Beneficiation of Hard Rock Lithium Bearing Minerals by ... · Bearing Minerals by Flotation By...

Beneficiation of Hard Rock Lithium Bearing Minerals by Flotation

By Daniel Dass

An Honours research thesis submitted to Murdoch University in fulfilment of the requirements for the degree of

Bachelor of Engineering (Honours) in Chemical and Metallurgical Engineering

Department of Chemical and Metallurgical Engineering

Murdoch University

December 2017

Supervisor: Dr Aleks Nikoloski

Author’s Declaration

I declare that this thesis is my own work and no part has been submitted for a degree at any

tertiary education institution.

(Daniel Dass)

IV

Abstract

In an effort to look for more efficient energy resources, we now see a re-emergence of lithium

as a viable energy alternative. Lithium has become a major contributor in the energy storage

industry, providing longer lasting, smaller batteries that are more powerful and more reliable

for consumers. This research project adds to the existing knowledge in the exploration of new

ways to extract lithium. Parameters such as collector dosage and type, activator usage, feed

size and pH level have been reported to have dramatically influenced the flotation recoveries

of lithium. These parameters are investigated in this project.

The aim of this research project was to improve the recovery of hard rock lithium bearing

minerals by flotation. This aim was achieved through a series of objectives that targeted

collector type and dosage, feed size, activator usage and pH. The project was separated into a

series of stages, leading to the eventual flotation of spodumene.

It was found that both feed size and pH level had significant impacts on the recovery of

spodumene. When compared to baseline tests, an approximate increase in recovery by 30%

was achieved. It was found that when both spodumene and lepidolite were floated at the

same conditions, both minerals were able to attain a similar grade and recovery.

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Table of Contents

Author’s Declaration .......................................................................................................................... V

Abstract ............................................................................................................................................ IV

List of Figures & Tables ....................................................................................................................... X

Abbreviations List............................................................................................................................ XIV

Acknowledgements ......................................................................................................................... XVI

1. Introduction ............................................................................................................................... 2

1.1. Overview..................................................................................................................... 2

1.2. Importance of Lithium ................................................................................................. 2

1.3. Global Dependence ..................................................................................................... 4

1.4. Project Aims & Objectives ........................................................................................... 5

2. Literature Review ....................................................................................................................... 6

2.1. Mineralogy .................................................................................................................. 6

2.1.1. Spodumene .................................................................................................... 6

2.1.2. Lepidolite ....................................................................................................... 7

2.1.3. Amblygonite ................................................................................................... 8

2.1.4. Petalite ........................................................................................................... 9

2.1.5. Zinvaldite .......................................................................................................10

2.2. Beneficiation of Lithium Ore .......................................................................................10

2.3. Flotation ....................................................................................................................11

2.3.1. Process ..........................................................................................................11

2.3.2. Reagents........................................................................................................12

2.3.2.1. Collector ...................................................................................12

2.3.2.2. Frother ......................................................................................12

2.3.2.3. Regulator ..................................................................................13

2.3.3. Spodumene Flotation ....................................................................................13

2.3.4. Lepidolite Flotation ........................................................................................14

VII

2.4. Previous Empirical Studies ..........................................................................................15

2.4.1. Collector Dosage/Type ...................................................................................15

2.4.2. Feed Size .......................................................................................................23

2.4.3. Use of Activator .............................................................................................27

3. Experimental Methodology .......................................................................................................34

3.1. Feed Materials ...........................................................................................................36

3.2. X-ray Diffraction (XRD) Measurements .......................................................................37

3.3. Sample Preparation ....................................................................................................38

3.4. Grind Establishments .................................................................................................38

3.5. Flotation Tests ...........................................................................................................38

3.6. Scanning Electron Microscopy (SEM) & Energy-Dispersive X-ray Spectroscopy (EDX)

Analysis .................................................................................................................................39

4. Results & Discussion ..................................................................................................................41

4.1. Baseline Tests ............................................................................................................41

4.2. Varying types of Collector ..........................................................................................43

4.3. Effect of Collector Dosage ..........................................................................................45

4.4. Activator Usage ..........................................................................................................46

4.5. Feed Size Alteration ...................................................................................................48

4.6. pH Variation ...............................................................................................................49

4.7. Optimized Conditions .................................................................................................51

4.8. Lithium Bearing Ore Comparison ................................................................................54

5. Limitations of Research Project .................................................................................................56

5.1. Equipment .................................................................................................................56

5.2. Time ...........................................................................................................................56

6. Conclusions ...............................................................................................................................58

7. Future Work & Recommendations ............................................................................................60

References ........................................................................................................................................62

Appendix...........................................................................................................................................66

VIII

Appendix A ............................................................................................................................66

Appendix B ............................................................................................................................68

Appendix C ............................................................................................................................72

X

List of Figures & Tables

Figure 2.1 – Spodumene sample......................................................................................................... 7

Figure 2.2 – Lepidolite Sample ........................................................................................................... 8

Figure 2.3 – Amblygonite sample obtained from Wikimedia commons ............................................... 9

Figure 2.4 – Petalite sample obtained from Wikimedia commons....................................................... 9

Figure 2.5 – Rougher flotation test results .........................................................................................18

Figure 2.6 – Flotation results utilizing different collectors ..................................................................20

Figure 2.7 – Flotation results utilizing different concentrations at pH 7 – 7.5 .....................................21

Figure 2.8 – Flotation results using single (NaOL) and mixed collector (NaOL/DTAC), in absence

and presence of regulators ................................................................................................................22

Figure 2.9 – Particle size influence using NaOL collector ....................................................................24

Figure 2.10 – Flotation recovery of wet and dry ground spodumene of different size fractions

(microns) as a function of pH in NaOL ...............................................................................................26

Figure 2.11 - Flotability of spodumene, albite and quartz as a function of pH with NaOL ...................29

Figure 2.12 – Flotability of spodumene, albite and quartz as a function of pH with NaOL and

Fe(III) ................................................................................................................................................30

Figure 2.13 - Flotability of spodumene, albite and quartz as a function of Fe(III) concentration

with NaOL .........................................................................................................................................30

Figure 2.14 – Flotability of spodumene, albite and quartz as a function of NaOL concentration

with Fe(III) .......................................................................................................................................31

Figure 2.15 – Particle size influence using NaOL collector (0.5mM) and presence of Fe(III)

activator (0.04mM) ...........................................................................................................................32

Figure 3.1 - Testwork Flowsheet ........................................................................................................35

Figure 3.2 – XRD analysis of Spodumene feed ...................................................................................37

Figure 4.1 – Baseline Test Grade & Recovery .....................................................................................42

Figure 4.2 - Baseline kinetic test on Grade .........................................................................................43

XI

Figure 4.4 – Baseline kinetic test on Recovery ...................................................................................43

Figure 4.5 – Varying collectors (*indicates baseline collector) ...........................................................44

Figure 4.6 – Varying Collector Dosage (*indicates baseline dosage) ...................................................46

Figure 4.7 – Activator Influence (*indicates baseline condition) ........................................................47

Figure 4.8 – Varying feed size (*indicates baseline feed size) .............................................................49

Figure 4.9 – Varying pH (*indicates baseline pH) ...............................................................................51

Figure 4.10 – Spodumene flotation at optimized conditions ..............................................................52

Figure 4.11 – SEM of optimized flotation concentrate at 100 microns ...............................................53

Figure 4.12 – SEM image of optimized flotation tails at 100 microns .................................................54

Figure 4.13 – Spodumene against Lepidolite at the same conditions .................................................55

Figure C1 – Points from SEM image of optimized concentrate for EDX Spectroscopy .........................73

Figure C2 – EDX Spectroscopy data for optimized concentrate ..........................................................73

Figure C3 – Points from SEM image of optimized tails for EDX Spectroscopy .....................................74

Figure C4 – EDX Spectroscopy data for optimized tails ......................................................................74

Table 2.1 – Results adapted from Vieceli et al. ..................................................................................17

Table 2.2 – Li2O Grades and Recoveries of the concentrates compared against mixed collector

SXQ and OPWNS ...............................................................................................................................23

Table 3.1 - Chemical Composition of Spodumene Ore .......................................................................36

Table B1 – Baseline & Kinetic Tests....................................................................................................69

Table B2 – Varying Collector Type .....................................................................................................69

Table B3 - Varying Collector Dosage ..................................................................................................70

Table B4 – Activator usage ................................................................................................................70

Table B5 – Varying Feed Size .............................................................................................................70

XII

Table B6 – Varying pH .......................................................................................................................71

Table B7 – Optimized conditions .......................................................................................................71

Table B8 – Lithium Bearing Minerals Comparison ..............................................................................71

XIV

Abbreviations List

Full Name Abbreviated

Australian Financial Review AFR

Sodium Hydroxide NaOH

Sodium Oleate NaOL

Sodium Sulfide Na2S

Lithium Oxide Li2O

Grams per tonne g/t

Stearyl Trimethyl Ammonium Chloride STAC

Dodecyl Trimethyl Ammonium Chloride DTAC

Iron (III) Chloride Hexahydrate Fe(III)

X-ray diffraction XRD

Hydrochloric Acid HCl

Scanning Electron Microscopy SEM

Energy-Dispersive X-Ray Spectroscopy EDX

XVI

Acknowledgements

I would like to acknowledge the support and guidance provided by my supervisor, Dr. Aleks

Nikoloski, throughout this project.

I would also like to thank my colleagues for sharing their invaluable experience and knowledge.

Last, but not least, thank you to my family for their encouragement and emotional support

through this journey.

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1. Introduction

In recent times, lithium has been highly sought after due to its energy capabilities, being an

emerging mineral. This research project will look into the viability of this emerging mineral and

investigate ways to improve its extraction.

1.1. Overview

Australia is sitting comfortably upon its vast reserves of lithium. This surge in lithium demand

has allowed the country to create thousands of jobs for workers as well as made itself known

as the biggest producer of lithium worldwide. With such high demand from other countries, it

is clear that Australia is at the heart of lithium production.

In an effort to look for more efficient energy resources, we now see a re-emergence of lithium

as a viable alternative energy resource. Lithium has become a major contributor in the energy

storage industry, providing longer lasting, smaller batteries that are more powerful and more

reliable for consumers.

As the demand for lithium rises, more research into new ways to obtain the valuable element

is needed. There are several lithium bearing minerals that have high economic value, primarily

spodumene and lepidolite. In order to extract these lithium bearing minerals, a common

process that is generally utilized by industry is the froth flotation method. This method

promotes the separation of the valuable spodumene and lepidolite from its unwanted gangue

minerals.

1.2. Importance of Lithium

The rise of the Internet and related technologies require better battery performance that is

longer lasting, more powerful and reliable. Lithium has proven to be most effective alternative

to lead-acid battery. Surging production of electric vehicles and lithium batteries that power

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them will drive the demand for lithium.

The two biggest lithium mines in Australia are set to grow dramatically, as sales from China

drive demand to meet the increase in sales of electric vehicles and mobile phones. Pilbara

Minerals in Western Australia has signed an offtake agreement with Chinese firm General

Lithium and plans to start mining next year, according to the Australian Financial Review (AFR,

2016). This trend is set to continue, with the soaring increases in the use of lithium globally.

AFR (2016) predicts five million new energy vehicles will be built in China by 2020 and the

conversion of millions of electric bikes to use lithium ion batteries. It was also reported that US

electric car maker Tesla has received 400,000 orders for its Series 3 model vehicle, which will

be released next year (AFR, 2016).

A direct result of this increasing interest in lithium has led to the expansion of Australia’s

lithium mines. An article reported by the Australian Broadcasting Corporation (2017) stated

that an estimated $500 million of investment suggests Australia’s economic engine room could

be home to as many as 7 lithium mines by 2018, compared to the single mine producing

lithium last year. In February 2017, Australia’s Mt Marion mine produced 15,000 tonnes of

lithium concentrate that was sent to China, primarily used for new-generation batteries. Mt

Marion is forecast to produce 400,000 tonnes of lithium concentrate per annum at full

capacity. Mark Cully, the chief economist from the Federal Government Department of

Industry, Innovation and Science has affirmed that ‘Lithium looks to have a bright future’

(Lucas 2017).

Australia’s Greenbushes lithium mine, the world’s biggest lithium mine, is also set to expand

due to the rise of popularity of lithium. The $320 million expansion was greenlit in March and

will set to more than double the current capacity (Williams 2017). With the introduction of a

second production facility at Greenbushes by 2019, the company will reach approximately 1.34

million tonnes of lithium concentrate produced annually.

