physical separation of thermally upgraded pyrrhotite

Physical Separation of Thermally Upgraded Pyrrhotite

by

Jaspreet Sandhu

A thesis submitted in conformity with the requirements for the degree of Masters of Applied Science

Graduate Department of Chemical Engineering University of Toronto

© Copyright by Jaspreet Sandhu 2019

ii

Physical Separation of Thermally Upgraded Pyrrhotite

Jaspreet Sandhu

Masters of Applied Science

Department of Chemical Engineering

University of Toronto

2019

Abstract

Over the past 40 years, the depression of pyrrhotite during the flotation of nickel sulphide

ores has led to a significant build-up of pyrrhotite tailings. These tailings are stored in

tailings ponds, where they have the potential to generate acid rock drainage. These

tailings contain up to 1% nickel, are easily accessible, and do not require grinding. A

pyrometallurgical method of nickel extraction is being developed which involves

converting pyrrhotite (Fe1-xS) to a ferronickel alloy (FeNi) and troilite (FeS). This thesis

focuses on physically separating the FeNi and FeS phases, by either froth flotation or

magnetic separation. The variables studied for froth flotation are pH, particle size, and

additive concentrations. Moreover, magnetic separation characteristics studied are

magnetic field intensity, particle size, and the production temperature of FeNi and FeS.

Results showed that magnetic separation produced concentrates of higher Ni grade and

Ni recovery compared to froth flotation.

iii

Acknowledgements

I would like to thank my supervisors Dr. Erin Bobicki and Dr. Mansoor Barati. Next, I would

like to thank Feng Liu and John Forster for all their help in the lab.

Next, I would like to thank Vale Mississauga, especially Andy Lee and Jie Dong for all

their help with analytical services and use of their facilities.

NSERC is thanked for providing the funding for this research.

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Contents

Abstract ............................................................................................................................ii

Acknowledgements ......................................................................................................... iii

Table of Figures ..............................................................................................................vi

Table of Tables .............................................................................................................. vii

Glossary ..........................................................................................................................ix

1 Introduction ............................................................................................................... 1

2 Literature Review ..................................................................................................... 3

2.1 Pyrrhotite Properties ........................................................................................... 4

2.1.1 Magnetism .................................................................................................... 7

2.1.2 Pyrrhotite Tailings ......................................................................................... 8

2.1.3 Galvanic Interactions .................................................................................. 11

2.1.4 Oxidation of Pyrrhotite ................................................................................ 13

2.1.5 Acid Mine Drainage Potential ...................................................................... 15

2.1.6 Zeta Potential .............................................................................................. 16

2.3 Historical and Current Pyrrhotite Treatment Methods ....................................... 17

2.4 Thermal Upgrading of Nickeliferous Po............................................................. 18

2.5 FeNi Properties ................................................................................................. 23

2.5.1 FeNi Passivation .......................................................................................... 23

2.6 Separation Methods ......................................................................................... 23

2.6.1 Comminution ............................................................................................... 24

2.6.2 Froth Flotation ............................................................................................. 24

2.6.3 Magnetic Separation ................................................................................... 33

2.6.4 Gravity Separation ...................................................................................... 36

3 Separation of Heat Treatment Analogue by Magnetic Separation and Froth Flotation

...................................................................................................................................... 39

3.1 Materials and Methods ...................................................................................... 39

3.1.1 HTA Preparation .......................................................................................... 39

3.1.2 Separation Procedure ................................................................................. 42

3.1.3 Characterization ............................................................................................. 45

3.2 Results ............................................................................................................... 47

v

3.2.1 HTA Magnetic Separation Results ............................................................... 47

3.2.2 HTA Microfloat Results ................................................................................ 49

3.3 Conclusions ...................................................................................................... 51

4 Magnetic Separation and Froth Flotation of Heat Treatment Product Produced at 950

°C with Iron Oxide Addition ........................................................................................... 52


4.1.1 Sample Preparation by Thermal Upgrading at 950°C with Iron Oxide ......... 52

4.1.2 Separation Procedure ................................................................................. 53

4.1.2 Characterization .......................................................................................... 55

4.2 Results ................................................................................................................. 59

4.2.1 Magnetic Separation Results ......................................................................... 59

4.2.2 Microflotation Results .................................................................................... 62

4.3.1 Conclusion ........................................................................................................ 62

5 Magnetic Separation of Heat Treatment Product Produced at 850 °C, 900 °C, and

950 °C with Iron Powder ................................................................................................ 64


5.1.1 Sample Preparation by Thermal Upgrading at Various Temperatures with

Iron Powder ............................................................................................................ 64

5.1.2 Separation Procedure .................................................................................. 65

5.2 Results .............................................................................................................. 65

5.3 Conclusion ........................................................................................................ 71

6 Conclusion and Summary ....................................................................................... 72

7 Recommendations for Future Work ......................................................................... 73

References ................................................................................................................... 74

vi

Table of Figures

Figure 1 a) Phase diagram for the Fe-S system (FeS to FeS2) [19] b) Phase diagram for

the Fe-S system [20] ....................................................................................................... 7

Figure 2: Net magnetic moment and Po’s superstructures. [19] ...................................... 8

Figure 3: Pourbaix diagram of water at 25 °C, 1 atm. ................................................... 12

Figure 4: Pourbaix diagram for the Fe-S-H2O system at 25 °C, 1 atm, 10-6 mol/L Fe and

S. Pyrrhotite is in yellow [4]. .......................................................................................... 12

Figure 5: Movement of nickel during thermal upgrading ............................................... 19

Figure 6: Equilibrium phase diagram of Fe-Ni-S at 900°C. The red arrow indicates the

initial Po concentration (star) and the change of composition during the thermal

upgrading process (γ denotes the ferronickel alloy and mss denotes the sulphide

solution). [45] ................................................................................................................. 20

Figure 7: Our research group’s proposed process to create nickeliferous alloy. ........... 21

Figure 8: Phase relations of the Fe-rich corner in (a) Fe-Ni-S system; (b) Fe-Co-S

system; (c) Fe-Cu-S system at 900 °C calculated by FactSage 6.4 [46] [47] ................ 22

Figure 9: Batch flotation cell .......................................................................................... 25

Figure 10: Davis Tube Tester ........................................................................................ 36

Figure 11: Concentration Criterion vs Specific Gravity of the Fluid Medium .................. 38

Figure 12: Microfloat cell (55 mL) .................................................................................. 42

Figure 13: DTT results on HTA after 1 run .................................................................... 48

Figure 14: DTT results on HTA after 2 runs .................................................................. 48

Figure 15: FeNi concentrate’s Ni Grade and Recovery for -74+45 µm size fraction (each

test is specified by pH level and collector addition (mL)/activator addition (mL)) .......... 50

Figure 16: FeNi concentrate’s Ni Grade and Recovery for -45 µm size fraction (each test

is specified by pH level and collector addition (mL)/activator addition (mL)) ................. 50

Figure 17: XRD results of heat-treated Po at 950°C with iron oxide addition, where 1)

iron oxide, 2) pyrrhotite concentrate, and 3) heat treated product. [45] ......................... 56

Figure 18: SEM image of heat-treated Po at 950°C with iron oxide addition, at a) x100

and b) x500 magnification. The lighter phase is the nickeliferous alloy and the dark gray

is the sulphide phase. .................................................................................................... 57

vii

Figure 19: The nickel concentration in the nickeliferous alloy phase and sulphide phase

from the EPMA analysis. [45] ........................................................................................ 57

Figure 20: SEM image of ground heat-treated Po, at 950°C with iron oxide addition, at

a) x200 b) 500x magnification. The lighter phase is the nickeliferous alloy and the dark

gray is the sulphide phase. ............................................................................................ 58

Figure 21: The grade vs. recovery of Ni in magnetic concentrates produced at 40 mT,

50 mT, and 100 mT ....................................................................................................... 60

Figure 22: FeNi microfloat concentrate’s Ni Grade and Recovery (each test is specified

by pH level and collector addition (mL)/activator addition (mL)) .................................... 62

Figure 23: SEM images of heat-treated product and composition of the metallic phase

(the lighter phase is the nickeliferous alloy and the dark gray is the sulphide phase) ... 67

Figure 24: Ni Grade vs Ni Recovery of magnetic concentrates at temperatures 850 °C,

900 °C, 950 °C .............................................................................................................. 69

Table of Tables

Table 1: Comparison of Glencore Po tailings and Vale Po Tailings from Sudbury. [2] .... 9

Table 2: Nickel distribution in Glencore Po tailings and Vale Po Tailings from Sudbury

[2]. ................................................................................................................................. 10

Table 3: Liberation and exposure of Po and pentlandite in Glencore Po tailings and Vale

Po Tailings from Sudbury [2]. ........................................................................................ 10

Table 4: Rest potential values of common minerals [24] ............................................... 13

Table 5: Physical properties of FeNi and FeS, and the technique used from separation.

...................................................................................................................................... 24

Table 6: Comparison of Pyrrhotite (Po) collectorless flotation in Ni ore deposits [66] [67]

...................................................................................................................................... 29

Table 7: The effects of mixtures of potassium amyl xanthate (PAX), sodium isobutyl

xanthate (SIBX),and isopropyl ethyl thionocarbamate (IPETC) collectors on grade and

recovery in the froth flotation of a nickel sulphide ore. [56] ............................................ 32

Table 8: Dependence on Concentration Criterion for separation [23] ........................... 37

viii

Table 9: Procedure for treatment of Voisey’s Bay Po Tailings ...................................... 40

Table 10: HTA magnetic separation testing variables ................................................... 43

Table 11: Procedure for microflotation of HTA .............................................................. 44

Table 12: HTA froth flotation testing variables ............................................................... 45

Table 13: XRD analysis of Voisey's Bay Tailings .......................................................... 46

Table 14: ICP-OES analysis of Voisey’s Bay Po Tailings (feed) material, Hexagonal Po

and Troilite Concentrate, and HTA at different size fractions ........................................ 46

Table 15: Procedure for microflotation of heat-treated material at 950 °C ..................... 54

Table 16: Testing variables for microflotation of heat-treated material at 950 °C .......... 55

Table 17: Composition of heat-treated material after grinding by ICP-OES analysis. ... 59

Table 18: The mass pull and grade vs. recovery of Ni, Co, and Cu from the magnetic

concentrates produced at 40 mT, 50 mT, and 100 mT.................................................. 61

Table 19: Passing size of initial and magnetic concentrates ......................................... 61

Table 20: Phase composition of heat-treated product at various temperatures (EPMA)66

Table 21: Particle size data of feed and magnetic concentrates produced from material

heat-treated at 850 °C .................................................................................................. 69


heat-treated at 900 °C ................................................................................................... 70


heat-treated at 950 °C ................................................................................................... 70

ix

Glossary

DTT = Davis Tube Tester

FeNi = ferronickel alloy

HTA = heat treatment analogue (described in 3.1.1 )

ICP-OES = Inductively coupled plasma - optical emission spectrometry

Po = pyrrhotite

SEM = Scanning Electron Microscope

XRD = X-ray Diffraction

1

1 Introduction

Over time the nickel grade in Canadian ores has declined; therefore pyrrhotite (Fe1-xS or

Po) is under investigation as a secondary source of nickel. Canadian mineral processing

facilities produce large amounts of waste Po that is stored subaqueously in tailings ponds

to limit their exposure to oxygen and prevent acid generation. Although subaqueous

deposition is best practice for sulphide minerals, tailing ponds require regular

maintenance and pose a long-term liability for mining companies. At the same time, Po

tailings represent a valuable source of nickel. Currently, there are 90-120 Mt Po tailings

stored in Canada, containing up to 1 wt% Ni and are valued at an estimated $4-11 billion

USD. Thus, it is desirable to process pyrrhotite tailings to not only reduce the quantity of

tailings stored, but also to extract the value.

A novel pyrometallurgical process to extract nickel from Po tailings is being developed.

The pyrometallurgical process provides an alternative for a common waste mineral,

attempts made in the past failed due to environmental and economic constraints. Firstly,

the tailings require upgrading to remove silicate minerals. The concentrated Po will be

heat-treated to convert Po to ferronickel (FeNi) and troilite (FeS). FeNi and FeS will

undergo physical separation. FeS will be treated to recover sulfur and iron. Depending on

the grade FeNi will likely be used as feed for a nickel smelter or used to produce nickel

pig iron. This research’s objective is to use traditional physical separation techniques on

the heat-treated product to maximize Ni recovery without sacrificing Ni grade.

