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UNIVERSIDADE FEDERAL DE SANTA CATARINA
CURSO DE GRADUAÇÃO EM ENGENHARIA DE MATERIAIS
GUSTAVO RHUAN PEREIRA
EVALUATION OF THE INFLUENCE OF FUEL OIL AND COLLECTOR ON
COPPER AND MOLYBDENUM RECOVERY IN THE BIGHAM CANYON ORE
FLORIANÓPOLIS
2011
UNIVERSIDADE FEDERAL DE SANTA CATARINA
CURSO DE GRADUAÇÃO EM ENGENHARIA DE MATERIAIS
GUSTAVO RHUAN PEREIRA
EVALUATION OF THE INFLUENCE OF FUEL OIL AND COLLECTOR ON
COPPER AND MOLYBDENUM RECOVERY IN THE BINGHAM CANYON ORE
Diploma thesis presented to the Undergraduate
Course of Materials Engineering of the
Universidade Federal de Santa Catarina as part of
the requisite to obtain the degree of Materials
Engineer.
Mentor: Professor Dylton do Vale Pereira Filho
FLORIANÓPOLIS
2011
GUSTAVO RHUAN PEREIRA
EVALUATION OF THE INFLUENCE OF FUEL OIL AND COLLECTOR ON
COPPER AND MOLYBDENUM RECOVERY IN THE BIGHAM CANYON ORE.
This Diploma Thesis was assessed adequate to
attainment of the title “Engenheiro de Materiais”
and approved by the Undergraduate Course of
Materials Engineering of the Universidade
Federal de Santa Catarina.
Prof. Fernando Cabral, Ph.D.
Coordinator
Assessment Committee:
Prof. Dylton do Vale Pereira Filho (Mentor)
Renan Muller Schroeder
Luiz Fernando Vieira
PEREIRA, Gustavo Rhuan, 1986 -
EVALUATION OF THE INFLUENCE OF FUEL OIL AND COLLECTOR ON
COPPER AND MOLYBDENUM RECOVERY IN THE BIGHAM CANYON ORE.
/ Gustavo Rhuan Pereira. – 2011
40 f.: il. color. 30cm.
Supervisor: Prof. Dylton do Vale Pereira Filho
Trabalho de Conclusão de Curso – Universidade Federal de Santa Catarina, Curso de
Engenharia de Materiais, 2011.
1. Froth Flotation 2. Copper 3. Lab Kinetic Test. I. Pereira Filho, Dylton do Vale. II
Universidade Federal de Santa Catarina. Curso de Graduação de Engenharia de Materiais.
III. EVALUATION OF THE INFLUENCE OF FUEL OIL AND COLLECTOR ON
COPPER AND MOLYBDENUM RECOVERY IN THE BIGHAM CANYON ORE.
What I cannot create, I do not understand.
(Richard Feynman)
ACKNOWLEDGES
Firstly, I would like to thank Rio Tinto Kennecott Utah Copper for this great
opportunity and support. I am very grateful to my supervisors Geraldine Lyons and Michael
Dunn, for the internship opportunity, and to my co supervisor Dylan Cirulis, who guided me
along the internship. Their continuous supervision and suggestions were very important.
I would like to say thank you to Rio Tinto HR People, Ann Rash, Mike Rodgers and
Jesse Roberts who helped me with all documentation.
Thanks to my family, Elise Hinz, Toni Haag and my brothers of Harmonia Itajaiense,
the most important people in my life, for the support and to always believe in me.
Thanks Phanindra Kodali, Kambi Pezeshki and Kamran Pezeshki, Ana Paula Boatto
my best friends in my American life!
(…) tenho aprendido que a noite do dia é apenas a precursora do dia eterno.
(Frank Marshall)
AGRADECIMENTOS
Como muitas das pessoas que participaram da minha formação não compreendem
inglês resolvi agradecer estas em português.
Primeiramente agradecer ao Pai Celestial que com sua infinita bomdade, tem
iluminando meus caminhos, minha família e meus amigos.
Aos meus pais, José Urbino Pereira e Sandra Maria Severino Pereira pelo amor e
carinho demonstrados ao longo da minha vida.
A Universidade Federal de Santa Catarina e aos Professores do Curso de Engeharia de
Materiais por proporcionar um curso dinâmico e de qualidade, formando Engenheiros com
alta adaptação e experiência.
Aos meus irmãos e tios do Capítulo Harmonia Itajaiense da Ordem DeMolay por me
apresentarem as 7 virtudes e ao longo destes anos nunca esquecerem que nossas amizades são
eternas e não tem idade.
Aos meus amigos “cimjectianos” do Laboratório Cimject por sempre me receberem de
portas abertas. Especial agradecimento para o Professor Gean Salmoria e os amigos: Peixoto,
Lelo, Pri, Caubi, Calouro, Germanovix, Fernandinho, Ruben, Michel, Fala Mansa, Ju, Pagi,
Testoni, Pedro e Jaca.
Aos grandes metalurgistas que a cada estágio influenciaram grandemente minha
formação cada qual contribuindo com sua experiência e amizade. São eles: José Francisco da
Silva, Fabiano Miranda, Renê Lelis, Sérgio Scherer, José Armando Campos, Caetano Nunes
da Silva, Denílson Aquino, Ângelo Campos Moreira, Sandro Marino, Michael Dunn,
Geraldine Lyons, Dylan Cirulis, Phanindra Kodali, Kambi Pezeshki e Kamran Pezeshki.
