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Metal Recovery from Steelmaking Slag
by
Qi Yu
A thesis submitted in conformity with the requirements for the degree of Master of Applied Science
Department of Materials Science and Engineering University of Toronto
© Copyright by Qi Yu, 2018
ii
Metal Recovery from Steelmaking Slag
Qi Yu
Master of Applied Science
Materials Science and Engineering
University of Toronto
2018
Abstract
An experimental study was carried out to recover metal values from a steelmaking slag through
air oxidation and magnetic separation. The oxidation of slag was carried out in both solid and
liquid states, attempting to optimize the conditions of each. The effects of oxidation time,
temperature, slag basicity, and magnetic field strength on recovery and grade of the concentrate
were investigated and optimized oxidation of iron mono-oxide (FeO) to magnetite (Fe3O4) and
other spinel ferrites in RO phase was observed. Solid state oxidation achieved overall optimum
results when roasted at 900 °C for 1 hour, and magnetically separated under 110 mT. Two products
were collected. An iron-rich concentrate containing 28% Fe, 3.9% Mn and an iron-lean tailing
containing 7% Fe and 0.4% Mn. In order to make the iron-rich product recyclable, its grade needs
to be improved. It was demonstrated that this is possible by grinding the roasted slag to a smaller
particle size. Liquid-state oxidation yielded poor results, likely due to rapid cooling of slag
following its oxidation that did not allow growth and segregation of iron-bearing minerals. It was
shown that slow cooling of such slag can improve the separation efficiency and grade.
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Acknowledgments
I must sincerely thank my supervisor and mentor, Prof. Mansoor Barati, who not only always
encouraged, guided and taught me in the academic and scientific fields, but also selflessly
supported me when I was lost and gave me enlightenments in life.
I am also earnestly thankful to Dr. Sina Mostaghel, who brings me the most interest to metallurgy,
and has always been a role model to me.
I must also offer heartfelt gratitude to Prof. Uwe Erb, who guided me in and outside the classroom
since undergrad. I wish I can always drink another beer with you.
I am indebted to all my colleagues in our research group, who tolerated all my shortcomings and
brought the two joyful years. I would like to particularly thank Dr. Karim Danaei for his
remarkable and selfless support in the technical and experimental aspects of this work.
I would also like to thank the colleagues from Department of Earth Science, Dr. Yanan Liu and
Dr. Mike Gorton for their technical supports of this work.
The financial support and technical advice from Hatch were elemental for completion of this study.
NSERC and OCE provided additional funding for this project and Arcelor Mittal Dofasco supplied
the slag samples and I would like to thank all of them for their support.
Lastly, I must thank my parents. This accomplishment would not have been possible without them.
Thank you!
Qi Yu
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Table of Contents
ACKNOWLEDGMENTS ........................................................................................................................ III
TABLE OF CONTENTS ......................................................................................................................... IV
LIST OF TABLES ................................................................................................................................... VII
LIST OF FIGURES .................................................................................................................................. IX
LIST OF ACRONYMS ......................................................................................................................... XIII
CHAPTER 1 ................................................................................................................................................ 1
INTRODUCTION .............................................................................................................................. 1
CHAPTER 2 ................................................................................................................................................ 3
LITERATURE REVIEW .................................................................................................................. 3
2.1 INTRODUCTION ................................................................................................................................. 3
2.2 SLAG PRODUCTION AND CHARACTERIZATION ................................................................................. 4
2.2.1 Ferrous Slags ........................................................................................................................... 6
2.2.1.1 Ironmaking and Steelmaking Slags ................................................................................. 6
2.2.1.2 Stainless Steel Slag ....................................................................................................... 10
2.2.1.3 Ferroalloy Slags ............................................................................................................ 11
2.2.2 Non-ferrous slag .................................................................................................................... 13
2.2.3 Slag Utilization ...................................................................................................................... 16
2.2.3.1 In-Plant Recycling of Slag ............................................................................................ 17
2.2.3.2 Slag as By-Product ........................................................................................................ 18
2.2.3.3 Recovery of Metals and Minerals from Slag ................................................................ 18
2.3 PHYSICAL PROCESSING TECHNOLOGIES FOR METAL RECOVERY .................................................. 19
2.3.1 Cooling and Sizing ................................................................................................................. 19
2.3.1.1 Slag Cooling Methods ................................................................................................... 19
2.3.1.2 Effect of Cooling Conditions on Metal Distribution ..................................................... 20
2.3.1.3 Size Reduction .............................................................................................................. 21
2.3.1.4 Screening as a Concentration Method........................................................................... 21
2.3.2 Gravity Separation ................................................................................................................. 22
2.3.3 Floatation .............................................................................................................................. 23
2.3.4 Magnetic Separation .............................................................................................................. 23
2.3.4.1 Magnet Ordering ........................................................................................................... 24
v
2.3.4.2 Spinel Ferrite ................................................................................................................. 27
2.3.4.3 Magnetic Separation ..................................................................................................... 30
2.3.4.4 Magnetic Separation of Metal in Slag ........................................................................... 32
2.4 PYROMETALLURGICAL METHODS .................................................................................................. 34
2.4.1 Carbothermic Reduction ........................................................................................................ 34
2.4.1.1 Direct Reduction ........................................................................................................... 35
2.4.1.2 Selective Reduction....................................................................................................... 36
2.4.2 Slag Settling ........................................................................................................................... 37
2.4.3 Slag Fuming ........................................................................................................................... 39
2.4.4 Sulfurization ........................................................................................................................... 39
2.4.5 Magnetizing Roast ................................................................................................................. 40
2.5 HYDROMETALLURGICAL METHODS ............................................................................................... 44
2.5.1 Leaching ................................................................................................................................ 44
2.5.1.1 Effect of Various Parameters on Leaching Kinetics and Efficiency ............................. 44
2.5.1.2 Acid Leaching ............................................................................................................... 45
2.5.1.3 Alkaline Leaching ......................................................................................................... 47
2.5.1.4 Salt Leaching ................................................................................................................ 48
2.5.2 Metal Recovery ...................................................................................................................... 52
2.5.2.1 Solvent Extraction ......................................................................................................... 52
2.5.2.2 Precipitation .................................................................................................................. 52
2.6 SUMMARY OF METAL RECOVERY METHODS FROM SLAG ............................................................. 53
CHAPTER 3 .............................................................................................................................................. 55
EXPERIMENTAL ............................................................................................................................ 55
3.1 OBJECTIVES .................................................................................................................................... 55
3.2 MATERIALS ..................................................................................................................................... 55
3.2.1 Slag ........................................................................................................................................ 55
3.2.2 Additives ................................................................................................................................ 58
3.3 EXPERIMENTAL SETUP ................................................................................................................... 58
3.3.1 Solid State Oxidation ............................................................................................................. 59
3.3.1.1 Oxidizing with CO2 ....................................................................................................... 60
3.3.1.2 Oxidizing with Air ........................................................................................................ 60
3.3.2 Liquid State Oxidation ........................................................................................................... 60
3.3.3 Magnetic Separation .............................................................................................................. 61
vi
3.4 EXPERIMENT PARAMETERS ............................................................................................................ 63
3.5 ANALYTICAL TECHNIQUES ............................................................................................................. 65
CHAPTER 4 .............................................................................................................................................. 66
RESULTS AND DISCUSSION ....................................................................................................... 66
4.1 CHARACTERIZATION OF RAW SLAG ............................................................................................... 66
4.1.1 RO Phase ............................................................................................................................... 66
4.1.2 Response of Raw Slag to Magnetic Separation ..................................................................... 69
4.2 THERMODYNAMIC ASSESSMENT .................................................................................................... 71
4.3 SOLID STATE OXIDATION ............................................................................................................... 76
4.3.1 Roasting with CO2 ................................................................................................................. 76
4.3.2 Roasting with Air ................................................................................................................... 79
4.3.2.1 Effect of Magnetic Field Strength ................................................................................. 79
4.3.2.2 Effect of Roasting Time ................................................................................................ 81
4.3.2.3 Effect of Roasting Temperature .................................................................................... 82
4.3.2.4 Phase Evolution in Roasting of Slag ............................................................................. 83
4.3.2.5 Magnetic Separation of Slag Roasted Under Optimized Conditions ............................ 85
4.3.2.6 Manganese Recovery .................................................................................................... 91
4.4 LIQUID STATE OXIDIZATION .......................................................................................................... 92
4.4.1 Effect of Temperature and Time ............................................................................................ 94
4.4.2 Effect of Basicity .................................................................................................................... 97
CHAPTER 5 ............................................................................................................................................ 100
CONCLUSIONS ............................................................................................................................. 100
CHAPTER 6 ............................................................................................................................................ 102
RECOMMENDATIONS FOR FUTURE WORK ....................................................................... 102
REFERENCES ........................................................................................................................................ 103
APPENDIX A .......................................................................................................................................... 114
APPENDIX B ........................................................................................................................................... 116
vii
List of Tables
Table 2-1: Chemical composition (wt%) and production rate of ferrous and non-ferrous slags
[2][3][4][5][6]. ................................................................................................................................ 5
Table 2-2: Typical metal values (ppm) in ferrous and non-ferrous slags [4][5] ............................. 5
Table 2-3: Steel slag composition from literature ........................................................................... 9
Table 2-4: Stainless steel slag composition from literature. ......................................................... 11
Table 2-5: Summary of copper slag composition from literatures. .............................................. 14
Table 2-6: Mineralogy of copper slags. Slag # refers to those in Table 2-5. ................................ 16
Table 2-7：Comparison of calculated spin-only magnetic moments with experimentally
observed data for some common transition metals, where µB is theoretical Bohr magneton, and
µeff is the observed Bohr magneton [74]. ...................................................................................... 29
Table 2-8: Cation distribution, structure type and saturation magnetization of spinel ferrites[73].
....................................................................................................................................................... 29
Table 2-9: Magnetic susceptibility of selected rocks, minerals and BOF slag [76][77]. .............. 30
Table 2-10: Summery of pyrometallurgical methods investigated on metal recovery from slag. 42
Table 2-11: A summary of leaching studies on slag for metal recovery ...................................... 50
Table 3-1: Chemical compositions of steel slag sample (wt%) from XRF analysis provided by
AM-Dofasco ................................................................................................................................. 56
Table 3-2: Experimental variables and their ranges. .................................................................... 63
Table 4-1: XRF results comparison between Dofasco and U of T analysis. ................................ 66
Table 4-2: Summary of WDS analysis of RO phase in the raw slag (42 points) ......................... 68
Table 4-3: Recovery, grade, and S.E. of iron recovery from raw (B=110 mT). ........................... 69
viii
Table 4-4: Comparison of magnetic separation performance (via 110 mT) between raw slag and
optimally oxidized slag. ................................................................................................................ 86
Table 4-5: XRF result of separation products under optimum conditions, comparing with raw
slag (wt%). .................................................................................................................................... 90
Table 4-6: Summary of separation performance under optimum conditions, and the effect of
particle size. .................................................................................................................................. 90
Table 4-7: Effect of silica addition (change in basicity) on grade of feed and concentrate ......... 99
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List of Figures
Figure 2-1: Annual production of various slags. ............................................................................ 4
Figure 2-2: Thermodynamically predicted and actual loss of copper in smelter slag [1] ............... 6
Figure 2-3: Schematic of slag generation from general ironmaking and steelmaking process [4],
[13]. ................................................................................................................................................. 8
Figure 2-4: World annual production of iron and steel slags from 1970 to 2017 [14] ................... 8
Figure 2-5: Schematic of stainless steel production process and slag generation points. ............. 10
Figure 2-6: Flowsheet of ferrochrome production process. .......................................................... 12
Figure 2-7: Flowsheet of ferromanganese alloy production and slag generation points .............. 12
Figure 2-8: Slag generation points in production of non-ferrous metals [4] ................................ 16
Figure 2-9: Schematic diagram of jigging operation[67]. ............................................................ 22
Figure 2-10: Magnetization versus applied magnetic field strength for idealized (a) paramagnetic,
diamagnetic and (b) ferromagnetic minerals [70]. ........................................................................ 25
Figure 2-11: Schematic diagram of spin interaction of several common magnetism types, namely
(a) paramagnet, (b) ferromagnetic, (c) antiferromagnetic, (d) canted antiferromagnetic (or weak
ferromagnetic), (e) ferrimagnetic ordering. These schematic diagrams are 2D illustrations of 3D
phenomena. ................................................................................................................................... 26
Figure 2-12: (a)Schematic diagram of A2+B3+2O4 normal spinel [72] (b) Anion X2- in the spinel
structure with its nearest cation neighbors [73]. ........................................................................... 28
Figure 2-13: Simplified schematic diagram for magnetic separation [75]. .................................. 32
Figure 2-14: An AC slag-settling furnace. A furnace of this size cleans 1000–1500 tonnes of slag
per day[1]. ..................................................................................................................................... 38
Figure 3-1: SEM BSE image of steel slag sample ........................................................................ 56
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Figure 3-2: SEM-EDS elemental mapping of steel slag sample ................................................... 56
Figure 3-3: XRD spectrum of the steel slag sample. .................................................................... 57
Figure 3-4: Particle size distribution of the pulverized slag. ........................................................ 58
Figure 3-5: Scheme of the experimental steps .............................................................................. 59
Figure 3-6: Schematic diagram of the experimental setup in solid state oxidation of slag. ......... 59
Figure 3-7: Schematic diagram of the setup used in liquid state oxidation of slag. ..................... 61
Figure 3-8: Schematic diagram of the Davis tube magnetic separator. ........................................ 62
Figure 4-1: BSE image of raw slags with phases identification of: points A & E, RO phase, their
compositions are given in the table (balanced with O wt% and minor Ca wt%); point B & F:
C4AF phase; C: C2S and C3S phases; D: metallic iron. .............................................................. 67
Figure 4-2: Processed image for phase distribution of Fe-species in raw steel slag. .................... 68
Figure 4-3: Schematic diagram of magnetic separation. .............................................................. 69
Figure 4-4: XRD patterns for the magnetic (a) and non-magnetic (b) fractions of the raw slag. . 70
Figure 4-5: Phase diagram of slag (composition given above), calculated by FactSage 7.0.
C3MS2: Ca2MgS2O8; C2AF: Ca2(Al, Fe)2O5; C3MA4: Ca3MgAl4O10. ...................................... 72
Figure 4-6: Phase diagrams of synthetic slags (composition given) constructed by Semykina et
al.[102] using FactSage 6.4. (Temperature unit is K, pressure unit is log(Pa) and air atmosphere
is represented by line A) ............................................................................................................... 73
Figure 4-7: Proportions of different minerals in slag equilibrated with air against temperature. . 74
Figure 4-8: Proportions of different minerals in slag with 40 wt% FeO. ..................................... 75
Figure 4-9: Proportions of different minerals in slag against basicity (pO2=0.21, T= 1000 °C). .. 75
Figure 4-10: XRD patterns for selected CO2-roasted samples at (a) 600°C and (b) 700°C. ........ 77
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Figure 4-11: XRD patterns of raw slag and slag roasted with CO2. ............................................. 78
Figure 4-12: Concentrate fraction as a function of time and temperature for slags roasted in air
and CO2 and subjected to 100 mT magnetic separation. .............................................................. 78
Figure 4-13: Concentrate fraction (% magnetic fraction of feed) vs. magnetic field strength
(roasted in air at 900°C for 1hr). ................................................................................................... 79
Figure 4-14: Effect of magnetic field strength on iron grade and recovery as well as separation
efficiency (roasted in air at 900°C for 1hr). .................................................................................. 80
Figure 4-15: Effect of roasting time on grade, recovery and S.E. of iron in the concentrate (air
roasting at 900°C, B=110 mT). ..................................................................................................... 81
Figure 4-16: Effect of roasting temperature on iron grade, recovery, and separation (air roasting
for 1 hour, magnetic field strength=110 mT)................................................................................ 82
Figure 4-17: Calculated percentage magnetite formation based on mass gain vs. temperature
(roasting time = 1 h). ..................................................................................................................... 83
Figure 4-18: XRD patterns for 900 °C, 1hr slag sample (a) and its magnetic fraction (b) and
nonmagnetic fraction (c), separated via 110 mT. ......................................................................... 84
Figure 4-19: XRD patterns of raw slag, and slag roasted in air at different conditions. .............. 85
Figure 4-20: Grade of iron, manganese and magnesium in concentrate with magnetic field
strength from 55 to 125 (roasting in air at 900°C, for 1hr). .......................................................... 86
Figure 4-21: Different particles generated during size reduction[104]. ....................................... 87
Figure 4-22: BSE image (200 magnification) of magnetic fraction of steel slag roasted in air at
900°C, 1hr and magnetically separated under 110 mT. Type 1 to 4 particles are marked. 1a
particles: RO phase rich in iron; 1b particles: RO phase rich in magnesium. .............................. 88
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Figure 4-23: BSE imaging (200 magnification) for nonmagnetic fraction of steel slag roasted in
air at 900°C, 1hr and magnetically separated under 110 mT. A: C2S; B: C3S; C: C2F; D: Free
lime. .............................................................................................................................................. 89
Figure 4-24: Effect of roasting temperature on Mn grade, recovery and SE (air roasting for 1
hour, magnetic field strength 110 mT). ........................................................................................ 91
Figure 4-25: Cross-sections of slag containing crucibles after treatment at different temperatures
(holding time 30 minutes). ............................................................................................................ 93
Figure 4-26: XRD pattern comparison of air treated slag at 900°C and 1550 °C. 2θ ranges from
35 to 44 degrees. ........................................................................................................................... 94
Figure 4-27: Effect of roasting temperature (a) and time (b) on iron grade, recovery, and
separation efficiency (air roasting, B=110 mT, roasting for 30 mins for (a)). ............................. 95
Figure 4-28: BSE imaging (2000 magnification) for steel slag oxidized by air at (a) 1550 °C for
15 minutes; (b) 1550 °C for 30 minutes; (c) 1400 °C for 30 minutes and (d) 1500 °C for 30
minutes. A (light gray region): magnetized RO phase; B (black region): C2S, C3S; C (dark gray
region): C2F, C4AF region. .......................................................................................................... 96
Figure 4-29: BSE images for steel slags oxidized with air at 1550 °C for 30 minutes (a)
C/S=2.91, original slag; (b) C/S = 2.5; (c) C/S = 2 and (d) C/S = 1.5. ......................................... 97
Figure 4-30: Effect of basicity on % concentrate, grade, recovery and S.E. of iron (air oxidation
at 1550 °C for 30 minutes and B=110 mT). ................................................................................. 98
xiii
List of Acronyms
BF = Blast furnace
BOF = Basic oxygen furnace
LD Slag = Slag from basic oxygen converter
EAF = Electric arc furnace
LF = Ladle metallurgy furnace
SAF = Submerged arc furnace
MOR = Manganese oxygen refining
C2S = Dicalcium silicate, 2CaO·SiO2
C3S = Tricalcium silicate, 3CaO·SiO2
C2F = Dicalcium ferrite, 2CaO·Fe2O3
C4AF = Tetracalcium alumino-ferrite, 4CaO·(Al,Fe)2O3
RO Phase = FeO–MnO–MgO monoxide solid solution
Chapter 1
Introduction
Metallurgical slags are produced at an immense rate of ~ 700 million tonnes annually. Traditionally
regarded as a waste, their disposal is increasingly difficult as a result of more stringent
environmental regulations, and the amount of valuable minerals that they contain. Metallurgical
slags can be divided into three large categories based on their composition and source: blast furnace
(BF) slags, steel slags, and non-ferrous slags.
Blast furnace slags constitute more than half of all the slag produced in the metals industry. Their
use in cement production however makes them less of a waste problem and more of a by-product.
On the other hand, due to the high value of non-ferrous metals and depletion of high grade ores of
such metals, slag is considered a secondary source of them. As a result, complex and costly
processes are becoming increasingly viable options to treat non-ferrous slags. However, effective
recycling of steel slags, as the second largest group of slags (after BF) remains a problem. On one
hand, sophisticated processes may not be economically justified for treating these slags and on the
other hand, presence of heavy metals, free lime, and iron oxide does not allow their use in typical
construction applications.
This research is part of a larger study on possibility and optimization of slag atomization as a
comprehensive method that not only fragments slag into small particles with little use of energy,
but also a facilitator to allow recovery of both heat and metallic values from slag. The present work
aims at understanding the oxidation of slag during air atomization and the effect on distribution of
iron species. In other words, it tries to answer the following question. Is it possible to manipulate
the atomization process, for example by controlling the atmosphere or cooling conditions, so that
iron species report to a magnetic phase, and can then be subsequently separated from the granulated
slag. The study explores this by oxidizing both solid and liquid slags, in an attempt to simulate the
atomization conditions where slag goes through a phase change.
This thesis contains five chapters. In Chapter 2, a survey of literatures with regard to production
and characteristics of slags, as well as methods of their processing is presented. The details of raw
materials and experimental procedures are discussed in Chapter 3. The results are presented and
2
discussed in Chapter 4. Finally, in Chapter 5, the major findings of the work are summarized and
suggestions for future works related to the current study are made.
3
Chapter 2
Literature Review
2.1 Introduction
Slag is a byproduct of metal extraction and refining processes. It is predominantly a mixture of
oxides generated in smelting or refining process, and originates from gangue materials of the ore,
added fluxes, products of chemical reactions, and erosion of refractories. There are mainly two
types of slags, distinguished by their primary metallic ore bodies: ferrous slags, generated in iron
and steelmaking processes or ferroalloy production processes; and non-ferrous slags, generated in
the extraction of other metals.
Pyrometallurgical processes produce massive amounts of slag, most of which are currently
stockpiled in slag dumps, pending on favorable economic condition and advancing technology for
utilization. Proper utilization of slag remains one of the major challenges of the metals extraction
industry.
In terms of minerology, slags contain valuable gangue materials: silicates, aluminosilicates, and
calcium-alumina-silicates, thus are an economical secondary feedstock to manufacture products in
cement industry and construction materials. In addition, many slags contain a quantity of metals,
which if recovered can constitute a valuable secondary resource. Thus, when it is practically
feasible, the metal values must be recovered before disposal or other recycling applications.
Over the years, an increasing number of studies have been carried out on the topic of metal
recovery from metallurgical slags. This report starts with reviewing chemical and mineralogical
characteristics of slags, followed by a discussion on the routes and scale of slag production
including ferrous and non-ferrous slags. The subsequent sections then summarize the major
processes and treatment options for metal recovery from slag. These treatments are generally
categorized into broader physical and chemical methods. The scope of this report is to discuss the
metal recovery methods for slags of steelmaking, stainless steel, ferroalloy, copper and nickel.
Blast furnace slags are not covered as there is little metal in such slags. Also, their complete
utilization in cement making is widespread.