Froth flotation and magnetic separation were investigated for the separation of the heat

treatment products, produced under varying temperatures and conditions. Initial tests

were conducted using a heat treatment product analogue (HTA), which consisted of

natural FeS, iron powder, and nickel powder. Tests will determine the recovery and

concentrate grade achievable with varying heat treatment product quality and particle

size.

Froth flotation was investigated to float FeS and sink FeNi. Tests were conducted via a

microfloat cell, across a range of pH values and reagent doses. Tests on HTA and on

2

heat-treated products show a failure to float all the sulphides leading to a poor concentrate

grade.

FeNi is highly magnetic and FeS is non-magnetic, making low-intensity magnetic

separation a viable option. Magnetic separation tests on HTA showed recovery of

sulphides even at low intensities and a failure to recover significant amounts of Fe and Ni

below 25 mT. Tests done on various heat-treated products showed a maximum of 8.8

wt% Ni grade and 68 wt% recovery of nickel. The poor nickel grade and recovery is

attributed to a lack of liberation and small particle size input. Particle sizes below 10-15

µm are difficult to separate by magnetic separation and liberation is difficult to achieve for

small FeNi particles produced in heat treatment.

The pyrometallurgical process provides an alternative for a common waste mineral;

however, work is still required to develop an effective physical separation technique for

the heat treatment products.

3

2 Literature Review

Canada is a major producer of nickel, producing 11-12% of the world’s supply in 2016 [1].

The mixed sulphide ore deposit in the Sudbury Basin has been mined for the past 100

years due to its high nickel content [2]. It is a major source of nickel, copper, cobalt, and

precious metals [1]. The sulphide minerals in the Sudbury Basin are primarily pentlandite,

chalcopyrite, pyrrhotite (Po), and pyrite [2].

Po is abundant and can be recovered along with valuable base metal sulphide ores;

however, it has little economic value by comparison. Nevertheless, Po could be recovered

if it contains platinum group metals (PGEs) or Ni or Co in solid solution [5] [7]. Po from

the Sudbury Basin contains nickel and cobalt, and due to high amounts of nickel it is

referred to as nickeliferous pyrrhotite.

Previously, Po was treated pyrometallurgically for its nickel content; however, the process

created large amounts of sulfur dioxide gas and slag [2] [5]. The recovery of Po also

diluted the concentrate feed and reduced the throughput of Ni units through the smelter.

Environmental legislations introduced in the 1970s required a reduction of sulfur dioxide

emissions, and Po contributed to greater than 75% of the feed sulphur at the time [4] [5].

Due to these stricter environmental regulations and economic constraints, operations

were altered to reject a majority of Po as waste [1] [2] [5] [7]. Depressed in froth flotation,

it is separated from the other sulphides and stored in tailings ponds, but this can still lead

to acid mine drainage potential. The degradation of pyrrhotite in tailings ponds produce

large amounts of acid, a costly problem that mining companies must monitor to avoid

harming the environment. Another risk for tailings ponds is dam failure, which can

severely harm the local community.

Currently, there are 90-120 Mt Po tailings stored in Canada, containing up to 1 wt% Ni

and is valued at an estimated 4-11.5 billion USD [1] [2]. Tailings can be reclaimed through

surface mining and an additional benefit is that the pyrrhotite is finely ground. Ore grades

4

have declined overtime, so mining companies are reassessing their tailings ponds to

determine how much value can be extracted.

A new pyrometallurgical process for the extraction of nickel from Po tailings is being

developed. First, the tailings are upgraded to remove silicate gangue minerals, and then

the concentrated Po is heat-treated to convert Po into ferronickel (FeNi) and troilite (FeS),

through the following reaction:

𝑭𝒆𝟏−𝒙𝑵𝒊𝒚𝑺 + 𝑭𝒆 → 𝑭𝒆𝑵𝒊𝒚 + 𝑭𝒆𝑺 (1)

This process does not create large amounts of SO2 as the previous methods. FeNi and

FeS then undergo physical separation. The FeS can be treated to recover sulfur and iron,

while the FeNi forms the feed for steelmaking or sent to a Ni smelter. This author’s project

focuses on the physical separation of FeNi and FeS via magnetic separation and froth

flotation.

2.1 Pyrrhotite Properties

Po is the second most common iron sulphide mineral in nature, after pyrite [4]. Po has

the form Fe1-xS, x=0-0.125 and is a non-stoichiometric iron-deficient compound with a

small amount of nickel substituting for iron in the lattice [1] [8]. It’s structure is based on

the NiAs crystal subcell. [4] [5] [9]. Nickel, cobalt, copper, and manganese have been

found to occur in pyrrhotite structure as trace elements substituting for iron [9].

Po has many superstructures of which only four exist naturally at room temperature.

These are separated into two main categories: magnetic/monoclinic (Fe7S8 (4C)) and

non-magnetic/hexagonal (Fe9S10 (5C), Fe10S11 (11C), and Fe11S12 (6C) [10] [11] [12]. Po

contains ordered Fe vacancies and this distribution is different for all types of Po, which

affects magnetism [4].

Monoclinic Po is ferromagnetic at room temperature and hexagonal Po is

antiferromagnetic or “non-magnetic” [4] [12] [11]. Natural pyrrhotite is generally found as

5

intergrown mixtures of Po superstructures and tends to contain contaminant

elements/minerals [13] [14] [15]. The Phase diagram for the Fe-S system (FeS to FeS2)

is below in Figure 1.

The stochiometric end member troilite, FeS, is also grouped with pyrrhotite, however in

mineral processing it is a distinct mineral [4] [16] [24]. Troilite has no iron deficiencies,

and is hexagonal and non-magnetic [1].

In the literature, experiments performed on Po were done on natural and synthetic

samples. However, high temperature synthesis methods create synthetic samples which

contain metastable pyrrhotite phases and other iron sulphides, such as pyrite, this led to

a disagreement over Po phase relations and crystallography below 350 °C [4] [22] [25]

[26].

7

b)

Figure 1 a) Phase diagram for the Fe-S system (FeS to FeS2) [19] b) Phase diagram

for the Fe-S system [20]

2.1.1 Magnetism

Pyrrhotite has many superstructures, each of which has different responses to a magnetic

field. Figure 2 shows the relationship between the net magnetic moment and Po

superstructures. Monoclinic Po (Fe7S8, or 4C) has the highest magnetic susceptibility of

known minerals after magnetite; thus, it can be separated by low-intensity magnetic

separators [21] [3]. The other natural Po superstructures (5C, 6C, and 11C) are grouped

as non-magnetic. The ordered iron vacancies yield a net magnetic moment for the 4C but

no net moment for 5C, 6C, 11C and troilite. However, non-stoichiometric forms of Po have

net magnetic moments which are all lower than that of 4C [4]. Non-stoichiometric Po does

not commonly occur in nature; however, the heat treatment process mentioned before

could form these superstructures.

8

Figure 2: Net magnetic moment and Po’s superstructures. [19]

2.1.2 Pyrrhotite Tailings

Currently, several Po rejection flowsheets exist: a) Po can be separated into monoclinic

and hexagonal Po through magnetic separation, after which the two types of Po are

independently depressed in flotation using different strategies; b) rejection of both types

of Po in the same flotation circuit for reasons such as maintenance/capital costs [4].

At present, Po is typically depressed in flotation. Po tailings are produced in slurry form

and are contain various waste minerals referred to as gangue.

Once sulphide ores are mined and brought up to the surface, they react with oxygen and

microorganisms; therefore, Po is stored in tailings ponds as a AMD prevention

mechanism [7]. This is explained in more depth in section 2.1.5 Acid Mine Drainage

Potential.

9

There are two sources of Po tailings that may be used in the proposed heat treatment

process: fresh tailings from flotation depression and accumulated tailings which have

been deposited in tailing ponds. Fresh tailings typically have experienced less oxidation

and are easier to recover. Additionally, some accumulated Po tailings have been mixed

with rock tails up to six times and are disposed of together [1] [22]. As a result, these two

inputs require different approaches in initial upgrading of the tailings. Most research has

been focused on fresh tailings.

The study by Duffy et al. [2] compares Glencore Po tailings and Vale Po tailings samples

from Sudbury as shown in Table 1. The main difference is the silicate content, which

affects the upgrading of the pyrrhotite tailings. Silicates are impurities that should be

removed because of energy losses in processes, such as heat treatment where the

silicates will go through unnecessary heating.

Table 1: Comparison of Glencore Po tailings and Vale Po Tailings from Sudbury.

[2]

Mineral Glencore Po tailings

Composition (wt%)

Vale Po tailings

Composition (wt%)

Pyrrhotite 61.3 86.2

Magnetite 7.4 4.5

Pentlandite 1.2 1.2

Chalcopyrite 0.2 0.6

Total silicates ~30 ~7

In these tailings nickel is mainly found in pentlandite and pyrrhotite, as shown in Table 2

[2]. Although only 1.2 wt % of the tailings is pentlandite, it contains almost half of nickel;

therefore, any processes employed should recover Ni from pentlandite as well as Po.

10

Moreover, fine pentlandite can be found intergrown in Po mass [1] [2] [7]; therefore,

recovery of Po should ensure the recovery of pentlandite.

Table 2: Nickel distribution in Glencore Po tailings and Vale Po Tailings from

Sudbury [2].

Mineral Glencore Po tailings

Composition (wt%)

Vale Po tailings

Composition (wt%)

Pentlandite 45 40

Pyrrhotite 55 59

Liberation and exposure data was also conducted by Duffy et al. [2], and is shown in

Table 3. The liberation percentage indicates the mass fraction of the mineral that is free

and liberated, while the exposed percentage indicates the mass fraction of the mineral

having an exposed surface greater than 80% [2]. Liberation affects any flotation

processes, since additives react with the surface, and the data shows that Po is well

liberated. Fine particles are not recovered well in most traditional physical separation

processes, so if pentlandite is embedded in Po and partially exposed it can be recovered

and processed in heat treatment.

Table 3: Liberation and exposure of Po and pentlandite in Glencore Po tailings and

Vale Po Tailings from Sudbury [2].

Glencore Vale

Liberated (%) Exposed (%) Liberated (%) Exposed (%)

Pentlandite 75 73 48 45

Pyrrhotite 94 92 97 96

11

2.1.3 Galvanic Interactions

For minerals in aqueous suspensions, such as flotation conditions, their thermodynamic

stabilities can be summarized by Eh-pH diagrams (Pourbaix diagrams). Electrochemical

reactions in aqueous environments that may involve Po include natural galvanic

interaction with other minerals, oxygen, and ferric iron. Figure 3 shows the Pourbaix

diagram of water at 25 °C, 1 atm and Figure 4 shows the Pourbaix diagram for the Fe-S-

H2O system at 25 °C, 1 atm, 10-6 mol/L Fe and S. The closer the Eh value is to the upper

water stability limit the more the mineral oxidizes by oxygen gas. Po is at the lower part

of the water stability zone, which means that the driving force of oxidation by oxygen is

thermodynamically favourable.

Eh is controlled by dissolved oxygen content where the more aeration increases Eh.

Additionally, the Eh can be reduced by the nitrogenation, addition of reductants, or the

mild steel [4]. Additionally, the higher the pH the higher the oxygen uptake [4]. At an

alkaline pH (8-11) and low pulp potentials (-450 to -650 mV), Po is expected to be stable

[4]. Any deviation leads to mineral dissolution and/or oxidation from ferric or ferrous iron.

Sulphides with a lower rest-potential acts as the anode and undergoes oxidation, while

the sulphide with the higher rest-potential acts as a cathode [5] [23] [24] [25]. Therefore,

when Po is locked up with other minerals such as pyrite, pentlandite, or chalcopyrite, it

causes Po’s oxidation rate to increase. Table 4 shows the rest potentials of common

minerals. Most minerals have an Eh lower than oxygen; therefore, oxygen is commonly

the final electron acceptor. Another electron acceptor is ferric iron, which reduces to

ferrous iron.

12

Figure 3: Pourbaix diagram of water at 25 °C, 1 atm.

Figure 4: Pourbaix diagram for the Fe-S-H2O system at 25 °C, 1 atm, 10-6 mol/L Fe

and S. Pyrrhotite is in yellow [4].