Aos amigos que me incentivaram e tornaram a distância da família menos dolorosa,
em especial: Felipe Augusto de Souza, Ricardo Matsukura, Camila Sato, André Lozovey, Igor
Branco, Ricardo Selke, Guilherme Dalmedico, Eduardo Hulse, Alexandre Werner Reis,
Ricardo Reis, Charles Max, Alisson Rizzi e Ana Paula Boatto.
A Elise Hinz por sua companhia especial e paciência nas minhas ausências.
ABSTRACT
The consumption of copper has increased greatly in recent years due to its use
principally in the energy and electronics. Approximately 80% of world production of copper
is done by flotation followed by pyrometallurgy. In this study the ore QMPSW of Bingham
Canyon Mine who will be used in the plant in 2014-2016 was studied by flotation tests in the
laboratory to maximize recovery values when applied industrially. After 18 tests, were
analyzed the levels of recovery of copper, molybdenum and calculated the value of the
constant of reaction kinetics. Mean values of collector S8989 and fuel oil presented best
results in 5 minutes of flotation in the laboratory, which parameter is used as the standard for
industrial scale.
Key-words: Copper, Molybdenum, Copper extractive metallurgy, Flotation, Collector.
RESUMO
O consumo de cobre vem aumentando muito nos últimos anos devido a sua utilização
principalmente na área de energia e eletrônica. Aproximadamente 80% da produção mundial
de cobre é feita através de flotação seguida de pirometalurgia. Neste estudo o minério
QMPSW de Bingham Canyon Mine que será utilizado na planta em 2014-2016 foi estudado
através de ensaios de flotação em laboratório para maximizar seus valores de recuperação
quando for aplicado industrialmente. Após 18 testes, foram analizados os teores de
recuperação de cobre, molibdênio e calculado o valor da constante de cinética da reação.
Médios valores de coletor S8989 e óleo combustível apresentaram melhores resultados em 5
minutos de flotação em laboratório, parâmetro que será utilizado como padrão para escala
industrial.
Palavras-chaves: Cobre, Molibdênio, Metalurgia extrativa cobre, Flotação, Coletor.
LIST OF FIGURES
Figure 1: USA geological map. (www.virtualamericas.net/usa/maps) .................................... 13
Figure 2: Bingham Canyon Mine. (www.kennecott.com) ........................................................ 14
Figure 3: Cross section depicting a hypothetical volcanic edifice over the Bingham Canyon
porphyry copper deposit. The reconstructed volcanic cover implies that some Bingham
intrusions vented to form domes that collapsed to form the existing debris avalanche deposits
(Deino & Keith, 1997) ............................................................................................................. 15
Figure 4: QMPSW Deposit Position. ....................................................................................... 16
Figure 5: Typical hydrometallurgical process for recovery from heap leaching. ................... 17
Figure 6: Typical pyrometallurgical process for copper recovery from sulfide ore. ............... 17
Figure 7: Principle of froth flotation. (http://en.wikipedia.org/wiki/Froth_flotation) ............. 19
Figure 8: Contact angle between bubble and particle in an aqueous medium. ....................... 20
Figure 9: Action of the frother. ............................................................................................... 22
Figure 10: Critical pH value for chalcopyrite. [7] .................................................................. 22
Figure 11: Classification of collectors. (after Glembotskii, 1972) .......................................... 23
Figure 12: Collector adsorption on mineral surface. .............................................................. 24
Figure 13: (1) Laboratory Flotation Cell; (2) Float Test. ....................................................... 26
Figure 14: Rod Mill. ................................................................................................................. 27
Figure 15: Grinding time curve. ............................................................................................... 27
Figure 16: Representative DOE table showing the dosages used for each test. ...................... 28
Figure 17: Filing the float cell (left) and Reagent addition (right). ......................................... 28
Figure 18: Collecting the concentrate. ..................................................................................... 28
Figure 19: Flotation Test. ........................................................................................................ 29
Figure 20: ANOVA plots illustrating the main effects on copper recovery (left) and the
interaction of effects on copper recovery (right) after 2 minutes. ............................................ 30
Figure 21: ANOVA plots illustrating the main effects on molybdenum recovery (left) and its
interaction of effects (right) after 2 minutes. ............................................................................ 30
Figure 22: ANOVA plots illustrating the main effects on copper recovery (left) and the
interaction of effects on copper recovery (right) after 5 minutes. ............................................ 31
Figure 23: ANOVA plots illustrating the main effects on molybdenum recovery (left) and the
interaction of effects on molybdenum recovery (right) after 5 minutes ................................... 31
Figure 24: ANOVA plots illustrating the main effects on copper recovery (left) and the
interaction of effects on copper recovery (right) after 10 minutes. .......................................... 32
Figure 25: ANOVA plots illustrating the main effects on molybdenum recovery (left) and the
interaction of effects on molybdenum recovery (right) after 10 minutes ................................. 32
Figure 26: K values and maximum recovery calculated for each test. .................................... 33