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Australia is sitting comfortably upon its vast reserves of lithium. This surge in lithium has

allowed the country to create jobs for workers as well as made herself known as the biggest

producer of lithium worldwide. With such high demand from other countries, it is clear that

Australia is at the heart of lithium production.

1.3. Global Dependence

Globally Australia’s successful lithium production techniques has drawn the attention of

international parties who are looking to tap into this popular resource. This has allowed

Australia to enter and develop strong partnerships with Europe and South America.

Positive results from Lithium Australia’s newfound Sileach technology has given them the

opportunity to enter into a joint-venture with Deutsche Rohstof subsidiary, Tin International

AG (Hoey 2017). This will see parties from both sides focus on the Sadisdorf project located in

Saxony, Germany where tin mining is popular. According to Lithium Australia, the real

potential for the project could lie in associated lithium mineralization that has not been fully

evaluated (Birney 2017). The orebody is speculated to contain 15% zinnwaldite, a lithium

bearing mineral which may be able to penetrate the European battery industry. With the

project being in close proximity to major industrial hubs in Germany, this would allow the

abundant lithium to thrive and create a ready market and incorporate Lithium Australia’s

successful Sileach technology for extraction. Zinnwaldite is ideal for the Sileach process, which

in turn can play a vital role in unlocking a new Lithium source for the European market (Birney

2017).

Australia is not only looking to mine lithium strictly in its own territory. Lithium Australia has

recently shown positive results from their Mexican clay mine, Agua Fria, reportedly extracting

99% lithium from its clays (Birney 2017). According to the article by Birney (25/5/2017) the

Agua Fria prospect in Mexico had extracted between 94% to 99% Lithium in only 4 hours,

5

without extensive energy sources such high temperature roasting. With the shift now towards

beneficiation of the clays, the company aims to produce a high-grade concentrate for further

metallurgical work.

There is a growing need to continually develop new ways of extracting lithium so as to make

the process more cost-effective and efficient. Australia needs to maintain her competitive

edge within the new emerging resources markets. This research project aims to add to the

existing knowledge in the exploration of new ways to extract lithium.

1.4. Project Aims & Objectives


minerals by flotation. This was achieved through a series of spodumene flotation tests

designed to evaluate the impacts of collector dosage and type, feed size, activator usage and

pH with respect to the lithium recovery. Once flotation tests were completed, conditions

resulting in the highest recovery were combined to conduct an optimized flotation test on the

spodumene. Two lithium bearing minerals were also compared at the same flotation

conditions to determine the influencing factors behind recovery.

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2. Literature Review

This chapter gives a summary of the analysis of the various types of lithium ore and ways of

concentrating them. Previous empirical studies conducted on the flotation of lithium bearing

minerals was also detailed in this section.

2.1. Mineralogy

The most common minerals used for the production of lithium currently and relevant to this

study are briefly introduced under the following subheadings.

2.1.1. Spodumene

Spodumene (LiAlSi₂O₆) is a pure, glossy yellowish to blue mineral that is insoluble in dilute acid

and is frequently found with an iron oxide coating. To be floated successfully, this surface

coating must be removed from the spodumene. Originating from the silicate group,

spodumene contains approximately 8% lithium oxide (Li2O), 64.5% silicon dioxide (SiO2) and

27.5% aluminium oxide (Al2O3) (Bulatovic 2014). Spodumene is normally found in granite-

pegmatities in the form of crystals, with a range of sizes and associations with quartz, mica

amblygonite, beryllium and tantalum. The image shown in Figure 2.1 reveals the physical

appearance of a spodumene sample following its liberation in a crushing circuit.

7

Figure 2.1 – Spodumene sample

2.1.2. Lepidolite

Lepidolite (K(Li,Al,Rb)2(Al,Si)4O10(F,OH)2), also known as lithium mica, is a pure mineral that is

attached to but not decomposed by acid. Existing as a purple mineral, lepidolite has a similar

crystal structure to that of muscovite. The interlayer potassium that exist is capable of ion

exchange, which frequently occurs with iron salts. This ion exchange can lead to anionic

activation of mica and could possibly, under acid flotation conditions, lead to activation of

other minerals in flotation pulp. Although containing less lithium than spodumene, lepidolite

contains approximately 11% to 28.8% of aluminium oxide (Al2O3), 46.9% to 60.1% silicon

dioxide (SiO2), 1.5% to 5% lithium oxide (Li2O), and 9% fluorite (CaF2) (Bulatovic 2014). The

mineral also contains up to 3.7% of rubidium and 1.5% of cesium, in the forms of rubidium

oxide (Rb2O) and cesium oxide (Cs2O), respectively (Bulatovic 2014). An image of a crushed

lepidolite sample is shown in Figure 2.2, highlighting the variation in physical appearance

existing between the mineral and that of spodumene as shown in the previous section.

8

Figure 2.2 – Lepidolite Sample

2.1.3. Amblygonite

Containing up to 10% Li2O, Amblygonite ((Li,Na)AlPO4(F,OH)) is considered a lithium

phosphate that has a secondary role in the production of lithium, due to its high phosphorus

content (Bulatovic 2014). The white to greenish colour mineral is made up of approximately

54% phosphorus pentoxide (P2O5), 34% aluminium oxide (Al2O3) and 12% fluorine (F2), with this

physical trait clearly evident in the amblygonite mineral sample shown in Figure 2.3.

9

Figure 2.3 – Amblygonite sample obtained from Wikimedia commons

2.1.4. Petalite

Petalite (LiAlSi4O10) is comprised of approximately 4.8% lithium oxide (Li2O), 16.7% aluminium

oxide (Al2O3) and 78.4% silicon dioxide (SiO2), with the physical appearance of the mineral

demonstrated in Figure 2.4 below. Being an aluminium silicate, it is found only in pegmatite

type deposits in the form of aggregates (Bulatovic 2014). Lepidolite, amblygonite and pollucite

are typical of these types of ores.

Figure 2.4 – Petalite sample obtained from Wikimedia commons

10

2.1.5. Zinnvaldite

As explained by Bulatovic (2014), zinnvaldite (KLiFeAl(AlSi3)O10(OH,F)2) lepidolite belongs to an

isomorph group of luotite-lepidolite complexes. Wolframite ((Fe,Mn)WO4), casssiterite (SnO2),

fluorite (CaF2) and topaz (Al2SiO4(F,OH)2) normally are evident within zinvaldite. The mineral is

made up of 12.5% iron oxide as well as poor magnetic separation success.

2.2. Beneficiation of Lithium Ore

The beneficiation of lithium ores has proved to be a challenging task with the lithium minerals

having similar physical-chemical characteristics to its gangue minerals (quartz, feldspar, etc.)

(Bulatovic 2014). Over the years, there have been several new ways of separation that have

been utilized throughout the world.

If the lithium minerals are in the form of large crystals, the hand sorting concentration method

is used. It highlights the colour differences between the lithium and its gangue minerals, and

allows for practical beneficiation of petalite ores. This method is primarily used in Zimbabwe,

Africa and China.

Heavy liquid separation is another method mainly used for spodumene ores (Bulatovic, 2014).

Exploiting the specific gravity differences between the spodumene and gangue minerals,

effective separation can be achieved.

In contrast to heavy liquid separation, thermal treatment is sometimes utilized for the

concentration of spodumene ores (Bulatovic 2014). This process sees the ore being heated to

95-1000 degrees Celsius, which allows the spodumene to be converted from its alpha form to

beta form. In this form, the spodumene exists as a powder which can then be separated from

other minerals by screening or air classification.

When it comes to the beneficiation of lithium minerals, flotation is still the principal method.

11

As time progresses, new and efficient technologies have been created to enhance the flotation

of spodumene and petalite. From Jessup and Bulatovic’s (2000) study, they were able to utilize

innovative new technology to produce a high grade petalite concentrate (4.6% Li2O).

2.3. Flotation

Flotation is considered the most dominant mineral concentration method, being utilized for

almost all sulfide minerals and widely used for non-sulfide metallic minerals. This process is

applied to low grade ores as well as ores that require fine grinding to achieve liberation. Due to

its selectivity, an important application is in the separation and concentration of minerals in

complex ores such as sulfide ores contacting copper, lead and zinc. This primary concentration

process, which is based on the interfacial chemistry of mineral particles in solution, can be split

into two groups; chemical conditions and physical-mechanical conditions (Kelly and

Spottiswood 1997). The chemical conditions involve interactions of chemical reagents with the

mineral particles to make one selectively hydrophobic. The Physical-mechanical conditions are

determined by the flotation machine characteristics.

2.3.1. Process

A major objective of every flotation machine is to make the particles that have been rendered

hydrophobic, contact and adhere to air bubbles, which allows those particles to rise to the

surface and form a froth product that is removed. Kelly and Spottiswood (1997) explain that to

achieve this, the flotation machine must be able to maintain all of the particles in suspension,

which requires upward pulp velocities exceeding the settling velocity of all particles present.

The machine must also ensure that all particles entering the machine are given the chance to

float. Promoting particle-bubble collision so that hydrophobic particles may attach to bubbles

and rise to the froth is an important aspect of any flotation machine. This is achieved through

vigorous agitation, counter-current flow or dissolved gas (air) precipitation. To ensure

minimization of pulp entrainment in the froth and turbulent disruption of the froth layer, a

12

quiescent pulp region is established. The machine must also provide a sufficient depth of froth

to permit drainage of entrained particles to occur.

Each machine with an individual cell, is fitted with an impeller that rotates within baffles. Air is

then introduced through the impeller, which in turn provides dispersion and sufficient mixing.

This results in particle-bubble collisions that are essential to particle-bubble attachment. To

provide particle-bubble contact, a highly turbulent region is required. This also allows for a

quiescent zone adjacent to the froth layer to be formed, which the mineral-laden bubbles can

rise without disruption of particles adhering to the bubble surface. The froth product may

either be removed via direct overflow or with mechanical assistance, such as paddles.

2.3.2. Reagents

The most important reagents that are utilized for flotation is discussed in the following section.

2.3.2.1. Collector

As explained in Bulatovic’s (2007) handbook, collectors are a large group of organic chemical

compounds that are different in chemical composition and function. The primary function of a

collector is to selectively form a hydrophobic layer on any given mineral surface, within the

flotation pulp. Providing conditions for hydrophobic particle attachment to air bubble is

another function of the collector. Collectors can be divided into 2 separate groups that differ

based on their ability to dissociate water. Oxyhydryl and sulfhydryl are known as anionic

collectors, based on their solidophilic property. Chemical compounds where the hydrocarbon

radical is protonized are considered to be cationic collectors. These collectors exist as amines,

from which the primary amine is known to be the most important flotation collectors.

2.3.2.2. Frother

Bulatovic (2007) describes frothers as heteropolar surface-active compounds which can lower

the surface tension of water. They also have the ability to adsorb on the air bubble-water

13

interface. Frothers increase the film strength of air bubbles, which lead to better attachment

of hydrophobic particles to bubbles. The effectiveness of frothers largely depend on the

flotation pulp pH. Optimum performance is achieved when frother is in molecular form.

Frother effectiveness can be divided into 2 groups, acidic and neutral. Acidic frothing ability is

reduced with an increase in pH from acidic to alkaline, while neutral frothing performance

does not depend on pH value of the pulp.

2.3.2.3. Regulator

Regulators are separated into activators, depressants and pH regulators, as explained by

Bulatovic (2007). The primary purpose of these regulators are to modify the action of the

collector on mineral surfaces, which leads to governing the selectivity of the flotation process.

When collectors and regulators are in the same system, the collector only adsorbs on particles

that are targeted for recovery. Activators are considered regulators that react directly with

mineral surfaces and provide suitable conditions for the interactions of the subjected mineral

with the collector. Regulators that reduce the conditions for hydrophobization, or the ability to

make surfaces hydrophilic are known as depressants. pH regulators are apparent to regulate

the ionic composition of the pulp by changing the concentration of the hydrogen ion within

the flotation pulp. As a result, an improvement in collector interaction with subjected mineral

is evident and reduction in collector interaction with undesirable minerals is apparent.