13

Table 4: Rest potential values of common minerals [24]

Mineral Formula Rest potential (V)

Pyrite FeS2 0.66

Chalcopyrite CuFeS2 0.56

Sphalerite ZnS 0.46

Pentlandite NiFeS 0.35

Pyrrhotite Fe(1-x)S 0.31

Galena PbS 0.28

2.1.4 Oxidation of Pyrrhotite

Sulphides are prone to oxidation when in contact with oxygen and water, and Po in

particular is highly reactive [5] [4] [8] [12] [26]. The oxidation of pyrrhotite is in the order

of 10-8 – 10-9 mol m-2 s-1 [12]. When freshly fractured Po is exposed to air for a few

seconds, half the iron in the first few layers oxidize [4] [27] [24]. Various factors affect

oxidation rates such as oxygen concentration (dissolved oxygen content, relative

humidity), pH, temperature, ferric iron concentration, iron vacancies in the lattice, trace

metal concentrations, galvanic effects, and surface area [4] [8] [9] [10].

Catalysts of oxidation mechanisms are ferric iron and bacteria, which require low pH

environments; therefore, they can be reduced by pH regulation [12]. Po follows similar

oxidation mechanisms as pyrite [12]. It is a multistep process involving an oxygen-

independent reaction (ferric iron attack on the mineral) and oxygen-dependent reactions

(re-oxidation of ferrous iron to ferric and oxidation of reduced sulfur compounds produced

as intermediates in the process, ultimately to sulfate) [8] [46] [47].

14

The oxidation reaction of Po surface is shown below [4] [9]:

𝑭𝒆𝟏−𝒙𝑺 + (𝟐 −𝟏

𝟐𝑿)𝑶𝟐 + 𝒙𝑯𝟐𝑶 → (𝟏 − 𝒙)𝑭𝒆𝟐+ + 𝑺𝑶𝟒

𝟐− + (𝟐𝒙)𝑯+ (2)

𝑤ℎ𝑒𝑟𝑒 0 < 𝑥 ≤ 0.125

Where Fe2+ is further oxidized:

𝟐𝑭𝒆𝟐+ +𝟏

𝟐𝑶𝟐 + 𝟐𝑯+ → 𝟐𝑭𝒆𝟑+ + 𝑯𝟐𝑶 (3)

At pH> 3, Fe3+ hydrolyzes and precipitates as iron hydroxide [9].

𝑭𝒆𝟑+ + 𝟑𝑯𝟐𝑶 → 𝑭𝒆(𝑶𝑯)𝟑 + 𝟑𝑯+ (4)

Ferric ions are much stronger oxidising agents than oxygen [5] [9]. These reactions

produce sulphates and hydroxides which render the surface hydrophilic. In flotation this

would result in a decrease of Po floatability.

Additionally, Po dissolves under acidic conditions and generates Fe2+ and H2S

𝑭𝒆𝟏−𝒙𝑺 + 𝟐𝑯+ → (𝟏 − 𝒙)𝑭𝒆𝟐+ + 𝑯𝟐𝑺 (5)

Under mildly oxidative conditions, incomplete surface oxidation can occur, where

elemental sulphur forms (6) and then iron oxidizes to Fe3+ (3) and hydrolyze, precipitating

as iron hydroxide (4) [5] [9].

𝑭𝒆(𝟏−𝒙)𝑺 +𝟏

𝟐(𝟏 − 𝒙)𝑶𝟐 + 𝟐(𝟏 − 𝒙)𝑯+ → (𝟏 − 𝒙)𝑭𝒆𝟐+ + 𝑺𝟎 + (𝟏 − 𝒙)𝑯𝟐𝑶 (6)

With partial oxidation, sulphur transitions to sulphur complexes: disulphide to

polysulphides to elemental sulfur to sulphites and eventually to sulphates [4] [5] [8] [9].

Sulphites and sulphates are hydrophilic, while the latter are hydrophilic, which is why

mildly oxidative conditions promote pyrrhotite flotation [4] [5] [8]. However, extensive

oxidation causes the formation of a layer of hydrophilic ferric hydroxides on the surface,

which additionally does not allow for xanthate adsorption [5].

15

When there are iron deficiencies in the matrix the charge imbalance causes the

surrounding Fe2+ ions to take the form of Fe3+ [18]. The Fe3+ ions are suspected to be an

internal oxidant in the structure and are believed to be most reactive towards oxygen,

therefore increasing oxidation rates [18]. In other words oxidation rates increases

proportionality to the amount of iron vacancies in Po; therefore, monoclinic Po

experiences more oxidation than hexagonal Po [5] [4] [8] [9].

The presence of trace metals such as nickel and cobalt are substituted for iron, which

creates a positive charge at that site [4] [8]. This restricts the movement of lattice electrons

which hinders mineral oxidation [4] [8]. As a result, higher trace metal content lowers the

oxidation rate [4] [9].

As mentioned before in section 2.1.3 Galvanic Interactions, Po naturally has galvanic

interactions with other minerals. When it is locked up with minerals such as pyrite,

pentlandite, or chalcopyrite, the oxidation rate increases.

Lastly, increases in temperature and ferric iron concentration increases oxidation rate.

Even freezer storage will allow some degree of oxidation [10].

2.1.5 Acid Mine Drainage Potential

Sulphide ores are brought above ground they react with oxygen, water, and

microorganisms; therefore, sulphidic mine tailings possess an acid mine drainage

environmental liability (AMD) [2] [28]. AMD causes acidification, ferric iron precipitation

and high concentration of dissolved metals in drainage waters [12]. Po is very reactive

and from reactions outlined in the previous section, the process creates acid.

Paper [30] follows the sequence of biogeochemical and mineral dissolution processes

leading to AMD. During the operational phase, no sulphidic oxidation should occur, but

once active operations are finished, the water level in the tailings pond drops, allowing

the sulphides to oxidize and the reaction produce acid [30]. Then, heavy metals

hydroxides, oxyhydroxides, and sulphates will form [31]. Ferric hydroxide has a red

16

precipitate, this is a visible sign of AMD formation [30]. The last stage is when the acid

concentration is very high and as a result metals such as copper, zinc, and lead are

mobilized through the tailings and infiltrate into the water; it is associated with bright

colours such as blue, green, yellow, white or red [29] [30].

In a tailings environment, nonoxidative dissolution is expected to be a significant

mechanism below the water table only [9]. Moreover, in tailings ponds the reactions below

create a cycle which continuously regenerate ferric ions which in turn produce more acid

[9]. This reaction is facilitated by the sulfur-oxidizing bacteria such as thiobacillus

thiooxidans and thiobacillus ferooxidans [29]. At pH values above 4, this may be mediated

chemically or biologically with iron-oxidizing bacteria such as gallionella ferruginea [32]

[33].

𝑭𝒆(𝟏−𝒙)𝑺 +𝟏

𝟐(𝟏 − 𝒙)𝑶𝟐 + 𝟐(𝟏 − 𝒙)𝑯+ → (𝟏 − 𝒙)𝑭𝒆𝟐+ + 𝑺𝟎 + (𝟏 − 𝒙)𝑯𝟐𝑶 (7)

𝟐𝑭𝒆𝟐+ +𝟏

𝟐𝑶𝟐 + 𝟐𝑯+ → 𝟐𝑭𝒆𝟑+ + 𝑯𝟐𝑶 (8)

𝑭𝒆𝟏−𝒙𝑺 + (𝟖 − 𝟐𝒙)𝑭𝒆𝟑+ + 𝟒𝑯𝟐𝑶 → (𝟗 − 𝟑𝒙)𝑭𝒆𝟐+ + 𝑺𝑶𝟒𝟐− + 𝟖𝑯+ (9)

𝑭𝒆𝟏−𝒙𝑺 + 𝟐𝒚𝑭𝒆𝟑+ → 𝟑𝒚𝑭𝒆𝟐+ + 𝑭𝒆𝟏−𝒙−𝒚𝑺 (10)

To inhibit the process, Po and other sulphides are deposited in shallow lakes or tailings

ponds, which limits the contact with oxygen. In tailings ponds pH control can limit Fe3+

concentrations and acidophilic bacteria, since they require low pH environments [12].

Additionally, in laboratory settings vacuum oven drying removes water, which also inhibits

the process.

2.1.6 Zeta Potential

A mineral in an aqueous solution has a natural surface charge, which causes an

imbalance at the mineral-water interface. At this boundary, the molecules arrange

themselves to satisfy the conditions in the bulk solution [4]. The absolute value of zeta

17

potential cannot be measured; therefore, isoelectric point (IEP) is used. This point is

where there is net zero potential. At low pH conditions, sulphide and oxide minerals will

have a net positive surface charge, due to H+ [4].

Freshly ground sulphide minerals generally have a pHIEP between 2 and 3; however, with

oxidation it changes to >6 [4] [23]. For pyrrhotite, Ni2+, Cu2+, Fe2+, Fe3+, and Ca2+ (via lime

addition) are common ions in flotation. The higher the ion concentration, the larger pHIEP

shift [4], where copper shows the largest shift [34]. These ions are used as activators as

they attach onto the surface of minerals through electrostatic interaction; therefore,

altering the surface charge of the mineral.

2.3 Historical and Current Pyrrhotite Treatment Methods

From the 1950s to the 1980s Falconbridge and Inco separated Po via magnetic

separation, which produced monoclinic Po, and tried various methods of processing it.

During the 1960s, the regulations applied to SO2 emissions changed. Inco responded by

constructing a superstack to disperse harmful gas and avoid damage to the local

environment [35]. The reverberatory and blast furnaces in use at the time were incapable

of producing gases suitable for acid production, and SO2 capture was not effective [1].

Additionally, constructing an acid plant was not feasible at the time, so companies

focused on rejecting Po without compromising pentlandite recovery. Overall Po rejection

was better for the environmental and more economical.

In the 1950s Falconbridge’s Iron Ore Plant used selective sulphation-roasting of Po,

which produced a calcine containing a water soluble Ni sulphate [7] [36] [37]. The plant

was closed due to the change in environmental regulations [38]. Additionally, the acid

generating capacity of the SO3 and sulphate products created corrosion, dust, and

hygiene problems in the plant [1].

The Inco Iron Ore Recovery Process was designed to recover all the valuable products

from monoclinic Po concentrate, producing a Ni oxide product, hematite, and H2SO4 [39].

18

The plant recovered 90 % of Fe, and 75 % of Ni at 1.25 Mt Po/year [40]. Slurry-fed fluid

bed roasters were used to dead-roast Po concentrate at temperatures up to 760 °C to

yield calcine with 0.2 % S [40]. The calcine was selectively reduced with partially

combusted natural gas at around 850 °C in a rotary kiln and then leached with NH2-CO2.

Cu was removed as a sulphide and Ni carbonate was precipitated and calcined to yield

an acid-soluble Ni oxide, containing 77 %Ni, 0.15 %Co, and 0.15 %S. The remaining

product, mostly magnetite, was processed to produce hematite. Fe recovery was 90%

and Ni up to 75% [41] [40]. Due to an economic downturn the reduction/leaching portion

of the process was closed in 1982. A scaled-down roaster operation was operated

until 1991 to maintain a market for sulphuric acid that would be generated when

SO2 capture and fixation began at the Copper Cliff smelter [12] [14] [42].

In 1967 Falconbridge created the Nickel Iron Refinery to treat Po and to capture all

valuable products. Po was dead roasted, and the off-gas was sent to a sulphur recovery

plant, where SO2 was converted in to H2S and then reduced to elemental sulphur. Nickel

containing iron oxide was reduced in a rotary kiln with coal to produce metallized NiFe

pellets [7]. On the pilot scale this worked well, but the scale-up was unsuccessful and

sulphur production was expensive [1].

In summary, the processes to treat Po for it’s Ni, Co, Fe, and S content created in the

past were not economically feasible or there were environmental issues [2] [1].

2.4 Thermal Upgrading of Nickeliferous Po

The purpose of the thermal upgrading process is to maximize the transfer nickel from the

Po phase into the iron rich phase; producing FeNi and FeS. The process is shown in

Figure 5 and in the equation below:

𝑭𝒆𝟏−𝒙𝑵𝒊𝒚𝑺 + 𝑭𝒆 → 𝑭𝒆𝑵𝒊𝒚 + 𝑭𝒆𝑺 (11)

19

Figure 5: Movement of nickel during thermal upgrading

In a study by Sridhar et al. [43] a pyrometallurgical process was developed to recover 75-

90 % of the nickel from nickeliferous Po, sourced form Sudbury, in a FeNi phase which is

4-7 wt % Ni [43]. The nickel concentrates varied from 0.68-1.5% Ni, where some of the

nickel was in the pyrrhotite solid solution and some in finely dispersed grains of

pentlandite [43].