Figure 27: Flotation recovery for copper graphic of the tests. ................................................ 33
Figure 28: Results of all tests. .................................................................................................. 36
CONTENTS
1. INTRODUCTION 11
2. OBJECTIVES 12
3. OVERALL REVIEW 13
3.1. GEOLOGY OF BINGHAM CANYON MINE 13
3.1.1. Tectonic Setting 14
3.1.2. QMPSW Deposit 15
3.2. COPPER PRODUCTION METHODS 16
3.2.1. Hydrometallurgy 16
3.2.2. Pyrometallurgy 17
3.3. FROTH FLOTATION 18
3.3.1. Principles of flotation 18
3.3.2. Frothers 21
3.3.3. The importance of pH 22
3.3.4. Collectors 23
3.3.5. Choice of collector 24
3.4. LABORATORY FLOTATION TESTING 25
4. EXPERIMENTAL PROCEDURES 27
4.1. GRINDING TIME CURVE 27
4.2. FLOTATION TESTS 28
5. RESULTS AND DISCUSSIONS 29
5.1. TWO MINUTES FLOTATION ANALYSIS 29
5.2. FIVE MINUTES FLOTATION ANALYSIS 30
5.3. TEN MINUTES FLOTATION ANALYSIS 32
5.4. KINETICS ANALYSIS 33
6. CONCLUSIONS 34
7. REFERENCES 35
8. APPENDIX A 36
11
1. INTRODUCTION
Copper is the third most used metal by man (iron and aluminum are produced and
consumed in greater quantities than copper). The critical need for copper began in 1850, with
the use of electricity. Given its malleability, conductivity of both heat and electricity, ability
to withstand corrosion, and its esthetic characteristics, copper has established crucial
importance in virtually all areas of development and newly developing economies most
notably in areas of construction, transport, and all kinds of electrical and electronic
applications. Since 1850 copper production has increased more than 300 times [1].
After two centuries of great use of natural resources, the economy of the twenty-first
century looks at sustainability as something essential for the future of the earth. Every day,
society increasingly demands that industries adopt a sustainable approach with a minimal use
of nonrenewable resources. To meet the demand of society, the mining sector increasingly
invests in innovative technologies that aim to make the most of the original ore products with
the least possible pollution.
The modern froth flotation process was independently invented in the early 1900s in
Australia by C. V. Potter and around the same time by G. D. Delprat. Initially, naturally
occurring chemicals such as fatty acids and oils were used as flotation reagents in a large
quantity to increase the hydrophobicity of the valuable minerals. Since then, this process has
been adapted and applied to a wide variety of materials to be separated, and additional
collector agents, including surfactants and synthetic compounds have been adopted for
various applications [2]. Today approximately 80% of world copper production uses forth
flotation [3].
In this background, the Ore Body Knowledge Team Kennecott Utah Mine in the state
of Utah, USA, is developing a project to improve the copper recovery in froth flotation. The
present academic work is part of this project and consists of testing one kind of ore that will
be used in 2014-2016.
12
2. OBJECTIVES
This work has as main objective to maximize the copper and molybdenum recovery
for QMPSW ore in froth flotation.
The specific objectives are:
Verify the influence of collector on recovery rates of copper and molybdenum.
Verify the influence of fuel oil on recovery rates of copper and molybdenum.
Calculate the Kinetic constant value for copper and molybdenum.
13
3. OVERALL REVIEW
This chapter contains an overview about the subjects which assists the understanding
of this work. Here it finds an introduction about the geology of Bingham Canyon Mine,
copper extraction process, reagents used in flotation, and froth flotation tests in laboratory.
3.1. GEOLOGY OF BINGHAM CANYON MINE
Bingham Canyon hosts one of the world’s largest porphyry copper deposits and is in
fact the world’s largest open pit mining excavation. The Figure 1 shows the position of
Bingham Canyon in USA geological map, and the Figure 2 shows the Bingham open pit. It is
primarily mined for copper but also produces substantial amounts of molybdenum, gold, and
silver which also contribute to the profitability of the mine. The Bingham pit is currently more
than 3 kilometers in diameter and more than 900 meters deep.
Figure 1: USA geological map. (www.virtualamericas.net/usa/maps)
Bingham Canyon has been explored since 1904 and your current owner is the Rio
Tinto Group. The Bingham stock porphyry is a classic giant porphyry copper ore body, with
low grade copper disseminated throughout the igneous host rock. Open pit reserves are
estimated to last until 2020. There are currently plans being made to mine underground
reserves by the block caving mining method estimated to keep the mine running for
approximately 15 years after that.
14
Figure 2: Bingham Canyon Mine. (www.kennecott.com)
3.1.1. Tectonic Setting
The Bingham Canyon mine is located in the central Oquirrh Mountains, and
approximately 50 kilometers long north-south oriented, fault-bounded mountain rage in the
eastern Great Basin. The range begins in northwest Utah County and stops at the shore of the
Great Salt Lake, separating Utah’s Salt Lake Valley from the Tooele Valley. The Oquirrh
Mountains (“Oquirrh” is a Ute Indian word meaning “The Shining Mountains”) represent the
eastern most fault block of the Basin and Range Province separated from the Colorado Plateau
by the Wasatch Fault [4].
The Oquirrh Mountains region has been tectonically active through much of geologic
time. Since early Paleozoic times it was until the late Carboniferous a passive continental
margin, where thin carbonates interbedded with clastic sediments were deposited. During late
Pennsylvanian time a slight deepening was accompanied by the rapid deposition of several
kilometers of shallow water carbonates and siliclastic sediments into the northwest-trending
Oquirrh Basin. In the Early Permian time followed the return to the passive margin
sedimentation [4].