4
2.2 Slag Production and Characterization
This chapter presents an overview of metallurgical slags with respect to their production process,
production rate, and chemical and mineralogical constituents. The overall annual production rate
of slags is shown in Figure 2-1. Since slag is not a mined material, accurate slags production data
are usually unavailable. They are often estimated using the metal production amounts and specific
slag rate of the given process. Based on this analysis, the estimated global slag production is nearly
680 Mt per year. According to this chart, ferrous slags constitute 90% of the total slag generation.
In comparison, 74 million tonnes of slags are from non-ferrous processes. However, it is to be
noted that due to the greater slag rate for production of non-ferrous metals, the production per plant
in non-ferrous operations is comparable to ferrous ones, making them equally or even more
attractive from slag utilization perspective.
Figure 2-1: Annual production of various slags.
The typical composition and generation rate of slags for the major metal extraction processes are
provided in Table 2-1and Table 2-2. Besides the main components (CaO, SiO2, MgO and Al2O3),
metallurgical slags usually contain metals in compounds form, which are commonly thought as
losses of the target product. Metal values in slags are generally present in two forms: (a) metal
alloy droplets which are mechanically trapped in the slag phase, and (b) metals oxides or sulfides
chemically dissolved in the slag. Using the copper extraction process as an example, the Cu in
smelting and converting slags exists as dissolved Cu in the form of oxide or sulfide, or as entrained
200 Mt
356 Mt
610 Mt
66 MtSteel Slag
Iron Slag
Stainless Steel Slag
Ferroalloy Slag
Copper Slag
Nickel
Other
Ferrous Slag
5
droplets of matte, as shown in Figure 2-2. The difference between operation data and chemical
copper loss is mechanically entrained copper matte or metals. The dissolved Cu is associated with
O2- ions (i.e., Cu2O), or with S2- ions (CuS) [1].
In the following sections, the slag generation points, annual production rates and applications of
slag from various metal extraction/refining processes are discussed.
Table 2-1: Chemical composition (wt%) and production rate of ferrous and non-ferrous
slags [2][3][4][5][6].
Slag Type CaO SiO2 Al2O3 MgO MnO FeOx
Typical Slagging Rate
(t slag /t crude metal)
BF Slag 34-43 27-38 7-12 7-15 0.15-0.76 0.2-1.6 0.3
Steel Slags
BOF 45-60 9-22 1-7 5-17 2-14 10-35 0.126
EAF 35-60 9-20 2-9 5-15 1-8 5-30 0.169
LF 30-60 2-35 5-35 1-10 0-5 0.1-15 0.030
Copper Slag 2-17 25-40 4-14 1-3 0.3-2 30-50 2.2
Ni-(Sulphide ore) 3 36 6 3 0.7 46 5.3
Ni-(oxide ore)/ FeNi - 54 2 32 0.06 11 14
FeCr 2 30 26 23 - 5 1.2
Table 2-2: Typical metal values (ppm) in ferrous and non-ferrous slags [4][5]
Slag Type Cu Cr Co Ni Zn
BF Slag 15.9 1032 9.5 3 79.5
Steel Slags 114 4798 8 153 748
Copper Slag 25088 445 3317 70.9 36314
Ni-(Sulphide ore) 140 - 1293 2762 187
Ni-(oxide ore)/ FeNi 2200 13400 700 6907 75.2
6
Figure 2-2: Thermodynamically predicted and actual loss of copper in smelter slag [1]
2.2.1 Ferrous Slags
2.2.1.1 Ironmaking and Steelmaking Slags
Figure 2-3 shows a simplified process flow diagram of iron and steelmaking processes. As seen,
slags are generated in each operating unit that handes liquid steel/iron. Figure 2-4 presents the
trend in global production of iron and steel slag from 1970s to 2017, calculated based on the World
Steel Association data. A rapid growth of slag generation was observed after 2000, as China’s
significant economic growth and development of infrastructure triggered high demand for iron and
steel. The total amount of iron and steel slag production adds up to more than 560 million tonnes
annually.
The future trends in production of iron and steel slag is not entirely certain due to rapid changes in
technology. While steel production will continue to rise, shift from BF-BOF to EAF will result in
lower slag generation per tonne of steel, a trend that has been seen in the US. A market report
titled “The Future of Ferrous Slag to 2022” from Smithers Apex in 2012 [7] shows an estimated
slag market of 447 Mt for BF, 140-145 Mt for BOF, and 50 Mt for EAF. The report comments that
the total output of ferrous slag is expected to increase slowly, or even to plateau or decline
marginally. The utilization of slag is however increasing due to stricter environmental regulations,
greater diversity in slag products, and more attractive economics. The ferrous slag market in 2014,
was estimated to be 30 billion USD.
7
The difference in chemical composition of each slag presents different environmental risk and
utilization options. Due to the very reducing atmosphere, BF slag contains little metal value. In
contrast, steelmaking slags are typically rich in Fe, Mn and Cr, mostly as oxides due to the
oxidizing atmosphere, and intense mixing between slag and metal. The composition of steel slags
from a number of operations is given in Table 2-3. These represent metal in the oxide form, noting
that varying amounts of metallic Fe (7-30% steel by weight [6] [8] [9]) is also present in the steel
slag. Furthermore, nowadays, steelmaking process increasingly uses scrap as a charge material
which leads to higher concentration of alloying elements (Ni, Mn, Cr, etc) in the slag. When
treating high phosphorus ores, the converter slags may end up with high levels of P, making them
suitable for feedstock of fertilizer [10].
Because the chemical composition of steel slags is highly variable, the mineralogy of slag is not
consistent. Olivine ((Mg,Fe)2SiO4), merwinite (Ca3Mg(SiO4)2), dicalcium silicate or C2S
(2CaO·SiO2), tricalcium silicate or C3S (3CaO·SiO2), tetracalcium alumino-ferrite or C4AF
(Ca4(Al, Fe)2O7), dicalcium ferrite or C2F (Ca2Fe2O5), RO phase (FeO–MnO–MgO solid solution)
and free lime are common in steel slag [9]. The most common phases found in LD slags are
metallic iron, C2S, C2F and wüstite (FeO) [11] Iron oxides in steel slag is usually in the form of
magnetite (Fe3O4) or wustite [9]. The presence of C3S, C2S, C4AF, and C2F renders cementitious
properties to the steel slag. The reactivity of steel slag increases with its basicity. The C3S content
in steel slag is much lower than in Portland cement. Thus, steel slag can be regarded as a weak
Portland cement clinker [12].
8
Figure 2-3: Schematic of slag generation from general ironmaking and steelmaking process
[4], [13].
Figure 2-4: World annual production of iron and steel slags from 1970 to 2017 [14]
9
Table 2-3: Steel slag composition from literature
Slag Type Constituent (wt%)
Reference Fe Total FeO CaO SiO2 Al2O3 MgO MnO P2O5 TiO2 C/S
BOF
15.8 20.3 38.6 19.3 2.7 8.1 7.5 - - 2.0 [9]
18.2 23.5 45.0 11.1 1.9 9.6 3.1 - - 4.1 [15]
23.6 30.3 39.4 12.0 2.2 9.7 2.7 1.0 - 3.3 [16]
16.2 20.8 52.3 15.3 1.3 1.1 0.4 3.1 - 3.4 [17]
Low C 24.6 31.6 42.3 10.8 1.2 6.9 2.3 1.2 - 3.9 [18]
High C 15.9 20.5 44.7 14.2 3.4 8.2 1.8 1.7 - 3.1 [18]
21.9 28.2 36.7 13.1 4.6 8.7 5.4 1.1 0.5 2.8 [19]
14.3 18.4 52.4 12.8 1.4 5.2 2.9 2.3 0.7 4.1 [20]
19.0 24.5 42.2 9.3 1.4 10.8 4.0 0.5 - 4.5 [21]
High V 17.9 23.1 51.0 11.3 1.6 1.2 3.7 - - 4.5 [21]
21.2 27.2 40.1 17.8 2.0 6.3 6.5 1.1 - 2.3 [22]
EAF
12.5 16.1 40.8 17.8 4.2 8.5 9.8 0.7 - 2.3 [22]
19.2 24.6 38.8 14.1 6.7 3.9 5.0 - - 2.8 [15]
18.4 23.7 35.7 17.5 6.3 6.5 2.5 - 0.8 2.0 [23]
LD
1.6 2.0 42.5 14.2 22.9 12.6 0.2 - - 3.0 [15]
16.0 20.6 52.7 12.6 1.4 6.1 3.0 2.7 0.6 4.2 [11]
0.6 0.8 58.1 9.5 15.7 4.8 0.1 - 1.2 6.1 [24]
17.0 21.9 47.7 13.3 3.0 6.4 2.6 1.5 0.7 3.6 [25]
12.3 15.8 44.3 6.9 1.2 1.0 4.5 1.9 1.0 6.4 [24] [27]
0.7 0.9 49.6 14.7 25.6 7.9 0.4 0.2 - 3.4 [20]
5.9 7.6 47.5 4.6 22.6 7.4 1.0 0.1 0.3 10.2 [16]
10
2.2.1.2 Stainless Steel Slag
Global stainless steel production increased from 29.8 million tonnes in 2007 to 34.7 million tonnes
in 2014, and to 45.8 million tonnes in 2016 [14].The significant growth is mainly driven by the
Chinese demand and production. In addition to electric arc furnace, stainless steel production
process has a few extra refining units including AOD (argon-oxygen-decarburization furnace) or
VOD (vacuum-oxygen-decarburization furnace), as shown in Figure 2-5. Approximately one
tonne of stainless steel slag (SS slag) is generated per 3 tonnes of stainless steel amounting to ~ 15
million tonnes of SS slag is 2016.
Figure 2-5: Schematic of stainless steel production process and slag generation points.
Table 2-4 shows the analysis of several SS slags. High concentration of chromium in these slags
(or in ferrochrome slag) is a major environmental concern. Chromium is alloyed into stainless steel
to provide the corrosion resistance. Converter in stainless steel making uses oxidation reaction to
refine molten metal, resulting in oxidation of Cr into the slag [28]. Due to the carcinogenic effect
of Cr6+ and the leaching properties of Cr-containing slags, Cr content are tightly specified in many
industrialized countries. It is reported that there is almost no utilization of slags from stainless
steelmaking, due to this limitation [21]. Thus, many investigations have been carried out on
chromium extraction from stainless steel (or ferrochrome slags), to not only recover the metal
values, but also to address this concern and allow the use of remaining slag as cement or
construction materials [28]- [29]. Mineralogical studies of SS slags by Shen et al. indicate that SS
slag contains about 2 to 4 % Cr, 1 to 2.6% Mn and 0.2% Ni. About 70% Fe and Cr is in the form
of oxide, whereas Ni and Mn in SS slag are all in the form of Cr containing alloys[30].
11
Table 2-4: Stainless steel slag composition from literature.
Constituent (wt%) Range
Cr2O3 2.66 1.72 3.4 2.9 1.8 6.74 0.81 0.8 – 2.9
FeO 5.2 0.93 0.8 1.4 1.7 2.31 0.54 0.5 – 5.2
CaO 44.57 42 42.1 46.9 54.1 38.8 56.0 38.8 – 56.0
SiO2 33.06 26 33.6 33.5 26.5 30.4 25.9 25.9 – 33.6
Al2O3 3.59 2.1 1.87 2.3 4.9 5.15 3.84 1.9 – 5.2
MgO 8.33 11 10.5 6.2 6.3 8.47 6.88 6.2 – 11.0
MnO 0.44 0.58 0.7 2.6 1.0 3.95 0.57 0.4 – 4.0
TiO2 2.15 0.83 1.23 0.2 1.0 - - 0 – 2.2
C/S 1.35 1.62 1.25 1.40 2.0 1.3 2.2 1.3 – 1.6
Reference [28] [29] [31] [32] [32] [21] [21]
2.2.1.3 Ferroalloy Slags
Ferroalloys are iron based alloys with high content of various elements and include FeCr,
ferrochrome; FeMn, ferromanganese; FeSiMn, silicomanganese; FeNi, ferronickel; FeSi,
ferrosilicon; FeTi, ferrotitanium; and FeV, ferrovanadium. The global production of all ferroalloys
increased from 46.8 million tonnes in 2012 [33] to 57.6 million tonnes in 2014 [34]. Most of the
ferroalloys serve as alloy feedstocks in production of steel, stainless steel, and other grades of alloy
steels. As a result, ferroalloys production has been growing since a large market is opening for
high quality steels and alloy steels.
Ferrochrome slag
Ferrochromium like other ferroalloys is mainly produced by direct carbothermic reduction in
submerged arc furnace, as shown in Figure 2-6. It is categorized by its carbon content into three
groups: high, medium, or low carbon FeCr. The high-carbon ferrochrome (HC FeCr) including
charge chrome (produced from low grade ores) is the most dominant form of ferrochrome,
accounting for 95% of the total production [35][36]. It is reported that the global charge chrome
(HC FeCr) production was 11.1 million tonnes in 2014 [37]. Given the slag rate of 1.2 t slag/t
crude metal [5], the estimated HC FeCr slag production was greater than 13 million tonnes in 2014.
12
Figure 2-6: Flowsheet of ferrochrome production process.
Ferromanganese slag
In the year 2014, world production of silicomanganese and ferromanganese production was 12.7
million tonnes and 6.7 million tonnes respectively [33]. Around 1.2–1.4 t of slag is typically
generated for every tonne of SiMn alloy produced; for FeMn this value is around 0.8 to 0.9 t of
slag for each tonne of FeMn alloy [38][39]. However, many conventional processes tend to recycle
Mn-rich slag back to the process. Thus, for slags of Mn alloys, only silicomanganese alloys are
considered, amounting to ~ 15 million tonne of slag in 2014. SiMn slag consists of oxides such as
SiO2, CaO, Al2O3, MgO and MnO. The MnO content of SiMn slag ranges between 6 and 10%.
The slags are mostly amorphous after solidification [38].
Figure 2-7: Flowsheet of ferromanganese alloy production and slag generation points
13
Ferronickel slag
Ferronickel is commonly produced from nickel lateritic ores through pre-reduction in rotary kilns
followed by smelting in EAF, a process flow similar to Figure 2-6. As the laterite ores are not
upgraded before smelting, the slag generation rate is enormous at around 14 t slag/ t FeNi alloy.
With global ferronickel alloy production of 3.9 Mt, estimation of slag production is found to be 54
Mt. [33]
2.2.2 Non-ferrous slag
Figure 2-8 shows a simplified process flow diagram with slag generation points in non-ferrous
metal production processes. Flash smelting was developed by Outokumpu and has been used in
the copper industry since 1949. It has replaced most reverberatory furnaces, and is now known as
the “conventional” copper smelting process. According to the World Copper Factbook 2014, in
the year of 2013 world, the copper smelter production reached 16.8 million tonnes of copper.
Considering the typical proportion of 2.2 tonnes of slag per ton of blister copper, about 37 million
tonnes of copper smelting related slags including smelter slag and converter slag are generated
[40]. Table 2-5 shows the composition of copper slags in several operations, ranging from 0.2 to
3.35% Cu with an average of 1.5% Cu. Converter slag has a wide range of composition and
generally much higher Cu content than smelter slags [41]. It also contains large contents of Co,
Ni, and Zn (some contain as high as 4% Co or 8.9% Zn), appreciable amounts of precious metal
and large amounts of iron and silica. In some cases, the molybdenum content is up to 0.4%,
representing the most valuable elements in the slags[42].
Table 2-6 lists the major mineral phases in copper slags. The dominant phases are fayalite
(Fe2SiO4) and magnetite. Some slags contain mainly glass phase, due to different thermal history.
Many studies conclude that most of the copper (80%) was in the form of copper sulfides (or
bornite, Cu5FeS4), and cobalt, zinc and nickel were uniformly disseminated as ferrites and silicates
[43][44][41]. However, some report that copper is mainly dispersed within the iron and silicate
phases[45]. This shows large variety of slags’ mineralogy, mainly due to their thermal and
chemical history. The effect of different cooling rates on slag mineralogy will be discussed later.
The difference in the mineralogy can affect the separation process of metals as will be discussed
later.
14
Table 2-5: Summary of copper slag composition from literatures.
Constituent Composition (wt%) (Continue to Next Page)
Slag 1 Slag 2 Slag 3 Slag 4 Slag 5 Slag 6 Slag 7 Slag 8 Slag 9 Slag 10 Slag 11 Slag 12
Cu 1.22 1.61 1.43 1.53 1.76 1.35 0.60 3.35 0.2 1.42 1.40 0.98
Fe 32.24 35.80 20.70 39.09 46.46 28.43 42.80 50.20 21.0 41.51 22.54 51.47
Co - - 0.72 0.04 0.19 4.09 - - 1.7 0.70 2.00 0.49
Ni - - - 0.02 0.23 0.04 - - NR 0.01 - 0.004
Zn - - 8.90 - - 1.70 - - 5.0 3.00 8.03 0.23
Pb - - - - - 1.16 - - 0.8 0.19 0.46 -
S - - 0.59 1.24 - 0.11 - 0.58 0.4 0.68 0.60 0.67
Si - - 15.37 14.45 33.57 15.38 13.30 13.80 14.4 11.53 14.49 9.94
Ca - - 6.26 2.82 2.09 5.13 - 0.15 6.1 3.77 7.15 -
Mg - - 2.53 1.70 0.97 2.15 - - 3.4 0.56 3.02 -
Al - - 2.56 1.22 1.58 3.22 1.06 - 3.2 1.51 1.51 -
Type Smelter slag
Slag 1: Reverberatory slag, Refimet, Chile[46]; Slag 2: Flash smelter Slag ,Codelco Chile Chuquicamata Division[46]; Slag 3: smelter
slag from Lubumbashi, Democratic Republic of Congo[44]; Slag 4: Smelter slag, Birla Copper, India [47]; Slag 5: smelter slag, Copper
Complex, Ghatsila, India[48]; Slag 6: smelter slag, Guangdong, China [43]; Slag 7, India[45]; Slag 8: smelter slag, Chile [42]; Slag 9:
Smelter slag, water granulated, GCM, Congo[49]; Slag 10: Flash smelter slag, Yanggu Xiangguang Copper Co., LTD., China[50]; Slag
11: Smelter slag, Lubumbashi plant, Congo[51]; Slag 12: Smelter slag, Küre, Turkey [52]; Slag 13: Flash smelter slag, Chile [53]; Slag
14: mix of smelter and converter slag, Black Sea Copper Works, Turkey[41]; Slag 15: convertor slag, Copper Complex, Ghatsila,
India[48]; Slag 16: converter slag, Chambishi Copper, Zambia[54]; Slag 17: converter slag, Legnica, Poland [55]; Slag 18: converter slag
Atlantic Copper Company, Spain [56]; Slag 19: converter slag, Ergani Copper Co. of Etibank, Turkey [57]. “-” as not reported.
15
(Table 2-5 Continued)
Constituent Amount (%)
Slag 13 Slag 14 Slag 15 Slag 16 Slag 17 Slag 18 Slag 19
Cu 2.27 2.64 4.03 22.87 1.70-5.60 9.13 2.40
Fe 41.30 47.20 38.32 32.35 30.40-
38.00 40.03 50.30
Co - 0.10 0.49 0.60 0.90-1.45 - 0.38
Ni - 0.07 1.97 - 0.045-
0.095 0.01 -
Zn - 0.67 - - 3.10-6.40 1.36 -
Pb - 0.13 - - 3.42-7.61 - -
S 0.83 1.30 - 3.49 0.17-0.78 1.07 2.92
Si 15.40 8.50 34.32 10.63 11.92-
15.35 - -
Ca 0.49 - 4.01 - 0.31-1.78 - -
Mg - - 2.66 - 0.13-1.17 - -
Al 1.60 - 0.08 - - - -
Type Smelter
slag Mixed slag Converter slag
16
Similar conditions apply to nickel extraction, since most nickel is present in similar form with
copper in sulfide ores. Nickel sulfide extraction process typically has a much higher slag rate.
Twomillion tonnes of primary nickel is produced, while 60% output is from nickel sulfide smelters.
At slagging rate of 5.3 t slag/t nickel, the nickel slag annual production is estimated to be 6.4
million tonnes. The estimation might have some discrepancy as the two metal processes can have
overlaps.
Figure 2-8: Slag generation points in production of non-ferrous metals [4]
Table 2-6: Mineralogy of copper slags. Slag # refers to those in Table 2-5.
Slag # Major Phases
Slag 1 & slag 2 Amorphous: Silicate of Fe, Ca and K with Al, Ti, Zn and traces of Cu;
fayalite: Iron silicate with Ca, K, Al, Ti and Zn
Slag 6 Fayalite Fe2SiO4
Slag 7 Fayalite Fe2SiO4; magnetite Fe3O4
Slag 10 Fayalite Fe2SiO4; magnetite Fe3O4(Some copper detected lost in
magnetite as lattice substituents); augite Ca(Fe,Mg)Si2O4
Slag 12 Fayalite Fe2SiO4 (60-65%); wustite FeO (15-17%)
Slag 14 (water granulated) Amorphous (94-96%); crystalline (4-6%): hedenbergite, magnetite and
fayalite
Slag 16 (converter slag) Fayalite Fe2SiO4, magnetite Fe3O4; metallic copper; chalcocite Cu2S
and quartz SiO2.
2.2.3 Slag Utilization
The traditional method of slags treatment is dumping which is not desired for several reasons: it
occupies land, is associated with loss of useful mineral resources, and damages the environment
[58]. For example, the exposure of non-ferrous slags of Penn Mine in Calaveras County
17
(California) to the ground and flood waters has resulted in concentrations of Cd, Cu and Zn in
surface waters in excess of USEPA (United States Environmental Protection Agency) chronic
toxicity guidelines for the protection of aquatic life [59]. It is therefore clear that treatment of slag
for complete utilization or conversion it to environmentally-benign material is highly desired, and
is increasingly an obligation. This has driven numerous investigations in the past few decades on
slag processing options. There are three major directions for slag treatments: recycling within the
plant, recovery of valuable elements/minerals from slag, and reuse in other applications.
2.2.3.1 In-Plant Recycling of Slag
The possibility of recycling slags inside metallurgical operations is an attractive proposition.
Ideally, if slag can be recycled at the source, i.e. the plant it is produced in, subsequent handling
and processing costs are minimized, also resource utilization is maximized. Slags from different
metallurgical extraction processes may be altered to make some low-cost substitute and secondary
sources for many other metallurgical processes. For example, steel slag contains valuable
components like iron, manganese, silica, magnesia and alumina. If containing low phosphorus, it
can be upgraded using gravity or magnetite separation to produce an iron concentrate which is then
recycled to sintering process and finally to the blast furnace. Due to the large content of basic
oxides in steel slags, such recycling process saves considerable amounts of iron ore, limestone and
dolomite. It has been suggested that using one tonne of converter slag in sintering process for pig
iron production, can save 300-480 kg iron-ore, 530-620 kg limestone, 110-120 kg dolomitic
limestone, 140-180 kg of manganese ore, and 80-120 kg coke[60]. However, particularlyin iron
and steel metallurgy, the viability of such method is limited by the acceptable phosphorus content
in the product.