The formation of ferronickel alloy from Po is possible at certain temperatures once Fe/S

ratio is shifted via addition of iron or removal of sulphur under non-oxidizing atmospheres

[44]. The phase diagram of the Fe-Ni-S system at 900 °C is shown in Figure 6. When the

S/Fe ratio is lowered to ≤1, it is theoretically possible to precipitate FeNi alloys [43]. The

two ways of lowering the S/Fe ratio is to 1) add iron into the matrix or 2) remove sulphur

from Po [43].

20

Figure 6: Equilibrium phase diagram of Fe-Ni-S at 900°C. The red arrow indicates

the initial Po concentration (star) and the change of composition during the thermal

upgrading process (γ denotes the ferronickel alloy and mss denotes the sulphide

solution). [45]

Iron can be added in one of several ways: 1) adding elemental iron to pyrrhotite and

heating the mixture to a temperature at which iron diffuses into the pyrrhotite lattice, 2)

adding iron oxide and a reductant, such that iron oxide is reduced to elemental iron in

situ. [43]

Removing sulphur can be done in two ways: 1) heating pyrrhotite to a temperature at

which the vapour pressure becomes significant and then a hydrogen atmosphere is used

to remove sulphur; 2) add lime and a reductant and then heat treat so it reacts with the

sulphur in pyrrhotite to form calcium sulphide; 3) particle oxidation to remove sulphur as

sulfur dioxide. [43]

Our research group’s process for nickeliferous Pyrrhotite upgrading is shown in Figure 7.

Where heat treatment process creates the ferronickel and troilite phase and then this

paper focuses on the grinding and investigation of separation process to separate the

phases.

21

Figure 7: Our research group’s proposed process to create nickeliferous alloy.

The additives used in this process are carbon, CaO, and Na2CO3. The Po concentrates

typically contain <5 wt% magnetite. Carbon is used as a reductant for magnetite (Fe₃O₄).

Additionally, if iron oxide is used instead of iron powder, then carbon can be used to

reduce iron oxide. However, at 900 °C, direct reduction is not kinetically favored, and thus

22

catalyst is needed. The additive Na2CO3 is used as catalyst for the reduction of iron oxide

by solid carbon. Moreover, CaO is used to increase the efficiency of sulphur removal,

where it acts as “sulfur scavenger” to react with the sulphur in Po and form calcium

sulphide.

The phase diagrams for the Fe-X-S systems at 900 °C are in Figure 8, where X represents

Ni, Co, and Cu. For Ni and Co they are concentrated in the Fe-rich phase, while Cu does

not show the same pattern and will be evenly distributed between the Fe-rich phase and

the sulphide-rich phase. [46]

(a)

(b)

(c)

Figure 8: Phase relations of the Fe-rich corner in (a) Fe-Ni-S system; (b) Fe-Co-S

system; (c) Fe-Cu-S system at 900 °C calculated by FactSage 6.4 [46] [47]

An investigation of the effects of temperature on thermal concentration were conducted

in [46]. When no liquid phase is generated, an increase in temperature from 800 to 950

°C produced a marginal effect on phase distribution, recovery, and concentrations of non-

ferrous metals. The thermodynamic driving force of extracting Ni and Ni recovery is

reduced with an increase in temperature; however, kinetics are favourable at higher

temperatures; it is also expected that larger particles of the alloy are produced at higher

temperatures. At higher temperatures the presence of S2 gas results in favourable

reaction which reacts with iron and produces FeS. Moreover, at temperatures above 950

°C cause a liquid phase to form and thus reduced Ni recovery and grade.

23

2.5 FeNi Properties

Nickel (Ni) is a transition metal, atomic number 28 with a specific gravity of 8.9. It is

ferromagnetic at room temperature. Iron (Fe) is a transition metal, atomic number 26 and

a specific gravity of 7.9. It is also ferromagnetic at room temperature.

2.5.1 FeNi Passivation

Nickel is naturally a passivating metal, yet iron is not [48]; therefore nickel is resistant to

corrosion while iron is not. A passive film is a thin oxide layer, which is nanometers thick,

and it acts as barrier between oxygen and other electrolytes [49] [50]. Passive films may

form spontaneously on metals and alloys in contact with air or in an aqueous environment

[49].

Iron-nickel alloys are widely used in industry, due to their unique properties which are

dependent on alloy composition [51] [52]. Those with high nickel concentration have

higher corrosion (in oxidizing solutions) and heat resistances [51].

Studies have shown that the top layer of FeNi3 is formed for iron-nickel alloys that have

more than 40% iron in the alloy [34] [36]. Raman spectroscopy results indicated that many

nanocrystalline boundaries were found in the Ni-Fe alloys that activated the diffusion of

nickel, which lead to the formation of more nickel oxides and nickel hydroxides [54].

For this study nickel content in the FeNi alloy is below 10 %; thus, the particles produced

will have limited corrosion resistance. Flotation requires air to pass through water, this

increases the rate of corrosion. In the Davis Tube Tester FeNi is kept in water with little

contact with air; thus, causing less corrosion.

2.6 Separation Methods

Froth flotation, magnetic separation, and gravity separation are all viable physical

separation techniques for the heat treatment product (FeNi and FeS). Table 5 outlines

the physical properties that can be exploited for separation.

24

Table 5: Physical properties of FeNi and FeS, and the technique used from

separation.

FeNi and FeS Properties Technique

FeS is more floatable than FeNi Froth Flotation

FeS is weakly magnetic

FeNi is magnetic

Magnetic Separation

SG FeS= 4.67-4.79

SG FeNi = 7.9

Gravity Separation

2.6.1 Comminution

Most minerals are finely disseminated in gangue and must be liberated before any

separation processes, this is done through crushing and grinding the ore. To determine

the degree of grinding, size analysis is used. In 2014 it was reported that about 2% of the

electrical energy generated worldwide is spent on comminution processes; therefore,

comminution is a costly process [55].

2.6.2 Froth Flotation

Froth flotation is a physical separation technique or a physicochemical process which

separates hydrophobic and hydrophilic minerals. Figure 9 shows a general flotation

process. Chemical additives can transition a mineral between hydrophilic and

hydrophobic states. Factors of flotation include: particle size, liberation, air flow rate and

bubble size [23]. The process works better on relatively fine particles. If they are too large

the force of adhesion will be less than gravity. Sulphides are reactive with oxygen

dissolved in the water [23].

25

Recovery can occur through 3 different processes:

o 1) selective attachment of hydrophobic species onto air bubbles

o 2) entrainment where water passes through the froth

o 3) aggregation, which is the physical entrapment of particles between air

bubbles

Regulators are used in flotation and the classes are activators, depressants, dispersants,

or pH modifiers. Collectors are surfactants which are added to the pulp. They are organic

compounds which render selected minerals hydrophobic, by adsorbing onto the mineral

surface. Activators are generally soluble inorganic salts and they are used alongside

collectors. Ions such as Cu2+ interact with mineral surfaces allowing collectors to bind.

Figure 9: Batch flotation cell

2.6.2.1 Xanthate and Pyrrhotite flotation

Xanthates are effective collectors [11] [56]. Potassium amyl xanthate (PAX) and sodium

isobutyl xanthate (SIBX) are traditional thiol collectors used in the bulk and selective froth

flotation of nickel and copper sulphide mineral ores [57].

26

Xanthates are negatively charged at pH>5; therefore, pyrrhotite’s surface potential is

lowered after xanthate adsorption [4]. The pKa of xanthic acid is approximately 5 [4].

Below pH 5, it is undissociated xanthic acid (HX); at pH 5 it is 50/50% mixture of HX and

X-; and above ph 5 it is mainly X- [4].

Xanthate attaches onto pyrrhotite first by physisorption (electrostatic) and then by

chemisorption (bond formation). Chemisorption requires oxygen and produces

dixanthogen [4] [25] [58] [59]. Dixanthogen is formed when oxidation potential is -200

mV/SHE [60]. For many sulphides their hydrophobicity is linked to their xanthate or

dixanthogen surface complexes [4]. The Po and xanthate reaction is below: [4] [60]

𝑷𝒐|| + 𝑯𝟐𝑶 ⟷ 𝑷𝒐||(𝑶𝑯)[𝑺]+ + 𝑯+ + 𝟐𝒆− (12)

𝑷𝒐||(𝑶𝑯)[𝑺]+ + 𝑿− ⟷ 𝑷𝒐||(𝑶𝑯)[𝑺][𝑿] (13)

𝑷𝒐||(𝑶𝑯)[𝑺][𝑿] + 𝑿− ⟷ 𝑷𝒐||(𝑶𝑯)[𝑺] [𝑿𝟐] + 𝒆− (14)

𝑶𝟐 + 𝟐𝑯𝟐𝑶 + 𝟒 𝒆− ⟷ 𝟒𝑶𝑯− (15)

Oxygen is often the final electron acceptor in flotation systems. Experiments performed

with nitrogen flotation found that dixanthogen does not form when air is replaced by

nitrogen; therefore, oxygen plays an important roles in xanthate adsorption [11] [25].

Paper [61] used a semi-empirical parameterized model to study the affinity to copper ions

of ethyl xanthate, propyl xanthate, iso-propyl xanthate, iso-butyl xanthate, and amyl

xanthate. In terms of the electron density map, heat of formation, and binding energy and

dipole moment. It was observed that compared to the other xanthates amyl-xanthate

strongly binds to the surface of copper ions. Batch flotation tests were used to verify this

finding.

Additionally, experimental work on Voisey’s Bay ore show that potassium amyl xanthate

gives the best nickel recovery compared to sodium ethyl xanthate and sodium isopropyl

xanthate [11].

27

At low pH conditions Po and xanthate salts can generate toxic gases. Xanthate salts

characteristically decompose in acid, creating toxic carbon disulphide gas [23] [62].

However, xanthate is relatively stable above pH 6. [23]

𝑹 − 𝑶 − 𝑪𝑺𝑺− + 𝑯+ ↔ 𝑹 − 𝑶 − 𝑪𝑺𝑺𝑯 → 𝑹 − 𝑶𝑯 + 𝑪𝑺𝟐 (16)

2.6.2.2 Copper Activation

Copper sulfate and nickel sulphate are commonly used as activators [11]. Copper

activation on sulphides is a one fast step involving Cu2+ adsorption onto reactive sulfur

sites, where it is reduced to Cu+ and the sulfide is oxidized, producing Cu2S [63]. In

alkaline conditions electrostatic interaction can occur producing CuS [63]. Copper

activation is a quick process, copper uptake has been observed to end after 5 minutes

[63]. Copper reacts chemically with collectors [63] [64].

At alkaline pH Cu2+ forms Cu(OH)2 precipitates [63] and then reacts with xanthate to form

CuX2. Overdosing copper sulfate can lead to excess copper hydroxide, which reacts with

xanthate and decreases floatability [60].

2.6.2.3 Frothers

Frothers are surfactants are used to keep froth stability constant [23]. Frothers functions

are to aid formation and preservation of small bubbles, reduce bubble rise velocity, and

aid the formation of froth [23]. Reducing rise velocity increases the residence time of

bubbles in the pulp, which increases collision rates and therefore increases flotation

kinetics [23]. Reduction of bubble size increases the surface area of bubbles, this

increases collision rates and therefore increases flotation kinetics [23]. However, too

much frother could lead to smaller bubbles that do not readily rupture and cause poor

drainage and decreases selectivity [57].

28

Bubble surface area flux is given by:

𝑺𝒃 =𝟔𝑱𝒈

𝑫𝟑𝟐 (17)

where Jg is the superficial gas velocity (volumetric air rate divided by cell cross-sectional

area), and Sb is bubble surface area flux, and D32 is the Sauter mean diameter [23].

2.6.2.4 Flotation of Po

Recent studies show some difference in flotation responses of different pyrrhotite

superstructures. Pyrrhotite exhibits some natural hydrophobicity, due to mild surface

oxidation, leading to collectorless flotation, although it is not enough in practise.

Compared to magnetic Po, non-magnetic Po tend to be easier to recover in flotation.

However, this is not always the case; Table 6 shows that mixed Po performed similarly to

magnetic ores. As a result, to understand the effect of Po’s superstructures, other factors,

such as mineralogy, pyrrhotite crystallography, activation/depression by surface

modifiers, pH, Fe3+ content, trace metal content, and oxidation should be assessed [4] [5]

[7] [10] [65].

Discrepancies in the literature tend to stem from inadequate mineral characterization and

preparation prior to the study which yields contradictory results on similar materials [4] [5]

[10].

