency, as determined by CSIRO Division of Process Science

and Engineering in long term tests, was approximately 8%. However, this was achieved in

column leach trials which do not necessarily expose all the surfaces of the sample to the leach

solution as channelling!!in the leach bed may occur.

!

Figure 3.1. Nickel extraction efficiency for “Ore Sample A” (1A) and “Ore Sample B” (1B) using 20% v/v H2SO4 over a 240h period

In the experiment undertaken in this thesis, the samples are rolled and consequently this

process ensures that all the surfaces of the sample are exposed to the leach solution. The

rolling process also ensures that the leached surface coatings on the sample grains can be

abraded and removed and that new surfaces are repeatedly exposed to the leach solutions.

Under these circumstances the effective final leach percentage, for a bottle rolled test, should

be higher than that of a column leach experiment.

If the problem with different leach efficiencies is simply a result of a build-up of a surface

barrier which is impenetrable to leach solutions under leach column conditions, rolling the

sample will overcome this. If not, then there should still be an equivalent variation between the

0.0!

5.0!

10.0!

15.0!

20.0!

25.0!

30.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample A: H2SO4

Ni!

Co!

Fe!

Mn! 0.0!

1.0!

2.0!

3.0!

4.0!

5.0!

6.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample B: H2SO4

Ni!

Co!

Fe!

Mn!

A B

! 47!

two samples when both sets of test results are compared. The extraction curves for iron,

cobalt and manganese for “Ore Sample A” (Figure. 3.1A) all seem to be parallel to each other

and maintain linearity over the entire extraction period, while the extraction curve for nickel is

considerably suppressed with respect to those of these other elements. The nickel extraction

curve is nevertheless linear. From the results for “Ore Sample B” detailed in Figure 3.1B, it is

apparent that while the slopes for the manganese and cobalt extraction curves parallel each

other, the slope for iron extraction seems to decrease after the first +/-100 hours of leaching

but at all times be greater than the extraction slope for nickel. The nickel extraction curve

seems largely to parallel those for manganese and cobalt. The implication here is that while

there is a greater extraction efficiency of nickel from “Ore Sample A” than from “Ore Sample

B”, there could be a different mechanism holding the nickel in the lattice of the goethite of the

two different samples as proportionately more iron is being extracted from “Ore Sample A”

than from “Ore Sample B”. However, at the end of the 240h extraction period, the relative

extraction efficiencies of nickel from the two samples are approximately 2.7:1 :: “Ore Sample

A”: Ore Sample B”, while their final relative extraction efficiencies based on CSIRO Division of

Process Science and Engineering data are 8.6:1. Consequently, it can be clearly

demonstrated that while differences do occur in the extraction efficiency of nickel from the two

samples, and while simple build-up of an impenetrable surface coating on sample grains is not

the reason for the variation in nickel extraction efficiencies between the two samples, their

relative extraction efficiencies after 240hours do not represent the final extraction efficiencies,

and hence the extractability of the ore, and the shorter term rolling trial using 20% v/v sulfuric

acid, cannot be used to predict results from the longer term test using the same medium.

3.1.2.2 Hydrochloric Acid Leach

One attractive alternative to extract nickel and other elements such as Fe and Co is

hydrochloric acid. Senanayake et al. (2011) reported that it is possible to effectively leach iron,

cobalt and nickel from limonitic ores by using hydrochloric acid (HCl). The reason for this is

that the HCl reacts with the iron to form an extremely soluble FeCl63- complex thus effectively

! 48!

dissolving the iron oxide and releasing co-associated elements into solution. Consequently an

experiment was undertaken where 20% v/v HCl was used as an extractants for “Ore Sample

A” and “Ore Sample B”. Results for the 240h extraction of these two samples are detailed in

Figures 3.2A and B.

It is apparent from these figures that the extraction curve for “Ore Sample B” starts to

asymptote the ordinate (time) axis after approximately 200h leaching (Figure. 3.2B) when

nickel extraction has reached approximately 3% efficiency. However, even after 240h the

nickel leach curve for sample “Ore Sample A” maintains linearity and slope. However, while

the relative extraction efficiency of both the 20% v/v H2SO4 leach and the 20% v/v HCl leach,

when applied to sample “Ore Sample B”, appear to give equivalent results for nickel extraction

efficiency, the equivalent extraction efficiency of nickel from “Ore Sample A” (Figure 3.2A) is

reduced from approximately 8% for the 20% v/v H2SO4 leach to 6% for the 20% v/v HCl leach,

implying that HCl is not quite as effective a medium for nickel extraction as H2SO4. One other

extremely important point is that whilst, for the 20% v/v H2SO4 leach, the relationships

between iron and nickel and between cobalt and manganese were not apparent, for the 20%

v/v HCl leach there appears to be an extremely strong relationship between the leach

efficiency of these two binary pairs. For both samples, the extraction curves for the analytes

differ significantly in slope; however, for both samples the binary relationship between nickel

and iron and between cobalt and manganese is strongly apparent.

However, as with the 20% v/v H2SO4 leach, the 20% v/v HCl leach cannot be used, over the

short term, to predict the relative differences in nickel leach efficiency of the long term, field-

based leach protocol.

! 49!

Figure 3.2. Nickel extraction efficiency for “Ore Sample A” (2A) and “Ore Sample B” (2B) using 20% v/v HCl over a 240h period.


The third leach solution investigated was Aqua Regia, a mixture of 3:1 parts by volume

concentrated hydrochloric and nitric acid. The choice of this acid mixture is that because of

the production of the OCl- radical, it is extremely efficient for the dissolution both ferrous and

non-ferrous metals and their oxides.

Figure 3.3. Nickel extraction efficiency for “Ore Sample A” (3A) and “Ore Sample B” (3B) using 20% v/v Aqua Regia over a 240h period.

0.0!

5.0!

10.0!

15.0!

20.0!

25.0!

30.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample A: HCl

Ni!

Co!

Fe!

Mn!0.0!

1.0!

2.0!

3.0!

4.0!

5.0!

6.0!

7.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample B: HCl

Ni!

Co!

Fe!

Mn!

0.0!5.0!10.0!15.0!20.0!25.0!30.0!35.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample A: Aqua regia

Ni!

Co!

Fe!

Mn! 0.0!1.0!2.0!3.0!4.0!5.0!6.0!7.0!8.0!9.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample B: aqua regia

Ni!

Co!

Fe!

Mn!

A B

A! B!

! 50!

From the data detailed in Figure 3.3, it is apparent that the extraction efficiency of nickel in a

20% v/v Aqua Regia solution is slightly better than for the equivalent 20% v/v hydrochloric

acid solution, and almost equivalent to using 20% v/v sulfuric acid. Again for this extraction

matrix there is a very lose association of the extraction of the binary pairs nickel and iron and

cobalt and manganese. However, as with all previous leach regimes, the extraction efficiency

of nickel from the two samples cannot be used to predict the extraction efficiency of this

element over the long term or suggest the relative extraction efficiencies of nickel from the two

samples.

3.1.2.4 Nitric Acid Leach

While nitric acid is not recommended for the dissolution of iron oxides, its ability to dissolve

nickel oxide is exceptional. Consequently, it was considered that the leaching potential of this

acid should be investigated. Results detailing the extraction curves for solutions of 20% v/v

nitric acid leach solutions are detailed in Figure. 3.4. The results for the extraction efficiency of

this solvent are considerably different from those of the previous three inorganic acid based

leaches. In the case of nitric acid the level of extraction of all analytes is significantly lower

than for all other extraction procedures tested.

Figure 3.4. Nickel extraction efficiency for “Ore Sample A” (4A) and “Ore Sample B” (4B) using 20% v/v nitric acid over a 240h period.

0.0!1.0!2.0!3.0!4.0!5.0!6.0!7.0!8.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample A: HNO3 Ni!

Co!

Fe!

Mn!

0.0!

0.5!

1.0!

1.5!

2.0!

2.5!

3.0!

3.5!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Ore Sample B: HNO3 Ni!

Co!

Fe!

Mn!

A! B!

! 51!

This may be because the dissolution chemistry of nitric acid results in it not being as efficient

for the dissolution of iron oxides as either sulfuric or hydrochloric acid and without the

dissolution of the sample matrix it is not possible to extract the analytes. Consequently, not

only does this extraction medium fail to extract nickel at anywhere near the efficiency of the

other extractants used, it also cannot be used to predict the final extraction efficiency of nickel

in the long term extraction regimes. This is clearly shown in Figures 3.4A and 3.4B where it

can be seen that there is little difference between the extraction efficiency of nickel from both

“Ore Sample A” and “Ore Sample B”, a situation that is not indicated in any of the previous

leach experiments and certainly not indicative of long term leach trial results.

3.1.2.5 Summary of Inorganic Leach Experiments

In general, the leach efficiency for the extraction of nickel from “Ore Sample A” and “Ore

Sample B” was very similar for all acid mixtures used, with the exception of nitric acid (Table

3.3 and Figure 3.5). As can be seen numerically by reference to Table 3.3 and graphically by

reference to Figure. 3.5, there is absolutely no relationship between the extraction efficiencies

of nickel, under any of the leach regimes tested, to the relative relationship of 8.6:1 :: “Ore

Sample A”: Ore Sample B” obtained for the long term leach experiment.

Table 3.3. Leach data for extraction efficiencies of all study analytes after a 240h leach period detailing all inorganic acid based leach solutions studied.