Due to high copper lockup in copper converter slag, slag cleaning to increase Cu recovery is often
required. Some plants practice matte settling in electric furnace and charge coke as reductant to
lower the viscosity of slag. Recycling copper converter slag back to the flash furnace has also been
practiced in some plants, leading to an increase in matte grade. In the Black Sea Copper Works
(Turkey), slag is disposed of to a slag area and left to cool for 24 hours. The cooled slag is then
crushed and treated with flotation. The copper concentrate produced this way is then charged into
flash furnace. Metals such as Co and Zn are not recovered in the process and report to the tailings
[41].
18
2.2.3.2 Slag as By-Product
Slags have found a wide range of applications outside metallurgical plants including cement
production, road construction, civil engineering work, fertilizer production, landfill cover, soil
reclamation etc. The utilization of slags varies with their type and physical or chemical properties,
as well as where they are produced. In the case of iron and steel slags, various metal and free lime
contents lead to different cementitious properties. It has been reported that BF slags are almost
100% utilized in various applications and about two thirds of them is granulated [1] [2] [58]. On
the other hand, the utilization of steel slags is more challenging due to their high content of iron
oxide and free lime. For example, the Chinese steel industry generated 100Mt of steel slags in
2014, of which only 10% was used. The unused portion has been accumulated in slag dumps
throughout the years and is estimated to be around one billion tonnes in this country [2]. Within
EU and the other industrialized countries, on average more than 70% of steel slags are utilized in
road building, cement production, and marine structures. However, there are still millions of tonnes
of discarded steel slags in the EU [21]. Nonferrous slags such as copper have been attempted for
several uses such as abrasive material, road base material, and in smaller amounts as insulating
material (slag wool) [42]. Nevertheless, their utilization is still low.
2.2.3.3 Recovery of Metals and Minerals from Slag
Different techniques for recovery of the mineral/metal value of slags have been studied and utilized
over many decades. This report categories them into two main sections, physical methods and
chemical methods. In physical methods, the difference in one or several physical properties of slag
and desired metal/mineral (e.g. density, magnetic susceptibility and wettability) is used for their
separation. The chemical methods however involve at least one chemical reaction in the process
and can be divided to pyrometallurgical and hydrometallurgical methods. Since BF slag has
extremely low metal content, and has well-established uses such as cement feedstock, the report
will not investigate the opportunity of metal recovery from this type of slag.
19
2.3 Physical Processing Technologies for Metal Recovery
Various phases in solidified slags possess different physical properties, which allows their
separation from one another, thereby recovering valuable metals/compounds. The current physical
separation methods include three common steps:
1. Cooling and solidifying the hot molten slag;
2. Size reduction and screening of solid slag;
3. Physical separation by gravity separation, floatation and magnetic separation.
2.3.1 Cooling and Sizing
2.3.1.1 Slag Cooling Methods
Cooling of slag is ordinarily done in a slag yard by air cooling or water spraying, leaving large
solids lumps of slag behind. There are two major drawbacks for this process. First, in order to
recover metal from slag or use the slag as a by-product, certain size reductions are required with
significant energy consumption, dust generation and operational costs. Second, air cooling method
is very slow, increasing the land requirement, and water spraying method consumes large amounts
of water [6]. To address these problems, several more efficient cooling processes have been put
into practice. The dominant approach is to reduce the size during prior to or along cooling; smaller
particles cool faster, also the subsequent comminution needs are reduced [2].
Hot-state Scooping Process
ArcelorMittal Americas employs mechanical stirring in their cooling process [61]. Slags are
dumped and sprinkled with water for initial cooling. Front-end loaders are used to scoop up the
slag to disintegrate it and accelerate cooling, hence it is a hot-state scooping process. This method
produced slag with particle size of under 300 mm in size.
Instantaneous Slag Chill (ISC) Process
A wet cooling process, known as instantaneous slag chill (ISC) process [6], [62], has been
employed in the Japanese steel industry and adapted by many other steelmakers around the world.
In this process, the fluid slag is poured into a steel box, and rapidly cooled by spraying and
immersing in water to 700 °C. Size reduction is facilitated due to the creation of cracks in the
solidified slag. Two more water cooling steps further lower its temperature to 200°C and to 100°C.
20
BSS Process
The BSS, also known as box tumbling method was developed by Russian Ural Steel and Chinese
Bao Steel. The process combines water spraying and steel ball cooling together, slag is poured into
a tumbling drum with high revolving speed, while water is sprayed into the drum. The inner
structure of the drum has a special design to release high pressure generated from steam formation.
Steel balls having high thermal conductivity are mixed in to accelerate the cooling process coupled
with the rotating drum, it also acts as a ball mill, introducing mechanical impact for size reduction.
Jin et al. [62] report that slags are broken into a small particle size with 90% of particles smaller
than 10 mm.
Wet and Dry Granulation Processes
The granulation methods employ a high-pressure jet of water or air to break up a falling stream of
slag into small granules, a few mm each. This creates a large surface area which significantly
improves the heat transfer from slag to water/air. Also, due to the small size of granules, their
handling is facilitated and subsequent grinding needs are minimized. Granulation has been applied
to many slag types, such as copper [47], ferroalloy [5][63], and steel [6][64], but widely used for
BF slags [65]. Wet granulation is an established method but it suffers from several disadvantages
including the consumption of an excessive amount of water (10 tonnes per tonne of slag) and
generating metal-contaminated effluent and gases such as H2S. For these reasons, in the recent
years dry slag granulation methods have gained a considerable attention and are being accepted by
industry [65].
2.3.1.2 Effect of Cooling Conditions on Metal Distribution
The cooling conditions of slag determine the slag structure and mineralogy of its constituents. For
example, BF slags are often quenched rapidly to produce amorphous slag which has proper
cementitious properties if used as cement feedstock. Forming an amorphous slag also often means
that the metals remain dissolved in a “glass” matrix, thus making their recovery difficult. In one
study [18], for example, four cooling conditions were tested for a BOF slag, namely water
granulation (rapid cooling), splashing (rapid cooling), air cooling (slow cooling), and furnace
cooling (slow cooling). The authors found that with slow cooling, more crystalline phases emerged
during solidification. For example, the XRD peaks of wustite became more prominent at slower
cooling rates indicating that more iron was concentrated into iron-containing phases. The study
21
concluded that rapid cooling of slag should be avoided when recovery of iron (oxide) from the
slag is the objective. Also, when it comes to application as an aggregate, slag of crystalline
structure is preferred [60].
Chemistry has a strong effect on the slag structure. For example recent research by Esfahani and
Barati [66] has shown that slag system with lower basicity have more tendency to form an
amorphous structure during cooling.
2.3.1.3 Size Reduction
Size reduction is necessary to liberate metal and metal-bearing minerals. This is traditionally
achieved by energy intensive processes of crashing and grinding. For example, there are two major
ways for pulverizing solid slag. One is multi-step crushing that may involve jaw crusher, cone
crusher, hammer/impact crasher, followed by grinding in ball or rod mills. Another approach
involves the one-stage dry autogenous grinding. Due to the uniform structure and hardness, slags
are generally tough (i.e. energy-consuming) to crush and grind. For example copper slag has the
grindability index of 26.8 kWh ton-1, [47] which is substantially higher than typical ores and
minerals (~10-15 kWh ton-1). Steel slags have shown high value in the Los Angeles abrasion test,
a test method indicating aggregate toughness and abrasion characteristics. All of these point to
difficulty of grinding slag to a particle size small enough for reasonable liberation of
minerals/metals. Any chemical or physical modification of slag to reduce the subsequent grinding
requirements can improve the metal recovery. These include growth of the mineral phases of
interest, size reduction of slag while liquid (e.g. granulation), and weakening the structure by
introducing cracks
2.3.1.4 Screening as a Concentration Method
A combination of comminution and sizing is known as a cost-effective way of splitting materials
into lean and rich streams. This is because fracture of particles often occurs at the interface between
the phases and due to the different brittleness and/or crystallite/grain size of various phases, sizing
the crushed product results in fractions with various concentrations of different elements. For
example Ma et al.[61] found that iron grade and sulfur content in steelmaking slag fines (-12.5mm)
are strongly affected by particle sizes. Large particles contained high iron and low sulfur, whereas
small particles contained low iron and high sulfur. This was mainly associated with metallic iron
22
droplets entrained in slag phase, which are more ductile, thus they remain as large particles. Sulfur
containing compounds, on the other hand, are more brittle and generate more fines. Therefore, a
simple screening treatment can produce two streams, one being rich in iron and low in sulfur,
which can then be recycled to the steelmaking vessels or sinter feed as a source iron.
2.3.2 Gravity Separation
Gravity separation has been applied to separate materials when phases have a significant density
difference. In slag recovery processes, gravity separation has been employed to separate suspended
metal alloys or high value minerals. Machines such as spirals, shaking tables and jigs have been
employed, the underlying principle in all these methods is the different sinking velocity of particles
with various density. In jigging operation, water is used as a carrier, the jigging action imparted
by water separates heavy and light minerals, as shown in Figure 2-9. The light tailing is lifted and
carried up with the tide generated by piston, while heavy concentrate passes through jig screen and
settles to the bottom of the tank.
Figure 2-9: Schematic diagram of jigging operation[67].
Indian researchers processed ferrochrome slags [60] by first crushing it to <10 mm and then
screening into 1-10 and <1 mm size fractions. The coarse fraction was subjected to jigging and the
fine fraction to tabling. The chromium recovery from jigging was 6.7% with a grade of 59%
chromium, while the tabling produced a 29% low grade chromium concentrate. In the late 1990’s,
a South African ferroalloy producer developed a process to recover ferrochrome from slag dumps
using pneumatic jigs [68]. The report showed metal recoveries of 76% and grades of over 90% for
FeCr and SiMn slags in the size range 0.15 to 3 mm. Tata Steel [69], another ferrochrome producer
23
in India, crushed and separated its slag into coarse and fine size fractions, with the coarse fraction
(-10 +1 mm) being treated in a two-stage Duplex mineral jig, and the -1 mm, treated in shaking
tables. The recoveries exceeded 85% and the grade was +58% Cr. From its granulated slag with
12% Cr grade, employing magnetic and jigging separation, the company can recover 90% of Cr
into a concentrate with 59% Cr grade [63]. The disadvantage of gravity separation is massive
consumption of water. There are also concerns of heavy metal leaching. Trace of Cr6+ was detected
in plant water during ferrochrome slag jigging in South Africa, reported as 1.6 mg L-1.
2.3.3 Floatation
Flotation is a viable method to separate hydrophobic mineral (e.g. sulfides) from slag. As
mentioned earlier, the floatation process is commercially used in some copper plants to upgrade
converter slag, the obtained concentrate is fed into smelting furnace. It can also be used to upgrade
low grade copper smelter slag.
Kas et al. [47] demonstrated that in alkaline pH good selectivity is obtained for copper sulfide
minerals. With particle size of d80 = 75 µm, in the presence of sodium isopropyl xanthate (SIX,
concentration: 200 g/t), copper grade of 11 to 17% in the concentrates with 82% recovery was
achieved at pH 9. However, a research done in the same region drew a different conclusion. Panda
et al. [45] used sodium silicate, sodium isopropyl xanthate (SIX), and methyl isobutyl carbinol
(MIBC) as dispersant, collector and frother respectively. Their experiment showed a limited
recovery despite high dosage of collector (600 to 2400 g/t); recovery of 42-46%, with copper grade
of 2-3%. The reason for such a significant difference in response to flotation of slag is copper
phase association: in the former study, Cu was present as Cu2O, Cu2S, CuO and free copper,
whereas in the latter work Cu was finely dispersed within the iron and silicate phases, thus not
liberated. Cu content in feed to floatation are also different (Slag 7 and Slag 13 in Table 2-5),
1.53% and 0.60%. These two cases show that the feasibility of floatation must be based on the
chemistry and history of the slag.
2.3.4 Magnetic Separation
Magnetic separation is an important method in mineral dressing. It exploits the different magnetic
properties of minerals to produce a valuable concentrate or to remove magnetic contaminants.
24
Since in the current study magnetic separation will be used for recovery of metals from steel slags,
the principles and applications of the technique are discussed in more detail.
2.3.4.1 Magnet Ordering
Placing any material close to a magnet, the applied magnetic field creates an induced magnetic
field inside the material. The applied magnetic field strength, which induces lines of force through
a material, is the magnetizing force, H (A m-1). The intensity of magnetization or the magnetization
induced in the material, M (A m-1) can also be thought of as the volumetric density of induced
magnetic dipoles in the material. The magnetic response of a material to an external magnetic field
has a simple relationship as:
M = H (1)
The ratio of the magnetization (per unit volume) to the magnetic field, , is the magnetic
susceptibility for the material. The dimensionless volume magnetic susceptibility, which is often
shown as v. For mineral processing propose, mass magnetic susceptibility (m) is more frequently
used. Measured in units of m3kg-1, mass susceptibility is used to describe the ratio of the material
magnetization per unit mass to the external field intensity [70].
All materials are affected by external magnetic field; thus, all have a certain magnetic susceptibility.
The magnetic susceptibility is a material property and can either be positive (paramagnetic) or
negative (diamagnetic). As shown in Figure 2-10 (a), chromite is a paramagnetic material with a
positive slop on M vs. H diagram, while quartz behave as diamagnetic with a negative slop.
Diamagnetic materials are repelled by an external magnetic field. Diamagnetism is a universal
behavior to all matters, whereas electrons try to shield the material from external magnetic field
by generating an extra opposite magnetic moment towards the external magnetic field. This will
cause the material to move away from regions of high magnetic field. The magnetic susceptibility
for diamagnetic materials is usually small, but temperature independent. Due to its weak effect,
diamagnetic substances are often referred as “non-magnetic”[71].
A paramagnetic material must have unpaired electrons in the molecular orbitals, introducing a net
magnetic dipole moment in the material. In the absence of an external field, individual magnetic
dipoles are randomly orientated in all directions, cancelling out each other’s effect and resulting
25
in a net zero magnetization. Application of external field promotes the relative orientation of
individual magnetic moments towards the same direction of the magnetic field, increasing the
magnitude of magnetization. At high value of magnetic field, all magnetic moments are aligned
parallel to the direction of field, and magnetization reaches its saturated value (Ms). However,
thermal energy works against the dipole alignment. With increasing temperature, the value of both
magnetization and susceptibility decreases[71].
(a) (b)
Figure 2-10: Magnetization versus applied magnetic field strength for idealized (a)
paramagnetic, diamagnetic and (b) ferromagnetic minerals [70].
In special cases of paramagnetic materials, magnetic moments can interact and couple with
neighbors and spontaneously form long range ordering magnetisms, or magnetic structures, such
as ferromagnetism, antiferromagnetism and ferrimagnetism.
Saturated ferromagnetic material has only one magnetic lattice, meaning its magnetic moments
(shown as little arrows in Figure 2-11) are spontaneously aligned in one certain direction. Such
material would maintain a certain degree of magnetization even in the absence of an external
magnetic field. The spontaneous magnetization is only maintained at low temperature and the
upper limit is Curie Temperature (TC); when the temperature is higher than TC the material shows
paramagnetic behavior. From Figure 2-10 (b), because of exchange coupling between cations
through their neighboring anions, ferromagnetic materials have much higher initial magnetic
susceptibility compared to paramagnetic materials. After all magnetic spins are aligned with the
applied magnetics force, the susceptibility decreases rapidly as a material reaching its saturated
magnetization (the slopes decrease from point 1 to point 3). Typical example of ferromagnetic
26
materials are some metals and alloys, such as iron (Fe) and FeNi alloy, and many iron bearing
oxides [70], [71].
Figure 2-11: Schematic diagram of spin interaction of several common magnetism types,
namely (a) paramagnet, (b) ferromagnetic, (c) antiferromagnetic, (d) canted
antiferromagnetic (or weak ferromagnetic), (e) ferrimagnetic ordering. These schematic
diagrams are 2D illustrations of 3D phenomena.
Antiferromagnetic materials are considered to have two magnetic sub-lattices: MA and MB, of the
same magnitude. Each of them is ferromagnetic. But the magnetic sub-lattice MA is oriented in
opposite direction to MB (MA = MB). As a result, MA and MB cancel each other out, and the
material has a total magnetic moment of zero when no magnetic field is applied. The critical
temperature limiting spontaneous ordering is named Neel Temperature (TN). As T TN,
magnetization remains zero and is not affected by varying temperature. A special case of
antiferromagnetism is canted antiferromagnetic materials, where the opposite magnetic dipoles are
slighted tilted towards a direction, producing a net weak magnetic moment. One classic example
for anitiferromagnetism is Hematite (Fe2O3). Due to opposite Fe3+ cations’ magnetic moments
arrangement, hematite exhibits antiferromagnetism below 250K. It becomes a canted
27
antiferromagnetic material at room temperature, and is usually described as weakly ferromagnetic
[71].
A ferrimagnetism material contains also two antiparallel magnetic sub-lattices. The difference is
that MA and MB have various magnetic moment (MA MB). As a result, it has a directional total
magnetic moment and behaves like ferromagnetic materials exhibiting a spontaneous
magnetization. Similar to ferromagnetism, the system loses spontaneous ordering above the Curie
temperature. A typical example is magnetite (Fe3O4). Most ferromagnetic and ferrimagnetic
materials have strong magnetization at room temperature, and are considered “strongly magnetic
materials” [71].
2.3.4.2 Spinel Ferrite
The metals indicated as losses in slag mainly belong to the group of transition metals with
incompletely filled d orbitals. The unpaired electrons have low reactivity, which results in the
formation of compounds in many oxidation states, non-stoichiometric compounds, and compounds
with mixed valence. Thus, transition metal compounds also exhibit a range of magnetic behaviors.
In the case of iron oxides, the oxidation state dictates the crystal structure and magnetic behavior
as following.
• Highest oxidation state, Fe2O3 hematite has rhombohedral crystal system, and is a canted
antiferromagnetic material;
• Magnetite Fe3O4 has spinel crystal structure, exhibits ferromagnetic behavior and
possesses high magnetic susceptibility;
• Wüstite (ferrous oxide – FeO) has rock salt crystal structure, is a paramagnetic material;
and is prone to non-stoichiometry, where small portion of Fe2+ replace two thirds of Fe3+.
As one of the first magnetic material discovered by human, magnetite is well known to have the
highest magnetic susceptibility among all oxidation states of iron. Magnetite belongs to the spinel
group. The oxide spinel group is a class of minerals with a general formulation of A2+B3+2O4. As
shown in Figure 2-12(a), A and B are divalent and trivalent cations, respectively, where A cations
fill 1/8th of the tetrahedral sites (known as A site) and B cations fills ½ of the octahedral sites (B
site). The structure described above is a normal spinel structure. Another closely related structure,
namely inverse spinel exists in which A cations and half of the B cations switch place. Thus,
28
inverse spinel is formulated as B[AB]O4. Finally, there are mixed spinels which are intermediate
states between the normal and inverse spinel structures with expression as: A1-xBx[AxB2-x]O4.
Magnetite is an inverse spinel, with half of Fe3+ occupying tetrahedral sites. When trivalent iron
forms spinel with other divalent metals, the oxides formulated as MeFe2O4 (Me = Fe, Mg, Mn,
Co, Ni, Zn), are named spinel ferrites. As iron is a major constituent of slag phase, magnetite is
expected in most ferrous slags. The magnetic properties of magnetite and other spinel ferrites
determine the possible recovery of each species in magnetic separation.
(a) (b)
Figure 2-12: (a)Schematic diagram of A2+B3+2O4 normal spinel [72] (b) Anion X2- in the
spinel structure with its nearest cation neighbors [73].
A ferrite spinel’s magnetic properties mainly depend on cation distribution (normal, inverse or
mixed), the nature of magnetic moment of cations, and exchange coupling between cations. Crystal
field stabilization energy (CFSE) determines cation distribution in A or B sites. Since Fe3+ ions
have no preference for octahedral or tetrahedral sites, ferrites can be normal, inverse or mixed
spinels, depending on the crystal field stabilization energy (CFSE) of MII ion. At room
temperature, based on CFSE of Fe2+, Co2+, and Ni2+, these cations generally all have a strong
preference towards octahedral sites, resulting in inverse spinel structure. ZnFe2O4 on the other hand
is a normal spinel, where Zn2+ ions have a preference in tetrahedral sites. Mg2+ and Mn2+ ions have
no preference for either site due to their CFSE value of zero. Thus, MgFe2O4 and MnFe2O4 are
mixed spinels.
As shown in Figure 2-12(b), the nearest neighbors of an anion suggest the fashion of exchange
coupling in spinel structure: A-B interaction and B-B interaction. Due to reasons such as angular
29
difference, A-B interaction is much stronger and predominant in spinel interaction, while B-B
interaction is much smaller and A-A interaction is negligible [73].
The comparison of spin-only magnetic moment with the effective magnetic moment for some
common transition metal cations are tabulated in Table 2-7. One can notice that Sc3+, Ti4+, Cu+
and Zn2+ have zero magnetic moment, and the highest value is found for Mn2+ and Fe3+.
Table 2-7：Comparison of calculated spin-only magnetic moments with experimentally
observed data for some common transition metals, where µB is theoretical Bohr magneton,
and µeff is the observed Bohr magneton [74].
Ion d shell electrons µB µeff(observed)
Sc3+, Ti4+ 0 0 0
Ti3+, V4+ 1 1.73 1.7-1.8
V3+ 2 2.83 2.8-3.1
V2+, Cr3+ 3 3.87 3.7-3.9
Cr2+, Mn3+ 4 4.90 4.8-4.9
Mn2+, Fe3+ 5 5.92 5.7-6.0
Fe2+ 6 4.90 5.0-5.6
Co2+ 7 3.87 4.3-5.2
Ni2+ 8 2.83 2.9-3.9
Cu2+ 9 1.73 1.9-2.1
Cu+,Zn2+ 10 0 0
This is shown in Table 2-8, transition metals induce various magnetic properties in the form of
spinel ferrites which is dependent on the element magnetic moment, cation distribution and
exchange coupling. As a way of modifying slag to improve metal recovery, some of these
transition metals can be added in to form oxide compounds with iron that have higher magnetic
susceptibility. The highest magnetic susceptibility is often associated with spinel structure, giving
the greatest chance for effective separation.
Table 2-8: Cation distribution, structure type and saturation magnetization of spinel
ferrites[73].
Cation distribution Type Ms (at 0 °K)
exp. (µeff)
MgFe2O4 Mg0.1 Fe0.9 [Mg0.9 Fe1.1] O4 Mixed 1.1
MnFe2O4 Mn0.8 Fe0.2 [Mg0.2 Fe1.8] O4 Mixed 4.6
FeFe2O4 Fe3+ [Fe2+ Fe3+] O4 Inverse 4.1
CoFe2O4 Fe [Co Fe] O4 Inverse 3.7
NiFe2O4 Fe [Ni Fe] O4 Inverse 2.2
CuFe2O4 Cu0.04 Fe0.96 [Cu0.90 Fe1.04] O4 Mixed 1.3
ZnFe2O4 Zn [Fe2] O4 Normal 0
30
2.3.4.3 Magnetic Separation
Depending on the different magnetic orderings and distinct magnetic susceptibility of minerals,
separation can be achieved by varying the induced magnetic field. Since most processes are
operated near room temperature, spontaneous magnetization and susceptibility at room
temperature is crucial. Table 2-9 presents magnetic susceptibility of some selected rocks and
minerals. Categorized by their ability of magnetic separation, there are generally three groups of
minerals: strongly magnetic minerals, weakly magnetic minerals and nonmagnetic minerals.