29

Table 6: Comparison of Pyrrhotite (Po) collectorless flotation in Ni ore deposits [66]

[67]

Pyrrhotite is depressed in the following ways:[4]

1) Collector starvation: xanthate preferentially binds to many other copper and

nickel sulphides

2) Aeration: this promotes ferric oxyhydroxides on pyrrhotite

3) Low Eh environment: prevents dixanthogen formation on pyrrhotite

4) High pH with sulphites and thiosulphates [4] [59]

Localized mild oxidation promotes the formation of hydrophobic species (refer to the

oxidation section); which in turn improves Po flotation. Additionally, minerals that have a

higher rest potential than Po, such as pyrite, chalcopyrite, and pentlandite, have improved

floatability [5] [25]. On the other hand, they can also decrease flotation [60].

In a low oxygen environment, xanthate adsorption is hindered, but with the addition of

pulp ions such as Cu2+, the adsorption is greatly improved [4] [60] [68].

Name Type of Po wt%Po Collectorless %Recovery of Po

Becker, et al Chimbganda,

et al

Sudbury CCN Non-mag Po 75.4 32

Sudbury Gert

West

Mag Po 85.2 3

Phoenix Mag Po 81.8 14

Nkomati MSB Mixed Po 83.8 6

Ore A Mag Po 11 - 30

Ore B Non-mag Po 62 - 71

30

Po floats well below pH 7 and in alkaline conditions lead to more hydrophilic ferric

hydroxide species on the surface, inhibiting xanthate adsorption; thus, decreasing the

floatability of all types of Po [8] [60] [65] [68] [69]. Additionally, at alkaline conditions the

zeta potential of Po is strongly negative; therefore, there is little chance of electrostatic

interaction between xanthate and the surface [60].

Moreover, copper activation of Po is reported to be not effective in the alkaline range [60].

Paper [70] investigated why pyrrhotite is more difficult to recover from alkaline slurries

than other base metal sulphide minerals. The formation of dixanthogen on pyrrhotite

surfaces is thermodynamically favourable in plant flotation slurries. However, the

activation of pyrrhotite by copper ions to form copper sulphide species does not occur in

alkaline solutions; this leads to the inhibition of xanthate interactions with the surface of

the mineral.

At low pH, pyrrhotite dissolves and generates H2S. The presence of oxygen leads the

liberated H2S to form elemental sulphur and with heat, the material can produce SO2. The

reaction is below: [71] [72]

𝑭𝒆𝟏−𝒙𝑺 + 𝟐𝒙𝒆− + 𝟐𝒙𝑯+ → (𝟏 − 𝒙)𝑭𝒆𝑺 + 𝒙𝑯𝟐𝑺 (18)

𝑭𝒆𝟏−𝒙𝑺 + 𝟐𝑯+ ↔ 𝑯𝟐𝑺 + (𝟏 − 𝒙)𝑭𝒆𝟐+ (19)

𝑯𝟐𝑺 +𝟏

𝟐𝑶𝟐 ↔ 𝑺𝟎 + 𝑯𝟐𝑶 (20)

𝑺 + 𝑶𝟐 ↔ 𝑺𝑶𝟐 (21)

The two different superstructures exhibit significant enough flotation responses [6] [65]

[73] [74].

Fresh 4C pyrrhotite is expected to contain more Fe(III) surface sites, which would promote

higher xanthate adsorption (with no pulp ions) [4]. Samples of freshly ground monoclinic

pyrrhotite are more floatable than hexagonal pyrrhotite, although the reverse occurred on

more oxidized samples [5]. However, 4C tends to oxidize more quickly, leading to 5C to

be more floatable at alkaline pH [5] [4] [8]. The general trend is that non-mag Po is less

31

reactive and thus more floatable [4]. Particle size also effects flotation. Non-mag Po is

floatable in the ranges 10-100 microns. Magnetic Po is recovered largely at <10 microns.

[4]

Microflotation investigations show that once surface oxidation is removed via sonification

and xanthate is present, magnetic and non magnetic Po superstructures behave similarly

[10]. Multani et al. shows that after 1 minute of sonication pre-treatment, the flotation

response for magnetic Po increased ~12% and non magnetic Po increased 20-80% (with

the use of collector) [10]. Additionally, various conditioning times were studied and 1

minute was optimal.

2.6.2.5 Alternative Collectors

Alternatives for xanthates are mercaptans and dithiophosphates, which are highly

effective for pentlandite and pyrrhotite flotation [11] . Mercaptons and lower-chain

xanthate can produce better nickel recovery than xanthate alone [11]. More selective thiol

collectors, such as dithiophosphates (DTP) and dithiocarbamates, are used as co-

collectors in mixtures with xanthates to improve selectivity. [57]

In study done by W. Maree et al., they found for an ore containing Pentlandite, pyrrhotite,

pyrite, and chalcopyrite, the nickel grade was highest for PAX, but the nickel recovery

was the lowest (Table 7) [57].

32

Table 7: The effects of mixtures of potassium amyl xanthate (PAX), sodium isobutyl

xanthate (SIBX),and isopropyl ethyl thionocarbamate (IPETC) collectors on grade

and recovery in the froth flotation of a nickel sulphide ore. [56]

Collector Mole

ratio

Cumulative

concentrate

mass (%)

Cumulative

nickel

recovery (%)

Cumulative

nickel grade

(%)

2 min 20 min 2 min 20 min 2 min 20 min

No collector 0 4.5 9.8 11.3 17.8 1.1 0.8

PAX 100 12.0 19.8 48.7 77.0 1.9 1.9

SIBX 100 11.5 20.8 48.9 79.1 2.1 1.9

IPETC 100 15.0 33.8 61.8 82.2 2.1 1.2

PAX:IPETC 95.5:4.5 12.3 23.2 50.3 84.6 1.9 1.7

SIBX:IPETC 95.5:4.5 12.7 23.5 50.8 82.0 1.8 1.6

PAX:SIBX 50:50 11.0 20.2 45.4 78.7 2.0 1.9

2.6.2.6 High Intensity Conditioning Flotation

Fine particles (6-50 micron) and ultra-fine (<6 micron) particles are difficult to recover

using traditional flotation techniques. It is estimated that recovering 15% of the metal lost

in the less than 10 µm fines in Canada would increase revenue by approximately $100 m

[75].

Conditioning is providing sufficient agitation and contact time for the reagents to react

with the minerals in the ground ore [76]. High intensity conditioning (HIC) exceeds this

minimum power input over a suitable length of time to induce aggregation of the finest

size fraction [76]. However, time alone is not sufficient to achieve aggregation, the

33

mechanical aspects and the chemical conditions must be matched to achieve good

aggregation [76]. Agitation speeds plays an important role in particle aggregation and

flotation, where it will peak and have positive impact on flotation kinetics [77]. The

threshold minimum power input, (kW/m3), depends on surface properties of the fine

particles and can vary considerably [76]. The mechanism of HIC is still unclear [77].

Hydrodynamic cavitation is one of the reasons why HIC can produce small size bubbles

[77].

High intensity conditioning has been adopted in many countries and has been reported

to improve nickel recoveries of nickel ores [11] [75] [78]. G. Chen et al. shows HIC

improved flotation behaviour on both whole and deslimed feeds of a Western Australian

nickel ore [78].

2.6.3 Magnetic Separation

Magnetic separators exploit the differences in the magnetic properties of minerals.

Commercially, belt or drum magnetic separators are often used, and they are typically

continuous. Flocculation or agglomeration may be a problem for smaller particle sizes

[23]. Gangue minerals can get entrained [23].

Paramagnetism in a material is caused by the presence of unpaired electrons which

create magnetic dipoles. The dipoles line up with the applied magnetic field.

Ferromagnetic is a special case of paramagnetism, where the magnetic dipoles of a

material undergo exchange coupling so that they can more rapidly align themselves with

an applied magnetic field. In this case ferronickel produced would be ferromagnetic.

Troilite is nonmagnetic [23]; however, as stated in the section, Magnetism of Po, Po

formed in the heat treatment process could be magnetic, which could be impurities in the

magnetic concentrate.

The types of magnetic separators: low-intensity vs. high-intensity and wet vs. dry. Low-

intensity magnetic separators, typically <~0.3 T, are used to separate ferromagnetic or

34

highly paramagnetic minerals [23]. Magnetic field intensity should be chosen carefully,

because higher field strengths may lead to an increased capture of weakly magnetic

particles. Dry magnetic separators treating fine material tend to have issues of

agglomeration; therefore, for particles below 5 mm, wet separation is used over dry [23].

In this study finer particle sizes are expected and ferronickel is strongly magnetic while

troilite is nonmagnetic; therefore, a wet low-intensity magnetic separation method is

preferred. However, low intensity wet magnetic separation is not effective for particles

less than 10 micron [79].

A previous test was done to separate a FeNi material and limonitic laterite ores, where

the FeNi phase was 10-20 micron with a WHIMS, wet high intensity magnetic separator.

Recovery of Ni was 91% but the grade was affected 4% (they expected 10 %). This was

likely due to poor liberation; however, there is a trade off between particle size/grinding

and recovery [80].

For the purposes of these experiments, batch tests will be conducted with a Davis Tube

Tester (DTT) (Figure 10). DTT is a wet magnetic separation method and is widely used

for concentration of fine magnetic particles [81] [79] [82] [83].

The magnetic force on the desired magnetic concentrate must be stronger than all the

sum of all the competing forces (gravitational, inertial, hydrodynamic, and centrifugal

forces). [23] [81] [79] [84]. The magnetic force (22) is a function of volume Vp, volume

magnetic susceptibility kp, magnetic field H, fluid medium of susceptibility kf, and µ0 is the

magnetic permeability of vacuum [84]. The volume magnetic susceptibility kp is positively

correlated with grain size [85]. The most significant forces that compete with the magnetic

force are the force of gravity (23), centrifugal force (24), and hydrodynamic force (25).

The particle density is ρp, density of the fluid is ρf, acceleration of gravity is g, radial

position of the particle is r, angular velocity is ω, reference area is A, drag coefficient is

Cd, and kinematic viscosity is ν. All the forces increase with grain size to varying degrees;

therefore, there will be an optimal grain size to maximize recovery. Moreover, magnetic

field intensity only affects the magnetic strength; therefore, it should increase magnetic

recovery. All the other variables will be kept constant.

35

�⃗⃗� 𝒎 =𝟏

𝟐𝝁𝟎(𝜿𝒑 − 𝜿𝒇)𝑽𝒑𝛁𝐇𝟐 (22)

�⃗⃗� 𝒈 = (𝝆𝒑 − 𝝆𝒇)𝑽𝒑�⃗⃗� (23)

�⃗⃗� 𝒄 = (𝝆𝒑 − 𝝆𝒇)𝝎𝟐𝑽𝒑�⃗� (24)

�⃗⃗� 𝒅 =𝟏

𝟐𝝆𝒇𝝂

𝟐𝑪𝒅𝑨 (25)

M. M. Ahmed performed a statistical analysis on DTT for the separation of magnetite from

quartz. The results show that in terms of recovery: [81]

• Particle size has the strongest positive effect. The larger the particle size, the larger

the magnetic force component.

• Current intensity is also positive. This increases the magnetic force.

• Wash water rate has a negative effect. The force caused by the washing of water

causes the magnetic material to fall.

• The slope of the Davis tube has a negative effect. The larger the slope the greater

the gravity component

• Tube oscillation has a low negative effect. The movement loosens some of the

magnetic particles from the magnetic field.

36

Figure 10: Davis Tube Tester

2.6.4 Gravity Separation

Gravity concentration uses the density difference between two minerals to separate them.

The motion of the particle is not only dependent on the SG but also the size of the particle.

Feed pulp density is an important factor, since relatively little deviation would rapidly

decrease efficiency [23]. These methods are relatively simple and work well on coarse

particles. Due to the simplicity of gravity separation, it tends to be the inexpensive

separation option.

Traditional techniques can separate minerals down to 50 microns [23]. To recover smaller

particle sizes, gravity concentrators with centrifugal force are used. Prior to separation

grinding is important to have adequate liberation. Successive regrinding should be

considered on the middlings.

Gravity separators are sensitive to slimes, which are ultrafine particles. Slimes increase

the viscosity of the slurry and hence reduce the sharpness of separation [23]. To deslime,

37

ultrafine particles are usually removed before gravity separation, resulting in the loss of

valuable material. [23]

Sulphides are usually removed by froth flotation due to their high SG, they tend to report

to the “heavy” product [23]. This should be taken into consideration when performing

tests, highly dense materials may be difficult to separate.