2.3.3. Spodumene Flotation

Bulatovic (2014) reported that the flotation of spodumene is affected by the presence of heavy

metal cations, such as iron. Studies conducted by Arbiter et al. (1961) saw the flotation of

spodumene-beryllium ores, in the presence of oleic acid. This acid was utilized in order to

separate the spodumene from beryllium. Arbiter et al. (1961) pretreated the pulp with sodium

hydroxide (NaOH), then deslimed the product before conditioning with collector at an elevated

temperature of 80 to 85 degrees Celsius. Utilizing this method, Arbiter et al. (1961) were able

14

to selectively float spodumene from beryllium and other gangue minerals.

Throughout the beneficiation of spodumene bearing ores, pulp pretreatment prior to flotation

is of vital importance, as confirmed by Browning and McVay’s (1961) report. Pretreatment is

conducted in 2 separate ways; through conditioning pulp at 50 – 60% solids with 0.5 to 1

kilograms per tonne of NaOH for 20 to 30 minutes, followed by desliming of the pulp. The

second method utilizes sodium sulfide (Na2S) in the pretreatment stage, which gave positive

metallurgical results. Bulatovic (2014) emphasized that the best collector for spodumene

flotation was oleic acid. This was further proved with pilot plant testing with South Dakota ore

(US), where NaOH was used for pretreatment. Desliming and spodumene flotation occurred

after the pretreatment, with oleic acid collector, resulting in 75% recovery and grade assaying

7.2% lithium oxide (Li2O).

2.3.4. Lepidolite Flotation

Bulatovic (2014) states that lepidolite, also known as lithium mica, has a large flotation area

normally with cationic collector, dodecylamines, which has a range of pH 2.5 to pH 11. In

contrast with spodumene, pretreatment of lepidolite with NaOH has no beneficial impact upon

flotation. Pretreatment with oleic acid improves the flotation of lepidolite, when utilizing

sodium oleate (NaOL) as a primary collector. Starches and sodium sulfide act as depressants

for lepidolite flotation. Considering amines that were examined in other studies, it was found

that hexadecyl amine acetate (Armac 16D – Akzo Nobel) proved to be the most effective

collector, when floating at pH 3.5.

Another important factor to consider when floating lepidolite is the pulp density. Having a pulp

density below 20% has a negative impact on lepidolite flotation. The optimum pulp density

reported by Bulatovic (2014) is approximately 25% solids. The South African National Institute

of Metals developed a collector mixture containing amine and petroleum sulfonate, at a 1 to 1

ratio, to be utilized for lepidolite flotation. This collector mixture, floating at a pH of 9, resulted

15

in a concentrate grade of 3.8% Li2O and a maximum recovery of 80% was achieved.

The research methods in this research project have been designed based on these factors.

2.4. Previous Empirical Studies

A number of research projects have been conducted which highlighted the importance of

collector dosage and type, the effect of feed size and the use of an activator, within the

flotation of lithium. Each of these variables and their respective impacts on the flotation of

lithium ores will be reviewed in the following section.

2.4.1. Collector Dosage/Type

Varying the collector dosage during the flotation of lithium ore was investigated by Vieceli et

al. (2016). One of the objectives of their research was to highlight the impact of 3 different

dosages (200, 350 and 500 grams per tonne) of Aeromine 3000C collector, on the flotation of

lepidolite.

Approximately 50 kilograms of lepidolite ore was ground to minus 500 microns through stage

grinding. Several sub samples of 750 grams was taken from the original crushed ore and used

for the flotation tests. To keep the flotation tests constant, a 3 litre float cell with a constant

impeller speed of 750 rpm and air flowrate of 200 litres per hour were used. MIBC was used as

a frothing agent and pH regulators, sulfuric acid and NaOH, were utilized. Due to high

consumption of reagents by ultrafine particles existing within the system, the flotation feed

was subjected to desliming via ‘beaker decantation’. Introduced by Wills and Munn (2006), this

process sees the combination of solid particle sedimentation in dilute pulp and siphoning the

liquid at the top layer. The flotation tests began with the deslimed product with 50% solids by

weight (1 Litre of water mix) being conditioned with Aeromine 3000C (5% vol. in water) for 10

minutes, with pH adjustments being made. The density of the pulp was reduced to 25% as

stated by Bulatovic (2014) as conditioning time came toward an end.

16

When maintaining pH and flotation time at a constant, Vieceli et al. (2016) found that with

increasing collector dosage, there was an increase in lithium recovery, but not lithium grade.

The first dosage tested was 200 grams per tonne (g/t) of Aeromine 3000C. It was found that at

a pH of 3 and flotation time of 10 minutes, this dosage resulted in a recovery of 67.44% and a

grade of 1.96%. Keeping the same pH and flotation time, the collector dosage was increased to

350 and 500 g/t. For the 350 g/t dosage, a recovery of 84.06% was found and a final grade of

1.93%. Similar figures were reported for the 500 g/t dosage, with recovery found at 86.62%

and grade at 1.82%.

Vieceli et al. (2016) found that with a dosage of 500 g/t, pH of 2 and flotation time of 12

minutes resulted in the highest recovery and grade, 91.5% and 1.96% respectively. Although a

higher lithium content was achieved with a lower dosage, it resulted in a lower recovery. Table

2.1 illustrates these findings and the influence of varied flotation conditions on the final

recovery and grade of lithium concentrates.

These findings highlight the importance of the collector, Aeromine 3000C, on lepidolite

flotation by successfully recovering lithium. Aeromine 3000C collector was considered for this

research project, however due to difficulty in obtaining it, other collectors were trialed.

17

Table 4.1 – Results adapted from Vieceli et al. (2016)

Independent Values Responses

Pulp pH Dosage of Collector

Flotation Time

Standard Order

Run Order

Original Units (-)

Coded x1

Original Units (g/t)

Coded x2

Original Units (min)

Coded x3

Li Recovery (%)

Li Content (%)

1 11 2 -1 200 -1 8 -1 70.99 2.07

2 8 4 1 200 -1 8 -1 46.68 1.97

3 3 2 -1 500 1 8 -1 89.91 1.96

4 5 4 1 500 1 8 -1 88.47 1.76

5 1 2 -1 200 -1 12 1 64.06 2.07

6 9 4 1 200 -1 12 1 47.03 1.9

7 4 2 -1 500 1 12 1 91.51 1.96

8 7 4 1 500 1 12 1 84.88 1.42

9 10 3 0 350 0 10 0 84.06 1.93

10 2 3 0 350 0 10 0 82.16 1.86

11 12 3 0 350 0 10 0 82.66 1.94

12 6 3 0 350 0 10 0 80.19 1.94

13 14 2 -1 350 0 10 0 82.72 1.93

14 15 4 1 350 0 10 0 78.53 1.82

15 16 3 0 200 -1 10 0 67.44 1.96

16 17 3 0 500 1 10 0 86.62 1.82

17 13 3 0 350 0 8 -1 79.51 1.96

18 18 3 0 350 0 12 1 84.34 1.93

Choi, et al. (2015) examined the influence of Stearyl Trimethyl Ammonium Chloride (STAC), as

a collector on the flotation behavior of lepidolite. Varying STAC collector dosages (50 – 200 g/t)

was an important objective of this study, which will aid in further understanding the most

efficient dosage for lepidolite flotation.

The lepidolite samples originated from South Korea, where it was crushed and ground to

minus 2.36 millimeters. Samples were crushed and ground further to minus 212, minus 150,

minus 106 and minus 75 microns. Modifiers that were utilized in this study include Aerofloat

65 (AF – 65) as a frother, and hydrochloric acid (HCl) as well as NaOH as pH regulators. The

18

flotation tests were conducted in a 1 Litre Denver flotation cell with an impeller speed set at

1200 rpm. 100 grams of lepidolite ore was slurried with 900 millilitres of deionized water

within the cell. Collector conditioning lasted 10 minutes with different dosages of STAC,

followed by frother conditioning with AF – 65 for a further 3 minutes. The flotation time for

each test was 10 minutes. The flotation tests were broken into 3 parts, the first part being a

rougher flotation, second and third being a first and second cleaner flotation, respectively. In

their first set of tests Choi et al. (2015) added STAC collector ranging from 50 to 200 g/t, to the

pulp at different feed sizes at a constant pH of 9. The study reports that the lithium recovery

increased with increasing STAC concentration and decreasing feed size. The lithium recovery

reached a plateau once the collector dosage was greater than 150 g/t and feed size less than

or equal to 150 microns. The yield of the floated products was greater than 90% by weight for

samples with STAC dosages of greater or equal to 150 g/t. The results from Choi et al. (2015)

tests can be seen in Figure 2.5.

Figure 2.5 – Rougher flotation test results adapted from Choi et al. (2015)

Once Choi et al. (2015) identified the optimal collector dosage and size fraction from the

rougher stage, they were able to further identify the conditions that would lead to the highest

recovery and grade of lithium. These tests were conducted in the first and second cleaner

stages.

0

20

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60

80

100

120

50 100 150 200

Li R

ecov

ery

(%)

Dosage of STAC (g/t)

-65 mesh -100 mesh -150 mesh -200 mesh

19

This study confirms that a high collector dosage is more favourable than a lower dosage,

resulting in a greater recovery. A comparison of collectors including STAC at high dosages was

carried out in this research project.

The study conducted by Xu et al. (2016) explores the selectivity in flotation of spodumene from

other pegmatic aluminosilicates such as feldspar and mica. It was found that a collector

mixture of NaOL and dodecyl trimethyl ammonium chloride (DTAC) demonstrates a high

selectivity for spodumene from feldspar. The study determined that the optimal molar ratio

between NaOL and DTAC collectors is approximately 9:1.

Spodumene and feldspar samples used for testing originated from the Jiajika Lithium mine, in

China. The samples were handpicked, crushed and ground in a laboratory porcelain mill to -

0.0074 millimeters. Xu et al. (2016) conducted both micro-flotation tests as well as batch

flotation tests. in the micro-flotation tests, NaOL and DTAC were used as the anionic and

cationic collector, respectively. Calcium chloride and sodium carbonate were used as

regulators, while HCl and NaOH were used as pH adjusters. A 40 millilitre hitch groove cell with

a constant impeller speed of 1700 rpm was employed for the micro-flotation tests.

Approximately 3 grams of the sample (1 gram of spodumene and 2 grams of feldspar) in 40

millilitres of deionized water was conditioned for 5 minutes, and then floated for a further 10

minutes. The batch flotation tests utilized a home-made mixed collector NaOL/DTAC, in

comparison with a mixed fatty acid soap, oxidised paraffin wax soap and naphthenic soap

(OPWNS) as a collector. Sodium carbonate, NaOH and calcium chloride were used as

regulators within the system. For the batch tests, 500 grams of ore was ground to 70% passing

at 74 microns. The pulp was then transferred to a flotation cell while agitating at 1500 rpm and

conditioned for 3 minutes.

The study began with a single mineral micro-flotation intended to compare the response of

spodumene and feldspar when floated with NaOL and DTAC. The first set of tests evaluated

20

flotation separation using the single collector NaOL (4.0 x 10-4 mol/litre) and DTAC (2.0 x 10-4

mol/litre). Xu et al. (2016) reported that the favourable pulp pH range, using NaOL as an

individual collector for spodumene flotation, is between 8 and 9. Figure 2.6 below highlights

that the flotation recovery of spodumene and feldspar increases with the increase of pH when

DTAC collector is used.

Figure 2.6 – Flotation results utilizing different collectors adapted from Xu et al. (2016) a) NaOL (4.0 x 10-4 mol/L) b) DTAC (2.0 x 10-

4 mol/L)

Following these tests, flotation recovery against collector concentration was tested within the

pH range of 7 to 7.5, which can be seen in Figure 2.7. The study shows that the recoveries of

spodumene and feldspar, with individual collector NaOL and DTAC, are dependent on collector

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Flot

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)

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Spodumene Feldspar

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)

pH

Spodumene Feldspar

21

concentration. An increase in concentration of DTAC resulted in an increase in recovery within

the low collector range, approximately less than 0.4 millimolar for NaOL and 0.2 millimolar for

DTAC. Concentrations above this range produced a flat horizontal, representing a maximum

flotation recovery. The maximum spodumene recovery reached an estimated 50% with NaOL

collector alone while recoveries for both spodumene and feldspar were above 80% using

DTAC. Xu et al. (2016) noted that selective spodumene flotation from feldspar using NaOL or

DTAC alone as a collector could not be achieved.