Strongly magnetic minerals can be easily removed, even with weak magnetic induction, up to 0.2
T. The specific magnetic susceptibility of this minerals is usually greater than 5 × 10−5 m3kg-1.
This category includes ferromagnetic and ferromagnetic materials such as magnetite, maghemite,
magnesioferrite. Weakly magnetic minerals (mostly antiferromagnetic or weakly paramagnetic
materials) such as wüstite, hematite, ilmenite and chromiteare are those that can be separated in a
magnetic induction higher than 0.6 T. Their specific magnetic susceptibility typically falls into a
range from1 × 10−7 to 5 × 10−5 m3kg-1. Nonmagnetic minerals belong to the category that
cannot be recovered by conventional magnetic separation, such as calcite and quartz; their
magnetic susceptibility is lower than 1 × 10−7m3kg-1. However, such definition may change as
the innovation and developments in magnetic techniques come about [75].
Table 2-9: Magnetic susceptibility of selected rocks, minerals and BOF slag [76][77].
Mineral Magnetic Susceptibility (10-8
× m3
kg-1
) Ref.
Calcite CaCO3 -0.3－1.4
[76]
Quartz SiO2 -0.5－0.6
Wüstite FeO 130
Magnitite Fe3O
4 20,000－110,000
Maghemite 𝛾-Fe2O
3 40,000－50,000
Hematite Fe2O
3 10－760
Jacobsite MnFe2O
4 500
Chromite FeCr2O
4 63－2,500
Magnesioferrite MgFe2O
4 8,030－80,700 [77]
BOF slag --- 24－136 [77]
31
In mineral processing, in terms of the surface magnetic “strength”, magnetic separators are divided
into three categories: high-intensity magnetic field (Up to 2T or greater), and low-intensity
magnetic field (lower than 0.3T). While the actual magnetic field intensity is measured in H, Am-
1, the magnetic “strength” of separation refers to the magnetic flux density, measured in Tesla (or
Gauss where 1T = 104G). Magnetic flux density is proportional to the magnetic field strength
outside magnets, thus Gauss is often used to describe magnetic field “strength”, because of the
ease of measurement.
In magnetic separation, the force that separates magnetic particles from nonmagnetic particles is
the magnetic force 𝐹𝑚⃑⃑ ⃑⃑ , given by following equation [75]:
𝐹𝑚⃑⃑ ⃑⃑ =
𝑣
𝜇0𝑉𝐵∇⃑⃑ 𝐵 (2)
where 𝑣 is the volumetric magnetic susceptibility of magnetic particle
𝑉 is the volume of magnetic particle
𝜇0 is the magnetic permeability of vacuum
𝐵 is the external magnetic induction
∇⃑⃑ 𝐵 is the magnetic field gradient of the magnetic induction
As shown in Figure 2-13, magnetic force is competing with gravity and hydrodynamic drag. In
most case, the hydrodynamic force 𝐹𝑑⃑⃑⃑⃑ acting on the particles may be written as [75]:
𝐹𝑑⃑⃑⃑⃑ = 6𝜋𝜂𝑑𝑣 (3)
where 𝜂 is the dynamic viscosity of flow
𝑑 is the radius of particle
𝑣 is the relative velocity of particle with respect to flow
Based on the above equation, many factors are affecting the result of separation. When the particle
size and flow rate is controlled, the main acting factors are magnetic susceptibility of the particle
and induced magnetic field intensity.
Increasing field intensity does not necessarily mean an improved separation, as higher field
strength may lead to increased capture of weakly magnetic minerals. For different minerals, there
is an optimum range for most efficient separation. There is a tradeoff between recovery and grade
in any separation process, the optimum affected by the magnetic field induction. Low magnetic
32
field strength is beneficial to obtaining particles with more magnetic fraction (i.e. high grade), but
a higher magnetic field strength is required for a better recovery [78].
Figure 2-13: Simplified schematic diagram for magnetic separation [75].
2.3.4.4 Magnetic Separation of Metal in Slag
Direct magnetic separation has been utilized in some steelmaking plants as one step of slag
processing technology. Nippon Steel Corporation processes their converter and ladle slags to
separate iron values and improve their slag quality. The steel slag is crushed and the iron is
recovered by magnetic separation. To increase the recovery of metallic iron, the crushing and
magnetic separation process is repeated several times [6]. The corporation has annual slag
treatment capacity of 2 Mt, reclaiming 0.18 Mt iron particles with more than 95% Fe content yearly.
In China, Anshan Iron and Steel Company recycles 0.28 million tonnes of steel with 60%-65% Fe
(mixed with gangue) and 0.4 million tonnes of iron concentrate with approximately 50% Fe
content annually from steel slag through a combination of sorting-magnetic separation-gravity
concentration processes. Benxi Iron and Steel Company produces 78,000 tonnes of slag steel and
89,000 tonnes of iron concentrate with the advanced “hot slag disintegrating treatment technique”
through magnetic separation [8].
It is usually the case that both the concentrate and tailing can be useful products. Lin et al. [78]
produced an iron-rich magnetic fraction and phosphorus-rich nonmagnetic fraction from
33
steelmaking slag. Adding reagents such as SiO2, Al2O3 and TiO2 increased P recovery rate. Under
optimum conditions, the phosphorus and iron recoveries were 80% and 60% respectively.
A combination of screening and magnetic separation can be very efficient at separating metallic
particles and metal bearing minerals from slags. ArcelorMittal Global R&D has developed a
technology of weak magnetic separation coupled with selective particle size screening for
recycling of steelmaking slag fines [61]. These materials are generated in routine slag processing
consisting of EAF, BOF, LMF slags and their particle sizes are usually smaller than 12.5 mm. The
study employed a pilot-scale magnetic drum separator, and found that effective separation of iron
from steelmaking slag fines occurs below 100 mT, and especially between 20 mT and 80 mT. The
slag fines (-12.5mm) are charged into a pilot-scale magnetic drum separator with 830G surface
magnetic field strength. The produced concentrate is than sized with 9.5 mm and 6.5 mm screens.
The +9.5mm product has an iron grade greater than 80%, which can be charged directly into EAF;
the 9.5-6.5 product has an iron content greater than 60%, which can be charged into blast furnace
for iron extraction; the -6.5 mm product with ~ 60% iron was deemed a suitable feed for iron ore
sintering process.
Menad et al. [11] studied and developed a magnetic separation operation for steel LD slag
consisting of dry and wet separation stages. LD slags were ground and divided into two size ranges:
+63 µm and -63 µm. Larger particles were charged into dry magnetic separator. The collected
magnetic particles were mixed with -63 µm slag powder and the mixture was subjected to wet
magnetic separation. In this system, both low intensity and high intensity magnetic separators were
employed. The recovered products of this integrated process were iron scrap, iron oxides,
paramagnetic materials and a non-magnetic fraction containing mostly calcium silicates.
Singh [79] designed a magnetic separation process for metal recovery from the ferrochrome slag.
Cooled slag was ground and screened into various particle size ranges: 0-1mm, 1-3mm, 3-6mm,
6-10mm and 10-25 mm. The obtained screen particles were separated in magnetic separation
process under various process conditions. He found that about 75% of metallic phases were in the
3-25mm size range. This optimized process can recover 87% of the metallic phases from the slag
achieving more than 90% purity. The recovered product is 48-60% Cr, 3-8% C, 4-9% Si and 26-
31% Fe.
34
2.4 Pyrometallurgical Methods
To recover metal values from slag via pyrometallurgy, there are two major directions: slag
metallization and slag magnetization.
a) Slag metallization aims to convert compounds (often oxides) into metallic elements/alloys
or sometimes matte, and later recover them. It is mainly achieved by completely reducing
metal oxides into metallic forms, and separate them using their distinctive physical
identities, such as density and volatility. Sometimes, reduction takes place to lower slag
viscosity, and facilitate settling of metallic phases from slag.
b) Slag magnetization is intended to use the magnetic properties of spinel oxides, whereas the
target metal bonded in spinel oxide can be recovered employing magnetic separation.
These magnetic oxides (mainly magnetite) can be produced under controlled conditions
(e.g. atmosphere, additives, etc).
2.4.1 Carbothermic Reduction
To reduce slag to metallic phases, carbon in the form of coal or coke is the most commonly used
reducing agent. Other materials such as ferrosilicon (FeSi), calcium carbide (CaC2), and gaseous
reductants (methane) are used because of their cost or better performance in reducing the slag. The
overall reactions in a C-containing system may be presented as below [80].
C(s) + MeO(slag) = Me (l) + CO(g) (4)
C(s) + MeO(slag) = Me (l) + CO(g) (5)
MeO(slag) + CO(g) = Me(l) + CO2(g) (6)
Various additives can be used to facilitate the subsequent separation. For example, Pyrite (FeS2)
or pyrrhotite (Fe1−XS) are added as a source of sulfurizing agent to supply sulfur and form matte
(more discussion in Section 2.4.4 Sulfurization). Slag modifiers such as CaO, CaF and TiO2 are
used to change the slag chemistry and fluidity. They can affect the Fe3+/Fe2+ ratio, and behavior of
iron, copper and other metals in molten slag. Lowering basicity decreases the solubility of copper
in the slag, also decreases the Fe3+/Fe2+ ratio. This promotes efficient copper recovery and low
slag viscosity, easing matte and slag separation [80][50][54]. Not only in on-site liquid-state slag
35
cleaning, reduction of metallic compounds can also take place in the solid state. Many studies have
been carried out to investigate techniques with high metal recovery and selectivity.
2.4.1.1 Direct Reduction
Direct carbothermic reduction of copper slag with coal was investigated in a laboratory scale
electric furnace by Maweja et al. [51], as shown in Table 2-10. The effect of coal-to-slag ratio on
recovery of Cu and Co showed a trade-off relationship: the reduction improves recovery of these
metals but also partially reduces iron oxide to iron which contaminates the metal phase. The lower
coal-to-slag ratio (0.0256) led to higher copper (35%) and cobalt (26%) contents in the metal
phases where the contamination by iron was kept low. The recovery was however low at less than
50% for copper and 60% for cobalt. These contents in metal phase dropped (12% for copper and
20% for cobalt) when the coal-to-slag ratio was 0.05, while the recoveries increased to 65% and
90% depending on the reduction time. The operating temperature is also a factor that enhances
the reaction kinetics at higher temperature and improves metal quality at lower temperature[1].
Maweja et al. [51] suggest that the temperature for carbothermic reduction should be under
1400 °C, as FeO reduction is favored at higher temperatures. In this study, zinc and lead were
recovered in dust phase. More than 90% Pb and 70% Zn could be reduced and vaporized under
optimum conditions.
Long et al. [93] introduced a carbothermic reduction method to produce an iron-chromium alloy
powder from chromium slag (metallic chromium production). Slag mixed with anthracite and
additives (Al2O3 and CaO) was heated to 1300 °C. The obtained solid was ground and underwent
magnetic separation, collecting a pulverized metal product of 72.5% Fe and 13.5% Cr. The
recovery was 80.3% for Fe and 80.7% for Cr.
Rudnik et al [55] reduced 900 kg of copper slag (Slag 17 in Table 2-5) in an industrial electric
furnace, as shown in Table 2-10. Fluxing agents (limestone 8.6%, dolomite 8.6%) and reducing
agents (coke breeze 2.7%, pig iron 2.0%) were added during the process. The reduction took place
at a temperature of 1425–1570°C for 90 mins, producing an alloy of Cu27–Co6–Fe64–Pb1.5.
A comprehensive study on reduction of steelmaking slag was carried out by Ye et al. [21] in a 3
MW DC furnace. Vanadium-rich BOF slag, normal BOF slag and SS slag were reduced in three
campaigns respectively. During the V-rich BOF campaign, various reductants (coke, pet-coke and
36
anthracite), slag feeding rates (1.0 -1.75 tonne/h), and slag temperatures (1550–1650°C) were
tested. The optimum iron alloy contained 10%V, 4%Mn, and 0.6%P. In the normal BOF campaign,
various slag modifiers and operating condition were tested based on the target product: sand
bauxite for clinker raw material; bauxite for metallurgical powder; scrap residue and bauxite for
hydraulic binder. On average, an iron alloy containing 4% Mn and 0.5-1.5% P was obtained.
During the SS slag campaign, SS EAF slag and AOD slag were treated. A metal containing
typically15–20% Cr and 4–5% Ni was obtained. The leachability of SS slag reduction residue was
also analyzed.
In the 1960s, a process for recycling high carbon ferrochromium slag using an electric furnace was
studied by Charvat [81]. Carbothermic reduction was utilized for production and separation of
metallic impurities from magnesium-aluminum spinel. Cold ferrochromium slag was finely
pulverized into less than 8 mesh and mixed with carbon powder. Iron in the form of scrap was
added to the mixture. On heating to 1850 °C, carbon vigorously reduced silicates, iron and
chromium oxides. After reaction, the mixture was superheated to 2050 °C for separation of a lower
phase containing ferrosilicon with minor quantities of chromium and an upper phase that contains
non-metallic phases of Al2O3 and MgO. The lower and upper phase could both be recovered as
ferrosilicon and magnesium-aluminum spinel.
2.4.1.2 Selective Reduction
To selectively reduce base metals (Cu and Co), CaO, CaF2 and TiO2 are the slag modifiers that
can be used. Banda et al. [80] tested the effect of three slag modifiers individually in a synthetic
copper slag system (2.5% Cu and 1.9% Co). The study suggested that TiO2 shows better selectivity
for Co recovery (higher Co/Fe recovery ratio), but lower overall recovery rate of cobalt compared
with CaO or CaF2. The selectivity of CaO and TiO2 was further investigated by Zhai et al [54], as
shown in Table 2-10. The group were looking at the possibility of selectively recovering Co and
Cu from the Chambishi copper smelter. The best result was achieved with 5% TiO2 addition. With
smelting temperature of 1350°C and time of 2.5 hours, a Co-Cu-Fe alloy (1.76% Co, 75.20% Cu,
and 12.85% Fe) was separated. The study reports recovery rates of more than 94% and 95% for
Co and Cu, and less than 18% for Fe.
37
A similar approach was developed by Pan et al. for nickel slag recycling: selective reduction
followed by magnetic separation [82]. The process is controlled by adjusting basicity (addition of
lime) to 0.15. The effect of reductant dosage was also discussed: the reduction of iron can be
effectively controlled via the addition of reductant. At the optimum condition, the recovery rate of
Cu and Ni were 80% and 82%, while the recovery of iron was only 42%. The obtained metal alloy
contained 3.25%Ni, 1.20%Cu and 75.26%Fe.
A two-stage direct carbothermic reduction process developed by Busolic et al. [53] can selectively
recover copper and iron from copper slag, as shown in Table 2-10. The first stage is de-
copperisation aiming at removing copper; the second stage is iron reduction. CaO was added to
obtain a final basic slag with CaO/SiO2=1.5 to ensure fluidity. A copper-rich and an iron-rich alloy
were produced in these two stages respectively. The copper rich-alloy contained 51.0% Cu and
48.2% Fe, as the iron-rich alloy contained 98.9% Fe and 0.84% Cu.
Palacios et al. [42] reported a study on selective reduction of copper slag in a reducing environment,
as shown in Table 2-10. A mixture of coke and slag was heated in a closed crucible at 1260°C.
Copper in the final slag decreased from 3.35% to 0.35% after 30 minutes, a superior performance
than the conventional slag cleaning process. The metallic phase separated contained 61.6% Cu,
28.0% Fe and 2.67% Mo.
2.4.2 Slag Settling
Slag cleaning was first proposed in the copper extraction industry in 1960s, specifically for
cleaning smelting and convertor copper (or nickel) slags. This operation keeps molten slag inside
a bath-smelting unit (known as settling furnace) for long periods, allowing mechanically
suspended matte droplets to settle to a molten matte layer.
As seen in Figure 2-2, a significant amount of copper is chemically dissolved as CuO and CuS in
the slag phase, which cannot be separated through settling or other physical methods. To optimize
recovery rate, additives are charged to promote conversion of Cu oxide to suspended CuS or
metallic Cu drops [1]. Figure 2-14 shows a typical AC slag settling furnace, which can process
1000 to 1500 tonnes of slag every day. Little turbulence is generated, resulting in a calm
environment for settling. An electric settling furnace has a power consumption from 15 to 70 kWh
per tonne of slag, depending on the target cleanness of the final slag (typically 0.6 to 1.3% Cu)[1].
38
Figure 2-14: An AC slag-settling furnace. A furnace of this size cleans 1000–1500 tonnes of
slag per day[1].
An alternative to electric furnace for slag cleaning is a fuel-fired slag cleaning furnace, where slag
is heated and melted using a fuel. In a Teniente slag-cleaning furnace for example [1], powdered
coke and air are injected to generate both heat and a reducing atmosphere. Operational data show
that it can clean the charged slag from 6 to 8% Cu to less than 1%, with a copper recovery of 85%.
Many innovations and improvements have been made. A Canadian invention by Degel et al. [83]
involves a two-step recovery of copper from molten slag, by charging slag into an AC and then a
DC furnace. The approach breaks the reduction process into half, thus shortening the reduction
time by half and increasing the operating capacity by two-fold. Slags are reported to be purified to
0.5% Cu and 4% magnetite after the two stages. The average power consumption is estimated to
be 107 kWh/t and the coke consumption is around 10 kg/t. Degel et al. claimed that this process
is suitable for recovery of metals in other processes such as those of ferroalloy industry.
The metallic alloys produced from previous processes often contains copper, iron and cobalt, thus
the isolation of metals has been of interest. For this purpose, Rudnik et al. [55] developed a process
that includes electrolytic dissolution of the alloy in an ammonia–ammonium chloride electrolyte,
ammoniacal leaching of the slime, and selective electrowinning of copper and cobalt. The products
are deposits of copper and cobalt with 99.9% and 92% purity respectively. Yin et al. [84] recovered
cobalt from Co-Cu-Fe alloys via magnetic separation and sulfuric acid leaching process. The main
cobalt bearing phase in the alloy is cobalt-iron alloy, which is a magnetic substance. After the
39
process, 95.6% cobalt was extracted from the alloy, and the tailing from magnetic separation and
the residue from leaching were recycled into copper smelting for near-complete recovery of copper.
Targeting only iron value, a study on copper slag deep reduction followed by magnetic separation
was carried out by Li et al. [85], as shown in Table 2-10. The iron recovery reached over 91%
under optimized conditions, and the iron content was over 96%.
2.4.3 Slag Fuming
Metallurgical slags generated in smelting of non-ferrous concentrates contain metallic values such
as zinc, tin, lead, cadmium, indium, arsenic, antimony and geranium. These metals can be
recovered in the slag fuming process. A fuming furnace is like Isasmelt furnace, with a vertical
body and one long blow tube which directly injects air into melt. Slags are fed into a fuming
furnace and are reduced by oxygen of the blast, a fuel and reductant. Off-gas of the process will
contain volatile constituents from the slag, which can be condensed and used for metal recovery.
In one case the slag starting with 17% Zn contained only 2% Zn after such treatment [86]. Utilizing
high velocity tuyere(s) (80 to 115m/s) entrains more small bubbles of fuel/reductants where the
actual reduction reaction of metal values occurs. This increases process efficiency thus saves fuel
and reductants, resulting in a time saving of 20% of the main fuming cycle and reduction of coal
consumption by about 20%.
2.4.4 Sulfurization
Sulfurization is another form of slag metallization, which is more suitable for converting Ni and
Cu value of slag into a metal-rich matte. Reactants such as FeS and CaS can be used for this
purpose. However, the shortcoming of sulfurization is that Co and Zn tend to remain in the slag
matrix [51].
Li et al. [50] developed a process to overcome the above challenge by using another waste
materials to recycle copper slag. They roasted copper slag with CaSO4-rich gypsum waste. CaSO4
is a source of both calcium and sulfur, automatically replacing iron sulfides and lime (as slag
modification agent). The following reaction shows the overall process of sulfurizing. It is clear
that a reducing agent is required, and lime is formed as a product.
(Cu2O) + CaSO4 + 4C [Cu2S] + CaO + 4CO(g) (7)
40
The optimum conditions for this process were found to be 12 wt.% coke, 20 wt.% CaSO4 at a
smelting temperature of 1350C for 3 h. The recoveries of Cu and Co were 92.0% and 95.6%
respectively. The concentrations of these metals in the matte were 13.6% Cu and 9.4% Co.
Selectivity ratios expressed as Cu/Fe and Co/Fe reached 6.00 and 6.24 respectively. Also, due to
the generation of lime, the cleaned slag showed an increase in basicity, from 0.21 to 0.36.
2.4.5 Magnetizing Roast
Magnetizing roast can be either oxidizing or reducing, depending on the initial and target oxidation
states of iron. When iron in the substance is mainly trivalent, partial reduction is used, which is
typically applied to upgrade iron ore. Its mechanism is to reduce hematite Fe2O3 to magnetite
Fe3O4. This allows easy separation of iron oxides from other oxides via magnetic separation with
low energy intensity. The reduction of hematite to magnetite is shown in reactions (8) and (9):
3Fe2O3 + C = 2Fe3O4 + CO (8)
3Fe2O3 + CO = 2Fe3O4 + CO2 (9)
When divalent iron is the dominant form of iron, oxidation is used to produce magnetite, as shown
in Eqs. (10) and (11):
4FeO + O2 = 2Fe2O3 (10)
FeO + Fe2O3 = Fe3O4 (11)
A study by Zhang et al. [87] developed a reducing roast process to recover iron concentrate from
pyrite process slag (the process to produce sulfur and sulfuric acid), as shown in Table 2-10. The
slag contains about 20% total iron mostly as Fe3+ and in the form of red limonite, magnetite, and
ferrosilite minerals. Under optimum conditions, the iron concentrate grade was 57% and recoveries
were up to 70%.
Semykina et al. [88] confirmed the possibility of recovering iron by oxidizing slag FeO to
magnetite in steelmaking slags. Steel slag in molten state is exposed to an oxidizing gas and then
quenched. Further study [89] by the same author demonstrated the feasibility of producing
nanoscale manganese ferrite from steel slag with high Mn content in similar oxidation conditions.
41
In both cases, increasing basicity promoted the activity of FeOx and MnOx, which is a strong
driving force for the formation of spinel phases.