Two orogenies with Mesozoic age affected the region: the mid Jurassic Elko Orogeny
and the late Cretaceous Sevier Orogeny. They formed in south of Bingham Canyon a series of
northwest-trending folds, and northeast-trending folds north of Bingham Canyon. Eastward
oriented compressional forces during the time of the Sevier Orogeny resulted in several major
thrusts and complex faulting and folding in the vicinity of the Cu-Mo-Au ore deposit [5].
Cenozoic activity at Bingham Canyon was dominated by an extensional regime. It
started with a minor Eocene extension associated with the emplacement of intrusions, dikes
and fissures and evolved into an intense period of Eocene intrusive and volcanic activity along
15
the Uinta Axis forming the west-trending Wasatch igneous belt with the Bingham Canyon
porphyry copper deposit [5].
The intrusion of igneous rocks in the Bingham Canyon mining district was controlled
by northwest-trending and northeast-striking faults. Some of the Eocene magmatic stocks and
dikes apparently vented to the surface and formed a composite volcano (Figure 3). Some parts
of this volcano are still preserved on the eastern side of the Oquirrh Mountains.
Figure 3: Cross section depicting a hypothetical volcanic edifice over the Bingham Canyon porphyry copper
deposit. The reconstructed volcanic cover implies that some Bingham intrusions vented to form domes that
collapsed to form the existing debris avalanche deposits [5].
In the Middle Eocene the Bingham and Last Chance Stock were emplaced, but only
the Bingham Stock is strongly mineralized. This appears to be due to the vicinity to the later
Quartz Monzonite Porphyry. This intrusive body acted as the source and center for the
hydrothermal activity.
Extension continued with major Miocene to Recent Basin and range extension which
opened the Salt Lake Valley and other north-south oriented basis along major listric faults.
Thereby the Bingham Canyon deposit was tilted eastwards by 15 to 20 degrees [5].
3.1.2. QMPSW Deposit
The QMPSW Deposit is one area who will be used in 2014-2016, this deposit has high
concentration of copper and molybdenum. The Figure 4 shows the position of this deposit on
Bingham Canyon Mine.
16
Figure 4: QMPSW Deposit Position.
3.2. COPPER PRODUCTION METHODS
Copper reserves are mostly present in the form of oxide and sulfide minerals. Copper
mining is performed either underground or in open pits. Laterite ores copper oxide minerals
such as cuprite and hydrous carbonates (malachite, azurite). Chalcopyrite, chalcocite, bornite,
cubanite, and enargite are common examples of copper sulfide minerals. Most of the copper
ores contain only a very small percentage of copper minerals, an even less molybdenum and
precious metals [6].
The remaining minerals, of little value, are discarded. Average head grade in most of
the mines is less than 1% copper. Depending on the ore type (oxide or sulfide) the copper
extraction process is designed. Hydrometallurgical processes, such as heap leaching is used to
extract copper from copper oxide ore and some copper sulfide ores. Subsequently copper is
extracted from the leach solution by solvent extraction and electrowinning. In the case of the
Pyrometallurgical process, a copper sulfide concentrate, produced by froth flotation is smelted
at high temperature and refined electrolytically [6].
3.2.1. Hydrometallurgy
Hydrometallurgical processes are used to extract copper from low grade ores,
especially copper oxide by heap leaching (Figure 5). Copper ore from the mine is crushed
typically with jaw crushers to pass about 1,27 cm top size. This crushed ore along with acid
solution is introduced into rotating agglomeration drums. In the agglomeration drums fine ore
particles are bonded to coarser ore particles via liquid bridges. The agglomeration product is
stacked on the heap leach pad to about 7 meters in height.
Sulfuric acid solution is introduced on the top of the heap leach pad and dissolves
copper as the solution as the solution passes through the heap. Copper recovery from the heap
17
leach pad depends on the particle size distribution, of the ore particle damage, and quality of
agglomerates. About 20% of world’s annual copper production is from leaching. Bioleaching
and autoclave leaching are also performed depending on ore type and grade in order to
improve the leaching efficiency.
Figure 5: Typical hydrometallurgical process for recovery from heap leaching.
Pregnant leach solution (copper rich solution) from the leach pads is concentrated and
purified by solvent extraction. During this solvent extraction stage, copper is separated from
the acid solution using extractant to stabilize copper in organic phase. After stripping copper
from the organic phase, copper metal is produced as cathodes during electrowinning.
3.2.2. Pyrometallurgy
About 80% of world’s annual copper production is from the pyrometallurgy of copper
sulfide ore (Figure 6). Copper ore from the mine is crushed with a gyratory or jaw crusher.
Discharge from crusher feeds a grinding circuit where, sag mills and ball mills further reduce
the ore particles to about 75 microns in size. This slurry of fine ore particles is conditioned
with chemicals to separate the copper sulfide mineral particles by flotation. In about 80 to
90% of the copper is recovered during flotation.
Figure 6: Typical pyrometallurgical process for copper recovery from sulfide ore.