Figure 2.7 – Flotation results utilizing different concentrations at pH 7 – 7.5 adapted from Xu et al. (2016) a) NaOL b) DTAC

To further examine the selectivity performance of these collectors, Xu et al. (2016) performed

micro-flotation tests of mixed mineral samples, with the results shown in Figure 2.8 below. A

0102030405060708090

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0 1 2 3 4 5 6 7

Flot

aion

Rec

over

y (%

)

Concentration (1.0 x 10-4 mol/L)

Feldspar Spodumene

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Flot

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Rec

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)

Concentration (1.0 x 10-4 mol/L)

Feldspar Spodumene

22

mixed mineral sample containing 2.62% Li2O was floated using NaOL collector and a

NaOL/DTAC collector mixture, with a molar ratio of 9:1, without the use of any regulators

(calcium chloride activator and sodium chloride depressant). These results were then

compared with grades and recoveries of Li2O, in the presence of the noted regulators. The

flotation tests reported that Li2O grade of a concentration using a single NaOL collector alters

from 4.13% to 3.33%, when in the presence of regulators calcium chloride and sodium

carbonate. The recovery of the concentration of Li2O also falls from 42.21% to 31.29% with and

without regulators present, respectively. When using the mixed collector, a concentration

containing 5.57% Li2O and 71.13% Li2O recovery was achieved. However, when compared to

the grade and recovery while in the presence of regulators, the grade drops to 4.47% and

recovery of 62.2%. Xu et al. (2016) made it clear that the NaOL/DTAC collector mixture has a

more selective flotation of spodumene from feldspar.

Figure 2.8 – Flotation results using single(NaOL) and mixed collector (NaOL/DTAC), in absence and presence of regulators adapted from Xu et al. (2016)

The primary objective in the batch flotation tests was to compare the mixed collector,

NaOL/DTAC (known as SXQ), to the oxidised paraffin wax soap and naphthenic soap (OPWNS).

It was found that the Li2O grade and recovery with the SXQ collector was higher than with

OPWNS collector. As illustrated in Table 2.2, the grade and recovery increased by 0.31% and

0102030405060708090100

0

1

2

3

4

5

6

NaOL NaOL/DTAC NaOL + Ca2+ +Na2CO3

NaOL/DTAC + Ca2+ +Na2CO3

Li2O

Rec

over

y (%

)

Li2O

Gra

de (%

)

Li Grade (%) Li Recovery (%)

23

4.93% respectively. Using 800 g/t of the SXQ collector resulted in a Li2O grade of 6.17% and a

Li2O recovery of 85.17%, which demonstrates the ability to be utilized in industrial application.

Xu et al. (2016) were successful in examining the use of mixed collectors in the flotation of

spodumene from feldspar. This project will also involve the mixing of different collectors to

explore its impact on flotation recovery.

Table 4.2 – Li2O Grades and Recoveries of the concentrates compared against mixed collector SXQ and OPWNS adapted from Xu et al. (2016)

System Products Ratio, w (%)

Li2O Grades (%) Li2O Recoveries (%)

SXQ (800 g/t) Concentrates 19.74 6.17 85.17 Tailings 80.26 0.26 14.83 Feed 100.00 1.43 100.00

OPWNS (2400 g/t) Concentrates 19.58 5.86 80.24 Tailings 80.42 0.35 19.76 Feed 100.00 1.43 100.00

The articles presented in this section provide information on effective mixture and dosage of

collectors as well as the range of collector types investigated during this research project.

2.4.2. Feed Size

An investigation conducted by Xu et al. (2016) examined the flotation and adsorption of NaOL

on spodumene with 4 different particle size fractions. Micro-flotation tests were conducted to

examine the effect of particle size on spodumene flotation.

Spodumene ore was handpicked from the Jiajika Lithium Mine, in China. XRD analysis showed

that the ore contained 7.86% lithium oxide with purity values greater than 90%. A laboratory

porcelain mill was used to crush and grind the samples into a powder, which was screened and

separated to 4 size fractions, 45 – 75 microns, 38 – 45 microns, 19 – 38 microns and 0 – 19

microns. NaOL was utilized as the anionic collector, while ferric trichloride (Fe3+) was used as

an activator. HCl and NaOH were employed as pH regulators. Approximately 3 grams of the

sample was placed in a 40 millilitre hitch groove flotation cell, and agitated for 3 minutes with

24

desired reagents. Once conditioning time ended, flotation carried out for 4 minutes.

A primary objective for Xu et al. (2016) was to investigate the flotation behavior of spodumene

with 4 different size fractions with NaOL collector in the absence and presence of the activator,

Fe3+ as a function of pH. This section is focused on the absence of the activator and its

influence on particle size. It was reported that the recovery of spodumene increases upon

decreasing the size fraction, from 45 – 75 microns to 38 – 45 microns. From Figure 2.9, the

maximum recovery was found at 38 – 45 microns, corresponding to approximately 40%

recovery. The recovery then decreased gradually as the size fraction shifts from 38 – 45

microns to 0 – 19 microns over the entire pH range. The study suggests that this decreasing

trend “indicated that fewer surface Al sites are available for chemisorption of NaOL when the

particle size of spodumene decreases. “(Xu et al., 2016).

Figure 2.9 – Particle size influence using NaOL collector (0.5mM NaOL) adapted from Xu et al. (2016)

The adsorption of NaOL collector on spodumene with different size fractions was explored to

confirm the flotability of spodumene with different sizes. The adsorption densities of NaOL

with the 4 different size fractions were measured as a function of pH. Xu et al. (2016) observed

that the trends of adsorption density of NaOL on spodumene agree with the micro-flotation

tests that were previously conducted, both tests reporting a peak at pH 8.5. The adsorption

density initially increases with decreasing size fractions, from 45 – 75 microns to 30 – 45

05

101520253035404550

0 2 4 6 8 10 12 14

Flot

aion

Rec

over

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)

pH

0-19 um 19-38 um 45-75 um 38-45 um

25

microns, attaining a peak value at 38 – 45 microns. After this increase, the adsorption density

decreases gradually with further decreasing size fractions, from 38 – 45 microns to 0 – 19

microns. This study summarized that as the particle sized altered, different crystal planes for

the chemisorption of NaOL on spodumene will play a different role.

These results highlight the importance of particle size upon spodumene flotation. From the

study, a fine particle size fraction of 38 – 45 microns was successful in floating spodumene

across a range of pH values. For this research project, a feed size of 38 – 45 microns was

considered, but due to the hardness of the ore, a minimum feed size of 53 microns was

chosen.

The influence of particle size on spodumene flotation was an important objective examined by

Zhu et al. (2015). Micro-flotation tests were conducted to compare the flotability of both wet

and dry ground spodumene in different grain sizes, as a function of pH.

Originating from China, the spodumene ore was handpicked and crushed to minus 3

millimeters. To eliminate iron contamination, dry strong magnetic separation was carried out,

resulting in iron impurities being removed. The crushed samples were ground to minus 105

microns and utilized for experimental work. Size fractions of -105 + 38 microns, -75 + 45

microns, -45 +38 microns and -38 + 23 microns were obtained from the ground spodumene

through screening. NaOL was employed as a collector with HCl and NaOH being added as pH

regulators. A 20 milliliter plexiglass cell was used for the micro-flotation tests, with a constant

impeller speed set at 1700 rpm. With the impeller speed set, 2 grams of spodumene was

slurried with deionized water and transferred into the cell, conditioned with 6.0 x 10-4 moles

per litre of NaOL collector and pH regulators for 2 minutes. After the 2 minutes of

conditioning, the froth was collected for 3 minutes.

In the study undertaken by Zhu et al. (2015), it was reported that there was a similar trend of

26

flotation recovery as a function of pH between the different feed sizes. There was a directly

proportional relationship between the recovery of spodumene and pH value, reaching a peak

value of 8 – 9. Once this value was reached, the relationship changed to an inversely

proportional relationship, with pH increasing with decreasing recovery. Zhu et al. (2015) the

highest recovery of approximately 67% reported to the size fraction -45 + 38 microns. -75 + 45

microns’ size fraction resulted in a slightly lower flotation recovery of 60%. Recoveries of -105

+38 microns’ and -38 +23 microns’ size fraction were less than the other 2 sizes, with 53% and

40%, respectively. The results can be seen in Figure 2.10.

This study supports the findings by Xu et al. (2016), where both studies conclude that -45 + 38

micron size fraction results in the highest flotation recovery. This similar trend reinforces the

choice of a fine grind size, which was achieved in this research project.

Figure 2.10 – Flotation recovery of wet and dry ground spodumene of different size fractions (microns) as a function of pH in NaOL adapted from Zhu et al. study (2015) (1) -75 + 45 dry (2) -75 + 45 wet (3) -45 + 38 dry (4) -45 + 38 wet (5) -38 + 23 dry (6) -38 + 23

wet (7) -105 + 38 dry (8) -105 + 38 wet

Varying STAC collector dosages (50 – 200 g/t) was an important objective of the study by Choi

et al. (2015). An investigation into optimal feed size was also conducted in the study, where

size fractions of -212 microns, -150 microns, -106 microns and -75 microns were used.

In the rougher flotation, the lithium recovery reached a plateau (100% recovery) with -106

0

10

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30

40

50

60

70

0 2 4 6 8 10 12 14

Reco

very

(%)

pH

4 2 8 6 1 3 5 7

27

microns and -75 microns, when STAC dosages were greater than 150 g/t. With STAC dosages

lower than 150 g/t, this high recovery was still attainable with -106 microns and -75 microns

yielding recoveries of 98% to 80%. It was also evident that -212 microns size yielded the lowest

recovery with approximately 30% to 40%. As the particle size decreased, an increase in

recovery was evident highlighting an inversely proportional relationship.

This trend is supported by Xu et al. (2016) and Tian et al. (2016), who both identified that finer

particles are more favourable for spodumene and lepidolite flotation, rather than coarser

particles. From these articles, it is clear that a finer feed size is recommended, ranging from -

106 microns to 38 microns. This further supports the use of a fine feed size.

2.4.3. Use of Activator

Jie et al. (2014) investigated the influence of Fe(III) ions on the flotation of spodumene, albite

and quartz minerals using NaOL. Micro-flotation tests were conducted with varying amounts of

NaOL collector and Fe(III) activator, with flotation recovery being measured.

Sample of spodumene, albite and quartz were hand-picked from the Lijiagou Lithium Mine, in

China. A final size range of -75 + 38 microns was obtained after several stages of crushing and

grinding. NaOL collector was utilized for the flotation tests with Iron (III) chloride hexahydrate

(Fe(III)) acting as the activator. HCl and NaOH were used as pH adjusters. For the flotation

tests, samples weighing 2 grams were placed in a 40 millilitre plexiglass cell, and made up with

30 milliliters of ultra-pure water. Conditioning with pH adjusters lasted for 2 minutes, followed

by a 2 minute conditioning with the activator and a further 3 minutes with the collector. Once

the pH was measured, the flotation tests were conducted for 4 minutes.

A primary objective of this study was to “investigate the underlying mechanism of the anionic

collectors on the preferential flotation of spodumene mineral from a mixture of albite and

quartz” (Jie, et al. 2014). Initial tests served to provide an insight into the flotability of

28

spodumene, albite and quartz as a function of pH. It was reported that the flotatbility of

spodumene, albite and quartz with the anionic NaOL collector (2.0 x 10-4 molar) acting alone,

was poor. The maximum recovery of spodumene was approximately 9.5%, at a pH value of 8.7.

Jie et al. (2014) attributed the poor flotability to spodumene having weak interactions with

NaOL collector, resulting in poor flotability.

The presence of Fe(III) metal ions was discovered to have a positive impact on flotation

separation, following several micro-flotation tests. The study highlights flotation recoveries

with the minerals as a function of pH, with 1.5 x 10-4 molar Fe(III) and 2.0 x 10-4 molar NaOL,

indicating a positive separation. The results show spodumene recovery increases with

increasing pH, up to a pH of 7, which after then decreases rapidly. Spodumene achieved a

maximum recovery of 70% obtained at a pH of 7, while albite and quartz achieved a maximum

recovery of 30% and 45% respectively.

Varying amounts of Fe(III) activator, with constant collector dosage and pH value, saw a

directly proportional relationship between activator and recovery. Keeping the collector

concentration at 2.0 x 10-4 molar at a pH value between 6 – 7, showed an increase in recovery

of spodumene and quartz with increasing Fe(III) concentration. However, there was no

reported change in the recovery of albite.