Edlinger [90] used oxidation to recover chromium and vanadium in spinel form from steelmaking
slags. In the process, oxygen was introduced to the liquid slag. The oxidization environment
allowed conversion of iron to iron oxide and about 5 to 20 wt% of the slag to magnetite. During
the formation of magnetite spinel phase, chromium in the slag is mostly associated with magnetite,
forming a mixed spinel (FeCr2O4.Fe3O4). It is essential to control slag basicity (CaO/SiO2) to
values greater than 2.5. The spinel phase could be magnetically removed from the cooled oxide.
The produced spinel comprises a high-quality chromium ore and may be recovered by
conventional reduction. This method is suitable for enrichment and recovery of Mn, Mg, Al, C, V
and Fe.
A study on oxidization behavior of copper smelting slags to selectively precipitate magnetite was
carried out by Fan et al. [91]. Oxidizing gas was blasted directly to molten copper slag. Under the
condition of lower oxygen partial pressure and lower gas flow rate, oxidation resulted in
precipitation of magnetite, rather than hematite. In the meantime, it was found that Cr and Zn tend
to be entrained in magnetite while Cu exhibits opposite phenomenon. The reaction mechanism
was analyzed, discussing that the mobility of O2- was relatively weak in the early stages and was
enhanced as oxidation progressed. The gas diffusion was assumed to be the “controlling step” in
the later stage of the reaction.
Ma et al. [92] carried out air oxidation followed by magnetic separation on nickel slag to produce
an iron concentrate, as shown in Table 2-10. Under optimum conditions, the final iron recovery
and grade reached 76% and 54.0%, respectively. During the study, they found that the average
size of the magnetite grains was 20 µm at 1350°C for 20 min oxidation. Air flow rate was related
to magnetite crystallite size, however when the injection rate exceeded 170 ml/min, the violent
stirring could break the dendrites and cause formation of tiny crystals.
42
Table 2-10: Summery of pyrometallurgical methods investigated on metal recovery from slag.
Slag Type Mothed Optimum condition Result Ref.
Copper Slag Carbothermic
reduction (coal)
Coal/Slag: 0.035-0.04; Time: 1 hr; T: 1200-1250°C Recovery: 70+% Zn, 80+% Cu,
80+% Co
[51]
Carbothermic
reduction (coke)
T: 1450°C; First stage: 100% of the coke required to reduce
Cu2O, CaO/SiO2 = 1.5, Time: 30 min; Second stage: 150% of
the coke required to reduce FeO, Time: 50 min
Two metallic alloy: 51.0% Cu and
48.2% Fe; 98.9% iron and 0.84%
Fe
[53]
Carbothermic
reduction (activated
carbon)
5% activated carbon; 5%TiO2; T: 1350°C; Time: 2.5 hr Recovery: 94% Co, 95% Cu, 18%
Fe; metallic alloy: 1.76% Co,
75.20% Cu and 12.85% Fe;
Residue slag: fayalite and hercynite
[54]
Carbothermic
reduction (coke)
Closed crucible; Excess stoichiometric amount coke; T:
1260°C; Time: 30 mins
Residual slag: 3.35% to 0.65% Cu
and 48.3% to 42.12% Fe; Metallic
alloy: 61.6% Cu, 28.0% Fe and
2.67% Mo
[42]
Carbothermic
reduction (coke)
coke 2.7%, pig iron 2.0%; electric furnace, voltages: 80–90
V, I: 900–950 A; T: 1425–1570°C; Time: 90 mins
Cu27–Co6–Fe64–Pb1.5 Metallic
alloy: 27% Cu, 6% Co, 64% Fe and
1.5% Pb
[55]
Carbothermic
reduction (coke) +
magnetic separation
Coke:14%; Ca/Si=0.2; T: 1300°C; Time: 3 hr
Grind 20 mins; magnetic field strength: 76 mT
Recovery:91.82% Fe; grade
96.21% Fe
[85]
43
Chromium
slag
Carbothermic
reduction (anthracite)
+ magnetic separation
Slag:antheracite:Al2O3:CaO=100:30:4:3; T: 1300°C; Time:
60 mins
PS: D80 = 74µm; magnetic field strength: 120 mT
Metal alloy: 72.54% Fe, 13.54%
Cr; Recovery rate: 80.34% Fe,
80.70% Cr
[93]
Nickel Slag Carbothermic
reduction (coal) +
magnetic separation
Coal: 5%; Ca/Si = 0.15; 1200 °C; 20 mins
PS: -75 µm; magnetic field strength 165 mT
Recovery: 80% Cu, 82% Ni and
42% Fe; Metal alloy: 3.25%Ni,
1.20%Cu and 75.26%Fe
[62]
Nickel Slag Magnetizing roasting
+ magnetic separation
CaO addition; T: 1350°C; Time: 20 min; Air: 170 mL/min
PS: -38 µm; magnetic field strength 120 mT (3 steps)
Recovery:75.99% Fe; grade
54.08% Fe
[92]
Pyrite Slag Magnetizing roasting
+ magnetic separation
Coke: 5%, PS: 1mm; Time: 20 min; T: 850 °C
PS: -74 µm; magnetic field strength: 290 mT
Recovery:70% Fe; grade 57% Fe [87]
44
2.5 Hydrometallurgical Methods
Hydrometallurgical methods are primarily employed for recovery of non-ferrous metals from
smelter/convertor or ferroalloy slags. In recovery of non-ferrous metals, iron is typically seen as
an impurity, thus its reporting to the product should be minimized. There are mainly two stages in
hydrometallurgical metal recovery processes: leaching and metal recovery.
2.5.1 Leaching
In the first stage of recovery process, lixiviants (such as strong acids) are employed to release
metals from solid material as ions into an aqueous solution. Many studies have investigated this
stage since the 1990’s. A summary of the leaching studies that is discussed in this section is shown
in Table 2-11.
2.5.1.1 Effect of Various Parameters on Leaching Kinetics and Efficiency
Leaching efficiency and rate of reaction are often optimized by the type and concentration of the
lixiviant, pH value of the solution, temperature, agitation intensity, and particle size. Many studies
reported that particle size reduction (also known as mechanical activation) and operating
temperature have the most significant impacts on metal recovery [46].
Influence of Temperature
Temperature of metal digestion is known to affect the reaction rate. However, different elements
in the same system can be influenced to different extents. In a Cl2/Cl- leaching study [46], copper
dissolution rate was not significantly dependent on temperature contrary to iron, suggesting that
the operating temperature can be optimized for the selectivity.
Influence of Particle Size
Particle size has a significant effect on the rate and extent of metal recovery; smaller particles give
rise to a greater surface area that improves the kinetics and also brings more metal in contact with
the lixiviant, hence larger recovery.
In a study by Herreros et al. [46], it was found that most of copper in their copper slag is in the
form of metallic copper, chalcocite, bornite and other complex sulfides of very small size, mainly
45
5–10 µm. On average, only 0.2–0.4% of the copper was present in dissolved form in the silicate
phases. This showed that no silicate reaction is needed for copper dissolution, and fine grinding is
beneficial for copper liberation. Carranza et al. [56] also confirmed this finding by leaching copper
slag in different size fractions. Their experimental results showed that high recovery rate of copper
(greater than 94%) is achieved for the -106 µm fraction.
Influence of Slag Cooling Rate
As discussed earlier, cooling rate affects the structure and mineralogical composition of the slag.
Tshiongo et al. [49] compared the response to leaching of several slags with a different thermal
history; water quenched slag, water granulated slag, air-cooled slag, slow-cooled slag. Leaching
of fast cooled slags, especially quenched slag in an ammonia solution (pH=12) shows a positive
impact on copper leaching (over 90% Cu extraction) and selectivity over Co, Zn, Pb and Fe. The
slow-cooled slag showed poor leaching of Cu and selectively.
2.5.1.2 Acid Leaching
Sulfuric acid (H2SO4) is considered as an efficient agent for leaching of metals. Many experiments
performed in metal recovery of slag using sulfuric acid show high values of metal extraction. Bulut
[52] tested direct digestion of copper slag in sulfuric acid; 78% Cu and 90.2% Co were extracted
(optimum conditions are given in Table 2-11). However, it also comes with the problem of iron
co-extraction; 71.5% Fe was dissolved.
To selectively extract base metals such as Cu, Co, Ni, oxygen pressure acts as an important factor.
In order to enhance leaching in an oxidizing environment, oxidants such as hydrogen peroxide,
potassium dichromate, chlorine and sodium chlorate have been used [46] [44], [43] [94].
In a study carried out for nickel recovery from slag [94], high selectivity was achieved by
controlling oxygen partial pressure to 600 kPa. Recovery of >97% for cobalt and nickel, >95% for
copper and <0.4% for iron was obtained by sulfuric acid leaching at 200 °C for 80 min.
Herreros et al. [46] investigated leaching reverberatory copper slag under Cl2/Cl-. The Cl2/Cl-
system was established from the reaction between sodium hypochlorite and hydrochloric or
sulfuric acid, as shown in reactions (12) and (13). Selective dissolution of 80 to 90% copper and
46
5% iron were obtained at room temperature Cl- concentration of 3.010-2 M and particle size <
20m. The extraction was very rapid, with the above recoveries achieved in only 5 mins.
NaOCl + 2HCl Cl2 + NaCl + H2O (12)
NaOCl + NaCl + H2SO4 Cl2 + 2NaCl + H2O (13)
Citric and nitric acids were used to selectively dissolve P2O5 in calcium silicate matrix for
steelmaking dephosphorization slag [10]. Dissolution ratio of P2O5 reached 60% and 80% in citric
acid and nitric acid leaching respectively, at pH 5. However, citric acid was highly selective with
no iron oxides dissolution.
In sulfuric acid leaching, formation of silica gel (H4SiO4) is a problem, as shown in reactions (14)
and (15). Silica gel complicates metal extraction and filtration processes, and it also causes
operational difficulties (such as lower metal extraction and pulp filtration) in the subsequent steps
of metal recovery [43], [44], [95].
2MeO·SiO2 + 2H2SO2 2MeSO2 + H4SiO4 (where Me = Fe, Co, Zn Cu) (14)
H4SiO4 2H2O + SiO2 (gel) (15)
A few studies have been carried out to address the silica gel problem. Banza et al. [44] have
demonstrated that silica suspension does not occur when the pH value is higher than 2, although
low pH value is beneficial for leaching of copper, cobalt and zinc. A process of leaching copper
slag with sulfuric acid with pH value controlled at 2.5 and in the presence of hydrogen peroxide
revealed high and selective extraction of base metals: 80% Cu, 90% Co, 90% Zn and only 5% Fe.
Yang et al. [43] leached copper slag with sulfuric acid and sodium chlorate oxidant and neutralized
silica gel and Fe(III) with calcium hydroxide addition. After leaching, calcium hydroxide was
added slowly to maintain a low pH level at 2.0 for 1 hour. The final recoveries were 98% Co, 97%
Zn and 89% Cu with only 3.2% Si and 0.02% of Fe dissolved.
In some investigations, acid roasting of the slag prior to leaching was tested, aiming to improve
the kinetics and/or recovery. For example Arslan et al. [41] proposed acid roasting followed by
hot water leaching to extract Cu, Co and Zn from copper slag. In acid roasting, crushed slag was
mixed with different amounts of acid, heated to a temperature between 150-300 °C and then cooled
and leached in 70 °C water. With the acid:slag ratio of 3:1, roasting time of 2 h and temperature
47
of 150°C, 87% of copper, 87% of cobalt, 93% of zinc, and 83% of iron were recovered. With
increasing roasting time to 3 h and temperature to 250°C, Cu recovery could reach 100%.
In another set of experiments [41], the same authors calcined the slag at 650°C for 2 h after acid
roasting, aiming to decompose metal sulfates, and then leached the product in hot water. The
recoveries were 79% Cu, 66%v Co, and 41% Zn with no iron extraction. Similar conditions were
tested by Bulut [52] for copper slag and the result agreed with Arslan et al. [38]. Ayala et al. used
acid roasting to leach Mn from SiMn slag. During water leaching stage, CaO and KOH were added
to control the pH value to 6, so as to inhibit the dissolution of silicon, aluminum and iron. Highly
selective recovery of 94% Mn was achieved
The oxidizing state of metals affects the solubility during leaching. For example, hematite and
magnetite show a low degree of leaching compared to wüstite. In Palacios et al.’s study [42],
molybdenum in copper slag was present as 2FeO.MoO2-Fe3O4 in magnetite. Molybdenum was
liberated from magnetite to a soluble phase by oxidizing roasting at 700°C in a fixed bed laboratory
roaster with O2/SO2 gas mixture, through the following reaction:
FeOx.MoO2 + yO2 Fe2O3 + MoO3 (16)
Leaching in a solution of 25 g/L H2SO4 was then performed to extract molybdenum, at room
temperature and L/S ratio of 10:1 in 2 hr. Molybdenum selective extraction has been increased
from 20% to 70% by the roasting which resulted in spinel decomposition.
In Mirazimi’s study [26], LD slag was roasted with sodium carbonate (Na2CO3) to transfer
vanadium to a more soluble form. Vanadium in the form of Ca2V2O7 from original LD slag was
converted to NaVO3 after 2 hr roasting at 1000 °C with 20% sodium carbonate. The optimum
conditions were L/S ratio of 20, temperature of 70 °C, sulfuric acid concentration of 2.05 M and
leaching time of 1 hour. Vanadium recovery rate was 96%.
2.5.1.3 Alkaline Leaching
Leaching of SS Slag for chromium extraction with alkaline lixiviant was investigated in multiple
directions by Kim et al [96]. One of their studies examined NaOH roasting of SS slag. They
reported 83% Cr extraction with limited dissolution of the matrix material (Al, Ca, Mg, Fe <0.01%,
Si 8.0%), while at optimum roasting condition (particle size <63 µm, 0.67 mass ratio of NaOH to
48
SS slag, 400 °C, 2 hr). By adding NaNO3 as an oxidizing agent (0.067 mass ratio of NaNO3 to SS
slag), Cr extraction rate improved to 90%.
Investigation of pressure alkaline leaching of SS slag with NaOH and pressurized O2 was also
carried out by the same group [29]. In a sealed reaction chamber, the oxygen pressure builds up
with increasing temperature. The maximum Cr leaching was 46% at 1M NaOH, 36 bar O2, SS slag
particle size <125 µm, at 240 °C for 6 hours. Al (2.88%), Si (0.12%), and Ca (0.05%) were also
dissolved during this pressure leach.
Roasting of SS slag with sodium hydroxide in the presence of NaOCl as an oxidant, followed by
water leaching was also investigated, aiming to lower the roasting temperature [31]. The optimum
conditions for this system were found to be: 105 °C, 6 hr, SS slag particle size <63 µm, mass ratio
of NaOH to SS slag 0.13, and 3.3 mmol NaOCl to 1 g SS slag. A selective Cr leaching of 68%
was reached, while dissolution of matrix materials was low (Al 0.3%, Ca 2.0%, Si 0.4%). The
residual slag was examined for strength test, reporting a compressive strength of 104 MPa. The
strength is much higher than high density limestone (55 MPa), indicating good performance in
various construction applications.
New approaches have been proposed for separating the copper slag into silica gel and an iron-rich
residue [97]. The process consists of air oxidation at 800 °C for 2 hours; hydrothermal treatment
of the oxidized slag with sodium hydroxide solution (140 g/l) at 190 °C for 3 hours to extract
silicate; and silica gel formation was avoided by lowing the pH of the alkaline silicate solution.
The oxidation process decomposes fayalite into silicate and iron oxides phases. The leaching
residue contained 79.8% of iron oxides and 8.9% of SiO2, where 70% of SiO2 was extracted.
2.5.1.4 Salt Leaching
Some methods use salts as lixiviants to extract metals from slag. One example is ferric chloride
(FeCl3) solution to dissolve sulfides and to some extent oxides of copper, nickel and cobalt (or
their metals if present) to give cuprous chloride [48]. Recoveries of copper 92.0%, nickel 28.0%
and cobalt 24.0% for converter slag (Slag 4 in Table 2-5), copper 54.0%, nickel 77.0% and cobalt
44.0% for smelter slag have been reported (Slag 5 in Table 2-5).
49
Ferric sulfate (Fe2S3) leaching was examined to extract copper from copper slag by Carranza et al.
[56]. Under optimum conditions, 11.5g/L Fe(III) could extract more than 93% Cu in 4 hours.
50
Table 2-11: A summary of leaching studies on slag for metal recovery
Slag Mothed Optimum conditions Result Ref.
Copper
Slag
#1: Acid leaching (H2SO4)
#2: Acid roasting (H2SO4) + hot
water leaching
#2: Acid roasting (H2SO4) +
Thermal decomposition + hot
water leaching
H2SO4: 120 g/ L; PS: <100 µm; T: 60 °C; LT: 2
hr; L/S = 10 mL/g;
H2SO4:Slag = 3:1; Acid roasting time: 1 hr, T: 200
°C; Water leaching time: 1 hr, T: 70 °C;
Thermal decomposition time: 2 hr; T: 600 °C
#1: Recovery: 78% Cu, 90.2%
Co and 71.5% Fe. Total
dissolution of slag: 72.2%.
#2: Recovery: 78.9% Cu, 73.7%
Co and 98.5% Fe. Total
dissolution of slag: 50%.
#3: Recovery: 63% Cu, 43% Co
and 4% Fe
[52]
Chlorine leaching (102 –103 M Cl2) PS: <20 µm; T: ambient; LT: 5 mins Recovery: 80 to 90% Cu and 5%
Fe
[46]
Sulfuric acid leaching + sodium
chlorate oxidant + calcium
hydroxide neutralization
Slag:H2SO4:NaClO3:Ca(OH)2 =
1:0.88:0.125:0.215; PS: <180 µm; T: 60 °C; pH=
2; Time: 3 hr (leaching) + 1 hr (neutralization)
Recovery: 98% Co, 97% Zn,
89% Cu, 3.2% Si and 0.02% Fe
[43]
Sulfuric acid leaching + hydrogen
peroxide oxidant
H2SO4: 500 kg/t; H2O2: 35 L/t; PS: <100 µm; pH:
2.5; T: 70 °C; LT: 3 hr
Recovery: 80% Cu, 90% Co,
90% Zn and 5% Fe
[44]
#1: Acid roasting (H2SO4) + hot
water leaching
#2: Acid roasting (H2SO4) +
Thermal decomposition + hot
water leaching
H2SO4:Slag = 3:1; Acid roasting time: 2 hr, T: 150
°C; Water leaching time: 1 hr, T: 70 °C;
Thermal decomposition time: 2 hr; T: 650 °C
#1: Recovery: 88% of copper
(100% when higher time and T),
87% of Co, 93%of Zn, and 83%
of Fe;
#2: Recovery: 79% Cu, 66%v
Co, and 41% Zn
[41]
51
Slag oxidization + Alkaline
(NaOH) leaching
Air oxidization, T: 800 °C, time: 2 hr;
Leaching: NaOH: 140 g/l; L/S = 5 mL/g; T: 190
°C; LT: 3 hr
Residue: 79.8% of iron oxides
and 8.9% of SiO2
70% SiO2 extraction
[97]
Ferric chloride leaching FeCl3: 1.25 times stoichiometric for Cu; PS: <53
µm; T: 85, 100 °C; LT: 2.5, 6 hr (respectively)
Recovery: 92% Cu, 28% Ni and
24% Co for converter slag (Slag
4 in Table 2-5); 54% Cu, 77.0%
Ni and 44.0% Co for smelter
slag (Slag 5 in Table 2-5).
[48]
Ferric Sulfate Fe2O3: 11.5g/L Fe(III), pH (1.67); PS: D80 = 47.03
µm; T: 60 °C; LT: 4 hr; plup density: 2%
Recovery: 93% Cu [56]
Nickel Slag Acid leaching (H2SO4) + Oxygen
pressure
PS: -150 + 74 µm; H2SO4: 0.3 mol/L; S/L = 7
mL/g; oxygen partial pressure: 600 kPa; T: 200
°C; LT: 80 min
Recovery: 97% Ni and Co, 95%
Cu, <0.4% Fe, <2% Si
[94]
SiMn Slag Acid roasting (H2SO4) + hot water
leaching
H2SO4:Slag = 1:9 (10% excess stoichiometric for
Mn); Acid roasting time: 0.5 hr, T: 200 °C;
Leaching: L/S = 5 mL/g; lime slurry: 15 w/v.%;
KOH 1 w/v%to reach pH=6.
Recovery: 94% Mn.
Leach solution: 31.2 g/L, other
elements <0.2%
[38]
LD Slag Salt roasting (Na2CO3) + acid
leaching (H2SO4)
Roasting: T:1000 °C; 20% Na2CO3; time: 2 hr
Leaching: L/S = 19.99; T: 70 °C; H2SO4: 2.05 M;
time: 1 hr
Recovery: 96% V [26]
52
2.5.2 Metal Recovery
Many well-established steps have been developed for recovery of the metals from the pregnant
leach solution, such as cementation and electrowinning. The solution may require purification and
concentration through solvent extraction prior to the recovery step.
2.5.2.1 Solvent Extraction
In solvent extraction, an organic solvent is used to extract metal ions of the interest (absorption),
and later release them into a more concentrated solution using a strong acid (stripping). Through
this process, the organic solvent is regenerated and recycled. [95].
Banza et al. [44] used LIX 984 for copper and D2EHPA (di-2-ethylhexyl phosphoric acid) for
cobalt and zinc extraction from a pregnant solution containing Fe. Copper was extracted and
stripped with LIX 984 and sulfuric acid (pH 1) at room temperature. More than 99% of Cu was
extracted with a less than 1% co-extraction of other metals. 200g/L CaCO3 was added to remove
99% of iron, with about 3% and 5% co-precipitation of Co and Zn. D2EHPA organic solvent was
then used to extract Co and Zn. Separation of Co from organic phase was achieved with sulfuric
acid at pH=2.5, followed by stripping Zn at pH=1. Extraction rate for Co and Zn were more than
96% and 99% respectively.
2.5.2.2 Precipitation
The precipitation of metal from the rich solution can be achieved by electrowinning or
cementation. For example, aiming to selectively recover chromium from SS slags leachate,
BaCrO4 precipitation was used in Kim et al.’s study [31]. The result shows >99.9% Cr recovery
with minor impurities such as Mn and V. The reaction occurs by adding BaCl2 according to
Equation (17).
BaCl2(a) + CrO4
2- BaCrO4(s) + 2Cl- (17)
Sulfide precipitation is carried out for the separation of impurities such as Zn, Ni, Co and Cu from
manganese solution, via a Na2S solution at pH=5–7 [38]. The collected solution is then
electrolyzed to obtain Mn metal. Electrowinning involves applying a low DC potential in an
53
electrolytic cell, resulting in deposition of metal in the cathode. This method is well established
and will not be reviewed here.
2.6 Summary of Metal Recovery Methods from Slag
Slag processing has numerous variants and can range from a simple comminution and sizing to
more sophisticated methods of chemical treatment. It is apparent that the processing methods are
evolving to yield higher metal recovery and produce less waste. The following is a summary of
overall findings in this literature survey.