The copper concentrate from flotation is sent to filtration to remove the water and to
dry the concentrate. Dry concentrate is introduced into the smelting furnaces. Smelting
furnace produces matte (high grade Cu/Fe sulfides). The matte is sent to the convertor where
blister copper is produced. The blister copper is cast into anodes and refined electrolycally as
final product (copper 99,99%).
In the present study, the influence of burner-oil and collector on copper and
molybdenum recovery has been studied.
18
3.3. FROTH FLOTATION
Flotation is undoubtedly the most important and versatile mineral processing
technique, and both its use and application are continually being expanded to treat greater
tonnages and to cover new areas [7].
Originally patented in 1906, flotation has permitted the mining of low grade and
complex ore bodies which would have otherwise been regarded as uneconomic. In earlier
practice the tailings of many gravity plants were of a higher grade than the ore treated in many
modern flotation plants [7].
Flotation is a selective process and can be used to achieve specific separations from
complex ores such as lead-zinc, copper-zinc, lead, copper-molybdenum, and zinc. The field of
flotation has now expanded to include platinum, nickel, gold-hosting sulphides, and oxides,
such as hematite and cassiterite, oxidized minerals, such as malachite and cerussite, and non-
metallic ores, such as fluorite, phosphates, and fine coal [6].
3.3.1. Principles of flotation
Flotation is a physico-chemical separation process that utilizes the difference in
surface properties of the valuable minerals and the unwanted gangue minerals. The theory of
froth flotation is complex, involving three phases (solids, water, and froth) with many
subprocesses and interactions, and is not completely understood [6].
The process of material being recovered by flotation from the pulp comprises three
mechanisms:
(1) Selective attachment to air bubbles (or ‟true flotation”);
(2) Entrainment in the water which passes through the froth.
(3) Physical entrapment between particles in the froth attached to air bubbles (often
referred to as ‟aggregation”.
The attachment of valuable minerals to air bubbles is the most important mechanism
and represents the majority of particles that are recovered to the concentrate. Although true
flotation is the dominant mechanism for the recovery of valuable mineral, the separation
efficiency between the valuable mineral and gangue is also dependent on the degree of
entrainment and physical entrapment. Unlike true flotation, which is chemically selective to
the mineral surface properties, both gangue and valuable minerals alike can be recovered by
entrainment and entrapment. Drainage of these minerals occurs in the froth phase and
19
controlling the stability of this phase is important to achieve an adequate separation. In
industrial flotation plant practice, entrainment of unwanted gangue can be common and hence
a single flotation stage is uncommon. Often several stages of flotation (called circuits) are
required to reach an economically acceptable quality of valuable mineral in the final product.
True flotation utilizes the differences in physico-chemical surface properties of
particles of various minerals. After treatment with reagents, such differences in surface
properties between the minerals within the flotation pulp become apparent and, for flotation to
take place, an air bubble must be able to attach itself to a particle, and lift it to the water
surface. The Figure 7 illustrates the principles of flotation in a mechanical flotation cell. The
agitator provides enough turbulence in the pulp phase to promote collision of particles and
bubbles which results in the attachment of valuable particles to bubbles and their transport
into the froth phase for recovery [6].
Figure 7: Principle of froth flotation. (http://en.wikipedia.org/wiki/Froth_flotation)
The process can only be applied to relatively fine particles, because if they are too
large the adhesion between the particle and the bubble will be less than the particle weight and
the bubble will therefore drop its load [6].
In flotation concentration, the mineral is usually transferred to the froth, or float
fraction, leaving the gangue in the pulp or tailing. This is direct flotation and the opposite is
reverse flotation, which the gangue is separated into the float fraction [6].
The function of the froth phase is to enhance the overall selectivity of the flotation
process. The froth achieves this by reducing the recovery of entrained material to the
concentrate stream, while preferentially retaining the attached material. This increases the
concentrate grade whilst limiting as far as possible the reduction in recovery of valuables. The
relationship between recovery and grade is a trade-off that needs to be managed according to
operational constraints and is incorporated in the management of an optimum froth stability.
20
As the final separation phase in a flotation cell, the froth phase is a crucial determinant of the
grade and recovery of the flotation process [6].
The mineral particles can only attach to the air bubbles if they are to some extent
water-repellent, or hydrophobic. Having reached the surface, the air bubbles can only continue
to support the mineral particles if they can form a stable froth, otherwise they will burst and
drop the mineral particles. To achieve these conditions it is necessary to use the numerous
chemical compounds known as flotation reagents [6].
The activity of a mineral surface in relation to flotation reagents in water depends on
the forces which operate on that surface. The forces tending to separate a particle and a bubble
are shown in Figure 8. The tensile forces lead to the development of an angle between the
mineral surface and the bubble surface. At equilibrium,
𝛾𝑠/𝑎 = 𝛾𝑠/𝑤 + 𝛾𝑤/𝑎 cos 𝜃
where 𝛾𝑠/𝑎 , 𝛾𝑠/𝑤 and 𝛾𝑤/𝑎 are the surface energies between solid air, solid and water, and
water and air, respectively, and 𝜃 is the contact angle between the mineral surface and the
bubble.
Figure 8: Contact angle between bubble and particle in an aqueous medium.
The force required to break the particle-bubble interface is called the work of
adhesion, 𝑤𝑠/𝑎 , and is equal to the work required to separate the solid-air interface and
produce separate air-water and solid-water interfaces.