Percent Extraction Efficiencies 240h

Ore Samples A B A B A B A B Leach solution Mn Mn Fe Fe Co Co Ni Ni 20% v/v H2SO4 24.2 4.5 21.4 5.6 21.2 5.4 7.7 3.2 20% v/v HCl 24.6 5.6 11.7 3 20.8 6.5 6.6 2.6 20% v/v HCl:HNO3 31.5 6.6 14.2 3.5 27 7.8 7.9 3.5 20% v/v HNO3 7.1 2.3 3.3 2.1 4.5 3.0 1.9 1.5

While the Aqua regia and sulphuric acid leach solutions have the highest efficiency for

nickel removal, the nearest relative extraction efficiency to 8.6:1::“Ore Sample A”: Ore

Sample B”, is approximately 2.5:1 for the 20% v/v hydrochloric acid leach. Consequently, it

must be concluded that it does not appear to be possible to use a short term leach, even

! 52!

under sample rolling extraction conditions, to predict the final long term leach results for

extractability of nickel from these ores.

Figure 3.5. Comparison of the nickel extraction efficiencies for all inorganic acid based leach solutions

3.1.3 Organic Acid Leaching


3.1.3.1. Oxalic Acid Leach

Results obtained for the oxalic acid leach trials are detailed in Figure 3.6.

Oxalic acid is a readily available, cheap, non-volatile (solid) organic acid which, while not

recognised as forming particularly soluble solid salts with the study analytes, certainly does

take them into solution and has found application in the field of mineral heap leaching

(Tarasova et al 2001). As with all the organic acid based leaching experiments detailed in this

thesis, three different acid strengths were used for all experiments, 0.2, 0.5 and 1.0M. Results

obtained for these leach trials are detailed in Figures 3.6A to 6F.

0.0!1.0!2.0!3.0!4.0!5.0!6.0!7.0!8.0!9.0!

0! 100! 200! 300!

Extrac'o

n*Effi

cien

cy*of*n

ickel*(%)*

Time*(h)*

Ore Sample A

Aqua!Regia!

HCL!

H2SO4!

HNO3!

A!

0.0!0.5!1.0!1.5!2.0!2.5!3.0!3.5!4.0!

0! 100! 200! 300!Extrac'o

n*Effi

cien

cy*of*n

ickel*(%)*

Time*(h)*

Ore Sample B

Aqua!Regia!

HCL!

H2SO4!

HNO3!

B!

! 53!

Figure 3.6. Analyte leach curves for “Ore Sample A” and “Ore Sample B” using selected molarity oxalic acid solutions

An extremely interesting aspect of these leach curves is that leaching of manganese appears

to be extremely fast initially with approximately 20% being leached from “Ore Sample B” within

the first 72 hours. This is in contrast to results obtained using inorganic acid leach solutions

0!

5!

10!

15!

20!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

0.2M Oxalic acid - Ore Sample A

Ni!

Co!

Fe!

Mn!

0!5!

10!15!20!25!30!35!40!45!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

0.2M Oxalic acid - Ore Sample B

Ni!

Co!

Fe!

Mn!

0!

5!

10!

15!

20!

25!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.5M Oxalic acid : Ore Sample A

Ni!

Co!

Fe!

Mn!

0!

5!

10!

15!

20!

25!

30!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.5M Oxalic acid : Ore Sample B

Ni!

Co!

Fe!

Mn!

0!

5!

10!

15!

20!

25!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

1M Oxalic acid - Ore Sample A

Ni!

Co!

Fe!

Mn!

0!

5!

10!

15!

20!

25!

30!

35!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

1M Oxalic acid - Ore Sample B

Ni!

Co!

Fe!

Mn!

C! D!

E! F!

A! B!

! 54!

where only approximately 2-6% of the available manganese was leached during the entire

240 hour experimental period. The leaching efficiency of this acid for nickel is, however, no

better than for any of the inorganic acid leachate solutions tested, while the leaching efficiency

for cobalt is significantly worse. This is surprising as cobalt is considered to be associated with

manganese in the goethite lattice. Extraction for iron from “Ore Sample A” follows a much

more conventional trend. However, unlike that or any of the inorganic acid leaches

investigated, none of the oxalic acid solutions investigated show any relation to the extraction

efficiency determined for nickel.

When results for leaching sample “Ore Sample A” are considered (Figures 3.6A, C and E), it is

apparent that the leach curves for all analytes are completely different from those for “Ore

Sample B” and follow much more conventional extraction amount vs. time curves. The relative

amount of iron leached is significantly higher for all concentrations of oxalic acid used than

determined for any of the inorganic acid leach solutions tested. Nickel extraction efficiency is

approximately 10% after 240 hours for all strengths of oxalic acid used. These variations may

be related to the relative amounts of refractory goethite present in the two ores, but if so it is

not an easily decipherable or understandable relationship. The comparison between the

relative extraction efficiencies for oxalic acid leaching of the two samples does not

approximate the required 8.6:1 relationship between “Ore Sample A” and “Ore Sample B”

respectively that is required if oxalic acid leaching is to be used to predict the final leach

characteristics of either ore. Consequently, the use of oxalic acid as a predictor for long term

leach efficiency of the ores under sulphuric acid leaching conditions cannot be considered as

appropriate.

3.1.3.2 Citric Acid Leach

This acid forms a slightly soluble citrate with iron and nickel but it can also from more soluble

mixed-metal complexes with these elements making it a potentially useful leaching agent for

! 55!

goethitic rich ores. Different studies have shown that this type of acid is effective for extracting

nickel from some laterite ores (Guang-hui et al. 2010). Leach curves for selected molarity citric

acid solutions are detailed in Figures 3.7A to F. These leach curves exhibit classical leach

percentage vs. time curves for both solids tested with further confirmation of the binary

association and behaviour of iron and nickel and of cobalt and manganese. The concentration

of acid seems to make very little difference to the extraction efficiencies of any of the analytes

over the 240 hour experimental period. Nickel extraction only reaches approximately 1% for

1M citric acid for both solids. An extremely interesting observation to be made from reference

to the extraction curves is that the relative extraction efficiency of the binary metal pairs (Fe/Ni

and Mn/Co) appears to be closely linked with the absolute metal extraction efficiencies of

each pair. This implies very strongly that as the iron is leached so too is the nickel and as the

manganese is leached so too is the cobalt. This extremely tight relationship has not been

seen before in any of the leaching experiments. While confirming the assumption of co-

consociation of these element pairs none of the leaching graphs show any relationship

between the extraction efficiencies for “Ore Sample A” and “Ore Sample B” that approximate

the 8.6:1 relationship for the relative long term sulfuric acid leachability of the ores.

Furthermore, any possibility of using this acid to predict long term nickel leachability of ores

will be further complicated by the fact that only approximately 1% of the available nickel

appears to be leached by any of the concentrations of citric acid used.

3.1.3.3 Tartaric Acid Leach

The extraction vs. time based curves for tartaric acid leach solutions are detailed in Figures

3.8A to 8F. The extraction curves for tartaric acid (Figures.3.8A-F) are similar to those for citric

acid (Figures 3.7A-F). This is not surprising as both acids behave similarly with the study

analytes. However, neither of these acids is as effective at leaching nickel as oxalic acid, with

the maximum nickel extraction being only approximately 1.5 % from either ore.

! 56!

Figure 3.7. Analyte leach curves for “Ore Sample A” and “Ore Sample B” using selected molarity citric acid solutions

0!2!4!6!8!10!12!14!16!18!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

Ni!

Co!

Fe!

Mn!

0.0!0.5!1.0!1.5!2.0!2.5!3.0!3.5!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

Ni!

Co!

Fe!

Mn!

0!2!4!6!8!10!12!14!16!18!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.5M Citric acid - Ore Sample A

Ni!

Co!

Fe!

Mn!

0.0!0.5!1.0!1.5!2.0!2.5!3.0!3.5!4.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.5M Citric acid - Ore Sample B

Ni!

Co!

Fe!

Mn!

0!

5!

10!

15!

20!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

1M Citric acid - Ore Sample A

Ni!

Co!

Fe!

Mn!

0.0!0.5!1.0!1.5!2.0!2.5!3.0!3.5!4.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

1M Citric acid - Ore Sample B

Ni!

Co!

Fe!

Mn!

A! B!

C! D!

E! F!

0.2M Citric Acid – Ore Sample A

!

0.2M Citric Acid – Ore Sample B

! 57!

Figure 3.8. Analyte leach curves for “Ore Sample A” and “Ore Sample B” using selected molarity tartaric acid solutions

0!

5!

10!

15!

20!

25!

30!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.2M Tartaric acid - Ore Sample A

Ni!

Co!

Fe!

Mn!0.0!0.5!1.0!1.5!2.0!2.5!3.0!3.5!4.0!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.2M Tartaric acid - Ore Sample B

Ni!

Co!

Fe!

Mn!

0!5!10!15!20!25!30!35!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.5M Tartaric acid - Ore Sample A

Ni!

Co!

Fe!

Mn!

0!1!1!2!2!3!3!4!4!5!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

0.5M Tartaric acid - Ore Sample B

Ni!

Co!

Fe!

Mn!

0!5!10!15!20!25!30!35!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

1M Tartaric acid - Ore Sample A

Ni!

Co!

Fe!

Mn!0!

1!

2!

3!

4!

5!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time (h)

1M Tartaric acid - Ore Sample B

Ni!

Co!

Fe!

Mn!

A! B!

C! D!

E! F!

! 58!