Size reduction is an obvious but most important step of slag treatment. Not only the most valuable
slag products are always in the form of aggregates or pulverized materials, but also the subsequent
treatment steps, such as magnetic separation, flotation, solid state roasting or leaching, also require
a certain particle size. Convectional crushing and grinding of air-cooled slag consume a great
amount of energy. Innovative approaches to combine cooling and size reduction with much lower
energy intensity are emerging to replace the comminution methods.
Cooling is a crucial step in slag treatment, since it determines the thermal history of slag which
effects the slag structure (crystalline/amorphous) and the distribution of metal in various
mineral/phases. For applications such as cement feedstock, amorphous slag with a silicate matrix
is desired whereas for magnetic separation a higher degree of crystallinity is beneficial. On the
other hand, a mixture of amorphous and crystalline phases is most suitable for selective leaching.
The most common physical treatments for copper and nickel slags are size-reduction and floatation.
Slag cleaning is also commonly used for converter slags to recover Cu. Metal values such as Co,
Zn, Ni or Mo, usually cycle or remain in slags phase. To recover those valuable elements, deep
selective reduction or complicated hydrometallurgical processes are often a must. Volatile
elements such as Zn can be recovered by vaporization and subsequent condensation. Also, leaching
or a combination of leaching and roasting have proven very effective in recovery of certain metals
including Cu, Co, Ni and Pb.
Chromium recovery is primarily done to meet the environmental requirement pertaining to slag
disposal. A variety of methods such as gravity and magnetic separation, as well as alkaline leaching
have been successfully employed for this purpose.
54
The ways to treat steelmaking slags mainly focused on separation of large metallic iron particles
through physical process. Microscale metallic iron and dissolved iron oxides are still reported to
waste. The status quo limits the usage of steelmaking slag. This research, as explained in Chapter
1, looks into oxidation treatment of steelmaking slags, aiming to investigate the feasibility of
converting the slag into two products, an iron-rich concentrate for recycling to sintering/pelletizing,
and an iron-lean product suitable for construction aggregate, cement, etc.
55
Chapter 3
Experimental
3.1 Objectives
The overall objective of the research is to investigate the feasibility of recovering an iron
concentrate from steelmaking slags using magnetic separation. A more specific objective is to
determine the conditions of slag treatment that leads to highest separation. This chapter will cover
the experimental aspects of the study including procedures, materials and equipment used.
3.2 Materials
3.2.1 Slag
The steelmaking slag used in this experiment, was obtained from ArcelorMittal Dofasco (AMD,
Ontario, Canada). The as-received slag arrived as gray aggregate with particle of ~2 cm in diameter,
with chemical composition given in Table 3-1 (provided by AMD). The composition is
comparable with steelmaking slags given in Table 2-3. Preliminary SEM imaging and elemental
mapping of slag are shown in Figure 3-1 and Figure 3-2. Entrained metal droplets were observed,
ranging from a few µm to more than 300 µm. In the slag matrix, light and dark phases are present.
With the aid of energy dispersive spectroscopy (EDS) elemental mapping, it is seen that the darker
phases are calcium silicates (with lower atomic mass) and oxides such as FeOx, MgO, and MnOx
are the lighter phases. Although small magnification may imply that Ca, Fe, Si and Mg are
distributed uniformly in the slag matrix, at higher magnification the separation of phases is clear
at the scale of 100 µm or less. The important understanding from this is that for liberation of oxides
from slag, pulverization to the range of 100 µm is necessary. XRD analysis confirmed the phase
identification from SEM imaging (Figure 3-3). In addition to C3S, C2S, R-O phase, wüstite and
Ca2Fe2O5 appeared to be present. According to XRD pattern database, the position (2θ) overlaps
with the spinel phases such as magnesium ferrite and magnetite, which constitute the RO phase
Therefore, XRD alone cannot distinguish the individual oxides from spinel phases and techniques
such as elemental mapping should be used for that purpose. The elemental mapping results indicate
that the light region in the SEM image is a combination of Ca2Fe2O5 and RO phase.
56
Table 3-1: Chemical compositions of steel slag sample (wt%) from XRF analysis provided
by AM-Dofasco
Total Fe FeO CaO SiO2 Al2O3 MgO MnO P2O5 TiO2 Cr2O3 C/S
18.9 24.3 43.6 15.0 4.4 9.0 3.7 0.9 1.2 0.5 2.9
Figure 3-1: SEM BSE image of steel slag sample
Figure 3-2: SEM-EDS elemental mapping of steel slag sample
57
Figure 3-3: XRD spectrum of the steel slag sample.
The slag was pulverized in a jar rolling mill for two hours. The machine has 805 W power, which
sustains a rotation speed of 180 rpm. About 2 kilogram of steel slag sample was charged into the
cylindrical jar with 22 cm diameter and 20 cm length. The mill was also loaded with 10 kg of steel
balls of 3.5 cm in diameter (slag to ball ratio = 1:5).
The particle size distribution of the pulverized slag is shown in Figure 3-4, showing d80=133 µm,
d20=11 µm, and a mean size of 80 µm. A double peak distribution is seen which is an indication
of presence of at least two phases with significantly different grindability. This in in agreement
with studies done by Wang et al. [98][99], who investigated cementitious properties of the finer
portion of their steel slag powder. The group separated the two size fractions and analyzed them
by XRD. The results showed that the finer fraction was richer in C3S, C2S but lower in RO phase,
the opposite was true for the coarser fraction. Higher metal content in RO phase might contribute
to resistance to grinding.
58
Figure 3-4: Particle size distribution of the pulverized slag.
3.2.2 Additives
For basicity adjustment (to CaO/SiO2 ratios of 1.5, 2.0 and 2.5), proper amounts of +99% quartz
served as slag addition.
3.3 Experimental Setup
The general idea and experimental scheme of this research are shown in Figure 3-5. Through an
oxidizing treatment, this study aims to convert iron species into magnetic spinel phases, which are
later separated by magnetic separation. An iron-rich product is then recovered which may be
recycled as s source of iron to processes such as sintering, and an iron-lean residue which because
of its high CaO and SiO2 content is a suitable feedstock for making cement.
A thermodynamic study was performed (Section 4.2), aiming to narrow down the experimental
conditions that lead to maximum spinel phase formation. Both solid and liquid state slags were
considered.
1 10 100 1000
0
10
20
30
40
50
60
70
80
90
100
0.0
0.5
1.0
1.5
2.0
2.5
3.0
3.5
4.0
4.5
Cu
mu
lati
veC
ou
nts
(%)
Vo
lum
e(%
)
Particle Size (μm)
Counts Cumulative counts
59
Figure 3-5: Scheme of the experimental steps
3.3.1 Solid State Oxidation
Solid state oxidizing was carried out in a horizontal tube furnace (Figure 3-6). The reaction
chamber was an 8 cm alumina tube heated externally with molybdenum disilicide elements. The
temperature was measured using a type B thermocouple placed in the uniform-temperature hot
zone of the furnace close to the slag samples. The working tube was fitted with water-cooled
aluminum caps on both ends that allowed sealing of the tube and applying a controlled gas
atmosphere. The influence of type of oxidizing gas, roasting temperature and time were
investigated in these experiments.
Figure 3-6: Schematic diagram of the experimental setup in solid state oxidation of slag.
A typical experiment began with preheating the furnace to the target temperature. When charging
sample, aluminum cap on the side of gas outlet was removed, and a base made from alumina foam
that held three MgO crucibles (about 20 g slag per crucible) was positioned carefully in. The foam
board was placed in the hot zone of the furnace where the temperature was uniform for ~ 20 cm.
The cap was then put back to seal the tube. The samples were heated in argon atmosphere for 30
minutes to reach the target temperature. Afterwards, the gas was switched to the oxidizing gas with
a flow rate of 100 mL/min for a controlled dwelling time. After completing the reaction, the
60
alumina foam base was pulled to the end of the tube for rapid cooling. The slag was then stored in
a desiccator for subsequent crushing and magnetic separation.
3.3.1.1 Oxidizing with CO2
Oxidation with compressed CO2 was carried out at 600 and 650 °C and with dwelling time of 10-
60 minutes. Mass loss instead of mass gain of the product was observed after oxidizing reaction.
On average, about 3.5% mass loss was observed. This was interpreted as dominant effect of
moisture loss and decomposition of Ca(OH)2 and carbonates[11].
3.3.1.2 Oxidizing with Air
In these experiments air was used for the oxidizing roast. Preliminary experiments again showed
mass loss. Therefore, a pre-roasting treatment (i.e. calcination) was added to remove moisture and
CO2. For this treatment, the pulverized slag rested in argon atmosphere at 1000 °C for one hour.
The obtained product was cooled and stored properly in a desiccator and then subjected to
oxidation with air.
Compressed air was used for oxidation at temperatures between 600 to 1000 °C with step of 100
°C and dwelling times of 15 mins to 4 hours were applied.
3.3.2 Liquid State Oxidation
Liquid state oxidation was carried out in a vertical tube furnace. The furnace consisted of an 8 cm
diameter alumina working tube, molybdenum disilicide heating elements and type B
thermocouples, as shown in Figure 3-7. Compressed air was used as the oxidant and the influence
of roasting temperature, time and slag basicity (by addition of SiO2) were investigated in this phase.
A typical procedure was as following. After heating the furnace to the set temperature (of 1400,
1500 or 1550 °C), about 25 g of calcined slag sample was placed in an MgO crucible. The crucible
was slowly raised to the hot zone of the furnace from the bottom using an alumina foam pedestal.
This was done in approximately 15 minutes to reduce the risk of crucible cracking due to thermal
shocks. The slag would then be left to melt for 30 minutes. Air was then introduced into the top of
the crucible at the rate of 600 mL/min through a 6mm alumina tube. The air flow rate was increased
slowly to the target level in ~ 1 min to reduce slag splashing. After the experiments, the slags were
rapidly cooled by quenching the crucible into a water bath. The slag was then removed from
61
crucible and dried in a muffle furnace. The obtained product was crushed to the same particle size
as raw slag by successive sieving (90% passing 150 µm and 60% passing 45 µm). The pulverized
slag was then subjected to magnetic separation.
Figure 3-7: Schematic diagram of the setup used in liquid state oxidation of slag.
3.3.3 Magnetic Separation
The magnetic separator used in this experiment was a Davis Tube (Figure 3-8), model CXG－
ZN50, by Tangshan Real and Pure Technologies Co. To separate magnetic content from slag,
about 10 g of slag sample was mixed with suitable amount of alcohol and 200 ml water. The
mixture was stirred thoroughly to fully wet the slag particles. The glass tube was filled with clear
water, about 5 cm above the magnetic poles, while turning on the electromagnet to a controlled
field intensity. Oscillating motion was applied to the tube using an electric motor. The slag slurry
was then slowly charged into the glass tube, while shaking the container to ensure the particles are
suspended. The magnetic particles adhere to the poles, nonmagnetic particles are washed away
with the discharging liquid. The magnetic fraction was washed with a gentle and constant-flow
stream of water until the discharge contained no visible solids. After filtration and drying, the iron-
62
rich magnetic fraction and iron-lean nonmagnetic fraction were recovered. The mass of
concentrate was recorded as C. The obtained tailing and concentrate were analyzed by XRF and
examined in a SEM. The metal grade in the concentrate was reported as c.
Figure 3-8: Schematic diagram of the Davis tube magnetic separator.
Magneticmaterial
Non-magnetic fluid mixture
Slag fluid mixture
Electromagnet
Glass tube
Front viewSide view
63
3.4 Experiment Parameters
The experiment variables and parameters are shown in Table 3-2. The effect of magnetic field
strength, oxidizing time, temperature and slag basicity were evaluated.
Table 3-2: Experimental variables and their ranges.
Variables Fixed
Solid State
Field Strength (mT)
50
1 hr, 900˚C
55
60
70
80
100
110
120
130
Roasting Time (hr)
0.25
110 mT, 900˚C
0.5
0.75
1
2
4
19
Temperature (˚C)
600
110 mT, 1hr 700
800
900
1000
Liquid State
Temperature (˚C)
1400
30 mins, 110 mT 1500
1550
Blow Time (min)
15
1400 °C, 110 mT 30
45
15
1550 °C, 110 mT 30
60
S/C
2.91
1550 °C,30 mins, 110 mT 2.5
2
1.5
64
From each magnetic separation, concentrate fraction is defined as following:
𝐶𝑜𝑛𝑐𝑒𝑛𝑡𝑟𝑎𝑡𝑒 𝐹𝑟𝑎𝑐𝑡𝑖𝑜𝑛 =𝐶
𝐹
(18)
where C = mass of concentrate,
F = mass of feed,
In the next chapter, the concentrate fraction is presented in % (=C/F x 100). The efficiency of the
oxidation and separation processes is typically evaluated by the recovery of the metal, Rm, and the
grade of iron in concentrate, c%. The calculation of Rm is shown in following equation:
𝑅𝑚 =𝐶 ∗ 𝑐
𝐹 ∗ 𝑓
(19)
where c = metal grade in concentrate,
f = metal grade in feed.
Although these are good indicators of the process performance for a given set of conditions, they
are inadequate when studying the effect of a variable. The reason is that generally grade and
recovery have an inverse relationship and one cannot simply decide between the better of high
recovery-low grade and low grade-high recovery scenarios. To address this, Separation Efficiency
(S.E.), introduced by Shukz [100], is used as a unified index that captures both the grade and
recovery. It is defined as:
𝑆. 𝐸. = 𝑅𝑚 − 𝑅𝑔 =100𝐶𝑚(𝑐 − 𝑓)
(𝑚 − 𝑓)
(20)
where Rm = % recovery of the valuable mineral,
Rg = % recovery of the gangue into the concentrate,
m = percentage of target metal in the valuable mineral[101].
Calculated parameters according to experimental conditions are summarized in Appendix A
65
3.5 Analytical Techniques
Hitachi SU3500 Scanning Electron Microscope (SEM) and JEOL 6610LV Scanning Electron
Microscope both with Energy Dispersive X- Ray Spectroscopy (EDS) system were employed for
micro scale imaging and EDS elemental analysis. Electron Probe X-Ray Microanalyzer (EPMA)
was employed for quantitative spot analysis with its Wavelength Dispersive Spectrometer (WDS)
system. Philips X-ray diffraction system (CuΚα radiation) was used for mineral phase analysis,
with slow scanning rate of 0.02 step size (2θ) and 2.0 s scan step time from 25° to 65°. X-ray
fluoroscopy (XRF) was employed for determining the slag composition (XRF analysis results are
summarized in Appendix B). The particles for both X-ray analyses should be very small, ideally
near 1 μm. Laser particle size analyzer (L-PSA) was used for particle size measurement.
66
Chapter 4
Results and Discussion
4.1 Characterization of Raw Slag
The chemical analysis of slag was performed using XRF and the results are compared with those
provided by the slag supplier, ArcelorMittal Dofasco. Table 4-1 shows the results, normalized to
100% total. The difference between the two analyses was around 0.1 to 0.2 percent and well within
the acceptable range of variation.
Table 4-1: XRF results comparison between Dofasco and U of T analysis.
Weight
%
Total
Fe FeO CaO SiO2 Al2O3 MgO MnO P2O5 TiO2 Cr2O3 Balance C/S
AM-
Dofasco 18.4 23.7 42.5 14.6 4.3 8.8 3.6 0.9 1.2 0.5 0 2.9
U of T 18.3 23.6 42.3 14.8 3.8 8.5 3.0 0.8 1.1 1.1 1.0 2.8
4.1.1 RO Phase
As mentioned in the literature review. RO phase is commonly found in steelmaking slag. RO is a
solid solution of FeO-MnO-MgO in a cubic structure similar to MnO, CaO and MgO [12]. Exactly
one divalent iron ion and two trivalent iron ions are required to construct one magnetite molecular
(Fe3+ [Fe2+ Fe3+] O4). Accordingly, only RO phase contains Fe2+ ions in steel slag that is suitable
to form magnetite under proper oxidizing atmosphere.
According to the EDS elemental mapping in Figure 3-2 and XRD analysis shown in Figure 3-3,
iron species are mainly present as dissolved wüstite and spinel ferrites in the RO and dicalcium
ferrite (C2F) phases. A BSE image gives a clearer picture of phase distribution (Figure 4-1). Based
on EDS analysis of various phases, narrow white regions correspond to a small amount of metallic
iron (point D). The lighter gray smooth region is a mixture of C2F, C4AF (where Al3+ and Fe3+
ions can switch sites, points B & E), and the RO phase is the rough lighter gray region (complete
MgO-FeO-MnO solid solution, points A & E). It is difficult to differentiate the various phases in
67
this region due to the mixture of constituents: Mg, Mn, Fe and Al, Ca, Fe sharing similar atomic
masses, giving the same grayscale contrast. The dark region (point C) is a mixture of C2S and C3S.
(wt%) Fe Mn Mg
Point A 35 9 20
Point E 47 10 13
Figure 4-1: BSE image of raw slags with phases identification of: points A & E, RO phase,
their compositions are given in the table (balanced with O wt% and minor Ca wt%); point
B & F: C4AF phase; C: C2S and C3S phases; D: metallic iron.
It was first thought that the RO phase will have a constant composition across. By comparing
points A and E in Figure 4-1, it is seen that Fe, Mn, and Mg can substitute one another in this
phase. As RO is the expected Fe-bearing phase for magnetic concentrate, WDS analysis was
obtained on 42 points in the RO phase with results provided in Table 4-2. The composition range
is quite wide with FeO varying from 21 to 77 wt%. In future analysis, when S.E. (defined on P63)
is to be estimated, the average Fe content of the RO phase obtained here (m=44.9 wt%) will be
used.
68
Table 4-2: Summary of WDS analysis of RO phase in the raw slag (42 points)
Mol% FeO MnO MgO
Average 48.2 8.2 40.3
Max. value 77.6 11.0 70.3
Min. value 20.9 4.5 13.4
STD 13.4 1.7 13.8
Grade (wt%) Fe Mn Mg
Average 44.9 7.4 16.5
Max. value 69.3 10.0 34.1
Min. value 23.3 1.9 3.2
STD 9.9 1.8 7.1
The distributions of RO phase and all iron-bearing species (RO + C2F phase), are shown in Figure
4-2, where the small amount of wüstite phase is considered to be RO. This image was obtained
based on EDS elemental mapping followed by image processing using ImageJ (by National
Institutes of Health). Between 18 and 23% of slag cross-section area (or approximately volume
percent) is occupied with RO phase (a), while RO and C2F phase (b) together comprise more than
46 vol% of steel slag sample. The remaining is calcium silicate phase. The average grain size for
the RO was also calculated to be 30 µm ± 10 µm.
Figure 4-2: Processed image for phase distribution of Fe-species in raw steel slag.
69
4.1.2 Response of Raw Slag to Magnetic Separation
Magnetic separation was first performed on the raw slag sample, as a baseline for slag upgrading.
The initial magnetic field strength was selected to be 100 mT as a recommended value for
separating magnetite. This yielded a concentrate fraction of 8.4%, increasing the field strength to
110 mT raised this value to 9.0%. The recovery, S.E. and Fe grade of the produced concentrate are
presented inTable 4-3. The m value is obtained assuming iron is totally contained in the RO phase,
then m = 44.9%. Figure 4-3 displays a pie-chart distribution of iron and other species in the
magnetic and non-magnetic fractions of the slag. It turns out that less than 10% of magnetic
fraction can be separated from raw slag (with particle size d80=133 µm, d20=11 µm). The recovery
was 13.6%, more than 85% of iron went to tailing. This indicates that either iron was not in
magnetic form (magnetite or spinel from RO), the separation process was inadequate, or both.
Table 4-3: Recovery, grade, and S.E. of iron recovery from raw (B=110 mT).
Feed Concentrate Recovery
(%)
S.E.
(%) Mass (g) %Fe Mass (g) %Fe
100 18.3 9 27.7 13.6 7.8
Figure 4-3: Schematic diagram of magnetic separation.
The XRD analysis for magnetic and nonmagnetic fractions are shown in Figure 4-4. The patterns
reveal that most of the spinel phase reported to the magnetic fraction with some wüstite carried
over, implying that spinel phase and small amount of RO phase are separated from the rest of slag.
Also, calcium silicates and calcium ferrites patterns are found in magnetic fraction but in small
70
quantities. Iron is present in the non-magnetic fraction in the form of wustite and in RO and
calcium ferrites. Calcium carbonate is formed possibly from free lime during magnetic separation
process, where the slag is mixed with water and later dried in oven at 100 °C with access to CO2
in atmosphere. The strong CaCO3 peak is also contributed by overlapping peak with C3S phase.
(a)
(b)
Figure 4-4: XRD patterns for the magnetic (a) and non-magnetic (b) fractions of the raw
slag.
71
4.2 Thermodynamic Assessment
A thermodynamic evaluation of the slag system oxidized under different conditions was carried
out in order to narrow down the range of experimental conditions. Based on the slag chemistry
obtained from XRF results, a phase stability diagram was produced using FactSage™ 7.0, Figure
4-5. The slag composition used to construct this diagram was obtained by normalizing the FeO,
CaO, SiO2, Al2O3, MgO and MnO contents to make up 100% of slag. They constituted 96% of the
actual slag mass. On the diagram, a vertical line labeled as “Air Line” indicates logpO2 = -0.68 the
partial pressure of oxygen in air.
This slag solidus temperature is approximately 1100 °C. Along the air line, many familiar phases
that were previously found in XRD and SEM analysis, such as C2S, C2AF are seen. However, the
phase such as C3MA4 (Ca3MgAl4O10) was not observed. This is probably due to the unknown
cooling history of the sample. The diagram also shows that spinel formation is possible in air below
1100°C. once melting begins, spinel turns into liquid. This means oxidation can still happen and
magnetic phases could precipitate out on cooling.
The phase diagram is also compared with that produced by Semykina et al. for synthetic CaO-
SiO2-FeO(+MnO) slags [102]. The phase diagram here is very different from their CSF slag with
C/S=1 and high concentration of FeO (or FeO + MnO), due to compositional difference. In the
CSA system, Figure 4-6 (a), FeO can be fully oxidized to Fe2O3 in air. In the presence of MnO, as
Figure 4-6 (b) depicts, spinel formation is strongly promoted in the entire temperature range in air.
In general, for all slag systems discussed here, low oxygen partial pressure is conductive to spinel
formation. High oxygen potential (e.g. air) might be preferred because of ease of use, cost, and
possibly faster kinetics of oxidation.
72
FeO CaO SiO2 Al2O3 MgO MnO
Wt% 23.6 42.3 14.8 3.8 8.5 3.0
Figure 4-5: Phase diagram of slag (composition given above), calculated by FactSage 7.0.
C3MS2: Ca2MgS2O8; C2AF: Ca2(Al, Fe)2O5; C3MA4: Ca3MgAl4O10.