𝑤𝑠/𝑎 = 𝛾𝑤/𝑎 + 𝛾𝑠/𝑤 − 𝛾𝑠/𝑎
Combining with Equation 3.1 gives
𝑤𝑠/𝑎 = 𝛾𝑤/𝑎 (1 − cos 𝜃)
(3.1)
(3.2)
(3.3)
21
It can be seen that the greater the contact angle the greater is the work of adhesion
between particle and bubble and the more resilient the system is to disruptive forces. The
hydrophobicity of a mineral therefore increases with the contact angle; minerals with a high
contact angle are said to be aerophilic, they have a higher affinity for air than for water. The
terms hydrophobicity and floatability are often used interchangeably. Hydrophobicity,
however, refers to a thermodynamic characteristic, whereas floatability is a kinetic
characteristic and incorporates other particle properties affecting amenability to flotation [6].
Most minerals are not water-repellent in their natural state and flotation reagents must
be added to the pulp. The most important reagents are the collectors, which adsorb on mineral
surfaces, rendering them hydrophobic (or aerophilic) and facilitating bubble attachment. The
frothers help maintain a reasonably stable froth [6].
3.3.2. Frothers
When mineral surfaces have been rendered hydrophobic by the use of a collector,
stability of bubble attachment, especially at the pulp surface, depends to a considerable extent
on the efficiency of the frother [8].
Ideally the frother acts entirely in the liquid phase and does not influence the
state of the mineral surface. In practice, however, interaction does occur between the frother,
mineral and other reagents, and the selection of a suitable frother for a given ore can only be
made after extensive test work.
In sulphides mineral flotation it is common practice to employ at least two
frothers and more than one collector. Specific frothers are chosen to provide adequate
physical properties to the froth, while the second frother interacts with the collectors to control
the dynamics of the flotation process. Froth build-up on the surfaces of thickeners, and
excessive frothing of flotation cells, are problems occurring in many mineral processing
plants. A good frother should have negligible collecting power, and also produce a froth
which is just stable enough to facilitate transfer of floated mineral from the cell surface to the
collecting launder [8].
Frothers are generally heteropolar surface-active organic reagents, capable of
being adsorbed on the air-water interface. When surface-active molecules react with water,
the water dipoles combine readily with the polar groups and hydrate them, but there is
practically no reaction with the non-polar hydrocarbon group, the tendency being to force the
latter into the air phase [8].
22
Figure 9: Action of the frother.
3.3.3. The importance of pH
One of most important parameters for best results in flotation is the balance between
reagent concentration and pH.
Flotation where possible is carried out in an alkaline medium, because the most
collectors, are stable under these conditions, and corrosion of cells, is minimized. Alkalinity is
controlled by the addition of lime or sodium carbonate.
These chemicals are often used in very significant amounts in almost all flotation
operations. Although they are cheaper than collectors and frothers, the overall cost is
generally higher with pH regulators per ton of ore treated than with any other processing
chemical [6].
Lime, being cheap, is very widely used to regulate pulp alkalinity, and is used in the
form of milk of lime (suspension of calcium hydroxide particles in a saturated aqueous
solution). Lime, is often added to the slurry prior to flotation to precipitate heavy metal ions
from solution in this sense, the alkali is acting as a “deactivator” [6].
The critical pH value depends on the nature of the mineral, the particular collector and
its concentration, and the temperature. The Figure 10 shows how the critical pH value for
pyrite depends on the concentration of sodium dithiophosphate.
Figure 10: Critical pH value for chalcopyrite. [7]
23
3.3.4. Collectors
Hydrophobicity has to be imparted to most minerals in order to float them. In order to
achieve this, surfactants known as collectors are added to the pulp and time is allowed for
adsorption during agitation in what is known as the conditioning period. Collectors are
organic compounds which render selected minerals water-repellent by adsorption of
molecules or ions on to the mineral surface, reducing the stability of the hydrated layer
separating the mineral surface from the air bubble to such a level that attachment of the
particle to the bubble can be made on contact [9].
Collector molecules may be ionizing compounds, which dissociate into ions in water, or
non-ionizing compounds, which are practically insoluble, and render the mineral water-
repellent by covering its surface with a thin film [6].
Ionizing collectors have found very wide application in flotation. They have complex
molecules which are asymmetric in structure and are heteropolar, the molecule contains a
non-polar hydrocarbon group and a polar group which may be one of a number of types. The
non-polar hydrocarbon radical has pronounced water-repellent properties, whereas the polar
group reacts with water [6].
Figure 11: Classification of collectors. (after Glembotskii, 1972)
Ionizing collectors are classed in accordance with the type of ion, anion or cation that
produces the water-repellent effect in water. This classification is given in Figure 11.
Because of chemical, electrical, or physical attraction between the polar portions and
surface sites, collectors adsorb on the particles with their non-polar ends orientated towards
the bulk solution, thereby imparting hydrophobicity to the particles (Figure 12). They are
usually used in small amounts, substantially those necessary to form a monomolecular layer
on particle surfaces (starvation level), as increased concentration, apart from the cost, tends to
24
float other minerals, reducing selectivity. It is always harder to eliminate a collector already
adsorbed than to prevent its adsorption.
Figure 12: Collector adsorption on mineral surface.