This is somewhat surprising as sample “Ore Sample A” is composed of an ore that is

extremely different in terms of its long term extractability to sample “Ore Sample B” and it

would be expected that at least there would be significantly greater nickel extraction from “Ore

Sample A” ore than from “Ore Sample B” ore material. As with the citric acid leach

experiment, the extraction curves for iron and nickel and for manganese and cobalt are

essentially equivalent. However, unlike the curves for these elements obtained using citric

acid, the extraction curves for these binary pairs appear to asymptote the abscissa far more

quickly than in equivalent leaching conditions using citric acid. The selective leaching of

manganese from both these ores using citric and tartaric acids could possibly be used as a

precursor to any long term nickel leaching protocols to remove cobalt at an early stage of

mine site extraction initiatives as approximately 25% of the available cobalt is removed with

only 1% of the iron and nickel and as the relative costs of these two metals is significantly

different this could be a profitable pre-leach action where cobalt concentrations in the ore are

relatively high.

Again the relative extraction efficiencies of the two ores, under tartaric acid leaching

conditions, in no way approximate the long term leach efficiencies of the ores under sulfuric

acid leaching conditions and therefore short term leach experiments using tartaric acid cannot

be used to predict long term leachability or refractiveness of either ore.

3.1.3.4 Summary Organic Acid based Leaching

Iron oxides, such as limonite and goethite are one of the main adsorbents for anions and

cations in soils and sediments. Consequently the secondary formation of these minerals,

following the weathering of metamorphically formed oxide based ores, will continue to have an

equivalent association of these precursor elements. It is not surprising therefore to find that

leaching characteristics of major elements, such as iron and manganese, are associated with

equivalent leaching characteristics for elements that were originally co-deposited with them.

! 59!

The potential use of organic acids is therefore important in that selective leaching of target

analytes may be possible, and relatively non-destructive leaching may be able to provide an

insight into the degree of association of analytes in refractory and non-refractory sites within

the final mineral assemblage of the ore. A summary of the results obtained for organic acid

based leaching of “Ore Sample A” and “Ore Sample B” is given in Table 3.4.

Table 3.4. Summary of results for leaching “Ore Sample A” and “Ore Sample B” ores using selected organic acids under pre-determined molarity concentrations

!! !! !! Percent!! Extraction! Efficiencies! After!240h! !! !!

Ore!Samples! A! B! A! B! A! B! A! B!

Leach!solution!! Mn! Mn! Fe! Fe! Co! Co! Ni! Ni!

Oxalic!0.2M! 12.5! 39.0! 19.0! 34.4! 15.4! 4.25! 9.14! 2.30!

Oxalic!0.5M! 13.4! 28.5! 19.8! 25.3! 17.4! 0.29! 9.69! 2.17!

Oxalic!1M! 15.8! 29.1! 22.4! 32.2! 19.2! 0.21! 11.3! 2.73!

Citric!0.2M! 13.9! 2.90! 0.45! 0.54! 14.9! 3.24! 0.97! 0.75!

Citric!0.5M! 15.4! 3.10! 0.53! 0.66! 16.6! 3.60! 1.10! 0.89!

Citric!1M! 16.6! 3.04! 0.57! 0.75! 17.9! 3.64! 1.23! 1.00!

Tartaric!0.2M! 27.6! 3.13! 1.13! 0.71! 23.2! 3.49! 1.99! 0.85!

Tartaric!0.5M! 27.8! 3.92! 1.42! 0.87! 25.0! 3.89! 1.94! 1.22!

Tartaric!1M! 32.3! 4.26! 1.77! 1.25! 26.2! 4.23! 2.60! 1.50!

From the data detailed in this table it is apparent that oxalic acid is by far the best leaching

agent of all the organic acids tested with up to 11.3% of all the available nickel leached from

the less refractory ore, “Ore Sample A”, and 2.7% leached from the more refractory ore “Ore

Sample B”. These figures equate to a 22.4% and 32.2 % extraction efficiency for iron from the

equivalent ores. Obviously the figures imply that the association of nickel and iron is different

in the two ores, a fact highlighted by the differences in their equivalent extraction efficiencies

for long term trials, but also that there is relatively an approximately 2:1 ratio of the

extractability for Fe:Ni for “Ore Sample A” while there is a 12:1 ratio for Fe:Ni extractability for

“Ore Sample B”. This points to the “Ore Sample B” ore being far more refractory than the “Ore

Sample A” ore, a fact further borne out by long term extraction tests. An inference could

perhaps be made that in using oxalic acid, as a short term Ore Sample B” would be some six

! 60!

times less than from “Ore Sample A”. While this figure is closer to the long term test ratio than

any before it is still some 50% out from the 8.6:1 ratio that has been experimentally confirmed

for the relative long term extractability of nickel from the two ores. There is an additional

interesting aspect of the oxalic acid leach experiments in that some 25% of the cobalt and

only 1.8% of the iron and equivalent nickel can be extracted from the relatively non-refractory

“Ore Sample A” ore using 1M oxalic acid. The consequent suggestion is that it may be useful

to pre-extract cobalt from the ores before undertaking the long term nickel extraction

procedure and thereby produce a relatively much less complicated leachate and one that

would contain a high proportion of the much more expensive cobalt rich fraction for

subsequent processing. In this way there may be a possibility to “value-add” to the more

cobalt rich ores. It is however obvious from these experiments that the use of any of the

organic acids tested in this study is unlikely to provide data that can be used as a predictive

tool for estimating the final leachability of either of the two study ores under long term leaching

conditions. Composite leach curves for nickel for “Ore Sample A” and “Ore Sample B” are

detailed in Figure 3.9.

!

Figure 3.9. Comparison of Leach Efficiencies of 1M Organic Acid Leach Solutions

0.0!

2.0!

4.0!

6.0!

8.0!

10.0!

12.0!

0! 100! 200! 300!NICKE

L*EX

TRAC

TION*EFFICIENCY

*(%)*

TIME*(H)*

Ore Sample A

Citric 1 M GSO1 Oxalic 1 M GSO1 Tartaric 1 M GSO1

0.0!0.5!1.0!1.5!2.0!2.5!3.0!3.5!

0! 200! 400!

NICKE

L*EX

TRAC

TION*

EFFICIEN

CY*(%

)*

TIME*(H)*

Ore Sample B

Citric 1 M K11 Oxalic 1 M K11 Tartaric 1 M K11

! 61!

3.1.4 Mixed Organic and Sulfuric Acid Leaching


Following the initial leaching experiments using inorganic acids and organic acids separately,

in was decided to run a simple trial by mixing sulfuric acid (10% v/v), which had been proven

to be the most cost effective and efficient leaching agent in the previous trials, with 0.5M

concentration of each of the three organic acids tested in this work. It was hoped that under

these conditions, the positive aspects of both types of leachate would be combined to provide

improved leachability for both ore types and by so doing the short term tests would give a

better indication of the relative leachability of both ores in long term trials. A leach mixture of

10% sulfuric acid and 0.5M organic acid was used in all experiments.

3.1.4.1. Sulfuric Acid and Oxalic Acid Leach

Details of the data produced for the oxalic acid/sulfuric acid leach trials using “Ore Sample A”

and “Ore Sample B” are detailed in Figure 3.10. When data generated for this mixed acid

extraction procedure are compared with equivalent data for the sulfuric acid leach

experiments, it is apparent that “Ore Sample B” extraction curves for Ni, Co and Fe are

significantly better using the mixed acid leachate, while there is a twofold better extraction for

manganese. The extraction curve for manganese is far more “classical” than it was in the

original 0.5m oxalic acid test but the absolute extraction efficiency of the 0.5m oxalic acid

leach has been reduced from approximately 30% to 5%. This is also true for iron where

extractability in 0.5 oxalic acid has been reduced from 25% to 5%. The extraction efficiency for

nickel from “Ore Sample B” has been increased from approximately 3% in 0.5M oxalic acid to

approximately 8% in the mixed acid leach. Cobalt extraction efficiency has increased from

<1% in 0.5% oxalic acid to over 14% in the mixed acid leach. For “Ore Sample A”, the

extractability of cobalt has increased from 17% in 0.5% oxalic acid to over 40% in the mixed

! 62!

acid leach, while nickel extraction efficiency has remained largely unchanged at approximately

10%.

Considering the mixed oxalic/sulfuric leach in general, there appears to be no advantage in

using mixed acid extraction in terms of increasing nickel extractability and certainly there is no

new relationship for the relative nickel extractability of the two ores that could be used as a

predictive model for determining the relative or actual long term extractability of either ore.

! ! Figure 3.10. Extraction efficiency for oxalic acid/sulfuric acid mixed acid leaching of “Ore Sample A” and “Ore Sample B”. 3.1.4.2 Sulphuric acid and Citric Acid Leach

Details of the data produced for the sulfuric acid/citric acid leach experiments are given in

Figure 3.11. There is a slight enhancement in the extraction efficiency of nickel from sample

“Ore Sample B” from 3% to 4% using a mixture of sulfuric and citric acids over the extraction

of nickel using sulfuric acid alone. With sulphuric acid, approximately 3% of the available

nickel is extracted from the ore while in the mixed acid leach approximately 10% of the nickel

is extracted. Iron extraction increased from approximately 5% to approximately 20% under the

same conditions for this ore. The increase is not due to the presence or absence of sulphuric

acid but could be to the complexing ability of the citric acid which assists in the maintenance

of the solubilized iron and consequently equivalent solubilization of the nickel. There is also a

disproportionate increase in the extractability of manganese and cobalt with an increase from

0!5!10!15!20!25!30!35!40!45!

0! 100! 200! 300!

Extrac'o

n*Effi

cien

cy*(%

)*

Time*(h)**

Oxalic+H2SO4 Leach :Ore Sample A

Nickel!