73
(a)
(b)
Figure 4-6: Phase diagrams of synthetic slags (composition given) constructed by Semykina
et al.[102] using FactSage 6.4. (Temperature unit is K, pressure unit is log(Pa) and air
atmosphere is represented by line A)
74
The mass fractions of phases produced in air are plotted in Figure 4-7. Solid circles indicating
mixed spinel phase consist of free MgO, Mn2O3 and iron rich spinel phases. It is found that the
RO weight fraction is very low under equilibrium conditions, throughout the entire temperature
range. Moreover, the weight ratio between silicates, RO phase and C2AF phase are relatively
stable at low temperature. The melting of individual phases is related to their iron content: the iron-
rich spinel phase melts first, followed by C2AF. The calcium silicates remain solid until 1400°C.
This melting behavior of steel slag is confirmed in the literature [103].
Figure 4-7: Proportions of different minerals in slag equilibrated with air against
temperature.
The effect of FeO content in slag was examined by increasing FeO to 40wt% while keeping the
ratio of other constitutes the same. The results are plotted in Figure 4-8. The spinel phase is much
greater here, reaching 20 wt% before melting. Adding FeO also significantly reduced the solidus
temperature to 900 °C.
0
20
40
60
80
100
600 800 1000 1200 1400 1600
Wei
ght
Per
cen
t (%
)
Temperature (°C)
C2AF R-O Silicates Slag Liquid C3MA4
75
Figure 4-8: Proportions of different minerals in slag with 40 wt% FeO.
The effect of basicity on phase evolution is shown in Figure 4-9. It is as expected that with decrease
in basicity decreases, CaO is unlocked from iron (C2AF), releasing FeO and Al2O3 which can then
form spinel phases. This is the reason for increase in spinel formation as basicity drops from 2.9
to ~ 2. Further decrease in basicity results in dissolution of FeO in the silicate slag, reducing the
spinel formation.
Figure 4-9: Proportions of different minerals in slag against basicity (pO2=0.21, T= 1000
°C).
0
20
40
60
80
100
600 800 1000 1200 1400 1600
Wei
ght
Per
cen
t (%
)
Temperature (°C)
C2AF R-O Phase Silcates Slag liquid
0
20
40
60
80
100
1.4 1.7 2 2.3 2.6 2.9
Wei
ght
Per
cen
t (%
)
Basicity C/S
C2AF Spinel Silicates MgO
76
4.3 Solid State Oxidation
4.3.1 Roasting with CO2
A set of preliminary experiments on CO2 roasting of slag were conducted as the thermodynamic
evaluation showed favorable conditions for spinel formation established by pure CO2 (pO2 =10-20
to 10-10 Pa from 600 to 1000 °C). Calcine from roasting was submitted for XRD phase
identification, and the results are shown in Figure 4-10. The most significant transformation taking
place as a result of this treatment is extensive development of spinel phase. In the meantime, the
dominant wüstite in raw slag almost disappears in calcine, implying the source of spinel formation.
Another observation of this process is the evolution of calcium carbonate at low temperature range,
as a product of excess lime and CO2 reaction. The carbonate is not stable at high temperature.
The 2θ range from 42 to 44 degrees was further analyzed (as shown in Figure 4-11) to investigate
the transformation of wüstite to spinel from the RO phase. As seen, on roasting, the strong wüstite
peak of the raw slag’s pattern disappears and another peak on its right, where spinel must be,
emerges. The higher the roasting temperature, the more complete the transformation.
A quantitative determination of magnetic phases in the roasted slag could not be made using XRD,
although attempts were made. This is because the composition of individual grain varies, and there
is no pure RO phase available to make an internal standard for quantitative XRD. Therefore, XRD
was only used for phase identification and a combination of XRF and magnetic separation data to
yield information on the magnetic fraction of the material. This is perhaps more interesting from
the practical point of view as it shows the effective formation and separation of spinel phases from
slag.
Magnetic separation was carried out for CO2 roasted calcine under a 100 mT magnetic field
strength. The effect of operating time and temperature on concentrate fraction is plotted in Figure
4-12. The concentrate fraction presented at time zero is that of raw slag. Increasing temperature
and roasting time both result in improved formation (and separation) of magnetic phases. This
could be a consequence of dual effects of generating more magnetic phases, as well as larger spinel
crystals that are more responsive to magnetic separation.
77
(a)
(b)
Figure 4-10: XRD patterns for selected CO2-roasted samples at (a) 600°C and (b) 700°C.
78
Figure 4-11: XRD patterns of raw slag and slag roasted with CO2.
The results of a preliminary test on air roasting of slag is also presented in the plot, which shows
improved separation. It appears that in air roasting, the reaction almost reaches the equilibrium in
approximately 60 minutes, while the slope of CO2 roasting at the same time is still high, indicating
incomplete reaction. The conclusion here may be that oxidation in air is faster, which is expected
noting the larger partial pressure of oxygen.
Noting that air proved to be more effective than CO2, the bulk of experiments were carried out
using air as the oxidant.
Figure 4-12: Concentrate fraction as a function of time and temperature for slags roasted
in air and CO2 and subjected to 100 mT magnetic separation.
79
4.3.2 Roasting with Air
The effect of roasting time and temperature as well as magnetic field strength on the formation
and separation of spinel phases was investigated and will be discussed in the following sections.
The particle size of the treated sample was fixed at d80=133 µm, d20=11 µm for all materials.
4.3.2.1 Effect of Magnetic Field Strength
At first, the optimum magnetic field strength of the Davis Tube separation was determined. As
shown in Figure 2-13, during magnetic separation, magnetic force competes with hydrodynamic
drag and gravity. Assuming that particle size and fluid flow are similar, the hydrodynamic drag
and gravity will be fixed. Thus, magnetic field strength can be the most important factor that
influences the separation process. An optimum value of field strength can be first determined and
applied to the materials treated under other conditions.
Figure 4-13: Concentrate fraction (% magnetic fraction of feed) vs. magnetic field strength
(roasted in air at 900°C for 1hr).
The magnetic fields strength (B) was varied from 50 to 130 mT. The concentrate fraction (magnetic
product as percent of the feed) increased with the increase of field strength, as shown in Figure
4-13. There is a gradual increase in the concentrate fraction for B up to 100 mT, at which point
there is an abrupt increase, followed by a plateau. The gradual increase in the first part may be
because of increased separation of those particles containing lesser amounts of magnetic phases.
0.0%
20.0%
40.0%
60.0%
80.0%
100.0%
50 60 70 80 90 100 110 120 130
Co
nce
ntr
ate
Frac
tio
n (
%)
B=Magnetic Field Strength (mT)
80
The sharp increase between B=100-110 mT is likely because another phase with lower magnetic
susceptibility gains enough magnetic force and responds to the separation at this field strength.
The maximum concentrate fraction reached in these tests is 54% at B=110 mT.
Figure 4-14 shows the recovery of Fe to the concentrate, the concentrate grade, and separation
efficiency calculated based on % Fe in the spinel RO mineral of (m = 44.9). Interestingly, the iron
grade remained steady at approximately 28%, an increase of ~10% from the starting material (%
Fe in slag =18.7%). The recovery and S.E. both increase for B up to 110 mT and then remain
constant. The highest Fe recovery achieved is 81%.
Figure 4-14: Effect of magnetic field strength on iron grade and recovery as well as
separation efficiency (roasted in air at 900°C for 1hr).
The increasing Fe recovery while grade is constant points to an interesting feature in the separation
process. It indicates that while with increase of B, more Fe is recovered, proportional amounts of
gangue also report to the concentrate. In other words, each particle reporting to the concentrate can
be assumed to have a fixed ratio of Fe and gangue, showing fine dissemination of Fe in the treated
slag. To improve the grade further, fine grinding or providing conditions that allow growth of
spinel phases are required. Similar trends in recovery were observed by Ma et al. for oxidized
nickel slags [92].
0%
20%
40%
60%
80%
100%
55 65 75 85 95 105 115 125
Per
cen
tage
(%
)
Magnetic Field Strength (mT)
Iron Grade
Iron Recovery
Iron Separation Efficiency
81
The optimum B for the conditions of these tests (particle size of d80=133 µm, d20=11 µm, 1 h
treatment in air at 900 °C) was selected based on the highest separation efficiency, being 48% at
B=110 mT.
4.3.2.2 Effect of Roasting Time
The effect of roasting time to iron upgrading from steel slag was examined from 15 minutes to 19
hours. The results from 0 minutes (untreated raw slag) to 4 hours are shown in Figure 4-15. As
seen, iron grade of the concentrate remained constant, while iron recovery reached a maximum
within 15 minutes, and remained constant afterwards. The results of 19 hours roasting are not
presented in this figure but followed the same trend. Comparing the results with those in the
preliminary experiments at lower temperature (Figure 4-12) show that the greatest recovery is
established much faster at higher temperature. Roasting time of 1 h was selected for future
experiments.
Figure 4-15: Effect of roasting time on grade, recovery and S.E. of iron in the concentrate
(air roasting at 900°C, B=110 mT).
0%
20%
40%
60%
80%
100%
0 0.5 1 1.5 2 2.5 3 3.5 4
Per
cen
tage
(%
)
Time (hr)
Iron Grade
Iron Recovery
Iron Separation Efficiency
82
4.3.2.3 Effect of Roasting Temperature
The effect of roasting temperature was tested from 600 to 1000 °C at 1 hour. The results are
provided in Figure 4-16. Increasing iron recovery and separation efficiency is realized with higher
roasting temperature while the grade is unaffected. As expected from the thermodynamic analysis,
roasting temperature below the solidus (~1100°C) should produce the highest amount of spinel
phases in steel slag, provided that equilibrium is reached. The amount of spinel phases based on
the system thermodynamics, as seen in Figure 4-7, should remain constant in the temperature range
of experiments (600-1000 °C). However, greater recovery at higher temperatures (Figure 4-16)
suggest that this is not the case. This could be due to faster oxidation kinetics at higher
temperatures.
Figure 4-16: Effect of roasting temperature on iron grade, recovery, and separation (air
roasting for 1 hour, magnetic field strength=110 mT).
The raw slag contained 23.6% FeO. If oxidized to Fe3O4, a mass gain of 1.77% should be
registered. Setting this as complete oxidation state to magnetite (100% of Fe species present as
magnetite), and measuring mass change of slag after oxidation, Figure 4-17 could be plotted. The
figure shows the effect of temperature on magnetite formation. Temperature has a strong effect on
the oxidation, pushing the magnetite formation to near completion at 1000 °C.
0%
20%
40%
60%
80%
100%
600 650 700 750 800 850 900 950 1000
Per
cen
tage
(%
)
Temperature (°C )
Iron Grade
Iron Recovery
Iron Separation Efficiency
83
Figure 4-17: Calculated percentage magnetite formation based on mass gain vs.
temperature (roasting time = 1 h).
4.3.2.4 Phase Evolution in Roasting of Slag
The XRD patterns of slag roasted at 900 °C for 1hr and its magnetic, and nonmagnetic fractions
are shown in Figure 4-18. Comparing with raw slag sample, cleaner peaks and higher peak
intensities are observed, illustrating higher crystallinity and grain growth during roasting. Also, as
discussed earlier, RO phase peaks emerge and become stronger with further oxidation. It is also
evident from Figure 4-18 that on magnetic separation, magnetized RO phases are concentrated in
the magnetic fraction.
Complete transition from wüstite to spinel phase is shown in Figure 4-19. The effect of temperature
and oxidation time is also shown in this figure. It is seen that at 700°C and 10 min roasting, the
shift from FeO to spinel from RO phase is half complete (from the position of peaks between 42-
43 deg.). Extending the time to 1 h seems to have completed the conversion. Higher temperature
produced stronger spinel peaks which is an indication of greater amounts of spinel and larger
crystals. This latter mechanism however has a smaller effect, if any, as will be discussed later that
segregation and growth of RO phases during roasting is not pronounced.
0.00%
0.25%
0.50%
0.75%
1.00%
1.25%
1.50%
1.75%
0%
20%
40%
60%
80%
100%
600 700 800 900 1000
Mas
s In
crea
se (
%)
Cal
cula
ted
Mag
net
ite
Form
atio
n (
%)
Temperature
84
(a)
(b)
(c)
Figure 4-18: XRD patterns for 900 °C, 1hr slag sample (a) and its magnetic fraction (b) and
nonmagnetic fraction (c), separated via 110 mT.
85
Figure 4-19: XRD patterns of raw slag, and slag roasted in air at different conditions.
4.3.2.5 Magnetic Separation of Slag Roasted Under Optimized Conditions
Magnetic separation was carried out for raw slag and slag roasted under optimized conditions (1h,
900 °C). Particle size and magnetic field strength were the same for both materials. Table 4-4
shows the results. The calculation of S.E. can be made assuming that magnetite is the iron species
separated (m = 72.4%). However, as discussed earlier, most of iron is present in a mixed spinel
phase of RO with much lower iron grade (mave = 44.9%). Using this value, the calculated S.E. for
raw and roasted slag are 7.8% and 47.7% respectively. Recovery of Fe also increased from 13.6%
to 82.3%, showing a significant improvement. The iron grade of the concentrate is however
unaffected by the treatment. These data reveal the primary effect of roasting: it converts non-
magnetic iron species to magnetic phases but does not improve the segregation or growth of such
phases enough so that they form liberated particles on crushing. This is at first appearing as
contradicting the XRD patterns. The patterns of magnetic and nonmagnetic fraction show much
sharper peaks compared to raw slag separation, which might be an indication of strong phase
separation. In addition, the spinel/C2S peak intensity ratio for raw slag was 2.4, lower than 3.0 in
roasted slag. These all indicate transformation of iron species to RO, which is then dissociated
from other slag constituents. This phase dissociation (or separation) is however expected to be in
86
microscale, too fine to allow sufficient liberation of the RO during grinding. In other words, iron
species become magnetic spinel phases, but their size and distribution are not affected so that on
separation, they still carry a large amount of gangue with them. As a result, more of such RO-
gangue “agglomerate” particles are separated (higher recovery) but the grade remains the same.
Table 4-4: Comparison of magnetic separation performance (via 110 mT) between raw slag
and optimally oxidized slag.
Mineral Sample Feed Concentrate
Fe Recovery (%) S.E. (%) Mass (g) %Fe Mass (g) %Fe
RO Phase
(m=44.9%)
Raw Slag 100 18.3 9 27.7 13.6 7.8
900 °C, 1hr 100 18.3 54 27.9 82.3 47.7
The above conclusion is more clearly seen by studying the effect of magnetic field strength on the
grade of recovered metals in the concentrate, Figure 4-20. The average values stand at Fe=28%,
Mn=3.7% and Mg=7.4%. As seen, deviations for each metal are small, suggesting that similar
proportions of RO and gangue report to the concentrate, regardless of the field strength. This is
consistent with the previous conclusion and shows that with increasing magnetic field strength,
only larger amounts of materials are separated as concentrate, but mineral proportions remain the
same, i.e. poor liberation. How stronger field results in collection of more particles of the same
mineral complex is perhaps related to the particle size. Those fine particles that would flow out
with the water in the Davis Tube are more likely to be captured when the magnetic field is stronger.
Figure 4-20: Grade of iron, manganese and magnesium in concentrate with magnetic field
strength from 55 to 125 (roasting in air at 900°C, for 1hr).
0.0
10.0
20.0
30.0
40.0
55 65 75 85 95 105 115 125
Per
cen
tage
(%
)
Magnetic Field Strength (mT)
Iron Grade
Manganese Grade
Magnesium Grade
87
For good liberation of minerals from gangues, Gaudin suggested d75 to be 1/10th of the size of
mineral grain size after comminution [104]. The RO grain size was estimated to be 30 µm ± 10
µm (from Figure 3-4), so the liberation size should be 3 µm indicating a relatively poor liberation
for the conditions of these experiments.
Further grinding of the slag should result in improved grade. In grinding, typically four types of
particles are generated, as schematically shown in Figure 4-21. Particles of type 1 have low-level
locking and are classified as concentrate. Particle type 4 would be classified as tailing. And
particles of types 2 and 3 are classified as middling, and might need further grinding. Figure 4-22
exhibits the magnetic fraction of roasted slag powder, which shows all 4 types of particles,
indicated as 1 to 4. In this figure, lighter region (white and light gray) are all RO phases, whereas
white regions (1a) are rich in iron and light gray regions (1b) are rich in magnesium. A small
fraction of type 4 particles was also carried over in this sample. A significant amount of middling,
especially type 3 with fine grains, was observed. This confirms why effective separation of RO
from other phases is difficult, they are highly intergrown, forming middlings that are carried to the
concentrate, hence the low grade.
Figure 4-21: Different particles generated during size reduction[104].
88
Figure 4-22: BSE image (200 magnification) of magnetic fraction of steel slag roasted in air
at 900°C, 1hr and magnetically separated under 110 mT. Type 1 to 4 particles are marked.
1a particles: RO phase rich in iron; 1b particles: RO phase rich in magnesium.
Figure 4-23 shows BSE images of the nonmagnetic fraction of the oxidized slag. Compared to the
magnetic fraction, a large population of fine particles is observed. This non-uniform distribution
of particles between magnetic and non-magnetic fractions shows preferred fracture during
commination and agrees with the double peak distribution as shown in Figure 3-4. No RO phase
was observed in this image, while it is rarely found globally. This phenomenon is also shown in
XRF analysis of the nonmagnetic fraction in Table 4-5. As seen, the Fe content of tailings is
substantially lower than that of the feed, hence an iron-lean product. This could potentially be a
useful material in cement production as in cement clinker, up to 8% Fe is desired. Therefore, one
of the objective of the research which was dividing the slag into an iron-rich concentrate and iron-
lean by-product, both as useful materials is achieved. In addition, C2S (point A), C3S (point B)
89
and free lime (point D) were observed as the abundant phases in this non-magnetic fraction. The
light regions were detected to be C2F phase (point C) with iron grade about 20%. This represents
iron loss to tailings. Based on thermodynamic analysis, C2F is always present at equilibrium and
the loss caused by C2F is inevitable. Enhanced presence of C3S, C2S. C4AF and C2F endorses
slag’s cementitious properties [12], implying the possibility of a Portland cement feedstock.
Figure 4-23: BSE imaging (200 magnification) for nonmagnetic fraction of steel slag
roasted in air at 900°C, 1hr and magnetically separated under 110 mT. A: C2S; B: C3S; C:
C2F; D: Free lime.
Based on previous results and discussion, one can conclude that the metal grade in concentrate are
independent of magnetic field strength, roasting time and temperature. Metal grades can be
optimized by further reduction of particle size which leads to higher liberation degree. Where
magnetic field strength controls the amount of magnetic fraction separated, roasting time and
temperature control the magnetic material formation. Under optimum conditions (roasting with air
90
at 900 °C for 1 hour), Fe recovery of 84% and 28% Fe grade and Mn recovery of 91% and Mn
grade of 4% were achieved. The final concentrate and tailing compositions are shown in Table
4-5, where concentrate is rich in iron, magnesium and manganese oxides, and the tailing is lean in
metal content. A summary of separation parameters is provided in Table 4-6.
Table 4-5: XRF result of separation products under optimum conditions, comparing with
raw slag (wt%).
Wt% FeO CaO SiO2 Al2O3 MgO MnO P2O5 TiO2 Cr2O3 Balance
Raw Slag 23.6 42.3 14.8 3.8 8.5 3.0 0.8 1.1 1.1 1.0
Concentrate 37.6 28.9 9.2 3.3 11.9 5.0 0.8 0.9 0.9 1.4
Tailing 9.1 56.5 20.9 4.3 4.0 0.6 0.7 1.4 1.3 1.2
In an exploratory experiment and in order to confirm the effect of particle size, an oxidation test
was conducted under the optimized conditions and then the roasted slag was subjected to size
reduction to reach d90=45 µm (compared to d80=133 µm, d20=11 µm prior to grinding). The results
of this experiment are compared with the base case in Table 4-6. The iron content in the
concentrate is improved to 34%, but recovery has dropped from 82% to 59%. The inverse grade-
recovery relationship is well established and expected. Noting this, future works are required to
obtain an optimum particle size for achieving the desired grade.
Table 4-6: Summary of separation performance under optimum conditions, and the effect
of particle size.
Particle Size
(µm) Products
Fraction
(%)
Fe Grade
(%)
Fe
Recovery
(%)
Mn Grade
(%)
Mn
Recovery
(%)
d80=133,
d20=11
Concentrate 54 28 82 3.9 91
Tailing 46 7 18 0.4 9
d90=45 Concentrate 32 34 59 5.0 69
Tailing 68 11 41 1.0 31
Ma et al. in oxidation of Ni slag obtained iron recovery and grade of 76% and 54%, respectively.
A higher recovery rate was achieved in the present study but the grade is lower. Their nickel slag
contained 39% iron compared with 18% here, which can result in larger RO grains and higher
grade [92]. In another study, Li et al. attempted to recover RO phase from steel slag by super
gravity. They have achieved grades of Fe=31.6%, Mn=2.5%, and recoveries of 60.4% and 72.24%
for Fe and Mn, respectively[103]. At larger particle size, with slightly lower iron grade in the
91
present study, the recovery values are significantly greater. With further reduction, the present
study is able to achieve similar recovery with higher iron grade. Noting the simplicity of solid
oxide roasting, a future study can look into optimizing the grade by for example fine grinding, and
establishing a cost-effective method of processing steel slags.
4.3.2.6 Manganese Recovery
Manganese is another valuable metal that is mainly present in the RO phase. Its grade in raw slag
according to XRF results, is 2.3%, which is lower that other slags in the literature. The separation
of Mn into the concentrate is showing an identical trend to iron separation, with good recovery and
relatively low grade. Small difference is found on the effect of roasting temperature. Figure 4-24
shows the results of Mn grade, recovery and SE against roasting temperature. The S.E. was
calculated based on m = 7.4% (Mn grade in the RO phase). Under optimum conditions of roasting
(1000 °C for 1 hour) Mn recovery reached 94% and S.E. was 56%. Mn grade in the concentrate is
consistent at an average value of 3.8%. Since Mn is only present in the R-O phase, one can
calculate the approximate amount of RO phase in the concentrate. Based on this, the concentrate
contains 51 wt% of RO phase and the remaining is the gangue carried over, whereas the raw slag
containing approximately 29% RO.
Figure 4-24: Effect of roasting temperature on Mn grade, recovery and SE (air roasting for
1 hour, magnetic field strength 110 mT).
0%
20%
40%
60%
80%
100%
600 650 700 750 800 850 900 950 1000
Per
cen
tage
(%
)
Temperature (°C )
Mn Grade
Mn recovery
Mn Separation Efficiency
92
4.4 Liquid State Oxidization
Solid-state roasting experiments revealed that although conditions are favorable for the formation
of magnetic spinel phases, growth and/or segregation of this phase from other minerals is very
limited. This resulted in poor liberation and subsequently low grade in the concentrate. It was
decided that performing the reactions at a temperature where phase separation is facilitated might
solve this problem. Therefore, the present study also investigated oxidation of slag in liquid state,
aiming to allow growth and separation of iron-rich minerals from other species.