For effective separation there must be enough of a density difference between the light

and heavy material. The difference can be calculated by the concentration criterion, ∆ρ:

[23]

∆𝝆 =𝝆𝒉−𝝆𝒇

𝝆𝒍−𝝆𝒇 (26)

where ρh is the density of the heavy mineral, ρl is the density of the light mineral, and ρf

is the density of the fluid medium. Table 8 gives a rough indication of whether or not

separation is possible using traditional methods.

Po has an SG of 4.58-4.65 and troilite has an SG of 4.67-4.79 and the density of a FeNi

with a concentration of Ni of 3.5 wt% is approximately 7.9 [86] [87]. Therefore, for using

a fluid medium such as water, separation is possible and with a SG >2.4, separation

should be relatively easy, this can be seen in Figure 11.

Table 8: Dependence on Concentration Criterion for separation [23]

38

Figure 11: Concentration Criterion vs Specific Gravity of the Fluid Medium

Dense Medium Separation (DMS) is a process where the fluid with a SG that is between

the two material’s SG is mixed in with the solids. The denser material are the sinks and

the floats are the less dense material. However, for this case the density of the medium

would need to be between 4.65 - 8.1. A heavy liquid within that range was not found,

therefore DMS is not an option.

To recover fine particles (< 45 µm), wet gravity separation methods which make use of

centrifugal force are used. The Kelsey Centrifugal Jig takes a conventional jig and spins

it in a centrifuge [88]. The Knelson Concentrator is a batch centrifugal separator with an

active fluidized bed to capture heavy minerals [88]. Lastly, the Falcon Concentrator is a

batch centrifugal separator [88]. These technologies typically require large amounts of

sample material for testing (at least 1 kg); therefore, these options were not investigated

further.

39

3 Separation of Heat Treatment Analogue by Magnetic

Separation and Froth Flotation

This section focuses on separation tests on a heat treatment analogue (HTA) via

magnetic separation and reverse flotation (concentrate is the sinks product). The

objective was to analyze the effects of particle size, pH, and collector/activator

concentrations on the efficiency of separation through magnetic separation and gravity

separation. Separation tests were done on an analogue due to lack of material. This

analogue consisted of iron and nickel powder, representing the FeNi product, and

hexagonal Po/troilite, representing troilite. This creates an idealized case, since this

material is 100% liberated. Also it assumes that all the Po was fully reacted in the thermal

upgrading step; therefore, there would be no magnetic Po.

3.1 Materials and Methods

The HTA was prepared to represent the heat-treated product. The HTA preparation

process is outlined in 3.1.1 and the procedures used to separate the HTA are outlined in

3.1.2.

3.1.1 HTA Preparation

A fresh tailings sample was collected from the Vale-owned Voisey’s Bay Concentrator.

Tailings from Voisey’s Bay were chosen for preparation of the HTA as they contain

primarily hexagonal, or non-magnetic Po with some troilite, and some monoclinic, or

magnetic Po. The sample was primarily composed of pyrrhotite but also contained

impurities, such as silicate minerals. The sample was upgraded to remove impurities via

flotation, followed by magnetic separation, to produce a hexagonal Po and troilite

concentrate.

40

Flotation tests were conducted in a 1.5 L Denver flotation batch cell for 26 minutes

following the procedure summarized in Table 9. When not in use, the slurry sample was

stored in a fridge to minimize oxidation. The sample was conditioned with activator

(copper sulphate), collector (PAX), and frother (Flottec F160-13). Since the objective was

to produce a product of the highest purity, rather than to obtain good recovery, only the

first two concentrate samples were used for further testing. Samples were then filtered,

and vacuum oven dried to minimize oxidation.

Table 9: Procedure for treatment of Voisey’s Bay Po Tailings

Step Time (min) Addition

Input - 600 g of Voisey’s Bay Po Tailings at approximately 30 % solids

Aeration = 10 mL/min

pH = 7 regulated by H2SO4

1 Sonication 3

2 Suspension 3

3 Activation 3 CuSO4 (12 g/L) = 1 mL

4 Collector 4 PAX (12 g/L) = 1.6 mL

5 Frother 4 Frother (10 g/L) = 0.5

6 Concentrate 1 collection 3

7 Conditioning 3 CuSO4 (12 g/L) = 0.2 mL

PAX (12 g/L) = 0.5 mL

8 Concentrate 2 collection 6 Frother (10 g/L) = 0.5

41

Manual magnetic separation was used to separate the magnetic Po and non-magnetic

Po in the upgraded sample. The sample was added to an agitated cell filled with water

and a hand magnet with a strength of 1000 mT was used to collect the magnetic portion

from the flotation concentrates. The samples were vacuum filtered and then vacuum-

oven-dried overnight.

The resulting hexagonal Po and troilite concentrate, meant to represent the troilite

produced in the heat treatment product, was separated into the size fractions -45 µm and

-75+45 µm using standard sieving techniques via Ro-Tap. Iron and nickel powder, meant

to represent the FeNi alloy in the heat treatment product, was sieved into the same size

fractions and mixed with upgraded hexagonal Po and troilite samples of the

corresponding size fraction. The iron and nickel powders were sourced from Fisher

Scientific, specifically “Alfa Aesar™ Nickel powder, -100 mesh, 99+%” and “Iron

(Electrolytic Powder), Fisher Chemical” from Thermo Scientific.

The mix formula, based on the phase diagram at 900oC, was 89% hexagonal Po and

troilite concentrate, 10% iron powder, and 1% nickel powder. This mixture comprised the

HTA. HTA was stored in a fridge when not in use to minimize oxidation.

If 100% of the nickel along with all the metallic iron were recovered from the nickel

powder, then we should expect Ni grade would be 10% with 70% overall Ni recovery. If

the Ni recovery is higher, the extra nickel would be sourced from the Po.

The elemental composition of the HTA and all solid samples was determined by ICP-OES

(inductively coupled plasma optical emission spectrometry) analysis at Vale. The sodium

peroxide fusion was used to prepare the sample for digestion. Quantitative mineralogy

was determined by XRD (X-ray diffraction) Rietveld Analysis at Vale. Where the samples

were micronized to -10 µm and artificially spiked with 5 wt. % fluorite (CaF2) as an internal

standard. Lastly, the laser particle size analysis (Laser Scattering) was done by Vale.

42

3.1.2 Separation Procedure

The HTA was separated to produce a nickel/iron concentrate and a sulphide phase.

Magnetic separation was performed via Davis Tube Tester (DTT) and froth flotation via a

custom-built microfloat cell (Figure 12). The microfloat cell design was from Dr. Kelebek’s

lab at Queen’s University.

Figure 12: Microfloat cell (55 mL)

Magnetic separation tests were conducted with 10 g of HTA. The HTA was first sonicated

for 3 minutes to fully wet the surface. The magnetic flux density was set to a

predetermined setting (25-100 mT). Experiments were initially run once across a range

of magnetic flux densities to understand the range that would provide a reasonable

recovery. Three replicates were preformed for magnetic flux densities of interest. In the

second set of experiments, the nonmagnetic concentrate was repeatedly passed through

the DTT until no more concentrate could be recovered. Samples were then filtered with

acetone and dried overnight to prevent oxidation. The experiment testing variables are in

Table 10.

Reverse flotation was used since the sinks product of the flotation test (tailings) was of

interest. Flotation tests were conducted for 28 minutes with 7 g of HTA. The samples were

43

sonicated for 3 minutes and the pH was held constant with H2SO4 and NaOH at a

predetermined level. The microflotation procedure is summarized in Table 11. Samples

were then filtered with acetone and dried overnight. The experimental testing variables

are summarized in Table 12.

Table 10: HTA magnetic separation testing variables

Magnetic Flux

Density (mT)

Particle size 1 run Multiple Runs

25 - 45 µm X X

30 - 45 µm X

35 - 45 µm X

40 - 45 µm X X

100 - 45 µm X X

25 - 75 µm +45 µm X X

30 - 75 µm +45 µm X

35 - 75 µm +45 µm X

40 - 75 µm +45 µm X X

100 - 75 µm +45 µm X X

44

Table 11: Procedure for microflotation of HTA


Input - 7 g of HTA

Aeration = 0.4 mL/min

pH regulated by H2SO4 and NaOH

1 Suspension 3

2 Activation 3 CuSO4 (12 g/L)

3 Collector 3 PAX (10 g/L)

4 Frother 2


6 Conditioning 1 2


8 Conditioning 2 2


10 Conditioning 3 2


12 Conditioning 4 2


45

Table 12: HTA froth flotation testing variables

pH PAX/ CuSO4

Addition Particle size

6 1.4/0.5 - 75 µm +45 µm

7 1.4/0.5 - 75 µm +45 µm

8 1.4/0.5 - 75 µm +45 µm

6 1.1/0.4 - 75 µm +45 µm

7 1.1/0.4 - 75 µm +45 µm

8 1.1/0.4 - 75 µm +45 µm

6 0.6/0.2 - 75 µm +45 µm

7 0.6/0.2 - 75 µm +45 µm

8 0.6/0.2 - 75 µm +45 µm

6 1.4/0.5 - 45 µm

7 1.4/0.5 - 45 µm

6 1.1/0.4 - 45 µm

7 1.1/0.4 - 45 µm

6 0.6/0.2 - 45 µm

7 0.6/0.2 - 45 µm

3.1.3 Characterization

Voisey’s Bay Tailings contain 63 wt% Po and 5 wt% troilite and contain 6 wt% Si. The

P80= 106 m, and the mineral characterization and ICP results are in Table 13 and Table

14 (performed by Vale Mississauga). The hexagonal Po and troilite concentrate and HTA

at different size fraction ICP results are also in Table 14.

46

Table 13: XRD analysis of Voisey's Bay Tailings

Mineral Formula Concentration (wt%)

Pyrrhotite Fe(1-x)S 62.68

Olivine (Mg,Fe)2SiO4 9.18

Plagioclase NaAlSi3O8 – CaAl2Si2O8 8.81

Amphibole RSi4O11 6.22

Troilite FeS 4.95

Mica AB2–3(X, Si)4O10 (O, F, OH)2 3.53

Chlorite (Mg,Fe)3(Si,Al)4O10(OH)2·(Mg,Fe)3(OH)6 3.18

Pentlandite (Fe,Ni)9S8 0.72

Table 14: ICP-OES analysis of Voisey’s Bay Po Tailings (feed) material, Hexagonal

Po and Troilite Concentrate, and HTA at different size fractions

Element Voisey’s Bay Po Tailings (wt%)

Hexagonal Po and Troilite Concentrate (wt%)

HTA

-75+45 (wt%)

HTA

+45 (wt%)

Ni 0.412 0.495 1.54 1.42

Fe 47.35 56.19 61.1 55.85

S 24.93 34.91 31.29 27.9

Si 6.2 2.09 1.08 2.77

Mg 2.91 1.04 0.44 1.31

Al 1.59 0.59 0.28 0.72

Ca 1.55 0.49 0.25 0.66

Co 0.018 0.024 0.023 0.024

47

3.2 Results

The results of the magnetic separation and froth flotation of the HTA are outlined in this

section.

3.2.1 HTA Magnetic Separation Results

If 100% of the nickel along with all the metallic iron were recovered from the nickel

powder, then we should expect Ni grade would be 10% with 70% overall Ni recovery. If

the Ni recovery is higher, the extra nickel would be sourced from the Po.

Figure 13 shows the results from one run through the DTT. With increasing magnetic flux

density, the Ni grade decreases but the recovery improves. The grade recovery curve for

the -45 µm material is shifted upward to the right, indicating better performance compared

to -75+45 µm material. As discussed in section 2.6.3 Magnetic Separation, all the forces

acting on the particles increase with grain size to varying degrees; therefore, there will be

an optimal grain size to maximize recovery. Above 45 µm, the force of gravity and

centrifugal force are stronger than the magnetic force. The optimal grain size in this case

would be less than -45 µm.

After 1 run through the DTT, there was some entrainment of Po. Also, in some cases

nickel was recovered at above 70%; therefore, this further shows that some Po was

entrained in the concentrate.

Figure 14 shows the results from several runs through the DTT. After 2 runs through the

DTT, no more magnetic concentrate was recovered. Similarly, to the 1 run results, -45

µm performed better than the larger particle size. Additionally, there was some sulfur in

the concentrate; therefore, some Po is entrained in the concentrate after several runs

through the DTT, even though the material was 100% liberated. In all cases nickel was

recovered at above 70%; therefore, this suggests all the nickel was recovered and some

Po was entrained in the concentrate even at very low magnetic flux densities. Lastly, the

target grade 10% Ni was achieved in the metallic concentrate. In conclusion, magnetic

48

separation performed very well, and the magnetic field intensity can be anywhere

between 25-100 mT.