Cobalt!

Iron!

Manganese!

0!2!4!6!8!

10!12!14!16!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Oxalic+H2SO4 Leach :Ore Sample B

Nickel!Cobalt!Iron!

! 63!

approximately 5% for both analytes in sulphuric acid to approximately 40% in the mixed acid

leachate for sample “Ore Sample B”. The increase in leaching percentage is slightly more than

when compared to the equivalent data for citric acid on its own. The results point to a

synergistically positive enhancement in the leachability of the “Ore Sample B” ore over single

acid leaching protocols. When data for the less refractory ore, “Ore Sample A”, are compared,

leach efficiencies for cobalt and manganese are equivalent for all three acid leaches used.

However, there is a significant increase in the percentage leaching of iron in the mixed acid

leach when compared to the citric acid data and a significant reduction from 20% extractability

in the sulfuric acid leach on its own. Nickel extraction is approximately 70% in the sulfuric acid

leach tests and 4% in the mixed acid leach but is enhanced from the approximately 1.5%

extractability in the citric acid leach alone. Again, as with all other leach regimes undertaken in

this thesis, there is no relationship is between the relative leachability of the two ores in the

short term trials and the final long term leach results.

! Figure 3.11. Extraction efficiency for sulfuric acid/citric acid mixed acid leaching of “Ore Sample A” and “Ore Sample B”.

3.1.4.3 Sulphuric and Tartaric Acid Leach

Details of the data produced for the sulfuric acid/citric acid leach experiments are given in

Figure 3.12.

0!

5!

10!

15!

20!

25!

30!

35!

0! 50! 100! 150! 200! 250! 300!

Extrac'o

n*Effi

cien

cy*(%

)*

Time*(h)**

Citric+H2SO4 Leach: Ore Sample A

Nickel!Cobalt!Iron!Manganese!

0!1!2!3!4!5!6!7!8!

0! 100! 200! 300!

Extrac'o

n*effi

cien

cy*(%

)*

Time*(h)*

Citric+H2SO4 Leach: Ore Sample B

Nickel!

Cobalt!

Iron!

Manganese!

! 64!

Using this mixed acid leach combination, the strong binary pair relationship of Mn/Co and

Fe/Ni is easily apparent. This relationship was shown strongly in the original tartaric acid leach

curves (Figure. 3.8) but were less distinct in the sulfuric acid leach curves (Figure. 3.1).

! ! Figure 3.12. Extraction efficiency for sulfuric acid/tartaric acid mixed acid leaching of “Ore Sample A” and “Ore Sample B”. The extraction efficiency of nickel shown in the mixed acid leach curves for sample “Ore

Sample B” reached approximately 3.5% while equivalent extraction efficiency for sulfuric acid

extraction was 3% and for tartaric acid was 1.5%. When data for “Ore Sample A” are

compared the mixed acid leach extracted approximately 9% of the available nickel while the

sulphuric acid leach extracted approximately 8% and the tartaric acid leach 2%. The relative

extraction efficiency of the mixed acid leach for “Ore Sample A” and “Ore Sample B” are

respectively 2.1:1. This is nowhere near the 8.6:1 required of this leach mixture to provide

sensible data for predicting relative long term leach characteristics for the two ores.

3.1.4.4 Summary for the Mixed Acid Leaches

A summary of the data produced during this series of experiments is detailed in Table 3.5 and

Figure 3.13. From this data it is apparent that there is an increase in nickel extraction

efficiency for both “Ore Sample A” and “Ore Sample B” over the 240 hour short term leaching

test for the mixed sulfuric/oxalic acid leaching regime over the single sulfuric acid leaching

0!1!2!3!4!5!6!7!8!

0! 100! 200! 300!

Extrac'o

n*Effi

cien

cy*(%

)*

Time*(h)**

Tartaric+H2SO4 Ore Sample A


0!

10!

20!

30!

40!

50!

0! 100! 200! 300!

Extrac'o

n*Effi

cien

cy*(%

)*

Time*(h)**

Tartaric+H2SO4 Ore Sample B


A! B!

! 65!

protocol. The mixed acid leach procedure, using sulfuric and oxalic acids, is significantly more

efficient

Table 3.5. Summary of the extraction efficiencies of all study analytes from “Ore Sample A” and “Ore Sample B” using mixed acid leaches.

!! Percent Extraction Efficiencies 240h

Ore Samples A B A B A B A B

Leach solution Mn Mn Fe Fe Co Co Ni Ni Oxalic + H2SO4 42! 11! 19.4! 14.2! 42.5! 14.2! 10! 8.0!

Citric + H2SO4 31! 5.1! 10.4! 4.9! 29.9! 6.77! 6.2! 4.0!

Tartaric + H2SO4 42! 5.4! 15.4! 3.27! 41.2! 6.81! 8.8! 3.4!

than any of the other mixed acid leaches studied. Relative increases in extraction efficiency

are also shown for most other analytes in both samples using the sulfuric acid oxalic acid

leach mixture. However, there appears to be no improvement in the extraction efficiency for

nickel in either sample using sulfuric acid / citric acid leach or a sulfuric acid / tartaric acid

leach.

Figure 3.13 Extraction efficiencies for nickel from “Ore Sample A” and “Ore Sample B” using three different mixed acid leach solutions

The synergistic effect of the mixed acid leaching protocols for sulfuric acid / oxalic acid is very

apparent for cobalt with a significant increase in extraction efficiency for this element when

0!

2!

4!

6!

8!

10!

12!

0! 100! 200! 300!

Nickel*Extrac'on

*Efficien

cy*(%

)*

Time*(h)*

Mixed acid Ore Sample A

Tartaric + H2SO4 Citric + H2SO4 Oxalic + H2SO4

0!1!2!3!4!5!6!7!8!9!

0! 100! 200! 300!

Nickel*Extrac'on

*Efficien

cy*(%

)*

Time*(h)*

Mixed acid Ore Sample B

Tartric + H2SO4 Citric + H2S04 Oxalic + H2SO4

! 66!

compared with single acid leaching. However, none of the mixed acid leach protocols provide

data for nickel extraction that gives any indication of the final leach potential of the ores and

consequently none of the mixed acid leaches investigated appear useful as a predictive model

for determining the refractiveness of the ore materials studied to nickel recovery.

3.1.5 Single inorganic acid Leaching following pre-oxidation of ore materials

Experiments associated with leaching the original ore samples failed to provide a set of

conditions that could be used to predict the final nickel leachability potential of the original

ores under long term leaching conditions. The variation in the leaching ability of the ores is

obviously related to the refractivity of the ore which in turn is related to the ability of goethite to

sequester nickel and then subsequently release the available nickel from the ore. This implies

that a variation in the crystal structure and crystallinity of the goethite, which is the primary

“carrier mineral” for the nickel in the ore may be responsible for the refractivity of the ore.

Goethite (!-FeOOH) is an important iron oxide with and exceptional chemical reactivity, in

particular, because of its wide variety of crystal structure it has a variety of lattice sites

available for the incorporation of nickel. Furthermore, under conditions of increased

temperature (Equation A), or by simple dehydration (Equation B) resulting in intermediates,

goethite can be oxidized into Hematite (!-Fe2O3), a process which involves and atomic

rearrangement of Fe+3, and the removal of hydrogen and some of the oxygen from the original

crystal. This process completely changes the mineralogy of the sample.

Equation A 2 α-FeOOH → α-Fe2O3 + H2O

Equation B α-FeOOH (goethite) ! Fe5/3(OH)O2 (protohaematite) !

Fe11/6(oh)1/2O5/2(hydrohaematite) ! !-Fe2O3(hematite)

The effect of heating samples “Ore Sample A” and “Ore Sample B” to 10000C for two hours

can be seen in the diffractograms of the starting and end products (Figure. 3.13). With

reference to Figure 3.14 it is possible to see that the goethite present in the sample has

! 67!

Figure 3.14. X-ray diffractograms of samples “Ore Sample A” (A and B) and “Ore Sample B” (C and D) before and after heating at 1000oC for 2 hours. Green lines correspond to Goethite and purple, to Hematite.

mostly been oxidised to haematite. The effects of various heating temperatures (Appendix A4)

on nickel extractability are presented graphically in Figures 3.15 to 3.18. In general it can be

seen that increasing the temperature to approximately 400–500oC for 2 hours increases the

nickel extraction efficiency of both “Ore Sample A” and “Ore Sample B” (Appendix table 5).

However, once a temperature of approximately 500oC has been reached, the nickel extraction

efficiency rapidly decreases to almost zero at 1000oC. At this point the sample is almost

entirely haematite. The extraction efficiency is first increased because the disordering of the

goethite lattice facilitates leaching. However, once recrystallisation of the disassociated

goethite into haematite starts to occur then the resulting haematite is completely refractory to

acid leach and the nickel becomes unavailable. The experiments reported in this section were

designed to determine the nickel extractability for both ore samples at all the pre-oxidation

temperatures studied and to assess if any one of these conditions provided a sample that,

A! B!

C! D!

! 68!

when leached under a series of per-determined regimes, could be used to predict the actual

long term efficiency of nickel removal.

3.1.5.1 20% Sulfuric Acid Leach of Pre-oxidised Ore Samples

From the data detailed in Figure 3.15, it is apparent that heating sample “Ore Sample B” to a

temperature of 500oC for two hours increases the nickel extraction efficiency from

approximately 3% to 43% over a 240h extraction period.

Figure 3.15. Variation in 20% v/v sulfuric acid based nickel extractability for ore samples “Ore Sample A” and “Ore Sample B” following pre-oxidation of the ores at selected temperatures for a two hour period.