The flotation of the 3 minerals was conducted with constant activator concentration between

a pH of 6 -7, with varying amounts of NaOL collector. A directly proportional relationship

between flotation recovery of spodumene and NaOL collector concentration discovered, until

a maximum concentration of 6.0 x 10-4 molar which resulted in the recovery remaining

unchanged. The maximum recovery of spodumene reached was approximately 90% at the

maximum concentration mentioned. Recovery of albite ranged from 5% – 30% until a collector

concentration of 4.0 x 10-4 molar was reached. Above this value, the recovery decreased. The

maximum recovery attained for quartz was reported at 70% with a corresponding NaOL

29

concentration value of 2.0 x 10-4 molar. Figures 2.11 through to 2.14 illustrate the results from

Jie et al. (2014) tests.

The study conducted by Jie et al. (2014) examined the positive impact of Fe(III) activator on the

flotation of spodumene, albite and quartz. It was reported that a maximum flotation recovery

of the spodumene mineral with iron activation was 90%, which is almost 10 times greater than

without the activator. Jie et al. (2014) were successful in highlighting the significance of an

activator within the flotation of spodumene. Hence for this research project, a ferric activator

was used.

Figure 2.11 - Flotability of spodumene, albite and quartz as a function of pH with NaOL adapted from Jie et al. (2014)

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Reco

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(%)

pH

Albite Quartz Spodumene

30

Figure 2.12 – Flotability of spodumene, albite and quartz as a function of pH with NaOL and Fe(III) adapted from Jie et al. (2014)

Figure 2.13 - Flotability of spodumene, albite and quartz as a function of Fe(III) concentration with NaOL at pH 6 -7 adapted from Jie et al. (2014)

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(%)

pH


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0.00005 0.0001 0.00015 0.0002

Reco

very

(%)

Fe(III) concentration (mol/L)


31

Figure 2.14 – Flotability of spodumene, albite and quartz as a function of NaOL concentration with Fe(III) at pH 6 -7 adapted from Jie et al. (2014)

A primary objective of Xu et al. (2016) was to investigate the flotation behavior of spodumene

with 4 different size fractions with NaOL collector, in the absence and presence of the activator

ferric trichloride (Fe3+) as a function of pH. The results regarding the presence of Fe3+ activator

will be discussed in this section.

According to Xu et al. (2016), the recovery of activated spodumene by Fe3+ is much higher than

that of un-activated spodumene for the same size fraction. Similar trends between activated

and un-activated sizes were evident, with both attaining maximum recoveries at pH 8.5. This

peak pH for all size fractions agrees with other oleate systems (Moon and Fuerstenau 2011.

Moon 1986). As seen in Figure 2.15, with the addition of 0.04 millimolar of Fe3+, the maximum

recovery of spodumene reached approximately 90% at pH 8.5, corresponding to a size fraction

of 38 – 45 microns. The study suggests that in terms of particle size of spodumene, the

activated and un-activated spodumene display the same flotation rule. Xu et al. (2016) imply

that the activator Fe3+ does not dictate the regularity of flotation of spodumene with different

sizes. The recovery increases with decreasing size fractions, from 45 – 75 microns to 38 – 45

microns, highlighting an inversely proportional relationship. The recovery begins to decrease

with decreasing particle size from 38 – 45 micron fraction to 0 – 19 microns, once the pH value

increases past 8.5. The study notes that the decrease in recovery is an indication that fewer

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0.00001 0.0001 0.001

Reco

very

(%)

NaOL concentration (mol/L)


32

surface Al sites are available for the chemisorption of NaOL, when the particle size of

spodumene decreases.

Figure 2.15 – Particle size influence using NaOL collector (0.5mM) and presence of Fe(III) activator (0.04mM) adapted from Xu et al. (2016)

The investigations presented in this section focus on the importance of activators on the

flotation of spodumene. Both papers found that recovery of spodumene increased greatly,

with the use of an iron activator. Xu et al. (2016) further reinforced the findings from Jie et al.

(2014), with both reporting 90% recoveries of spodumene, in the presence of an activator.

These positive results of flotation recovery support the use of a ferric activator.

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0-19 um 19-38 um 45-75 um 38-45 um

34

3. Experimental Methodology


minerals by flotation. This was achieved through the following objectives:

• To investigate the effects of collector type and dosage on the flotation of spodumene.

The collectors that will be investigated are Sodium Oleate (NaOL), Stearyl Trimethyl

Ammonium Chloride (STAC) and Dodecyl Trimethyl Ammonium Chloride (DTAC) at 50,

100, 150 and 200 g/t.

• To investigate the effects of feed size on spodumene. Feed sizes that will be

investigated are 212 microns, 106 microns, 75 microns and 53 microns

• To investigate the effects of the absence and presence of an activator on spodumene.

The activator that will be considered in this project is Iron (III) Chloride Hexahydrate

(FeCl3), referred to as Fe3+.

• To investigate the effects of pH on the flotation of spodumene. The pH was ranged

from 6 to 12.

• Once optimal conditions for collector dosage and type, feed size, activator type and pH

were determined through these investigations, they were used for the determination

of an optimized spodumene flotation test.

• To compare the mineral recoveries between spodumene and lepidolite. Both ores

were floated at the same conditions to identify the influencing factors in their flotation

process.

Figure 3.1 visually represents the projects testwork scheme.

35

Figure 3.1 - Testwork Flowsheet

This research project was separated into a series of stages, leading up to the eventual flotation

of spodumene. After each flotation test was conducted, optimum conditions were attained

based on the recovery of lithium. Once these conditions were found, an optimized float was

conducted on spodumene. There was a comparative test between the spodumene ore and

lepidolite. Both were floated at the same conditions to investigate the differences in lithium

recoveries.

The first stage of processing dealt with sample preparation. Once the ore was at a reasonable

size, grind establishments was conducted in the second stage. The final stage involved

flotation tests with varying conditions as outlined in Chapter 3. Appendix A details the steps

taken within each stage.

To gain insight into the chemical composition of the feed material, the resulting froth

concentrate and tails, samples were sent to an off-site laboratory to undergo X-ray fluoresces

(XRF) and inductively coupled plasma mass spectrometry (ICP-MS) analysis. The ICP-MS

analysis was used to identify the lithium grade within the samples by fusing it with sodium

peroxide (Na2O2) following a digest with dilution hydrochloric acid (HCl). The XRF analysis was

used to identify the remaining associated elements, such as aluminium, silicon and potassium

Spodumene Flotation

Collector Type

- NaOL

- STAC

- DTAC

Collector Dosage

- 50 g/t

- 100 g/t

- 150 g/t

- 200 g/t

Activator Usage

- FeCl3

pH Level

- 6

- 8

- 10

- 12

Feed Size

- 53 µm

- 75 µm

- 106

µm

- 212

36

by fusing the sample in lithium borate flux with lithium nitrate additive.

3.1. Feed Materials

The spodumene samples originated from Talison’s Greenbushes mine site, located south east

of Bunbury in Western Australia. The spodumene ore was crushed, ground and split for

flotation test work. The lepidolite ore originated from Lepidolite Hill at a P100 of 3.35

millimeters.

The spodumene samples were primarily at 4 millimeters, which had to be dried and

homogenized before sample preparation could take place. X-ray diffraction (XRD) techniques,

as well as chemical composition analysis (ICP-MS & XRF), were utilized to study the chemical

and mineral compositions of the spodumene ore. Table 3.1 highlights the chemical

composition of the ore.

Table 1.1 - Chemical Composition of Spodumene Ore

Sample Li (%) Fe2O3 (%) K2O (%) Al2O3 (%) SiO2 (%) Other (%) Total (%)

Spodumene 1.21 1.05 1.91 16.43 71.57 6.43 100.00

The 3 different collectors were investigated throughout the flotation tests to determine which

of these could attain the highest recovery. Stearyl Trimethyl Ammonium Chloride (STAC) and

Dodecyl Trimethyl Ammonium Chloride (DTAC) collector was prepared at approximately 2%

volume in water. These collectors were made up prior to their respective tests to minimize

contamination within storage vessels. NaOL was used as a comparative collector and made up

in the same manner as the other 2 collectors.

The pH of the pulp was controlled by NaOH and HCl. These pH regulators were both prepared

at approximately 10% volume with distilled water. The frothing agent utilized for the flotation

37

tests was W22 Polyfroth.

Iron (III) Chloride Hexahydrate will be utilized as an activator for this stage. Previous tests

conducted by Jie et al. (2014) have highlighted the significance of this activator. In light of this,

4.0 x 10-2 millimoles of activator was used when conditioning the cell.

3.2. X-ray Diffraction (XRD) Measurements

X-ray Diffraction (XRD) was utilized to identify the main minerals within the spodumene feed.

The GBC – EMMA Spellman DF3 x-ray machine was utilized for this measurement. The machine

has an output of approximately 35 kV and 28 mA with a step width of 0.02°. The spodumene

samples were pulverized to an acceptable size for the machine. Once at the optimal size, the

samples are placed in individual slits and placed within the machine which scans between 5°

and 60°. Figure 3.2 illustrates the XRD image.

Figure 3.2 – XRD analysis of Spodumene feed

This XRD pattern confirms the feed samples mineralogy, identifying the lithium bearing

38

mineral as spodumene and the associated gangue mineral quartz. This identification is

demonstrated on the pattern with both red and blue lines, representing the respective

minerals above, accounting for the ore sample peaks produced.

3.3. Sample Preparation

The samples received were crushed using a conventional cone crusher. Samples were stage

crushed to a desired P100 size of 3.35 millimetres. During stage crushing, the sample went

through the cone crusher and then screened over 3.35 millimetres, with the oversize being re-

crushed and re-screened. Once samples were at 3.35 millimetres, a rotary sample splitter was

used to homogenize the sample. Samples flowed through the top of the machine and were

split into fractions, which normalized the sample. The sample was cycled through

approximately 3 times to ensure it was well mixed.

3.4. Grind Establishments

A laboratory steel rod mill was utilized to grind the sample to optimal size. The maximum

capacity of the mill is 0.5 kg dry solids, therefore 0.5 kg charges were placed within the mill,

and made up to 50% solids. Grind establishments were conducted to obtain grind times for P80

sizes of 212 microns, 106 microns, 75 microns and 53 microns. After grinding was completed,

samples were wet screened over lowest screen size (53 microns) to avoid fine buildup within

screens. After wet screening was completed, sample was filtered, dried and then screened to

obtain size fractions using a dry screen stack machine.

3.5. Flotation Tests

A 1.5 litre flotation cell was utilized for the flotation tests with the impeller speed fixed at

approximately 1200 rpm. The ground spodumene sample was transferred into the cell and

made up to 50% solids with Perth tap water. Once the cell was at the desired volume, pH

regulators were added to attain a pH of 10. The selected collector was added to the cell and

39

conditioned for 5 minutes, followed by the frother addition which was conditioned for a

further 2 minutes. The froth was scraped off the top of the cell every 6 seconds, for a total of

10 minutes. The froth concentrate and tails were then filtered, dried and prepared separately

for assaying.

3.6. Scanning Electron Microscopy (SEM) & Energy-Dispersive X-ray

Spectroscopy (EDX) Analysis

The optimized flotation tests were subjected to SEM/EDX analysis where electron beams were

projected onto the sample and were detected via an electron detector to produce an image.

The backscattered electrons provide a contrast based on atomic number, whereby the larger

the atoms, the higher energy electrons are bounced back out. This then creates an image of

the particles within the sample.

To analyse the samples, they were first made into resins by mixing approximately 2 grams of

desired sample with a mixture of Epoxy Resin and EpoFix Hardener. The resin was polished and

buffed before they were analysed.

The SEM machine operates in vacuum conditions and samples must be rendered conductive.

To do this, samples were carbon coated by the SPI Supplies Sputter Coater. The following steps

briefly outline the process of coating the sample:

- Samples were placed within sample holder along with sample mount of similar height

- A piece of carbon fibre was cut out and attached between the 2 fibre holders within

the carbon coater head

- The chamber was sealed with holder and carbon fibre inside

- Once the air within the chamber is deposited, the carbon coat was pulsed onto the

sample

Once the sample was coated with carbon, it was placed within the SEM chamber and analyzed.

40

An EDX spectroscopy analysis is simultaneously conducted, whereby a spectrum is plotted and

elements within the sample are detected. Characteristic radiation spikes are plotted, which

correspond to the matching element and can be identified. Due to the machine’s

specifications, lithium was not visible in the EDX, however its associated elements were

identified.