Figure 13: DTT results on HTA after 1 run

Figure 14: DTT results on HTA after 2 runs

49

3.2.2 HTA Microfloat Results

Figure 15 and Figure 16 show the ICP-OES analysis of the reverse flotation concentrate

(sinks product) Ni Grade and recovery for the -74+45 µm and -45 µm size fractions. The

feed grade for the -74+45 µm was 1.54 wt% and the feed grade for the -45 µm size

fraction was 1.42 wt%; therefore, the results were unsatisfactory.

As expected from the literature, at pH 8 flotation appears to have been compromised due

to oxidation. Pyrrhotite flotation at pH 6 usually produces the best result, as the sample

is in a more reducing atmosphere. As expected, pH 6 at reagent additions (PAX/CuSO4)

1.1/0.4 and pH 7 at 1.4/0.5 produced the best results. Higher PAX/CuSO4 does not always

produce better results, section 2.6.2.2 Copper Activation explains how excess copper

reacts with xanthate; thus lowering floatability.

Overall froth flotation does not appear to be a satisfactory method to separate HTA;

however, microfloat results should only be used to understand trends. Additionally, from

literature (section 2.6.2.4 Flotation of Po) we know that different materials and

compositions affect flotation, so it is still worth testing on the real heat treatment product.

50

Figure 15: FeNi concentrate’s Ni Grade and Recovery for -74+45 µm size fraction

(each test is specified by pH level and collector addition (mL)/activator addition

(mL))

Figure 16: FeNi concentrate’s Ni Grade and Recovery for -45 µm size fraction

(each test is specified by pH level and collector addition (mL)/activator addition

(mL))

51

3.3 Conclusions

Magnetic separation and froth flotation were conducted on an analogue material. This

material creates an idealized case, since this material is 100% liberated and assumes all

the Po was fully reacted in the thermal upgrading step.

The magnetic separation tests were very promising, they show that the target grade 10%

Ni was achieved and all the nickel was recovered. However, even at low magnetic flux

densities and 100% liberation, small amounts of Po were entrained. Moreover, -45 µm

performed better in terms of Ni grade and recovery compared to -75+45 µm; indicating

the optimal particle size for magnetic separation is below 45 µm. Lastly, the magnetic field

intensity can be anywhere between 25-100 mT, although lower magnetic field intensities

are desirable since they are less energy intensive.

HTA froth flotation produced low grades and recovery. The lowest performance was at

pH 8 due to oxidation, while tests at pH 6 and 7 produced only slightly better results.

However, microfloat results should only be used to understand trends and a Denver cell

would produce more accurate results. Additionally, from the literature we know that the

mineralogy, pyrrhotite crystallography, activation/depression by surface modifiers, pH,

Fe3+ content, trace metal content, and oxidation affect flotation, so it is worth testing out

on heat treatment product.

52

4 Magnetic Separation and Froth Flotation of Heat Treatment

Product Produced at 950 °C with Iron Oxide Addition

This chapter outlines the physical separation of the heat treatment product formed at 950

°C with iron oxide addition. The heat treatment procedure is also described. The results

of separation tests on heat treatment product are compared to the separation tests on

HTA as described in Chapter 3.


The heat-treated material was produced by Feng Liu (PhD candidate) and the procedure

used is outlined in section 4.1.1. Grinding, magnetic separation, and froth flotation

procedures are outlined in section 4.1.2. The input material for separation was limited;

therefore, only a small number of parameters were explored, and optimization of grinding

was not possible.

4.1.1 Sample Preparation by Thermal Upgrading at 950°C with Iron Oxide

The Po feed material used in this study is flotation-upgraded pyrrhotite tailings from

Strathcona mill (~1.59 %Ni, ~0.043 %Co, ~0.302 %Cu, and 33.7 %S). The mixture for

thermal upgrading consisted of Glencore flotation concentrate Po (at 89% purity), iron ore

fines (>95% purity), activated carbon (90% purity), CaO (chemical grade), and Na2CO3

(chemical grade). Activated carbon was added at a stoichiometric addition of 1.5 (C:O),

while CaO was 2% and Na2CO3 was 1%. Activated carbon and Na2CO3 are added to

reduce any magnetite and iron oxide in the Po concentrate. Lime is mixed in, such that

the sulphur in pyrrhotite forms calcium sulphide; thus, removing sulphur. After

53

compressing the mixture into pellets, the samples underwent isothermal heating at a

predetermined temperature for 2 hr under Ar atmosphere. The sample preparation was

performed by Feng Liu (PhD candidate). This heat-treated material was used as the input

for the physical separation methods.


The procedures for comminution, magnetic separation, froth flotation are outlined in this

section.

4.1.2.1 Comminution

The heat-treated product was crushed and ground by a puck mill for 2 minutes. The P80

was 34 µm.

4.1.2.2 Magnetic Separation

The dry sample was mixed with water and sonicated for 3 minutes to fully wet the surface.

Magnetic separation tests were conducted on 10 g using a DTT at a predetermined

magnetic flux density and water as the medium. The test was repeated until little to no

more magnetic concentrate was collected. The magnetic concentrate and non-magnetic

concentrate were collected, filtered, and dried with acetone, which minimizes the

oxidation of the sulphides and iron. The samples were characterized by ICP-OES and

Laser Particle Size Analysis at Vale Mississauga. XRD and SEM images were produced

by Feng Liu.

54

4.1.2.3 Reverse Flotation

Since the tailings is of interest this method used is reverse froth flotation. Flotation tests

were conducted for 28 minutes with 7 g of HTP. The sample was sonicated for 3 minutes

prior to the test and the pH was kept constant with H2SO4 and NaOH at a predetermined

value. The sample procedure is summarized in Table 15. Samples were then filtered with

acetone and dried overnight. The experimental variables are summarized in Table 16.

Table 15: Procedure for microflotation of heat-treated material at 950 °C


Input - 7 g of heat-treated product

Aeration = 0.4 mL/min

pH regulated by H2SO4 and NaOH

Sonication 3

Suspension 2

Activation 3 CuSO4 (12 g/L)

Collector 3 PAX (10 g/L)

Frother 2

Concentrate 1 collection 1

Conditioning 1 3 CuSO4 (12 g/L)

PAX (10 g/L)



PAX (10 g/L)



PAX (10 g/L)



PAX (10 g/L)


55

Table 16: Testing variables for microflotation of heat-treated material at 950 °C

pH PAX/ CuSO4 Addition Total PAX/

CuSO4 Addition Initial

Activation

and

Collector

addition

Conditioning

1

Conditioning

2

Conditioning

3

6 0.6/0.2 0.3/0.1 0.3/0.1 0.3/0.1 1.5/0.5

7 0.6/0.2 0.3/0.1 0.3/0.1 0.3/0.1 1.5/0.5

8 0.6/0.2 0.3/0.1 0.3/0.1 0.3/0.1 1.5/0.5

6 0.3/0.1 0.3/0.1 0.3/0.1 0.3/0.1 1.2/0.4

7 0.3/0.1 0.3/0.1 0.3/0.1 0.3/0.1 1.2/0.4

8 0.3/0.1 0.3/0.1 0.3/0.1 0.3/0.1 1.2/0.4

6 0.3/0.1 0.3/0.1 0.6/0.2

7 0.3/0.1 0.3/0.1 0.6/0.2

8 0.3/0.1 0.3/0.1 0.6/0.2

4.1.2 Characterization

The XRD results in Figure 17 shows that during heat treatment Po transformed into troilite

and nickeliferous alloy. The SEM images shown in Figure 18 show two distinct phases,

where the nickeliferous phase is embedded in the sulphide structures, with few particles

above 10 µm. The EPMA results represented in Figure 19, where the %Ni in the sulphide

phase is 0.61 and %Ni in the alloy phase is 3.61.

56

Figure 17: XRD results of heat-treated Po at 950°C with iron oxide addition, where

1) iron oxide, 2) pyrrhotite concentrate, and 3) heat treated product. [45]

57

a) b)

Figure 18: SEM image of heat-treated Po at 950°C with iron oxide addition, at a)

x100 and b) x500 magnification. The lighter phase is the nickeliferous alloy and the

dark gray is the sulphide phase.

Figure 19: The nickel concentration in the nickeliferous alloy phase and sulphide

phase from the EPMA analysis. [45]

58

As a result of grinding, 90% of the particles were <45 µm. Figure 20 show the SEM images

of the ground particles, where the P80 of the particles is 34 µm. Liberation was difficult to

achieve due to the size of the nickeliferous particles produced from heat treatment.

Further grinding to achieve 100% liberation is expensive on the industrial scale, also once

the particles are too small they become difficult to recover through traditional physical

separation. Furthermore, as a result of low liberation, sulphide impurities appear in the

magnetic concentrates and decrease the grade.

The ICP results after grinding are in Table 17. Ni is at 1.28 %, Co is at 0.034 %, and Cu

is at 0.244 %. The target grade is 3.61 %Ni (from the EPMA results).

a) b)

Figure 20: SEM image of ground heat-treated Po, at 950°C with iron oxide addition,

at a) x200 b) 500x magnification. The lighter phase is the nickeliferous alloy and

the dark gray is the sulphide phase.

59

Table 17: Composition of heat-treated material after grinding by ICP-OES analysis.

Element wt%

Fe 59.52

Ni 1.28

S 26.84

Si 2.51

Co 0.034

Cu 0.244

4.2 Results

The results of the magnetic separation and froth flotation experiments on the heat

treatment product formed at 950 °C with iron oxide addition are below.

4.2.1 Magnetic Separation Results

The Ni recovery vs grade of the magnetic concentrates are shown in Figure 21, and the

mass pull and recoveries and grades of Co and Cu are in Table 18. With increasing

magnetic flux density, the Ni grade decreased while the Ni recovery increased. Although

increasing the magnetic flux density resulted in only small changes to the Ni grade, it

significantly increased the Ni recovery. The Ni grade of the feed was 1.3%. The Ni grade

of the concentrates were 0.6-0.7 % Ni less than the theoretical maximum grade of 3.61

% Ni. Particles which had lower levels of liberation required higher magnetic force for

recovery; furthermore, this added more sulphides into the magnetic concentrate and

thereby decreases the Ni grade. The maximum recovery achieved was 65%; however, it

is suspected from the trend of the curve, that higher recoveries can be achieved with

60

higher magnetic flux densities, with little impact on grade. Additionally, at higher magnetic

flux densities more of the Co and Cu were also recovered in the alloy phase. The highest

amount achieved was 25% Co and 65% Cu, again given the trend this is expected to

increase with higher magnetic forces. Compared to Chapter 3 results, where target grade

and recovery was achieved, these tests did not perform as well. This is likely due to poor

thermal upgrading and grinding was not optimized.

Particle size analysis results from the feed sample and the magnetic concentrates are

shown in Table 19. The P80 of the magnetic concentrates was much higher than the feed

sample, indicating that larger particles were preferentially recovered. It appears that for

particles below 10-15 µm, the recovery was very low. This could mean that the system

was not able to recover small particle sizes, and/or the smaller particles were non-

magnetic. Additionally, increasing the magnetic flux density only slightly increased the

recovery of smaller particles.

Figure 21: The grade vs. recovery of Ni in magnetic concentrates produced at 40

mT, 50 mT, and 100 mT

61

Table 18: The mass pull and grade vs. recovery of Ni, Co, and Cu from the magnetic

concentrates produced at 40 mT, 50 mT, and 100 mT

Element Magnetic Field

Strength (mT)

Magnetic

concentrate mass

recovery (g)

Recovery

(%)

Grade (%)

Ni Initial 0.128 - -

40 0.059 45.9 2.98

50 0.068 53.3 2.89

100 0.083 65.0 2.88

Co Initial 0.024 - -

40 0.004 17.4 0.216

50 0.005 20.2 0.209

100 0.006 25.0 0.211

Cu Initial 0.003 - -

40 0.002 46.9 0.081

50 0.002 52.8 0.076

100 0.002 64.6 0.076

Table 19: Passing size of initial and magnetic concentrates

Magnetic Concentrate

Initial 100 mT 50 mT 40 mT

P10 2.7 14.4 15.7 16.3

P50 14.6 35.4 35.6 36.3

P80 33.6 55.7 55.4 56.3

P90 45.6 68.5 68.0 69.3

62

4.2.2 Microflotation Results

The nickel concentration of the feed was 1.3 wt% and the recovery vs grade of the tails

from the microfloat tests are shown in Figure 22. Very limited upgrading was observed

and the trends found in section 3.2.2 HTA Microfloat Results were not seen here. As

mentioned before, thermal upgrading and grinding were not optimized; this has a negative

impact on the efficiency of separation. Additionally, due to the low amount of sample

material, different parameters such as concentration of collector and activator were not

explored.