The long term extraction efficiency of this sample is only approximately 8% and consequently

pre-oxidation significantly improves the nickel recovery. Data for “Ore Sample A” indicates that

nickel leachability has also been improved from 7.7% to 58.9% over the same 240 hour

period. However, the long term leachability for nickel from this sample is 68% which has not

been reached over the 240 hour period. The ratio of 8.6:1 for the nickel leachability ratio for

long term leaching of samples “Ore Sample A” and “Ore Sample B” respectively has not been

approximated in any of these experiments. Consequently, pre-oxidation of the samples to

temperatures up to 10000C and subsequent leaching of the ores in 20% v/v sulfuric acid

0

10

20

30

40

50

60

70

0! 100! 200! 300!

Nic

kel E

xtra

ctio

n Ef

ficie

ncy(

%)

Time (h)

Ore Sample A: H2SO4

Normal!

250°C!

400°C!

500°C!

600°C!

750°C!

1000°C!

0 5

10 15 20 25 30 35 40 45 50

0! 100! 200! 300!

Nic

kel E

xtra

ctio

n Ef

ficie

ncy

(%)

Time (h)

Ore Sample B: H2SO4

Normal!

250°C!

400°C!

500°C!

600°C!

750°C!

1000°C!

! 69!

produces no leach scenarios that can be used as a predictive experiment to estimate long

term leachability of any of the ores.

3.1.5.2 20% Hydrochloric Leach of Pre-oxidised Ore Samples

Leaching the pre-oxidised ores using 20% v/v hydrochloric acid produced data (Figure. 3.16)

that are essentially similar to those for the 20% v/v sulfuric acid leach experiments.

Figure 3.16. Variation in 20% v/v hydrochloric acid based nickel extractability for ore samples “Ore Sample B” and “Ore Sample A” following pre-oxidation of the ores at selected temperatures for a two hour

The maximum percentage nickel leachability for “Ore Sample B” in 20% v/v hydrochloric acid

is essentially the same for both the 400oC and 500oC pre-oxidised materials and is recorded

as 35.5% and 37% respectively. This value is slightly lower than the 43% obtained when 20%

v/v sulfuric acid was used. The same generic pattern of leachability is however achieved with

hydrochloric acid in that the maximum nickel leachability from the ore is for material that has

been pre-heated to approximately 400-500oC. There is a very significant reduction in nickel

leachability from ores that have been heated to higher temperatures. The same general trend

can be seen for sample “Ore Sample A” with a maximum nickel leachability of approximately

0!

10!

20!

30!

40!

50!

60!

0! 100! 200! 300!

Nic

kel E

xtra

ctio

n Ef

ficie

ncy

(%)

Time (h)

Ore Sample A: HCl

Normal

250°C

400°C

500°C

600°C

750°C

1000°C

0!5!

10!15!20!25!30!35!40!45!

0! 100! 200! 300!

Nic

kel E

xtra

ctio

n Ef

ficie

ncy

(%)

Time (h)

Ore Sample B: HCl

Normal

250°C

400°C

500°C

600°C

750°C

1000°C

! 70!

50% for a sample that has been pre-heated to a temperature of 500oC. However, no data from

any of the experiments approximates values for the relative long term nickel extraction

efficiency relationship between “Ore Sample A” and “Ore Sample B” of 8.6:1 and

consequently none of the hydrochloric acid leach experiments can provide data that could be

used to predict long term leach efficiencies of the study ores.

3.1.5.3 20% Aqua Regia Acid Leach of Pre-oxidised Ore Samples

The nickel extraction curves for samples “Ore Sample A” and “Ore Sample B”, using Aqua

Regia as a leach agent, are almost idnetivcal to those for the same samples using

hydrochloric and sulfuric acid (Figure. 3.17).

Figure 3.17 Variation in 20% v/v Aqua Regia acid based nickel extractability for ore samples “Ore Sample A” and “Ore Sample B” following pre-oxidation of the ores at selected temperatures for a two hour period.

There is an increase in the nickel extractability when compared to the original single acid

extraction data, by using ores that have been pre-heated. The increase is from 3.5% to 43%

for “Ore Sample B” and 8% to 53% for “Ore Sample A”. However, as with the previous

experiments none of the experiments produce results that give relative nickel extraction ratios

for “Ore Sample A”:” Ore Sample B” that approach the 8.6:1 required. Consequently none of

0!

10!

20!

30!

40!

50!

60!

0! 100! 200! 300!

Nic

kel E

xtra

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n Ef

ficie

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(%)

Time (h)

Aqua Regia Ore:Sample A

Normal

250°C

400°C

500°C

600°C

750°C

1000°C

0!5!

10!15!20!25!30!35!40!45!50!

0! 100! 200! 300!

Nic

kel E

xtra

ctio

n Ef

ficie

ncy

(%)

Time (h)

Aqua Regia: Ore Sample B

Normal

250°C

400°C

500°C

600°C

750°C

1000°C

! 71!

the hydrochloric acid leach experiments appear to be able to provide data that could be used

to predict long term leach efficiencies of the study ores.

3.1.5.4 20% Nitric Acid Leach of Pre-oxidised Ore Samples

While the leach characteristics of “Ore Sample A” are essentially similar for the other three

acid leaches, data for the 20% v/v nitric acid extraction of nickel from pre-oxidised “Ore

Sample B” ore is significantly different from that of these acids Figure 3.18. This is probably

because the ability of nitric acid to leach oxides of iron is extremely poor and even in the initial

experiments with unheated ore, the ability of nitric acid leach solutions to remove nickel from

the ores was extremely poor and worse than any of the other acids used. Nevertheless, the

extraction efficiency for nickel from the pre-heated ores is significantly greater than that from

the unheated ore, increasing from 1.9 to 21.9% for “Ore Sample A”, and 1.5 to 14.5% for “Ore

Sample B”.

!

Figure 3.18 Variation in 20% v/v nitric acid based nickel extractability for ore samples “Ore Sample A” and “Ore Sample B” following pre-oxidation of the ores at selected temperatures for a two hour period.

There is no discernible relationship between the extraction efficiency of nickel from either of

these pre-heated ores, when using nitric acid as a leaching agent. This is somewhat difficult to

0!

5!

10!

15!

20!

25!

0! 100! 200! 300!

Nic

kel E

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n Ef

ficie

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(%)

Time (h)

Nitric: Ore Sample A

Normal

250°C

400°C

500°C

600°C

750°C

1000°C 0

2

4

6

8

10

12

14

16

0! 200!

Nic

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(%)

Time (h)

Nitric: Ore Sample B

Normal

250°C

400°C

500°C

600°C

750°C

1000°C

! 72!

explain but is probably associated with the inability of nitric acid to dissolve iron oxide based

ores and the variation of crystallinity and mineral composition in which the nickel and iron are

associated in the two different extractability ore samples used in this study. Again, none of the

experimental data appear to be useful in providing a predictive model for long term leachability

estimation in either “Ore Sample A” or “Ore Sample B”.

3.1.5.5 Comparative Leachability for all Study Analytes from “Ore Sample A” and “Ore

Sample B” Pre-oxidised Ore

Data for the leachability of all study analytes under the four study inorganic acid extraction

regimes is given in Table 3.6. The data in this table describe a very wide range of

extractabilities for the study analytes and points to the wide range of mineralogical

transformations that have taken place during the various pre-oxidation stages that the sample

have been subjected to. However, these changes in mineralogy are outside the scope of this

thesis and are therefore not dealt with here.

In general, it can be concluded that up to temperatures of approximately 500oC there is a

steady increase in the percentage of analytes that can be extracted. Above this temperature,

and probably because of recrystallisation of the oxide species of iron, from goethite to

haematite, and their simultaneous incorporation of nickel into an extremely refractory matrix,

extraction efficiencies drop to nearly zero. In spite of being able to leach pre-oxidised ore over

a number of temperatures and using a variety of acid attacks, none of the experiments have

yielded any data that facilitates using short term leach experiments as predictive models for

their determination of the ultimate long term leachability of the ores.

From the data obtained it has been able to be concluded that it would be possible to devise a

better leach solution than the sulfuric acid currently being used as the industry standard.

! 73!

Table 3.6. Leach data for all study analytes from samples “Ore Sample A” and “Ore Sample B” following pre-oxidation at selected temperatures for two hours.