41

4. Results & Discussion

This chapter discusses the findings generated from the outlined tests that were described in

the previous chapters. Primary baseline tests were initially conducted to create a reference

point that was later averaged and compared against other tests. For more detailed figures,

such as weights and ratios, refer to Appendix B.

An off-site laboratory conducted ICP-MS tests on sub-samples of the feed, froth concentrate

and tails of each flotation test to identify the lithium grade. The lithium recovery of each test

was conducted on a mass basis, which can be seen in Appendix B. The starting mass,

concentrate mass and tails mass were recorded after each test. Using the corresponding grade

of each mass sample, the lithium recovery was determined within the feed, froth concentrate

and tails.

4.1. Baseline Tests

Baseline tests were conducted in order to obtain a reference point to compare with all other

flotation tests. This reference point represents a typical spodumene flotation test, and all

other tests conducted were altered from this point. The tests were conducted at a pH of 10

with 150 g/t of NaOL collector, and a P80 size of 75 microns. Figure 4.1 highlights the lithium

grades and recoveries of the baseline tests.

It should be noted that due to equipment issues, there were some samples lost which resulted

in mass recovery data not equating correctly in Appendix B1. Due to experimental limitations,

such as time, no repeat tests were conducted. Refer to section 5 for other limitations that

affected the project.

42

Figure 4.1 – Baseline Test Grade & Recovery

Two baseline tests were conducted to provide a reference point. Both tests produced similar

grades of approximately 1.2% lithium, while recoveries differed slightly. In the first test, a

recovery of 14% was achieved while in the second test a recovery of 10% was achieved. It

should be noted that approximately 1 gram of sample was lost in both tests due to equipment

issues, which may in turn have affected the lithium recoveries of the baseline tests. Refer to

Appendix B for mass balance data.

A kinetic test was also conducted to obtain information on the residence time of the lithium

concentrate within the system. The kinetic test was conducted at baseline conditions, however

after each minute of float time the concentrate was collected and processed separately. As

time progressed, the grade decreased and reached a plateau at approximately 0.8% lithium

after 10 minutes of flotation. The decrease in grade may be attributed to the entrainment of

gangue minerals within the froth concentrate, which lowers grade. The recovery of lithium

reached a plateau after approximately 4 minutes of flotation time. This indicates that after 4

minutes, majority of the floatable lithium minerals have already reported to the froth and

flotation can cease. Figures 4.2 and 4.3 illustrate the lithium grades and recoveries over time.

0

2

4

6

8

10

12

14

16

0

0.2

0.4

0.6

0.8

1

1.2

1.4

BASE 1 BASE 2

Li R

ecov

ery

(%)

Li G

rade

(%)


43

Figure 4.2 - Baseline kinetic test on Grade

Figure 4.3 – Baseline kinetic test on Recovery

4.2. Varying types of Collector

The primary objective of this test was to investigate the effects of different collector types.

Both STAC and DTAC collectors were tested against the selected baseline collector, NaOL. 150

g/t of each collector was used in separate flotation tests at a pH of 10 and a P80 75 microns.

The STAC collector reported an approximate grade of 0.84% lithium and a recovery of 4.2%,

while the DTAC reported a slightly higher grade and recovery of approximately 0.86% and

0.0

0.2

0.4

0.6

0.8

1.0

1.2

1.4

1.6

1.8

0 1 2 3 4 5 6 7 8 9 10

Li G

rade

(%)

Time (min)

0

0.5

1

1.5

2

2.5

3

3.5

0 1 2 3 4 5 6 7 8 9 10

Li R

ecov

ery

(%)

Time (min)

44

5.14% respectively. Comparison between the three collectors can be seen in Figure 4.4 below.

When compared to the baseline collector, NaOL, it was evident that the two collectors tested

were less effective in recovery of lithium. The NaOL collector was able to achieve a grade of

1.22% lithium and a recovery of approximately 11.4%. A lithium mass balance of the system

can be seen in Appendix B2, which indicates a mass difference of approximately 0.02 grams

between the starting mass and combined mass of the froth concentrate and tails.

Figure 4.4 – Varying collectors (*indicates baseline collector)

Having similar active Al sites for interactions with anionic collectors, flotation separation

between spodumene and its gangue minerals has proven to be quite difficult (Rai, et al. 2011).

It has been reported that the recovery of spodumene by flotation can be accomplished by both

cationic and anionic collectors (Menendex, et al. 2004). Both STAC and DTAC are amine

cationic collectors, while NaOL is an anionic, fatty acid collector. From previous spodumene

flotation studies conducted by Mason, Banks, & Phillip (1953), it was noted that amine

collectors are able to float off impurities and recover the spodumene within the tails residue.

This supports the low recovery values achieved by the two amine collectors when compared to

the baseline NaOL collector.

0

2

4

6

8

10

12

0.0

0.2

0.4

0.6

0.8

1.0

1.2

1.4

STAC DTAC NaOL*

Li R

ecov

ery

(%)

Li G

rade

(%)


45

4.3. Effect of Collector Dosage

Various dosages of the baseline collector, NaOL, were tested on the spodumene ore to

determine an optimum dosage that would generate the highest recovery of lithium. The

collector dosage ranged from 50 g/t to 200 g/t, while all other flotation conditions were kept

at baseline conditions.

Figure 4.5 highlights the varying collector dosages at baseline conditions. As evident from the

figure below, the lithium grades slightly decrease with decreasing dosage. A dosage of 200 g/t

was able to achieve a grade of approximately 1.28% lithium, while baseline dosage achieved a

lower grade of 1.21% lithium. This trend continued between dosages of 100 g/t and 50 g/t,

with attained grades of 1.15% lithium and 1.03% lithium, respectively.

A similar trend was evident in terms of lithium recovery. At baseline dosage, 11.4% of lithium

was recovered, while 100 g/t of NaOL collector was able to recover 6.8% and 50 g/t was able

to recover 6.4%. However, this trend does not follow suit between the dosages of 200 g/t and

baseline dosage of 150 g/t. From Figure 4.5, it can be seen that at 200 g/t of collector, the

recovery was achieved at 7.8%. The recovery then increased by 3.7% once at baseline dosage,

however continued to decrease once at a lower dosage.

46

Figure 4.5 – Varying Collector Dosage (*indicates baseline dosage)

The decrease in recovery may be due to slimes present within the system. Slimes typically

interfere with selective flotation as well as consume reagents (Michaud 2017). These slimes

could have possibly consumed a large amount of the collector, resulting in lower recoveries.

Low dosages may also be a contributing factor in the overall low recoveries attained by each

test. When conducting flotation experiments with approximately 0.5 kilograms at 50-60%

solids, collector dosages range between 250 – 1000 g/t in order to achieve optimum

recoveries. This is evident in batch tests conducted by Xu, Hu and Tian, et al. (2016), where

approximately 500 gram spodumene samples were floated with 800 g/t of collector which

produced significantly higher recoveries.

4.4. Activator Usage

The aim of this test was to investigate the influence of a ferric activator on the flotation of

spodumene. Approximately 2 x 10-4 molar Iron(III) Chloride Hexahydrate was added to the cell

prior to flotation, while at baseline conditions.

It was found that the presence of the ferric activator had a significant impact on the recovery

of lithium. Figure 4.6 highlights the grades and recoveries of the tests which were conducted in

the presence and subsequent absence of the activator, respectively. From the figure below, it

0

2

4

6

8

10

12

0

0.2

0.4

0.6

0.8

1

1.2

1.4

200 g/t 150 g/t* 100 g/t 50 g/t

Li R

ecov

ery

(%)

Li G

rade

(%)


47

can be seen that the lithium grade between the two conditions are similar. The activated test

attained a grade of 1.25%, while the baseline test attained 1.22%. However, the difference

between the two is evident when comparing lithium recoveries. In the presence of the ferric

activator, lithium recovery was approximately 9.8%, while the absence of the activator yielded

a recovery of 11.4%.

Figure 4.6 – Activator Influence (*indicates baseline condition)

The presence of the ferric activator was to reduce the negative charge on the spodumene

surface, which supports a mechanism where the Fe3+ adsorbs onto the mineral surfaces

resulting in an expected increase in recovery (Jie, et al. 2014). Although as previously

discussed, this is not the case. Reports by Xu, Hu and Tian, et al. (2016) mention that ion

activation by Fe3+ results in poor selective flotation of spodumene as the gangue minerals are

simultaneously activated. This increases their flotation by promoting collector adsorption

(Ejtemaeia, Irannajad and Gharabaghi 2012). In this case, the ferric activator has seemingly

activated more gangue than the lithium, resulting in a lower recovery.

From this test, it was evident that the presence of the ferric activator had a negative impact on

the flotation of spodumene by reducing the lithium recovery.

9

9.5

10

10.5

11

11.5

12

0

0.2

0.4

0.6

0.8

1

1.2

1.4

Fe(III) No Fe(III)*

Li R

ecov

ery

(%)

Li G

rade

(%)


48

4.5. Feed Size Alteration

The objective of this test was to determine the most effective feed size that would attain the

highest recovery of lithium. Feed sizes that ranged from 212 microns to 53 microns were

tested at baseline conditions to explore the effect of particle size on flotation. Figure 4.7

illustrates the grades and recoveries achieved by each size fraction.

The grade between the 212 microns and 106 microns slightly decreased, which can be seen in

Figure 5.7. A drop of approximately 0.15% was evident between the sizes, however there was

a minor increase in grade of 0.11% between 106 microns and the baseline size of 75 microns.

This trend continued as the size fraction decreased, with an increase of 0.34% lithium between

75 microns and 53 microns.

When comparing lithium recoveries between the size fractions, there appears to be an

inversely proportional trend. As evident in Figure 5.7, the decrease in particle size led to an

increase in lithium recovery. A recovery of approximately 1.9% was obtained at a size of 212

microns, while recoveries of 5.2% and 11.4% were achieved at feed sizes of 106 microns and

75 microns, respectively. The particle size of 53 microns obtained the highest recovery when

compared to the other sizes, reaching approximately 12.5%. Appendix B5 highlights the mass

balance data of the test. It can be seen that there was a loss of approximately 8 grams of

sample with the 212 micron test. A possible reason behind the loss may be attributable to

faulty equipment, specifically the batch filter press. During the test, it was evident that there

were some leaks in the filter press which resulted in sample being drained out.

49

Figure 4.7 – Varying feed size (*indicates baseline feed size)

With NaOL being an anionic collector, it is noted that the collector is not adsorbed on the

negatively charged spodumene by physical adsorption (Chernyshova, Ponnurangam and

Somasundaran 2011). Instead, the anionic NaOL collector is adsorbed by chemisorption onto

the spodumene surface. The main contributor of the chemisorption onto NaOL is the active Al

sites on the spodumene surface (Xu, Hu and Wu, et al. 2016). The data above suggests that

there are more active Al sites within the finer particles for the collector to adsorb onto,

resulting in a higher lithium recovery. This finding indicates that finer particle sizes of 53

microns allow for improved liberation of spodumene particles within the flotation system,

increasing their interaction with collecting species and subsequently improving lithium

recovery, The results indicate that particle size does have an impact on the flotation of

spodumene. From the tests conducted it can be seen that a finer particle size yields a higher

recovery in lithium. The trend outlined above is supported by studies conducted by Xu, Hu and

Wu, et al. (2016) where the decrease in particle size led to an increase in lithium recovery.

4.6. pH Variation

To determine the ideal pH of the flotation system, multiple pHs were tested with the aim of

obtaining the highest recovery. pHs of 6, 8 and 12 were all tested against the baseline pH of 10

0

2

4

6

8

10

12

14

0

0.2

0.4

0.6

0.8

1

1.2

1.4

1.6

1.8

212 um 106 um 75 um* 53 um

Li R

ecov

ery

(%)

Li G

rade

(%)


50

whilst all other parameters were kept at baseline conditions.

From Figure 4.8, it is evident that the grades between pH 8, 10 and 12 did not change

significantly. At pH 8 and 10, the approximate grade was 1.2% while at pH 12 the grade was

1.1%. At a pH of 6 the grade slightly decreased to 0.9%.

It is clear that when comparing recoveries with varying pH there is a relationship, to an extent.

As the pH increases from 6 to 10, there is a decrease in lithium recovery. As seen in Figure 5.8,

the recovery at pH 6 is approximately 26.3% while at pH 8 and pH 10, the recovery reaches

11.4%. However, this trend then disappears between pH 10 and pH 12, where the recovery

drastically increased to approximately 56%.