An excessive concentration of collector can also have adverse effect on the recovery of
valuable minerals, possibly due to the development of collector multi-layers on the particles,
reducing the proportion of hydrocarbon radicals orientated into the bulk solution. The
hydrophobicity of the particles is thus reduced, and hence their floatability. The flotation limit
may be extended without loss of selectivity by using a collector with a longer hydrocarbon
chain, thus producing greater water-repulsion, rather than by increasing the concentration of a
shorter chain collector. However, chain length is usually limited to two to five carbon atoms,
since the solubility of the collector in water rapidly diminishes with increasing chain length
and, although there is a corresponding decrease in solubility of the collector products, which
therefore adsorb very readily on the mineral surfaces, it is, of course, necessary for the
collector to ionize in water for chemisorption to take place on the mineral surfaces. Not only
the chain length but also the chain structure, affects solubility and adsorption, branched chains
have higher solubility than straight chains [8].
It is common to add more than one collector to a flotation system. A selective collector
may be used at the head of the circuit, to float the highly hydrophobic minerals, after which a
more powerful, but less selective one, is added to promote recovery of the slower floating
minerals [8].
3.3.5. Choice of collector
In most porphyry copper molybdenum operations, xanathe or dithiophosphates are used
as the primary collector, while a variety of secondary collectors are used including xanthogen
formats, thionocarbamates, xanthic esters and mercaptobenzothiazole. In some cases, only
dithiophosphate collectors are used. There is no general rule by which secondary collectors
25
are selected. There are, however, several factors that influence the selection of this collector
such as presence of clay minerals in the ore, presence of middlings and type of frother used.
Approximately 50% of world’s molybdenum production comes from copper-
molybdenum ore as a by-product. The floatability of molybdenum during copper flotation
also depends on many factors, including type of collector, type of frother, flotation pH and
type of hydrocarbon oil used. During the flotation of copper molybdenum ores, fuel oil or
kerosene is added to the grinding to enhance molybdenum recovery [8].
In this study, Cytec S8989 and fuel oil were used as collector.
3.4. LABORATORY FLOTATION TESTING
One of most widely used techniques to determine the amount and what reagents
(collectors and frothers) will be used on industrial scale is the laboratory flotation test.
Flotation testing is also carried out on ores in existing plants to improve procedures and for
development of new reagents [9].
It is essential that test is carried out on ore which is representative of that treated in the
commercial plant. Samples for test work must be representative, not only in chemical
composition, but also degree of dissemination. A mineralogical examination of drill cores or
other individual samples should therefore be made before a representative sample is selected.
Composite drill core samples are ideal for testing if drilling in the deposit has been extensive;
the cores generally contain ore from points widely distributed over the area and in depth. It
must be realized ore bodies are variable and that a representative sample will not apply
equally well to all parts of the ore body. It is used therefore for development of general
flotation procedure. Additional tests must be made on samples from various areas and depths
to establish optimum conditions in each case and to give design data over the whole range of
ore variation [9].
Having selected representative samples of the ore, it is necessary to prepare them for
flotation testing, which involves comminution of the ore to its optimum particle size.
Crushing must be carried out with care in order to avoid accidental contamination of the
sample by grease or oil, or with other materials which have been previously crushed. Even in
commercial plant, a small amount of grease or oil can temporarily upset the flotation circuit
[9].
Storage of the crushed sample is important, since oxidation of the surfaces is to be
avoided especially with sulphides ores. Not only does oxidation inhibit collector absorption,
26
but it also facilitates the dissolution of heavy metal ions, which may interfere with the
flotation process. Sulphides should be tested as soon as possible after obtaining the sample
and ore samples must be shipped in sealed drums in as coarse a state as possible.
Wet grinding of the samples should always be undertaken immediately prior flotation
testing to avoid oxidation of the liberated mineral surfaces. Batch laboratory grinding, using
ball mills, produces a flotation feed with a wider size distribution than that obtained in
continuous closed-circuit grinding; to minimize this, batch rod mills are used which give
products having a size distribution which approximates closely to that obtained in closed-
circuit ball mills. True simulations is never really achieved, however, as overgrinding of high
density minerals , which is a feature of closed-circuit grinding, is avoided in a batch rod mill.
The optimum grinding size of the particles depends not only on their grain size, but
also on their floatability. Initial examination of the ore should be made to determine the
degree of liberation in terms of particle size in order to estimate the grinding time. Tests
should then be carried out over a range of grinding sizes in conjunction with flotation tests in
order to determine the optimum mesh of grind. In certain cases, it may be necessary to over
grind the ore in order that particles are small enough to be lifted by air-bubbles [9].
The bulk of laboratory test is carried out in batch flotation cells, usually with the same
density of the production (26-38% of solids). The cells are mechanically agitated, the speed of
rotation of the impellers being variable, and simulate the large-scale models commercially
available. Introduction of air to the cell is normally via a hollow standpipe surrounding the
impeller shaft.
Figure 13: (1) Laboratory Flotation Cell; (2) Float Test.
27
4. EXPERIMENTAL PROCEDURES
These experiments could be divided in two parts: grinding and flotation. In the first
moment, the blast ore was received and characterized by size and grinded up to 32% of
particles were greater than 100 mesh. After grinded, the ore was floated in laboratory to
analyze the best combination of fuel oil and collector who gives the best recovery.