!! !! !! Percent Extraction Efficiencies 240h !! !!Ore Samples A B A B A B A B

Analytes Mn Mn Fe Fe Co Co Ni Ni

H2SO4 Leach solution !! !! !! !! !! !! !! !! Original sample 24.2 4.49 21.4 5.59 21.2 5.36 7.69 3.21

250oC 45 6.12 32.4 5.94 37.6 8.29 15.7 4.51

400oC 45.1 33.5 101 97.2 33.6 40.2 49.1 24.9

500oC 97.8 58.3 102 92.6 76.7 73 58.8 43

600oC 30.4 21.2 84.3 40.2 25.7 30.2 15.9 15.9

750oC 30.7 20.4 11.6 8.52 38.2 42.4 11.2 15.4

1000oC 2.44 1.24 0.6 0.17 2.01 0.88 0.47 0.22

!! !! !! !! !! !! !! !! !!HCl Leach solution !! !! !! !! !! !! !! !!Original sample 24.6 5.56 11.7 3.02 20.8 6.49 6.57 2.56

250oC 33.5 4.73 20.1 4.09 24.7 6.16 9.46 2.99

400oC 56.5 48.7 97.2 85.6 40.3 58.7 42.5 37

500oC 79.2 49.7 96 67.2 66.2 65.6 50.2 35.5

600oC 55.4 24.4 64.5 23.5 49.6 39.5 27.2 20.1

750oC 24.6 18 7.69 6.34 32.9 37.8 9.28 13.5

1000oC 1.74 1.43 0.34 0.17 1.37 0.93 0.27 0.29

!! !! !! !! !! !! !! !! !!Aqua Regia Leach solution !! !! !! !! !! !! !! !!Original sample 31.5 6.65 14.2 3.48 27 7.82 7.93 3.46

250oC 28.4 4.79 16.7 3.72 19.1 6.15 7.89 2.52

400oC 77.5 57.5 98.3 84.8 57.6 68.6 52.7 42.9

500oC 81.7 51.8 96.4 77.5 67.4 73.1 52.8 42.2

600oC 61.5 31.1 66 32.2 54 46.7 31.4 24.9

750oC 35.8 20.3 9.52 6.86 47.1 42.2 9.2 14.7

1000oC 1.88 1.43 0.35 0.15 1.48 0.93 0.28 0.36

!! !! !! !! !! !! !! !! !!HNO3 Leach solution !! !! !! !! !! !! !! !!Original sample 7.13 2.26 3.28 2.07 4.47 3.02 1.9 1.48

250oC 15.1 3.33 6.41 2.3 9.34 4.31 4.71 2.17

400oC 28.6 6.78 48.6 18 19.7 12.3 18.6 8.91

500oC 40.6 13 38.4 15.8 48.7 28.1 21.9 14.5

600oC 19.6 9.95 11.3 9.13 29.1 23 8.81 13

750oC 27 17.5 4.22 4.05 42.7 43.6 8.2 13.8

1000oC 1.51 0.93 0.21 0.1 1.36 0.67 0.24 0.29

! 74!

Consequently, a final set of experiments was devised using study ores that had been held in

reserve to test any other appropriate leaching solutions investigated throughout this study, as

“calibration standards” and on the basis of their short term leachability try to develop a set of

standards that in the short term will possibly mimic the long term leaching behaviour of a

variety of partially leachable ores. This experiment was undertaken using 20% v/v sulfuric

acid.

3.1.6 Estimation of long term leachability of nickel based on short term calibration

against “standard samples”

In addition to “Ore Sample A” and “Ore Sample B”, provided by CSIRO Division of Process

Science and Engineering for the extraction trials detailed in this thesis, a series of other

samples of known long term nickel extractability were also provided (Appendix Table 6).

Unfortunately, only a small mass of material was avalible for these samples, and it was not

possible to undertake repeat trials. Nevertheless, six additional samples, “Ore Samples C

– H”, were provided with sufficient mass (approximately 50-100g) to be used in an

additional extraction experiment (Table 3.7). The concept behind this experiment was that

if the leach characteristics of ores with known nickel extractability could be determined

over a period of up to 240 hours, it may be possible to use these characteristics to plot a

“calibration curve” of extractability over the long term against extractability over the short

term and use this calibration curve to calculate the extraction efficiency of unknown ores

over the long term.

In this way it was envisioned that providing a series of various known long term

extractability standards to companies dealing with refractory ores would allow them to run

these standards and unknown samples together in short term leach experiments and

thereby estimate the long term extractability of the unknown ores by reference to a

calibration curve produced using the provided standards. Both ICP-AES and ICP-MS

! 75!

instrumentation was used to produce data for this study. Results are detailed in Figures

3.19 to 3.22 with the final composite data detailed in Figure 3.23.

Table 3.7. Long Term Nickel Extractability of Analytes from the “Standard Samples”

!! Long!term!field!trials! Analytes! Mn Fe Co Ni

Ore!Sample!Number! Ni!extraction!percent! !! ppm! %! ppm! ppm!C 50! 7650 51 815 10500

D 25! 776 42 140 7180

E 40! 10400 51 834 6600

F 40! 5620 51 757 7400

G 70! 4390 33 1270 8900

H 4! 472 34 163 2880

Data detailed in Figure 3.10 and 3.20 represent the production of a calibration curve based

on samples “Ore Sample C” (50% nickel extractability), “Ore Sample D” (25% nickel

extractability), “Ore Sample F” (40% nickel extractability), “Ore Sample G” (70% nickel

extractability) and “Ore Sample H” (4% nickel extractability). The reason for using both

ICP-AES and ICP-MS is that many laboratories will have either one or the other technique

available but possibly not both and the developed technique should be useable in either

type of laboratory. The concentration of the analytes is significantly above the detection

limits for both techniques so either technique can be used to provide appropriate data for

this experiment. The analytical solutions used in this study were made up to be identical for

both the ICP-AES and ICP-MS analyses. This approach, while not being used throughout

the thesis for the analysis of samples as more solution was available for these studies, was

undertaken here to be certain that reanalysis of solutions was possible with the limited

amount of leach solution available. While this represented a compromise it was the only

appropriate solution for the experimental design used here. The main problem with his type

of analysis is that the iron matrix gives rise to variable bias in both techniques. This is

because the calibration curves produced are derived from the analysis of aqueous (slightly

acidic) standards with low concentrations of all analytes. Consequently, analysis of

! 76!

relatively high matrix salt content samples will produce physical interference and give rise

to a generic difference, of a few percent relative, between data sets for the two different

instruments. While this is a problem with respect to absolute accuracy of the data, it is not

a problem with respect to comparison of the data from a single technique. The

development of an analytical technique to overcome this bias, and produce absolute

accuracy of data from both techniques, can be an extremely long and costly process and is

beyond the scope of the investigation covered by work in this thesis. As can be seen by

reference to data the effects of bias on the results are relative and not absolute and

consequently all data can be accurately compared within the data sets for each analytical

technique to provide valid conclusions.

Data obtained from the 20% v/v sulfuric acid leaching of the six samples of ore that were to

be used as calibration standards is shown in Figures 3.19 and 3.20 for the analysis by ICP-

AES and ICP-MS respectively. From these figures it is obvious that there is a comparability

of the data for short term leachability of these ore with that for the long term leachability

experiments. The close correlation of the two sets of data for both instrumental techniques

becomes more apparent with increased periods of leaching. The fit of data to a regression

line becomes markedly better the longer the ore samples are exposed to the leach

solutions. After a leaching period approaching 144 hours the R2 value for the regression

slope of a line drawn to represent the relationship between the two sets of data (regression

line), is approximately 0.96 or above implying that the regression line represents some

96% of the variation of the data and that consequently there is an excellent relationship

between the two parameters being plotted, the short and long term extraction efficiencies

for nickel from the ores.

When the data for “Ore Sample E” (40% nickel extractability) is added to the graphs as an

additional yellow point (Figures 3.21 and 3.22), it can easily be seen that in both the ICP-

AES and ICP-MS sets of data the closeness of fit of this new data to the calibration curves

produced from the “standard ores’ is excellent for all times above 144 hours leaching. It is

! 77!

certainly probable that the time between 72 hours leaching and 144 hours leaching, which

represented a weekend during which time analysis was not possible due to an instrument

failure, that equivalent comparability could easily have been achieved. However, due to

circumstances, which could not be controlled, this suggestion will have to be confirmed in

a further set of experiments at a late stage. Nevertheless, the use of “standard ores” for the

construction of a “calibration graph” as described here obviously provides a very strong

indication of the potential of this concept to provide long term accurate extractability

figures, for refractory ores, over a very short time span. It will obviously be necessary to

repeat and expand this study using a greater variety of ore types with different mineralogy

and a variety of different nickel extractabilities to ensure that the technique is robust. In

addition, it will be necessary to develop a final analytical technique in which there is no

bias and this will probably ultimately involve the use of the ICP-MS technique where it will

be possible to dilute the sample to such an extent that there is relatively no interference

from iron.

The relative relationship of all the “calibration ores” together with the “test ore” is detailed in

Figure 3.23. Here it can be seen that throughout the experiment there is a close

relationship between the two 40% nickel extractable ores (“Ore Sample F” and “Ore

Sample E”), and that the “calibration slope” is linear.

One obvious problem is that there is a gap in the data between 72 hours and 144 hours,

representing a Saturday and Sunday over which period it was not possible to analyse samples

because of instrumental problems. Nevertheless, there is an extremely close relationship

between data for samples “Ore Sample E” and “Ore Sample F” from approximately 72 hours

of extraction onwards.

! 78!

y!=!0.0221x!+!0.8111!R²!=!0.79226!

0!

0.5!

1!

1.5!

2!

2.5!

3!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*

Long*Term*Ni*Extrac'on*(%)*

AES!data!24h!Leach!

y!=!0.0342x!+!0.9262!R²!=!0.80525!

0!

0.5!

1!

1.5!

2!

2.5!

3!

3.5!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


AES!data!48h!Leach!

y!=!0.0399x!+!1.069!R²!=!0.86768!

0!0.5!1!

1.5!2!

2.5!3!

3.5!4!

4.5!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


AES!data!72h!Leach!

y!=!0.1545x!+!0.3696!R²!=!0.96027!

0!

2!

4!

6!

8!

10!

12!

14!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


AES!data!144h!Leach!

y!=!0.1726x!+!0.3566!R²!=!0.96299!

0!

2!

4!

6!

8!

10!

12!

14!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*



y!=!0.1975x!+!0.2293!R²!=!0.96276!

0!2!4!6!8!

10!12!14!16!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*



y!=!0.2206x!+!0.1255!R²!=!0.96791!

0!2!4!6!8!10!12!14!16!18!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*



y!=!0.2502x!+!0.0542!R²!=!0.97177!