Liquid-state slag oxidization was carried out according to the experimental parameters in Table
3-2. Three temperatures were examined, from 1400, 1500 to 1550 °C. As the slag samples were
melted at high temperature and then quenched in water, a few difficulties were encountered.
Crucibles would crack during transfer in and out of the furnace, and liquid slag could corrode the
crucible and leak out (Figure 4-25: Cross-sections of slag containing crucibles after treatment at
different temperatures (holding time 30 minutes).). As a result of severe interaction of slag with
the crucible that sometimes causes slag penetration to the outside of the crucible, mass change
during oxidation could not be reasonably measured. The slag however could be subjected to
grinding and magnetic separation.
Gas bubbles were seen in the slag for all the temperatures examined (Figure 4-25), suggesting
good mixing and gas-liquid interaction. The differences are in the size and number of bubbles; the
1400 °C sample contains numerous small bubbles, 1500 °C contains less but larger bubbles, and
1550 °C sample had only one bubble as a large cavity. It is known that lower viscosity contributes
to smaller bubble size [105] and slag with higher temperature would have lower viscosity. The
contradiction could be explained by the bubble release during cooling. When slowly removing
slag from furnace away from the air nozzle, lower temperature slag has less time for the bubbles
to merge and escape. As quenched in water, small bubbles are frozen in place. On the other hand,
for the slag with higher temperature (and lower viscosity), bubbles have a greater chance to merge
and escape. Those large bubbles seen at higher temperature are frozen in place due to quenching.
93
Figure 4-25: Cross-sections of slag containing crucibles after treatment at different
temperatures (holding time 30 minutes).
The XRD pattern of oxidized slag features characteristics of small grain size and low crystallinity
(Figure 4-26). The spinel phase exhibited a broad peak at 1550 °C, that according to Scherrer
equation is an indication of small grain size. Lower crystallinity of calcium silicate was also
revealed in the XRD pattern, with much lower intensity but broader peak. However, the intensity
of the spinel phase is strong and comparable to those slags oxidized at 900°C, pointing to
successful magnetizing treatment.
94
Figure 4-26: XRD pattern comparison of air treated slag at 900°C and 1550 °C. 2θ ranges
from 35 to 44 degrees.
4.4.1 Effect of Temperature and Time
The effect of temperature and oxidation time was examined, and the results are shown in Figure
4-27. Increasing temperature seems to have a small and positive effect on recovery and S.E.
between 1400 and 1500 °C while grade remains constant around 22%. The increase in recovery is
expected as kinetics of oxidation reactions, hence spinel formation, are greater at higher
temperature, as also seen in another work [102]. The activity of FeO is also greater at higher
temperature which can lead to larger amount of spinel phase. The recovery is much greater at 1550
°C. Increasing the oxidation time seems to improve the recovery and S.E. but only to a small
extent. The optimum conditions are oxidizing at 1550 °C for 1 hour, with iron recovery of 71%
and grade of 22% Fe. However, these figures are far from desired, as it was expected that liquid
phase oxidation improves the grade and recovery. The trends seen here are believed to be caused
by the small size of magnetized RO grains due to rapid cooling of the slag, as seen in Figure 4-28.
The RO grains are essentially sub-micron at 1500-1550 °C temperatures and ~ 1-10 µm for the
1400 °C-30 min treatment, which is far too small for liberation with particle sizes produced here
(d90=150µm). An evidence to this is higher grade at 1400 °C, which is believed to have been caused
by the larger RO phase grain size, as shown in Figure 4-28(c). The magnetized RO grains in this
95
sample are also more developed, i.e. separated from the matrix, which improves their liberation on
crushing. The larger grain size at 1400 °C might be an artifact of the rate of cooling of the slag: as
seen earlier, this slag contains a large number of air bubbles that decrease the overall thermal
conductivity of the slag and slow down the cooling of the slag on quenching. For slags at higher
temperature, such small air bubbles were not present, so a faster cooling and less time for growth
and segregation of RO phases is expected.
(a)
(b)
Figure 4-27: Effect of roasting temperature (a) and time (b) on iron grade, recovery, and
separation efficiency (air roasting, B=110 mT, roasting for 30 mins for (a)).
0%
20%
40%
60%
80%
100%
1400 1450 1500 1550
Pre
cen
tage
(%
)
Temperature (˚C)
Iron Grade
Iron Recovery
Iron Separation Efficiency
0%
20%
40%
60%
80%
100%
15 25 35 45 55
Per
cen
tage
(%
)
Time (min)
1550 °C, Fe Grade
1550 °C, Fe Recovery
1550 °C, Fe S.E.
1400 °C, Fe Grade
1400 °C, Fe Recovery
1400 °C, Fe S.E.
96
Figure 4-28: BSE imaging (2000 magnification) for steel slag oxidized by air at (a) 1550 °C
for 15 minutes; (b) 1550 °C for 30 minutes; (c) 1400 °C for 30 minutes and (d) 1500 °C for
30 minutes. A (light gray region): magnetized RO phase; B (black region): C2S, C3S; C
(dark gray region): C2F, C4AF region.
Comparing the results of liquid and solid state oxidations show lower recovery for the liquid state.
This can be explained as follows. The thermodynamic assessment of slag oxidation showed that
for oxidation in air, some iron should appear as non-spinel phases such as C2AF. Further, liquid
slag can dissolve some of Fe species, which when quenched, remain as non-magnetic phase. At
97
higher temperature, approach of equilibrium is more likely, hence reporting of Fe to these non-
magnetic phases is more pronounced.
4.4.2 Effect of Basicity
Region A White Magnetized RO phase (~40% Fe)
Region B Black C2S or C3S
Region C Gray C2F or C4AF
Region D White Spinel Mixture of Fe, Mg, Mn and Al (~35% Fe)
Region E Black Mixture of Si, Ca and Al oxides
Figure 4-29: BSE images for steel slags oxidized with air at 1550 °C for 30 minutes (a)
C/S=2.91, original slag; (b) C/S = 2.5; (c) C/S = 2 and (d) C/S = 1.5.
The basicity of slag, defined as the CaO/SiO2 ratio was varied in the range of 2.9 (raw slag) to 1.5
by addition of silica. Figure 4-29 shows the BSE images of slag together with phase identification
98
based on EDS analysis. The mineralogical compositions are consistent with the previous
thermodynamic assessment of the system. As silica is added, the solubility of calcium in the slag
increases, resulting in dissolution of free lime and C2F. With decrease in basicity, CaO in
dicalcium ferrite (C2F) started to form calcium silicate with silica, leaving iron and aluminum to
enter the RO phase as a mixed spinel (region D). As more silica was added, the solubility of
alumina in silicate was also increased. It is observed that in the slag with C/S =2, alumina is
distributed in both regions E and D. When the basicity was reduced to C/S = 1.5, most of alumina
is dissolved in the silicate matrix.
The results of magnetic separation on slags with different basicity oxidized at 1550°C are
presented in Figure 4-30. At lower basicity, the percent of feed reporting to concentrate
(concentrate %) is greater. Iron recovery follows the same trend, whereas grade and S.E. go
through a slight increase at higher basicity. The decrease in recovery at higher C/S is justified
using the thermodynamic assessment made earlier. As seen, with increase in basicity, the amount
of C2AF phase increases, resulting in less spinel formation, thus yielding less material to
concentrate.
Figure 4-30: Effect of basicity on % concentrate, grade, recovery and S.E. of iron (air
oxidation at 1550 °C for 30 minutes and B=110 mT).
0%
20%
40%
60%
80%
100%
1.5 2 2.5 3
Per
cen
tage
(%
)
Basicity (C/S)
Iron Grade
Iron Recovery
Iron Separation Efficiency
Magnetic Fraction
99
The grade of iron in the feed drops with addition of silica because of dilution as Table 4-7 shows.
Consequently, the concentrate grade follows the same trend. In other words, the variation in grade
with basicity is an artifact of slag dilution and not a fundamental change in phase evolution or
segregation.
Table 4-7: Effect of silica addition (change in basicity) on grade of feed and concentrate
Basicity Feed grade (%) Concentrate grade (%)
2.91 18.3 21.8
2.5 17.9 19.0
2 16.5 17.5
1.5 15.1 16.2
In summary, liquid-phase oxidation did not improve the recovery of Fe into an Fe-rich concentrate.
The recovery and grade were lower than in solid state oxidation. This is because of the very fine
grains of spinel phases which was likely caused by rapid cooling of the slag, as also reported by
other researchers [89], [106]. Improvements can be made by allowing slag to cool slowly or
grinding it further. For the RO grain sizes obtained under fast cooling conditions, grinding to a
particle size of -5 µm is required to have a reasonable liberation. For a process just to recover iron
and manganese from steel slag, intensive grinding is not economical. The alternative solution is to
have a slow cooling rate after slag oxidization, encouraging grain growth.
As a proof-of-concept and guideline for future research, slow cooling of one slag was attempted.
The slag was melted and oxidized under the optimized conditions (1500 °C for 1 hour), then cooled
at 5.3 °C/min (as opposed to quenching) and separated at B=110 mT while the particle size was
kept the same. The iron grade of the concentrate was 24.9% Fe and recovery was 57%. The
separation efficiency is improved to 25%. The increase in grade and S.E. both point to improved
segregation of the iron-bearing phases. This improvement is however not significant, indicating
that much slower cooling rates should be examined. In a comparable study done by Ma et al., the
cooling rate for oxidized nickel slag was 5 °C/min and they obtained magnetite grain sizes ranging
from 10 to 50 µm [92].
100
Chapter 5
Conclusions
The present study focused on possible concentration of iron species from steel slags in air-
atomization processes into an iron-rich by-product for recycling into ironmaking processes. This
will also produce an iron-lean product suitable for applications such as cement making. In order
to simulate the conditions pertaining to slag atomized by air, oxidation of slag in both liquid and
solid states was investigated. Parameters such as oxidation time and temperature, as well as the
field strength in magnetic separation were evaluated and optimized. The characteristics of slag
prior to and after the treatment were studied. The following conclusions were drawn.
Solid State Oxidation
1. The RO phase of the slag is the only phase capable of being magnetized, where FeO is
transformed into magnetic spinel ferrites under air or pure CO2 atmosphere. The kinetics of the
transformation is faster in air atmosphere due to the higher oxygen potential.
2. Induced magnetic field strength has a significant effect on iron recovery rate. With increasing
magnetic field intensity, a larger magnetic fraction is obtained. For slag particles of d80=133
µm, the highest separation efficiency was reached at B=110 mT.
3. Increasing roasting temperature promotes magnetite formation, where at 1000 °C the
transformation is most complete.
4. Oxidizing kinetics was shown to be fast so that the reaction reached steady state in < 15 minutes.
Extending the reaction time beyond this time did not influence magnetic fraction or grade.
5. The grade of iron in the concentrate was essentially constant ~ 28% (compared to feed of 18%)
regardless of the oxidation conditions. This was attributed to the very fine particles of magnetic
phases and that with the particle size employed here, sufficient liberation could not be achieved.
6. Iron grade was found to be constant with increasing recovery rate under different conditions,
due to small size and embedding distribution of RO phase.
7. The optimum conditions for oxidation and separation of this slag (d80=133 µm, d20=11 µm)
were roasting in air atmosphere at 900 °C for 1 hour, and magnetic separating under 110 mT.
The grade of iron, and manganese in concentrate are 28% and 3.9%; the grade of iron and
101
manganese in tailing are 7% and 0.4%. 82% iron recovery and 91% manganese recovery were
achieved.
8. Grinding the roasted slag to d90=45 µm, improved the grade at the expense of lower recovery
(grades were 34% Fe, 5% Mn and recoveries were 59% Fe and 69% Mn).
Liquid Stage Oxidation
1. Magnetization takes place while oxidizing liquid slag. Increasing temperature and oxidizing
time show positive effect on separation efficiency. Decreasing basicity also promotes magnetic
phase formation.
2. Rapid cooling of slag limited the growth and segregation of magnetic iron phases, leading to
poor separation efficiency. This effect may have a negative impact on metal separation after air
atomization process.
3. The optimum conditions were oxidation at 1550 °C for 1 hour. Rapidly cooled slag (in water)
yielded a concentrate with Fe grade of 22% and recovery of 71%.
4. Slow cooling (5 °C/min) of a slag oxidized under the same conditions improved the grade to
25% Fe while the recovery dropped to 57%.
102
Chapter 6
Recommendations for Future Work
The study here pointed to several interesting observations which can lead future research towards
improvement of the oxidation-separation process.
First, it was found that magnetized slag grains are very small (tens of microns), therefore their
effective separation requires further grinding of the slag. A study on the effect of grinding after
roasting can determine what the optimum particle size should be, how the performance of the
magnetic separation step changes with particle size, and whether it will be economical to grind the
slag to that size.
Second, a fundamental study on possible growth of spinel phases both in solid and liquid states
(effect of time, temperature, and slag composition) will be useful to explore the possibility of
producing larger magnetic grains, hence less demand for grinding.
The iron grade in tailings of the current work were relatively low. The effective utilization of this
material in cement making should be investigated in consultation with cement producers.
Finally, slag oxidation can take place even at room temperature but slowly, this is how iron ores
(hematite and magnetite) were formed in the first place. Although this was not investigated in the
present work, weathering of granulated slag in ambient for extended periods of time is suggested.
This can be a low-cost approach in which granulated slag is first weathered for long periods of
time by storing it in shallow beds, followed by grinding and magnetic separation.
Furthermore, similar recovery methods can be explored on copper/nickel slag, ferroalloy slags,
as well as their compatibility with air atomization technique.
103
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114
Appendix A
Experimental conditions and data obtained.
Conditions Data (iron recovery)
Variables Fixed Feed (F)
in gram
Feed
Grade (f)
Concentrate
(C) in gram
Concentrate
Fraction
Concentrate
Grade (c)
Recovery Mineral
Grade (m)
SE
Raw Slag Field Strength
(mT) 110 N/A 11.8 18% 1.1 9% 28% 14% 45% 8%
Solid State Field Strength
(mT)
50 1 hr, 900˚C 8.3 18% 0.5 6% N/A N/A 45% N/A
55 8.5 18% 1.2 14% 28% 21% 45% 12%
60 8.3 18% 2.1 26% 27% 38% 45% 21%
70 8.4 18% 2.7 32% 27% 48% 45% 27%
80 9.2 18% 3.1 34% 28% 53% 45% 32%
100 8.8 18% 3.6 41% 30% 66% 45% 43%
110 8.1 18% 4.4 54% 28% 82% 45% 48%
120 8.4 18% 4.6 54% 27% 81% 45% 45%
130 8.2 18% 4.5 55% 27% 81% 45% 45%
Roasting Time
(hr)
0.25 110 mT, 900˚C 8.7 18% 4.6 53% 27% 79% 45% 44%
0.5 8.7 18% 4.6 53% 28% 79% 45% 45%
0.75 8.4 18% 4.5 53% 27% 78% 45% 42%
1 8.1 18% 4.4 54% 28% 82% 45% 48%
2 8.7 18% 4.4 51% 27% 75% 45% 42%
4 9.3 18% 4.8 52% 28% 78% 45% 44%
19 10.1 18% 5.3 52% 28% 78% 45% 45%
Temperature
(˚C)
600 110 mT, 1hr 9.8 18% 3.4 35% 29% 56% 45% 36%
700 8.9 18% 4.1 47% 28% 72% 45% 43%
115
800 9.0 18% 4.5 50% 28% 77% 45% 45%
900 8.1 18% 4.4 54% 28% 82% 45% 48%
1000 7.6 18% 4.2 55% 27% 82% 45% 44%
Size Reduction
(µm) <45
110 mR, 900˚C,
1hr 9.22 18% 2.95 32% 34% 59% 45% 46%
Variables Fixed Feed (F)
in gram
Feed
Grade (f)
Concentrate
(C) in gram
Concentrate
Fraction
Concentrate
Grade (c)
Recovery Mineral
Grade (m)
SE
Liquid
State
Temperature
(˚C)
1400 30 mins, 110
mT
7.6 18% 1.6 21% 24% 27% 45% 11%
1500 8.6 18% 2.8 33% 22% 39% 45% 10%
1550 8.9 18% 4.5 50% 22% 60% 45% 16%
Blow Time
(min)
15 1400 °C, 110
mT
8.8 18% 1.7 20% 23% 25% 45% 9%
30 7.6 18% 1.6 21% 24% 27% 45% 11%
45 9.0 18% 1.9 21% 25% 29% 45% 13%
15 1550 °C, 110
mT
10.1 18% 4.6 46% 21% 53% 45% 12%
30 8.9 18% 4.5 50% 22% 60% 45% 16%
60 10.0 18% 5.9 59% 22% 71% 45% 20%
S/C 2.91 1550 °C, 30
mins, 110 mT
8.9 18% 4.5 50% 22% 60% 45% 16%
2.5 8.7 18% 6.4 74% 19% 78% 45% 8%
2 10.4 16% 9.0 87% 17% 92% 45% 8%
1.5 10.6 15% 8.8 83% 16% 89% 45% 9%
Cooling Rate
(°C/min) 5
1550 °C, 1 hr,
110 mT 11.0 18% 4.56 42% 25% 57% 45% 25%
116
Appendix B
XRF results for solid state experiments:
Variables Magenitic Filed Strength Temperature
Temp ( C ) 900 900 900 900 900 900 900 900 1000 900 800 700 600
Time (hr) 1 1 1 1 1 1 1 1 1 1 1 1 1
MFT (mT) 55 60 70 80 100 110 120 130 110 110 110 110 110
comment M M M M M M M M M M M M M
Fe3O4 38.30 37.16 37.63 39.24 41.00 38.50 37.79 37.61 37.20 38.50 38.63 39.11 40.57
CaO 28.33 29.70 28.96 27.99 25.16 28.93 28.59 28.74 28.63 28.93 28.33 28.43 27.34
SiO2 9.95 10.58 9.97 9.73 7.72 9.22 9.43 9.97 9.52 9.22 9.81 9.93 9.86
Al2O3 3.71 3.62 3.84 3.47 3.38 3.25 3.29 3.41 3.09 3.25 3.39 3.43 3.46
MgO 11.65 11.17 11.72 11.55 14.49 11.94 12.90 12.53 13.90 11.94 11.95 11.47 11.02
MnO 4.61 4.32 4.50 4.54 4.89 5.05 5.03 4.96 4.89 5.05 5.04 4.93 5.08
P2O5 0.68 0.68 0.66 0.69 0.57 0.79 0.81 0.74 0.82 0.79 0.79 0.71 0.74
TiO2 0.86 0.88 0.83 0.80 0.78 0.91 0.84 0.76 0.73 0.91 0.82 0.78 0.74
Cr2O3 1.53 1.54 1.47 1.60 1.69 0.87 0.87 0.89 0.88 0.87 0.83 0.87 0.85
V2O5 0.18 0.19 0.22 0.21 0.17 0.18 0.14 0.15 0.14 0.18 0.17 0.16 0.16
SO3 0.1 0.1 0.1 0.1 0.1 0.2 0.2 0.1 0.1 0.2 0.1 0.1 0.1
Cl 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Ni 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Cu 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Sr 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Zr 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Nb 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
W 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
117
(Continued) Variables Duration
Temp ( C ) 900 900 900 900 900 900 900
Time (hr) 0.25 0.5 0.75 1 2 4 19.5
MFT (mT) 110 110 110 110 110 110 110
comment M M M M M M M
Fe3O4 37.93 38.09 37.18 38.50 37.58 38.02 38.10
CaO 29.13 28.70 29.02 28.93 28.42 28.59 29.01
SiO2 9.86 9.57 10.31 9.22 9.62 9.55 9.59
Al2O3 3.29 3.28 3.31 3.25 3.25 3.27 3.61
MgO 12.00 12.67 12.48 11.94 13.21 12.81 11.85
MnO 4.92 5.00 4.89 5.05 5.01 4.96 4.89
P2O5 0.80 0.81 0.80 0.79 0.79 0.78 0.87
TiO2 0.82 0.75 0.81 0.91 0.85 0.78 0.85
Cr2O3 0.87 0.79 0.86 0.87 0.91 0.85 0.81
V2O5 0.15 0.15 0.14 0.18 0.16 0.17 0.17
SO3 0.11 0.10 0.09 0.21 0.09 0.10 0.14
Cl 0.01 0.00 0.02 0.01 0.01 0.01 0.02
Ni 0.00 0.00 0.01 0.03 0.00 0.01 0.02
Cu 0.01 0.01 0.01 0.03 0.02 0.01 0.00
Sr 0.02 0.02 0.02 0.02 0.02 0.02 0.02
Zr 0.01 0.01 0.01 0.01 0.01 0.01 0.01
Nb 0.04 0.04 0.04 0.04 0.04 0.04 0.04
W 0.02 0.00 0.00 0.00 0.03 0.00 0.00
118
XRF results for liquid state experiments:
Variable Temperature Duration Basicity
C/s 2.9 2.9 2.9 2.9 2.9 2.9 2.9 2.9 2.9 2.5 2 1.5
Temp ( C ) 1400 1500 1550 1550 1550 1550 1400 1400 1550 1550 1550 1550
Time (hr) 30 30 30 15 30 60 30 45 30 30 30 30
MFT (mT) 110 110 110 110 110 110 110 110 110 110 110 110
comment M M M M M M M M M M M M
Fe3O4 33.08 30.06 30.16 29.32 30.16 30.35 33.08 34.64 30.16 30.16 28.12 26.45
CaO 33.97 32.95 36.47 36.41 36.47 37.18 33.97 32.62 36.47 35.69 33.06 31.22
SiO2 10.42 9.96 10.74 10.75 10.74 10.74 10.42 10.26 10.74 12.95 14.44 19.11
Al2O3 3.24 2.51 2.70 2.73 2.70 2.69 3.24 3.03 2.70 2.83 2.45 2.27
MgO 13.66 13.62 13.69 14.37 13.69 13.72 13.66 13.92 13.69 14.08 14.55 15.57
MnO 3.74 3.11 3.12 3.09 3.12 3.11 3.74 3.88 3.12 3.16 2.90 2.63
P2O5 0.89 0.72 1.08 1.04 1.08 1.02 0.89 0.85 1.08 0.98 0.90 0.78
TiO2 1.09 0.97 1.00 0.99 1.00 1.14 1.09 1.05 1.00 0.98 1.01 0.86
Cr2O3 0.77 0.57 0.60 0.55 0.60 0.60 0.77 0.51 0.60 0.56 0.53 0.50
V2O5 0.18 0.15 0.16 0.21 0.16 0.23 0.18 0.11 0.16 0.17 0.17 0.16
SO3 0.3 0.0 0.1 0.1 0.1 0.1 0.3 0.2 0.1 0.0 0.0 0.0
Cl 0.0 0.0 0.1 0.0 0.1 0.0 0.0 0.0 0.1 0.0 0.0 0.0
Ni 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Cu 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Sr 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Y 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.1
Nb 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
W 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.1 0.0 0.0