Looking at the varying pH tests, it is apparent that at the lower pH range a lower recovery was

attained. However, when flotation is being conducted at a high pH range, a high recovery is

achieved.

HCl was added into the system to attain a pH of 6, while NaOH was used to increase the pH to

10 and 12. The system’s pH was naturally at 8, so no regulators were added. From Figure 4.8,

it is clear that the changes in recovery are evident within the tests where a regulator was used.

It can be said that pH regulators do have a significant impact on the flotation of spodumene.

51

Figure 4.8 – Varying pH (*indicates baseline pH)

4.7. Optimized Conditions

After all flotation tests were completed, the altered conditions resulting in the highest

recovery were selected in order to conduct an optimized flotation test on spodumene. The test

was operated at a pH of 12 with a particle size of 53 microns. The collector NaOL was selected

at 150 g/t, without the presence of a ferric activator. Figure 5.9 illustrates the comparison

between the optimized and the baseline flotation recoveries and grades.

As illustrated in the plot of Figure 4.9 below, there is a notable difference in both grade and

recovery when operating at optimized conditions. At optimized conditions, the lithium grade

and recovery was recorded at 1.6% and 40.2%, respectively. While grade only slightly

increased from the baseline conditions (1.2% at baseline), the recovery increased significantly

by approximately 38% from baseline conditions.

With the flotation conditions outlined, it is evident that the lithium grade and recovery can be

increased. From the optimized test conducted, it can be noted that to attain a high recovery a

small feed size, in combination with a high pH, is required.

0

10

20

30

40

50

60

0

0.2

0.4

0.6

0.8

1

1.2

1.4

pH 6 pH 8 pH 10* pH 12

Li R

ecov

ery

(%)

Li G

rade

(%)


52

Figure 4.9 – Spodumene flotation at optimized conditions

To further study the optimized conditions, SEM and EDX analysis were undertaken. This

allowed for visual representations of the flotation concentrate and tails, as well as elemental

analysis within each sample.

The SEM image of the optimized flotation concentrate can be seen in Figure 4.10. From the

image below, it can be noted that the particle size appears to be sufficient as there is no visible

entrainment of particles. This suggests that at 53 microns, the particles are well dispersed and

broken apart from any gangue minerals, making it acceptable for flotation.

0

5

10

15

20

25

30

35

40

45

0

0.2

0.4

0.6

0.8

1

1.2

1.4

1.6

1.8

Optimized Baseline

Li R

ecov

ery

(%)

Li G

rade

(%)


53

Figure 4.10 – SEM of optimized flotation concentrate at 100 microns

To identify the specific elements within the optimized concentrate, EDX analysis was

conducted. It was found that the concentrate primarily contained silicon and aluminium, with

some particles containing small amounts of potassium, iron and calcium. The particles

containing silicon and aluminium resembled that of spodumene particles, as silicon and

aluminum are the associated elements within the lithium bearing mineral. The potassium and

calcium bearing particles suggest that K-feldspar exists within the concentrate. Refer to

Appendix C to see specific points identified and the associated elemental plot.

The SEM image of the optimized flotation tails can be seen in Figure 4.11. Similar to the

previous figure, the particle size appears to be sufficient as there are no visible entrainment of

particles. Referring to Figure 5.11, the particles are broken up and disseminated throughout

the image.

After EDX analysis was conducted on the flotation tails SEM image, it was evident that silicon

54

and aluminium were the dominant elements within each of the selected particles. It was also

clear that minor potassium was apparent within some particles. As discussed earlier, the

aluminium and silicon combination suggests that these particles are spodumene, while the

potassium bearing particles appear to be K-feldspar. Appendix C highlights the selected

particles associated with the EDX plot.

Figure 4.11 – SEM image of optimized flotation tails at 100 microns

4.8. Lithium Bearing Ore Comparison

To gain an understanding in the flotation behaviours within lithium bearing minerals, a

flotation test was conducted on two different lithium ores. Both spodumene and lepidolite

were tested at a pH of 10 with a feed size of 106 microns. The collector utilized was NaOL at

150 g/t and no activator was used. Figure 4.12 highlights the lithium grades and recoveries of

each mineral.

55

Figure 4.12 – Spodumene against Lepidolite at the same conditions

As illustrated in the above figure, both spodumene and lepidolite share similar grades and

recoveries to each other when floated at the same conditions. Spodumene was able to attain a

grade and recovery of approximately 1.09% and 5.21%, respectively. Similar grades and

recoveries were reported for the lepidolite. Referring to Figure 4.12, a grade of approximately

1.45% was achieved for the lepidolite, with a lithium recovery of 5.77%.

The low recovery values may be attributed to the feed size. At 106 microns, the particles are

not fully liberated from the gangue particles. The gangue particles surround the valuable

lithium, making it harder for effective flotation to take place. This would result in less lithium

being collected in the froth, resulting in lower recoveries.

From this comparison, it can be seen that these specific flotation conditions slightly favour

lepidolite recovery rather than spodumene. Overall these results suggest that both of the

lithium bearing minerals have similar flotation responses under the tested conditions.

0

1

2

3

4

5

6

7

0.000.200.400.600.801.001.201.401.601.802.00

Spodumene Lepidolite

Li R

ecov

ery

(%)

Li G

rade

(%)


56

5. Limitations of Research Project

Two factors which impacted the results of this study were limitations of equipment and

constraints of time.

5.1. Equipment

Traces of impurities were evident within samples. This was unavoidable due to faulty

equipment.

The flotation process requires a constant air flow rate in order to collect the valuable minerals

within the system. However, at times the airline connected to the flotation cell was clogged,

which provided sparges of air to the system. This resulted in both concentrate and gangue

minerals overflowing to the concentrate. Random sparges were also apparent, which resulted

in flotation samples being spilt and lost. This created mass balance errors within the data.

Desired particle size could not be achieved for optimized tests on the alternate lithium bearing

minerals due to limitations of equipment. To overcome this, the test was altered to a higher

particle size at baseline conditions.

These limitations of equipment were noted and taken into account in the analysis of results.

5.2. Time

Power outages that occurred during the time of the study contributed to the delay of

completion.

During the assaying stage, toxic hydrofluoric gas was involved in the dissolution process

required for each of the samples. Due to safety concerns, samples were sent out for analysis

during the assaying stage to an external site. This resulted in the delay of data analysis, which

meant tests could not be analysed or altered to ensure correct flotation conditions were met

57

and to improve testing.

Taking into account these limitations, the results achieved were satisfactory.

58

6. Conclusions

Empirical studies have shown that there are a wide range of variables that greatly impact the

flotation of lithium. Parameters such as collector dosage and type, activator usage, feed size

and pH level have been reported to have dramatically influenced the flotation recoveries of

lithium.

When varying different collector types, it was evident that both the amine cationic collectors

recovered less lithium when compared to the baseline fatty acid anionic collector. STAC and

DTAC recovered approximately 4.2% and 5.14% lithium, respectively, while NaOL recovered

11.4% lithium. This suggests that the anionic collector is more favourable when compared to

cationic collectors and should be utilized.

Varying collector dosage from baseline conditions highlighted a drop in both grade and

recovery of lithium. With respect to lithium grade, it was clear that as the dosage decreased,

the grade also decreased. Grades ranging from 1.28% to 1.03% was achieved with dosages

starting at 200 g/t to 50 g/t. A similar trend is evident in terms of lithium recovery, where a

decrease in dosage results in a decrease in recovery. However, between the dosages of 200 g/t

and 150 g/t, the recoveries increased from 7.8% to 11.4%. A possible reason for the drop in

recovery from 150 g/t to 200 g/t may be attributed to slimes present in the system. If more

collector is within the flotation system, more of it would be consumed by the slimes which

would lead to less lithium being recovered.

The presence of a ferric activator within the system proved to have a negative impact on the

flotation of lithium. A recovery of approximately 9.8% was achieved with a ferric activator

being utilized, which was 1.6% lower than baseline conditions. It was found that the activator

not only activated the valuable spodumene, but also its associated gangue minerals. This

resulted in poor selectivity of the spodumene and achieved a lower recovery.

59

The alteration of particle size highlighted significant impacts upon lithium recovery. It was

clear that the decrease in particle size leads to an increase in lithium recovery. The recoveries

increased between 212 microns to 75 microns, ranging from approximately 1.9% to 11.4%

lithium. The particle size of 53 microns obtained the highest recovery when compared to the

other sizes, reaching approximately 12.5%. The presence of active Al sites on the spodumene

mineral was the suggested reason as to why the recoveries increased with decreasing size. As

the particles get broken up further, it allows for more Al sites to be apparent and aid in NaOL

chemisorption, resulting in greater recovery.

One of the major influences of spodumene flotation was the pH. It was found that the

recovery decreases from pH 6 to pH 10 by approximately 14.9%. The recovery then increases

dramatically from 11.4% at pH 10 to 56% at pH 12. The presence of pH regulators has had a

significant impact on the flotation of spodumene. Both HCl and NaOH have increased the

recoveries of lithium, in comparison with the absence of these regulators.

Optimized tests showed that a combination of a feed size of 53 microns, a pH of 12, 150 g/t of

NaOL without any ferric iron can attain recoveries approximately 30% higher than baseline

recoveries. This suggests that a fine particle size as well as a high pH should be set parameters

to attain optimal recoveries of spodumene.

It was found that when both spodumene and lepidolite are floated at the same conditions,

both minerals were able to attain a similar grade and recovery. Spodumene was able to attain

a grade and recovery of approximately 1.09% and 5.21%, respectively. Lepidolite achieved a

grade of approximately 1.45% and recover 5.77% lithium. By attaining similar grades and

recoveries, it can be said that both minerals share similar flotation responses under the given

conditions.

60

7. Future Work & Recommendations

The investigation into the beneficiation of hard rock lithium minerals by flotation has

presented conditions for optimal recoveries. These conditions can be taken into account when

optimizing flotation recoveries and implementing them on a larger scale. The investigation has

also uncovered conditions that negatively impact the flotation and hinder lithium recovery.

Further research on the factors that hindered recovery can be explored to improve product

qualities.

It is recommended that a deslime and an increase in collector dosage should be applied to

further work on lithium flotation. The deslime would potentially reduce the consumption of

reagents by the ultrafine particles and could increase the recovery of lithium.

A baseline collector dosage of approximately 800 g/t – 1200 g/t could significantly increase the

lithium recovery and establish a firmer baseline to alter from.

To further understand the particle interactions between the mineral surface and reagents, a

zeta potential test may be conducted. This test would reveal at what pH the surface charge is

0, and could further suggest which collector types will be effective.

62

References

911Metallurgist. 2017. Froth Flotation Spodumene Processing Lithium Extraction.

https://www.911metallurgist.com/blog/froth-flotation-spodumene-processing-

lithium-extraction.

Arbiter, N, J.R Abshiev, and J.W Crawford. 1961. "Attritioning and conditioning in flotation of

spodumene ore quartz." 321-322.

Birney, M. 2017. Lithium Australia extracts 99% Lithium form Mexican clays. May 25.

http://www.businessnews.com.au.

—. 2017. Lithium Australia's Sileach process wins approval of new German partner. May 25.

http://www.businessnews.com.

Browning, J.S, and T.L McVay. 1961. Beneficiating spodumene from pegmatites. Gatson

County: USBM .

Bulatovic, S. 2014. Handbook of Flotation Reagents: Chemistry, Theory and Practise.

—. 2007. Handbook of Flotation Reagents: Chemistry, Theory and Practise.

Chernyshova, I.V., P Ponnurangam, and P Somasundaran. 2011. "Adsorption of fatty acids on

iron (hydr) oxides from aqueous solutions." Langmuir 27 10007-10018.

Choi, J, J Hong, K Park, Y Han, W Kim, and B Kim. 2015. "Lepidolite floration from low grade

ores using cationic surfactant."

Ejtemaeia, M, M Irannajad, and M Gharabaghi. 2012. "Role of dissolved mineral species in

selective flotation of smithsonite from quartz using oleate as a collector." Int. J. Miner.

Process 114-117, 40-47.

63

Hoey, T. 2017. Lithium Australia enters Europe and reveals promising exploration results near

Mt Cattli. http://www.finfeed.com.

Jessup, T. 2000. United States of America Patent 6.