0!

5!

10!

15!

20!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*



Figure 3.19 Calibration leach curves for “Ore Samples C, D, F, G and H” in 20% sulfuric acid for selected time periods – ICP-AES data.

! 79!

y!=!0.0359x!+!1.2314!R²!=!0.72263!0!

0.5!1!

1.5!2!

2.5!3!

3.5!4!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!24h!Leach!

y!=!0.0454x!+!1.2549!R²!=!0.70888!

0!

1!

2!

3!

4!

5!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!48h!Leach!

y!=!0.0568x!+!1.1197!R²!=!0.87226!

0!

1!

2!

3!

4!

5!

6!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!72h!Leach!

y!=!0.2027x!+!0.8109!R²!=!0.95797!

0!2!4!6!8!10!12!14!16!18!

0! 20! 40! 60! 80!

Expe

riman

tal*N

i*Extrac'on

*(%)*


MS!data!144h!Leach!

y!=!0.2165x!+!1.4866!R²!=!0.93552!0!

5!

10!

15!

20!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!168h!Leach!

y!=!0.263x!+!0.4718!R²!=!0.96597!

0!

5!

10!

15!

20!

25!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!192h!Leach!

y!=!0.2959x!+!0.268!R²!=!0.96662!

0!

5!

10!

15!

20!

25!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!216h!Leach!

y!=!0.3303x!+!0.088!R²!=!0.96631!

0!

5!

10!

15!

20!

25!

30!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!240h!Leach!

Figure 3.20 Calibration leach curves for “Ore Samples C, D, F, G and H” in 20% sulfuric acid for selected time periods – ICP-MS data.

!

! 80!

y!=!0.0228x!+!0.9501!R²!=!0.5298!

0!

0.5!

1!

1.5!

2!

2.5!

3!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


AES!data!24h!Leach!

y!=!0.0344x!+!0.9774!R²!=!0.78375!

0!0.5!1!

1.5!2!

2.5!3!

3.5!4!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


AES!data!48h!Leach!

y!=!0.0402x!+!1.1181!R²!=!0.85026!

0!0.5!1!

1.5!2!

2.5!3!

3.5!4!

4.5!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


AES!data!72h!Leach!

y!=!0.1545x!+!0.3855!R²!=!0.96021!

0!

2!

4!

6!

8!

10!

12!

14!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*



y!=!0.1726x!+!0.3458!R²!=!0.96296!

0!

2!

4!

6!

8!

10!

12!

14!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*



y!=!0.1972x!+!0.1734!R²!=!0.96142!

0!2!4!6!8!

10!12!14!16!

0! 20! 40! 60! 80!

Expe

rimetna

l*Ni*Extrac'on

*(%)*



y!=!0.2202x!+!0.0576!R²!=!0.96632!

0!2!4!6!8!10!12!14!16!18!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*



y!=!0.2497x!W!0.0513!R²!=!0.96878!

0!

5!

10!

15!

20!

0! 20! 40! 60! 80!

Expe

rimetna

l*Ni*Extrac'on

*(%)*



Figure 3.21 Calibration leach curves for “Ore Samples C, D, F, G and H” and unknown “Ore sample E”, in 20% sulfuric acid for selected time periods – ICP-AES data.

!

! 81!

y!=!0.0365x!+!1.3445!R²!=!0.65091!

0!0.5!1!

1.5!2!

2.5!3!

3.5!4!

4.5!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!24h!Leach!

y!=!0.0461x!+!1.3832!R²!=!0.65277!

0!

1!

2!

3!

4!

5!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!48h!Leach!

y!=!0.0573x!+!1.2089!R²!=!0.8433!

0!

1!

2!

3!

4!

5!

6!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!72h!Leach!

y!=!0.2033x!+!0.9249!R²!=!0.95331!

0!2!4!6!8!10!12!14!16!18!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!144h!Leach!

y!=!0.2164x!+!1.4825!R²!=!0.9356!

0!

5!

10!

15!

20!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MSdata!168h!Leach!

y!=!0.2631x!+!0.4866!R²!=!0.96599!

0!

5!

10!

15!

20!

25!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MSdata!192h!Leach!

y!=!0.2956x!+!0.2168!R²!=!0.96612!

0!

5!

10!

15!

20!

25!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!216h!Leach!

y!=!0.3301x!+!0.0525!R²!=!0.96614!

0!

5!

10!

15!

20!

25!

30!

0! 20! 40! 60! 80!

Expe

rimen

tal*N

i*Extrac'on

*(%)*


MS!data!240h!Leach!

Figure 3.22 Calibration leach curves for “Ore Samples C, D, F, G and H” and unknown “Ore sample E”, in 20% sulfuric acid for selected time periods – ICP-MS data.

!

! 82!

Figure 3.23. Calibration curves produced for the relative relationship between the extractability of nickel using both ICP-AES and ICP-MS and describing the relationship between long and short term extractability of this metal from various refractory ores

Consequently, it is valid to suggest that starting a leach regime on a Monday morning would

provide data by Friday afternoon that could be used to give an accurate estimation of the long

term extractability of refractory ores. This claim obviously needs to be validated by further

study. However, the study would seem worthwhile especially as there is a solid indication that

within a period of five days it would be possible to grade the refractivity of ores for which it has

previously been necessary to take a period of up to six months to provide equivalent data.

0!2!4!6!8!10!12!14!16!18!20!

0! 100! 200! 300!

Nic

kel E

xtra

ctio

n (%

)

Time (h)

ICP-AES data C-50% D-25% E-40% F-40% G-70% H-4%

0!

5!

10!

15!

20!

25!

30!

0! 100! 200! 300!

Nic

kel E

xtra

ctio

n (%

)

Time (h)

ICP-MS data C-50% D-25% E-40% F-40% G-70% H-4%

! 83!

CHAPTER 4

CONCLUSION & FUTURE WORK

The research undertaken in this thesis was designed to try and establish a leaching protocol

that would be able to estimate the long term extractability of nickel from refractory and non-

refractory ores. A range of lateritic (oxide) ores, having nickel extractability ranging between

4% and 70%, were tested under a variety of single and mixed acid leaching conditions over

leaching periods of 240 hours. The relative nickel extractability is largely the result of

variability in the refractiveness of the goethite, an iron hydroxy-oxide mineral in the ore, which

primarily contains the nickel. Research was divided into two parts. In phase one, test samples

A and B (8% and 66% nickel extractability respectively) were tested, while in phase two, six

ores with nickel extractabilities between 4% and 70% were used. The phase one tests involve

exposing the two ore samples to a variety of inorganic and organic acids, and to combinations

of sulphuric acid and organic acids, to try and develop relative leach curves that would give an

approximate 1:8 extraction relationship between the two ores over a leaching period of 240h

or less. The ratio factor has been calculated as that which the final six month long term leach

tests indicate to be the relative nickel leachabilities of the two ores.

Results indicate that 20% v/v Aqua Regia and 20% v/v H2SO4, have the highest nickel

extraction efficiency, followed by 20% v/v HCl and 20% v/v HNO3. Even though there is a

significant difference in all cases between the extraction efficiency of the two samples

(refractory and non-refractory) the relative nickel extraction efficiency of the two ores never

approaches the 1:8 ratio required for these leaching conditions to indicate the long term

relative extraction efficiency of the two ores. Consequently, none of the inorganic based acid

leach regimes investigated appear useful as predictors to the long term extraction behaviour

of the ores.

! 84!

The second part of phase one research involved the use of organic acids (tartaric, citric and

oxalic) which can form complexes with those ions and cations already in solution and

therefore increase the ultimate extractability of the nickel. Again, none of the organic acids

leaching regimes used produced the 1:8 ratio required to indicate long term nickel

extractability of the two ores initially investigated in this study. However, one important aspect

that transpired from these experiments was that the selective leaching of manganese from

both these ores using citric and tartaric acids could possibly be used as a precursor to any

long term nickel leaching protocols to remove cobalt at an early stage of mine site extraction

initiatives as approximately 25% of the available cobalt is removed with only 1% of the iron

and nickel. As the relative costs of these two metals are significantly different this could be a

profitable pre-leach procedure where cobalt concentrations in the ore are relatively high.

In the third stage of phase one, ores were leached using a mixture of 10% H2SO4 and 10% v/v

of the three organic acids (citric, tartaric and oxalic). Again, none of these acid combinations

resulted in comparison ratios of 1:8 required for these mixtures to be used as short term

indicators of the ultimate leaching potential of the two ores.

In order to study the effect of heat on nickel extractability, samples were calcined at various

temperatures up to 1000°C for a period of two hours, and then leached using a 20 % v/v

sulphuric, Aqua Regia, nitric and hydrochloric acid solutions. The changes in mineralogy,

associated with heating the samples at these temperatures, were investigated using XRD.

The nickel extraction efficiency of both ores increased significantly, up to calcining

temperatures of 500oC, after which temperature there is a rapid decline in nickel extraction

efficiency of the ores to almost zero. While the lower calcining temperatures directly affect the

crystal structure of the goethite, causing lattice rearrangement and consequent increase in the

ease with which nickel can be leached, the higher temperatures cause any goethite present to

be changed, by removal of the OH- radical, into haematite which effectively binds the nickel

into a refractory crystal and almost completely eliminates extractability of this metal. In all of

! 85!

these experiments, none of the extraction protocols investigated produced the required

relative ratio of nickel extractability of 1:8.