Figure 22: FeNi microfloat concentrate’s Ni Grade and Recovery (each test is

specified by pH level and collector addition (mL)/activator addition (mL))

4.3.1 Conclusion

The thermal upgrading of Po performed at 950 °C with iron oxide addition was not

sufficient enough to produce large enough particles for separation by flotation or magnetic

separation. Additionally, liberation was compromised, resulting in poor mineral liberation.

Therefore, sulphides were present in the nickeliferous alloy concentrate, thereby reducing

63

the Ni grade. However, optimization of grinding would increase separation efficiency for

both magnetic separation and flotation.

A maximum recovery of 65%, with a Ni grade of 2.9%, was achieved by magnetic

separation at 100 mT. Higher recoveries are expected with higher magnetic flux densities.

Increasing the magnetic flux density resulted in only small changes to the Ni grade, while

significantly increasing the Ni recovery. Low particle sizes (<10-15 µm) were difficult to

recover and increasing magnetic flux density does little to recover them. The trace

amounts of Co and Cu were also partially recovered with magnetic separation. Moreover,

similar to the microflotation results achieved for HTA, the heat-treated material was

difficult to separate using microflotation.

64

5 Magnetic Separation of Heat Treatment Product Produced

at 850 °C, 900 °C, and 950 °C with Iron Powder

This chapter focuses on the grinding and magnetic separation of thermally upgraded

nickeliferous Po produced with iron powder and at different temperatures: 850 °C, 900

°C, and 950 °C. Based on the unsatisfactory results from Chapter 4, microflotation was

not explored in this chapter.


The heat-treated material was produced by Feng Liu (PhD candidate) and the procedure

he used is outlined in 5.1.1. The grinding and magnetic separation procedures are

outlined in 5.1.2. Similarly, to Chapter 4, there was limited amount of material; therefore,

grinding was not optimized to increase separation efficiency.

5.1.1 Sample Preparation by Thermal Upgrading at Various Temperatures with Iron Powder

The Po-containing material used in this study was flotation-upgraded pyrrhotite

concentrate from Strathcona mill (~1.59 %Ni, ~0.043 %Co, ~0.302 %Cu, and 33.7 %S).

The mixture for thermal upgrading consisted of Glencore flotation concentrate Po (at 89%

purity), iron powder (<75 μm), activated carbon (90% purity), CaO (chemical grade), and

Na2CO3 (chemical grade). Activated carbon was added at a stoichiometric addition of 1.5

(C:O), while CaO was 2% and Na2CO3 was 1%. After compressing the mixture into

pellets, the sample underwent isothermal heating at a predetermined temperature for 4

hrs under an Ar atmosphere. Sample preparation was performed by Feng Liu (PhD

candidate).

65


The heat-treated product was crushed and ground by a puck mill for 1.5 minutes. The dry

sample was mixed with water, and a sonicated for 3 minutes to fully wet the surfaces and

remove any iron oxides from the surface of the sulphides. Magnetic separation tests were

conducted on 10 g using a DTT at a predetermined magnetic flux density with water as

the medium. The test was repeated until little to no more magnetic concentrate was

collected. The magnetic concentrate and non-magnetic concentrate were collected,

filtered, and dried with acetone, which minimizes the oxidation of the sulphides and iron.

The samples were characterized by ICP-OES and Laser Particle Size Analysis at Vale

Mississauga. XRD and SEM images were produced by Feng Liu.

5.2 Results

Table 20 outlines the concentration of each species in the alloy and sulphide phase, while

Figure 23 shows the SEM images of heat-treated material before and after treatments.

The average nickel concentration in the alloy phase is 12.6 % at 850 °C, 9.71 % at 900

°C, and 5.29 % at 950°C. Therefore, the reporting of nickel into the alloy phase increases

with decreasing temperature. This is expected, as reported in section 2.4 Thermal

Upgrading of Nickeliferous Po, the thermodynamic driving force is much stronger at a

lower temperature. Cobalt will follow the same trend as nickel, although copper will not;

this phenomenon is explained in section 2.4 Thermal Upgrading of Nickeliferous Po.

A rough estimate of the grain size of the alloy phase can be made from Figure 23. 900 °C

produces larger grain sizes (as large as 50 µm) compared to 850 °C and 950 °C (the

largest was around 25 µm). At higher temperatures, the kinetics are better leading to

higher grain size. However, at 950 °C, the phase diagram shifts and producing a single

liquid phase; therefore, decreasing the amount of alloy present and the grain size. Also,

at 900 °C, the alloy phase shows some degree of separation from the sulphide phase;

this would result in less energy for grinding, which is an energy intensive process.

66

Moreover, although the nickel concentration is highest at 850 °C, the grain size is quite

small.

As observed in Chapter 4, the DTT starts to fail at particle sizes <10 µm. Since many of

the FeNi grains produced in the heat treatment were fine, achieving high liberation could

be detrimental for magnetic separation. Lower liberation would result in higher amounts

of sulphur in the magnetic concentrate; furthermore, recovery of nickel takes higher

priority if used in a nickel smelting process.

Table 20: Phase composition of heat-treated product at various temperatures

(EPMA)

T

°C

Concentration in Sulphide, wt% Concentration in Alloy, wt%

Ni Co Cu Ni Co Cu

850 0.30–0.37

(avg. 0.33)

~0.09

(avg. 0.07)

0.29–0.32

(avg. 0.31)

2.55–29.3

(avg. 12.6)

0.09–0.72

(avg. 0.32)

0.09–0.55

(avg. 0.27)

900 0.49–0.57

(avg. 0.53)

~0.09

(avg. 0.07)

0.27–0.41

(avg. 0.32)

9.24–14.0

(avg. 9.71)

0.30–0.43

(avg. 0.36)

0.06–0.20

(avg. 0.10)

950 0.54–0.77

(avg. 0.65)

0.05–0.10

(avg. 0.07)

0.21–0.50

(avg. 0.31)

1.85–9.31

(avg. 5.29)

0.15–0.34

(avg. 0.28)

~0.17

(avg. 0.08)

67

Fe Particle: 0 %Ni, 0.08–0.113 %Co, 0 %Cu

(a) Initial

(b) heat-treated at 850 °C for 4 h

(c) heat-treated at 900 °C for 4 h

(d) heat-treated at 950 °C for 4 h

Figure 23: SEM images of heat-treated product and composition of the metallic

phase (the lighter phase is the nickeliferous alloy and the dark gray is the sulphide

phase)

68

The grade vs recovery curves achieved by the magnetic separation of heat treatment

product produced at temperatures of 850 °C, 900 °C, and 950 °C are shown in Figure 24.

The highest grades were obtained from material produced at 850 °C, followed by 900 °C,

and 950 °C. The recovery was still quite low ranging from 50-68 %. Increasing the

magnetic flux density resulted in an increase of recovery of 2-10%, with little impact on

grade, this is consistent with 4.1.2.2 Magnetic Separation. Particles which had lower

levels of liberation required higher magnetic force for recovery; furthermore, this added

more sulphides into the magnetic concentrate and thereby decreases the Ni grade. 850

°C had the highest range of recovery of nickel vs magnetic flux density, where the

maximum recovery was 68% at 200 mT. This shows that higher thermal upgrading

efficiency has a strong positive effect on magnetic separation. The heat-treated product

produced at 900 °C had larger grain sizes; however, they played a lesser role in

increasing magnetic separation efficiency.

The passing size of the sample is shown in Table 21, Table 22, and Table 23. The P80 of

the concentrates increases significantly compared to the feed material, almost doubling

in some cases; this may indicate that much of the finer material is lost. Additionally, 900

°C produced much larger particles compared to the other temperatures, this is consistent

with Figure 23. The P80 decreases with increasing magnetic flux density (about 10 µm),

indicating that higher magnetic flux densities can recover finer particles, but only to a

certain extent.

69

Figure 24: Ni Grade vs Ni Recovery of magnetic concentrates at temperatures 850

°C, 900 °C, 950 °C

Table 21: Particle size data of feed and magnetic concentrates produced from

material heat-treated at 850 °C

Input Magnetic Concentrate at 850 °C

Passing Size

(µm)


P10 2.3 8.6 13.8 15.3

P50 11.1 30.5 35.1 39.4

P80 28.9 52.7 55.9 62.0

P90 32.7 66.8 68.7 76.0

70




Passing Size

(µm)


P10 3.3 25.3 26.7 28.5

P50 21.9 48.9 53.5 55.8

P80 44.3 71.9 78.6 83.1

P90 57.5 85.6 95.1 100.4




Passing Size

(µm)


P10 2.5 12.1 11.9 15.8

P50 15.8 34.1 32.6 36.7

P80 36.0 55.2 52.0 57.4

P90 49.2 68.1 63.5 69.9

71

5.3 Conclusion

The effect of the thermal upgrading temperature on the efficiency of magnetic separation

was analyzed. Thermal upgrading at 900 °C produced the largest alloy grain sizes, while

the highest thermal upgrading efficiency was at 850 °C. Magnetic separation of the heat

treatment product at 850 °C, produced higher nickel grades overall. The best Ni grade

and recovery was 8.8 % Ni grade and 68 % Ni recovery, which was produced from thermal

upgrading at 850°C and magnetic separation at 200 mT. This shows that higher thermal

upgrading efficiency has a strong positive effect on magnetic separation. At 900 °C the

grain size was larger; however, the grain size played a lesser role in increasing magnetic

separation efficiency. Additionally, thermal upgrading at 850 °C is economically

favourable. Furthermore, increasing magnetic flux density increases nickel recovery

significantly with little impact on nickel grade. Also increasing magnetic flux density

recovers some of the smaller particles to an extent. Overall, the nickel recovery achieved

was low (below 70%); however, an economic analysis is required to know what recoveries

would be acceptable. Moreover, optimization of grinding would increase the separation

efficiency.

72

6 Conclusion and Summary

The FeS and FeNi were separated via microfloation and wet low-intensity magnetic

separation to determine which one has sufficient separation efficiency. An analogue

material was used for initial testing. It is an ideal material, since it has 100% liberation

and assumes all of the Po was reacted in the heat treatment stage. Magnetic separation

tests via a DTT showed that target nickel grades and recoveries were achieved at even

low magnetic intensities. However, even at low magnetic flux densities and 100%

liberation, small amounts of Po was entrained. Moreover, the optimal particle size for

magnetic separation is below 45 µm. Microflotation of the HTA showed low separation

efficiency.

Secondly, material, which was heat-treated at 950 °C with iron oxide, was also separated

via microfloation and wet low-intensity magnetic separation. The thermal upgrading step

had poor nickel transfer efficiency and had small grain sizes. The small grain sizes led to

poor mineral liberation, resulting in lower grades in the final alloy concentrates. Magnetic

separation fails at low particle sizes, 10-15 µm. Increasing the magnetic flux density had

little negative impact on the nickel grade but greatly improved nickel recovery. Overall,

the recovery of FeNi material was low, with the highest achieved being 65% at 100 mT.

Microflotation results on this material produced unsatisfactory results with little to no

separation achieved.

Lastly, heat treated materials produced at 850 °C, 900 °C, and 950 °C were magnetically

separated. Although heat treatment at 900 °C produced larger FeNi grain sizes, magnetic

separation of the heat treatment product at 850 °C produced higher nickel grades overall.

Magnetic separation at 200 mT of material heat treated at 850 °C produced the best grade

and recovery, which were 8.8 % Ni grade and 68 % Ni recovery. Overall, higher thermal

upgrading efficiency has a stronger positive effect compared to positive effect of larger

particle sizes.

73

7 Recommendations for Future Work

These are the recommendations for advancing this area of research:

• The 850 °C heat treated material had the best magnetic separation efficiency. This

can be further improved by optimizing the comminution step and possibly looking

at even higher magnetic flux densities. This can also be improved by focusing on

the thermal upgrading step to create larger particle sizes, one possible method is

to implement a two-stage process, where it’s heated at a higher temperature and

then at 850 °C.

• The amount of material was limited in this study, if more material is obtained then

flotation tests via a Denver cell can be investigated. A Denver cell is a more

accurate depiction of how flotation would perform in a mineral processing plant.

Additionally, with more sample material, gravity separation via a centrifugal

concentrator can also be investigated.

• This paper focused only on traditional methods of separation; however, newer

technologies, such as high intensity flotation, can be investigated.

74

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