Jie, Z, W Weiging, L Jing, H Yang, F Qiming, and Z Hong. 2014. "Fe(III) as an activator for the

flotation of spodumene, albite and quartz minerals."

Kelly, E.G, and D.J Spottiswood. 1997. Introduction to Mineral Processing.

Lucas, J. 2017. Lithium mines pegged to grow sevenfold as Chinese investment propels WA

boom. http://www.abc.net.au/news/2017-02-08/chinese-investment-driving-wa-

lithium-boom/8252068.

Mason, L, William T. Banks, and N Phillip. 1953. "A method for concentration of North Carolina

spodumene ores." Miner Eng. 2 181-186.

Menendex, M, A Vidal, J Torano, and M Gent. 2004. "Optimization of spodumene flotation."

Eur. J. Miner. Process. Environ. Protect. 130-135.

Moon, K.S. 1986. "Surface and crystal chemistry of spodumene and its flotation behaviour."

Diss. Abstr. Int. 46 277.

Moon, K.S, and D.W Fuerstenau. 2003. "Surface crystal chemistry in selective flotation of

spodumene ([LiAl[SiO3]2) from other aluminosilicates." Int. J. Miner. Process. 11-24.

Rai, B., P Satish, J Tanwar, Pradip, K.S. Moon, and D.W. Fuersteanau. 2011. "A molecular

dynamics study of the interaction of oleate and dodecylammonium chloride

surfactants with complex aluminosilicate minerals." J. Colloid Interface Sci. 362 510-

516.

Review, Australian Financial. 2016. Why Australia will be at the centre of Lithium Boom.

64

http://www.afr.com.

Vieceli, N, F.O Durao, C Guimares, C A Nogueira, M F C Pereira, and Margarido. 2016. "Grade-

Recovery modelling and optimization of the froth flotation process of a lepidolite ore."

Williams, P. 2017. Biggest lithium mine doubles with Greenbushes expansion.

https://thewest.com.au/business/lithium/biggest-lithium-mine-doubles-with-

greenbushes-expansion-ng-b88417791z.

Wills, B A, and T N Munn. 2006. Will's Mineral Processing Technology, Seventh ed.

Xu, L, Y Hu, H Wu, J Tian, J Liu, Z Gao, and L Wang. 2016. "Surface crystal chemistry of

Spodumene with different size fractions and implications for flotation."

Xu, L, Y Hu, J Tian, H Wu, Y Yang, X Zeng, Z Wang, and J Wang. 2016. "Selective Flotation

separation of Spodumene from Feldspar using new mized anionic/cationic collectors."

Zhu, G, Y Wang, X Liu, F Yu, and D Lu. 2015. "The Cleavage and Surface Properties of wet and

dry ground Spodumene and their flotation behaviour."

66

Appendix

The following section details specific processes as well as raw data that have been previously

mentioned.

Appendix A

The section highlights the step by step methodology taken to ensure results would be

consistent and accurate.

Sample Preparation was broken up into crushing and homogenizing. Crushing was operated in

a cone crusher, while homogenizing made use of a rotary splitter.

Crushing:

1. Sample will be pre-screen over 1 millimeter screen

2. Oversize sample will be weighed

3. Oversize will be passed through crusher, and screen crusher product

4. Oversize sample will be weighed

5. Steps 1. To 4. will be repeated until P100 1 millimeter

6. -1 millimeter sample will be bagged up (labelled as A1) and mass will be obtained

Homogenizing:

1. Sample collected from 1A

2. The sample will be passed through Rotary Splitter

3. After the pass, the sample will be recombined and passed through splitter two more

times

4. 1 kilogram charges will be made up from homogenized sample and the remainder will

be reserved (labelled as 1B)

Grind Establishments were conducted to attain specific grind times for desired feed sizes.

67

1. Collect 1 kilogram charge from 1B

2. Sample will be placed in Rod mill with 1 kilogram of tap water (50% solids)

3. Grinding will be conducted in Rod Mill for desired time

4. The sample will be washed out of mill into 20L bucket

5. Small portion of sample will be poured over vibrating wet screen of size 38 microns

6. Sample will be washed over screen ensure all fines pass through into separate bucket

underneath and collect coarse on screen and wash into aluminium tray

7. Repeat step 5 & 6. until all sample is passed through 38 micron screen

8. Fines will be transferred into bucket and then to batch filter press.

9. Sample will be filtered and dried in oven

10. Coarse tray will be placed in oven to dry

11. Once coarse sample is dry, sample will be placed over RoTap machine and dry

screened

12. Size fractions will be weighed up and bagged

13. Grind times will be determined from mass percentages relating to size

The flotation steps outlined was used for each test, however multiple steps were altered to

match the desired test parameters. Step 2 was altered to correspond with desired feed size.

Steps 5 to 7 were altered to correspond with their corresponding variability tests (eg. Step 5

was altered to pH 6, 8 and 12).

1. Initial charge will be obtained from 1B

2. The sample will be ground from times obtained from establishments

3. Sample will be transferred into 3L float cell and made up to desired level with

deionized water (50% solids)

4. Impeller will be switched on to ensure proper agitation of sample without air

5. The pH will be adjusted using NaOH regulator and brought up to 9-10

68

6. The desired collector at desired dosage will be added into cell and conditioned for 5

minutes

7. After collector conditioning, activator may be added if required for test (0.2 millimoles

of ferric is added and conditioned for 2 minutes)

8. After the collector conditioning, Polyfroth W22 will be added and conditioned for 2

minutes

9. Once conditioning is done, airflow is added into the cell and froth is scrapped at a rate

of 1 scrape every 6 seconds

10. Froth will be collected into aluminium trays for approximately 10 minutes

11. Once all valuable froth is collected, it will be vacuum filtered and dried in oven

12. The cell tails will be transferred into batch filter press, filtered and dried in oven

Appendix B

This section details the raw data obtained from each flotation test. The data highlights the

specific lithium grades and recoveries obtained from each test.

69

Table B1 – Baseline & Kinetic Tests

Mass (g) Ratio (%) Li Grade (%) Li Mass Rec (g) Li Recovery (%) Baseline 1 Con 65.87 13.20 1.25 0.82 13.63

Tail 433.28 86.80 0.99 4.31 71.31 Feed 499.15 100 1.21 6.05 100

Baseline 2 Con 47.08 9.43 1.19 0.56 9.20 Tail 455.62 91.28 1.00 4.55 74.81

Feed 502.70 100 1.21 6.08 100.00 Kinetic Con 1 11.224 2.242 1.62 0.182 3.004

Con 2 7.584 1.515 1.56 0.118 1.948 Con 3 3.928 0.785 1.47 0.058 0.955 Con 4 1.79 0.358 1.39 0.025 0.410 Con 5 1.712 0.342 1.24 0.021 0.352 Con 6 1.831 0.366 0.99 0.018 0.299 Con 7 1.115 0.223 1.01 0.011 0.186 Con 8 2.018 0.403 0.86 0.017 0.286 Con 9 1.726 0.345 0.84 0.015 0.240

Con 10 1.848 0.369 0.82 0.015 0.249 Tails 10 465.82 93.053 1.23 5.734 94.667

Feed 500.6 100 1.21 6.057 100.000 Baseline Con Rec. Av 10.254

Table B2 – Varying Collector Type

Mass (g) Ratio (%) Li Grade (%) Li Mass Rec (g) Li Recovery (%) STAC Con 30.24 6.05 0.84 0.25 4.20

Tail 469.76 93.95 1.24 5.83 96.12 Feed 500 100.00 1.21 6.06 100.00

DTAC Con 36.12 7.21 0.86 0.31 5.13 Tail 464.84 92.79 1.25 5.79 95.32

Feed 500.96 100.00 1.21 6.07 100.00

70

Table B3 - Varying Collector Dosage

Mass (g) Ratio (%) Li Grade (%) Li Mass Rec (g) Li Recovery (%)

200 g/t Con 37.15 7.37 1.27 0.47 7.76

Tail 466.74 92.63 1.16 5.40 88.54

Feed 503.89 100 1.21 6.10 100.00

100 g/t Con 35.68 7.09 1.16 0.41 6.77 Tail 467.78 92.91 1.22 5.71 93.81

Feed 503.46 100.00 1.21 6.09 100.00

50 g/t Con 37.73 7.48 1.03 0.39 6.38

Tail 466.42 92.52 1.26 5.87 96.25

Feed 504.15 100.00 1.21 6.10 100.00

Table B4 – Activator usage

Mass (g) Ratio (%) Li Grade (%) Li Mass Rec (g) Li Recovery (%) Fe(III) Con 47.74 9.58 1.24 0.59 9.85

Tail 450.81 90.42 1.18 5.34 88.52 Feed 498.55 100.00 1.21 6.03 100.00

Table B5 – Varying Feed Size

Mass (g) Ratio (%) Li Grade (%) Li Mass Rec (g) Li Recovery (%) 212um Con 9.42 1.88 1.24 0.12 1.94

Tail 490.58 98.12 1.18 5.81 96.05 Feed 500.00 100.00 1.21 6.05 100.00

106um Con 29.01 5.80 1.09 0.32 5.21 Tail 471.20 94.20 1.23 5.80 95.83

Feed 500.21 100.00 1.21 6.05 100.00 53um Con 48.55 9.71 1.56 0.76 12.52

Tail 451.52 90.29 1.18 5.35 88.39 Feed 500.07 100.00 1.21 6.05 100.00

71

Table B6 – Varying pH

Mass (g) Ratio (%) Li Grade (%) Li Mass Rec (g) Li Recovery (%) pH 6 Con 162.13 32.43 0.98 1.59 26.27

Tail 337.87 67.57 1.30 4.38 72.38 Feed 500 100 1.21 6.05 100

pH 8 Con 56.52 11.30 1.22 0.69 11.37 Tail 443.48 88.70 1.21 5.36 88.53

Feed 500.00 100.00 1.21 6.05 100.00

pH 12 Con 289.85 57.97 1.17 3.38 55.86 Tail 210.15 42.03 1.27 2.67 44.21

Feed 500.00 100.00 1.21 6.05 100.00

Table B7 – Optimized conditions


OP 1 Con 155.72 31.14 1.58 2.45 40.23 Tail 330.79 66.16 1.05 3.47 56.97

Feed 500.00 100 1.21 6.05 100.00

Table B8 – Lithium Bearing Minerals Comparison


SPOD Con 29.01 5.80 1.09 0.32 5.21 Tail 471.2 94.20 1.23 5.80 95.83

Feed 500.21 100.00 1.21 6.05 100.00 LEP Con 28.71 5.73 1.45 0.42 5.77

Tail 472.31 94.27 1.45 6.85 94.92 Feed 501.02 100.00 1.44 7.21 100.00

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Mass In (g) Mass Out (g)

Li in Feed 6.05 Li in Con 0.82 Li in Tail 4.31

Total 6.05 5.13

Variability (%) 15.23

Appendix C

This section details the Energy-Dispersive X-ray Spectroscopy data collected from the JCM-

6000 JEOL SEM machine.

Figure C1 highlights the selected points used for EDX analysis within the optimized flotation

concentrate, while Figure C2 illustrates the elemental spectroscopy of the image. An area scan

of the SEM image was also conducted to provide a general overview as to what elements were

within the system. As evident in Figure C1, there were 6 spots selected for analysis. However,

upon inspection of the EDX data it was apparent that Spot 2 had extremely high amounts of

iron, nickel and chromium. This suggests that small amounts of steel grinding media had

entered the system and contaminated the concentrate. Because of this, Spot 2 data was

removed from the EDX Spectroscopy analysis.

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Figure C1 – Points from SEM image of optimized concentrate for EDX Spectroscopy

Figure C2 – EDX Spectroscopy data for optimized concentrate

The optimized flotation tails SEM image and EDX plots can also be seen in Figures C3 and C4,

respectively. Spot 1 was removed from the data set as it was producing irregular results that

would have a very low probability of being accurate.

0 1 2 3 4 5 6 7

X-ray energy (keV)

Spot 1

Spot 3

Spot 4

Spot 5

Spot 6

Areascan

Fe K Al

Ca

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Figure C3 – Points from SEM image of optimized tails for EDX Spectroscopy

Figure C4 – EDX Spectroscopy data for optimized tails

0 1 2 3 4 5 6 7

X-ray energy (keV)

Spot 2

Spot 3

Spot 4

Spot 5

Spot 6

Spot 7

Areascan

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