At this stage of the research it can be concluded that it is not possible to develop a short term

leaching protocol that will give an indication of the relative extraction efficiency of the two ores

and consequently the ability to use this concept for any other ore must also be considered

improbable. However, CSIRO Division of Process Science and Engineering provided six

additional samples that had enough mass to be used in a further investigation, phase 2.

In phase 2, five laterite ore samples were used to generate nickel extraction curves under

20% v/v sulfuric acid leaching conditions over a 240h period. The data for these curves were

plotted against equivalent long term leach test data provided by CSIRO Division of Process

Science and Engineering to give graphs describing short term vs. long term nickel

extractability for the five ores (Comparable Extraction Graphs). These graphs indicate that

there is a linear relationship between the long and short term nickel extraction efficiencies of

the five ores and that after a period of a maximum of 144 hours the R2 values for the

regression lines have a value of approximately 0.96. This means that the regression line

represents at least 98% of the distribution of the points from which it has been constructed

and the resulting plot can therefore can be used as an exceptionally good “calibration” graph

for determining the nickel extractability of refractory and non-refractory nickel ores. When this

theory was tested using the remaining ore (40% nickel extractability) the results completely

confirmed this opinion (using all graphs representing extraction times between 75 and 240h)

giving nickel extraction efficiencies that varied by no more than +/-5% relative from the correct

value. The use of “Comparable Extraction Graphs” has the potential to save companies

significants amounts of money and time by being able to identify not only refractory and non-

refractory lateritic ores but also to quantify the relative long term (6 month) extractability of

these ores. However, before the idea can be adopted it is essential that more ores, with

markedly different mineralogy, be tested to confirm the theory and that, if correct, bulk ores be

provided to enable a relative nickel extraction efficiency series of standards to be prepared

! 86!

that can be circulated throughout the industry, to facilitate the adoption of this approach as

the industry standard.

FUTURE WORK

1) Undertake extraction experiments under different chemical (different acids and

different concentration) and physical (temperature, rolling speed and grain size)

conditions.

2) Use of a greater number of ores, of more diverse mineralogy, to test the ruggedness of

the Comparable Extraction Graph theory.

3) Reduce the extraction time required to produce useable short term extractability

graphs that indicate the long term extractability of ores.

4) Compare results to other experiments that involves different minerals and the

extraction of other analytes to try to find out similarities.

5) Acquire bulk samples of a wide range of refractory ores that can be used to produce a

standard series that can be used on-site to construct Comparable Extraction Graphs

for the determination of long term nickel extractability.

!

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!

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!

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!

! 87!

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APPENDIX

Table 1 XRD Instrumentation parameters.

Scan!Axis! Gonio!

Start!Position![°2Th.]! 10.0033!

End!Position![°2Th.]! 79.9893!

Step!Size![°2Th.]! 0.0070!

Scan!Step!Time![s]! 8.6700!

Scan!Type! Continuous!

PSD!Mode! Scanning!

PSD!Length![°2Th.]! 3.35!

Offset![°2Th.]! 0.0000!

Divergence!Slit!Type! Fixed!

Divergence!Slit!Size![°]! 0.1250!

Specimen!Length![mm]! 10.00!

Measurement!Temperature![°C]! 25.00!

Anode!Material! Cu!

KZAlpha1![Å]! 1.54060!

KZAlpha2![Å]! 1.54443!

KZBeta![Å]! 1.39225!

KZA2!/!KZA1!Ratio! 0.50000!

Generator!Settings! 40!mA,!40!kV!

Diffractometer!Type! 0000000011092810!

Diffractometer!Number! 0!

Goniometer!Radius![mm]! 240.00!

Dist.!FocusZDiverg.!Slit![mm]! 100.00!

Incident!Beam!Monochromator! No!

Spinning! No!

! 92!

Table 2 Inorganic Leach.

Type of Acid for Sample A

% of Extraction of Manganese

% of Extraction of Iron

% of Extraction of Cobalt

% of Extraction of Nickel

H2SO4 24.2 21.4 21.2 7.7 HCl 24.6 11.7 20.8 6.6 Aqua Regia 31.5 14.2 27 7.9 HNO3 7.1 3.3 4.5 1.9 Type of Acid for Sample B





H2SO4 4.5 5.6 5.4 3.2 HCl 5.6 3 6.5 2.6 Aqua Regia 6.6 3.5 7.8 3.5 HNO3 2.3 2.1 3 1.5

Table 3. Organic Leach.

Type of Acid for Sample A





Oxalic 0.2M 12.5 19 15.4 9.14 Oxalic 0.5M 13.4 19.8 17.4 9.69 Oxalic 1M 15.8 22.4 19.2 11.33 Citric 0.2M 13.9 0.45 14.9 0.97 Citric 0.5M 15.4 0.53 16.6 1.1 Citric 1M 16.6 0.57 17.9 1.23 Tartaric 0.2M 27.6 1.13 23.2 1.99 Tartaric 0.5M 27.8 1.42 25 1.94 Tartaric 1M 32.3 1.77 26.2 2.6 Type of Acid for Sample B





Oxalic 0.2M 38 34 4.4 2.14 Oxalic 0.5M 27.4 25.8 2.4 3.69 Oxalic 1M 29.8 32.4 2.2 3.33 Citric 0.2M 2.9 0.45 3.3 0.77 Citric 0.5M 3.2 0.56 3.6 0.9 Citric 1M 3.1 0.7 3.6 1.03 Tartaric 0.2M 3.2 0.6 3.5 0.9 Tartaric 0.5M 4.4 1.42 4.4 1.6 Tartaric 1M 4.3 1.2 4.2 1.3

! 93!

Table 4. Mixture Leach.

Type of Acid for SAMPLE A




Oxalic + H2SO4 41.5 19.4 42.5 Citric + H2SO4 31.3 10.4 29.9 Tartaric + H2SO4 42.3 15.4 41.2 Type of Acid for SAMPLE B




Oxalic + H2SO4 10.8 14.2 14.2 Citric + H2SO4 5.12 4.9 6.77 Tartaric + H2SO4 5.35 3.27 6.81

Table 5. Oxidation Leach.

!! % Mn % Fe % Co %Ni !! % Mn % Fe % Co %Ni

Extraction Extraction Extraction Extraction Extraction Extraction Extraction Extraction Aqua Regia

Aqua Regia

Sample A

Sample B

250 ºC 28.4 16.7 19.1 7.89 250 ºC 4.79 3.72 6.15 2.52

400 ºC 77.5 98.3 57.6 52.7 400 ºC 57.5 84.8 68.6 42.9

500 ºC 81.7 96.4 67.4 52.8 500 ºC 51.8 77.5 73.1 42.2

600 ºC 61.5 66 54 31.4 600 ºC 31.1 32.2 46.7 24.9

750 ºC 35.8 9.52 47.1 9.2 750 ºC 20.3 6.86 42.2 14.7

1000 ºC 1.88 0.35 1.48 0.28 1000 ºC 1.43 0.15 0.93 0.36

HCl 24.6 11.7 20.8 6.57 HCl 5.56 3.02 6.49 2.56

Sample A Sample B

250 ºC 33.5 20.1 24.7 9.46 250 ºC 4.73 4.09 6.16 2.99

400 ºC 56.5 97.2 40.3 42.5 400 ºC 48.7 85.6 58.7 37

500 ºC 79.2 96 66.2 50.2 500 ºC 49.7 67.2 65.6 35.5

600 ºC 55.4 64.5 49.6 27.2 600 ºC 24.4 23.5 39.5 20.1

750 ºC 24.6 7.69 32.9 9.28 750 ºC 18 6.34 37.8 13.5

1000 ºC 1.74 0.34 1.37 0.27 1000 ºC 1.24 0.17 0.93 0.29

H2SO4 24.2 21.4 21.2 7.69 H2SO4 4.49 5.59 5.36 3.21

SAMPLE A SAMPLE B

250 ºC 45 32.4 37.6 15.7 250 ºC 6.12 5.94 8.29 4.51

400 ºC 45.1 101 33.6 49.1 400 ºC 33.52 97.15 40.21 24.89

500 ºC 97.8 102 76.7 58.8 500 ºC 58.27 92.62 73.03 43

600 ºC 30.4 84.3 25.7 15.9 600 ºC 21.17 40.24 30.21 15.9

750 ºC 30.7 11.6 38.2 11.2 750 ºC 20.36 8.52 42.44 15.4

1000 ºC 2.44 0.6 2.01 0.47 1000 ºC 1.24 0.17 0.88 0.22

HNO3 7.13 3.28 4.47 1.9 HNO3 2.26 2.07 3.02 1.48

SAMPLE A SAMPLE B

250 ºC 15.1 6.41 9.34 4.71 250 ºC 3.33 2.3 4.31 2.17

400 ºC 28.6 48.6 19.71 18.6 400 ºC 6.78 18 12.3 8.91

500 ºC 40.6 38.4 48.7 21.9 500 ºC 13 15.8 28.1 14.5

600 ºC 19.6 11.3 29.1 8.81 600 ºC 9.95 9.13 23 13

750 ºC 27 4.22 42.7 8.2 750 ºC 17.5 4.05 43.6 13.8

1000 ºC 1.51 0.21 1.36 0.24 1000 ºC 0.93 0.1 0.67 0.29

! 94!

Table!6.!Long Term Nickel Extractability of Analytes from the “Standard Samples

Long term field trials Analytes Mn Fe Co Ni

Ore Sample Number

Ni extraction percent ppm % ppm ppm

C 50 7650 51 815 10500

D 25 776 42 140 7180

E 40 10400 51 834 6600

F 40 5620 51 757 7400

G 70 4390 33 1270 8900

H 4 472 34 163 2880

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