4.1. GRINDING TIME CURVE
Approximately 1500g of ore was used in each test. This ore was received in the
laboratory with a particle size higher than considered optimal to do a flotation test.
Two tests were done in the laboratory rod mill to take the optimal particle size. In one
test the grinding time was 5 minutes and in the other test grinding time was 12 minutes. After
grinded, each sample was dried and screened to take the percent of particles higher than 100
mesh.
Figure 14: Rod Mill.
With these two points, it was possible to take the grinding time, 7 minutes and 12
seconds, corresponding to 32% of +100 mesh particles.
Figure 15: Grinding time curve.
28
4.2. FLOTATION TESTS
After determining the grinding time, 18 tests were done. The configuration of each test is
showed in the Figure 16. In each test the ore sample was prepared in the rod mill with 1.5g of
lime (for pH control), and 22 µl of frother Dow Chemical’s X-237.
Figure 16: Representative DOE table showing the dosages used for each test.
Flotation tests were performed after each grinding cycle. In these tests, samples were
collected at 0.5, 2, 5, and 10 minutes and sent for assay analysis. The results will be discussed
below.
Figure 17: Filing the float cell (left) and Reagent addition (right).
The parameters of Denver Flotation Machine were: 4 liters of solution, 1500 rpm and 4
m³/h of air flow.
Figure 18: Collecting the concentrate.
29
5. RESULTS AND DISCUSSIONS
The recovery of copper and molybdenum was calculated for every test after 30
seconds, 2, 5 and 10 minutes respectively. Appendix A shows the calculated recovery tables.
Figure 19: Flotation Test.
To determine the effect of the reagents on the overall recovery of the minerals,
ANOVA was used to identify what reagent dosages would give the best results, as well as
what combinations could give the best recovery. Although the analysis was done for every
test after 2, 5 and 10 minutes, the main focus was on the 5 minutes recovery since it is
representative of the plant process and can be scaled up to fit the plant requirements. The
analysis at different times was done to determine how the flotation kinetics in the system
worked and to determine if there was a relationship between them. Also the data could be
used to decide whether a longer residence time could be applied to the process or not.
5.1. TWO MINUTES FLOTATION ANALYSIS
The ANOVA showed that after 2 minutes there was a better recovery with a
combination of 20 µl of S8989 collector and 33 µl of fuel oil dosages respectively (Figure 20).
This shows that after 2 minutes increasing the amount of fuel oil provides a better copper
recovery but it is not the same when the S8989 dosage increases. However, if analyzed
separately it is shown that only a medium level of S8989 and burner oil improves copper
recovery. Thus, ANOVA shows that the two variables are not linearly related and only a
combination of high burner oil dosage and an intermediate level of S8989 would provide good
results.
For the molybdenum recovery (Figure 21), the trend was not the same as in the copper
recovery interaction plot. Molybdenum had a better recovery with a low burner oil dosage
and a low S8989 (13µl and 14µl respectively). However, the difference between the recovery
using these dosages and using the dosage that gave the best copper recovery is only 1%. After
30
the 2 minutes it can be assumed that if at a combination of 20µl S8989 and 33µl fuel oil
dosages respectively, the recovery of copper and molybdenum is significantly higher. The
two variables, S8989 and fuel oil, are not linearly related and only a combination of high fuel
oil dosage and an intermediate level of S8989 would provide good results. At this point
neither the S8989 nor the burner oil plays a significant effect on the recovery of valuable
mineral.
Figure 20: ANOVA plots illustrating the main effects on copper recovery (left) and the interaction of effects on
copper recovery (right) after 2 minutes.
Figure 21: ANOVA plots illustrating the main effects on molybdenum recovery (left) and its interaction of
effects (right) after 2 minutes.
5.2. FIVE MINUTES FLOTATION ANALYSIS
After 5 minutes the recovery patterns stayed almost the same as the patterns observed
after the 2 minutes time frame. If analyzed individually, only a medium level of S8989 and
fuel oil helped improve copper recovery. If the interaction plot is analyzed, there was a better
recovery with a combination of 14µl of S8989 and 23µl fuel oil dosages respectively (Figure
22). Thus, the dosage of reagents decreased but the recovery was increased by approximately
7 % (also attributed to the residence time). Moreover, when performing a t-test, to identify the
variable that contributes more significantly to the result, the test indicated that S8989 is the
31
most significant variable and has a higher effect on the copper recovery. The test only
accounts for marginal or partial contribution but it is a good indicator of which variable would
have the most significant effect on overall recovery. The test reveals important information
since the 5-minute refloat data is representative of the plant process and is usually used for
scale-up. Nonetheless, more experiments should be carried out to support these statements.
Figure 22: ANOVA plots illustrating the main effects on copper recovery (left) and the interaction of effects on
copper recovery (right) after 5 minutes.
In the case of MoS2, when the reagents were analyzed individually, it was observed
that increasing the level of fuel oil also increased the recovery of Molybdenum. The opposite
occurred with S8989, since the lowest the level of S8989 conferred the highest overall
recovery. The t-test showed that the most significant effect on recovery was attributed to the
level of S8989. Thus, changing the dosage of S8989 will have a detrimental effect on the
recovery of Molybdenum. However, if the interaction plots are analyzed, the combination of
14µl of S8989 and 13µl of fuel oil provides the best recovery.
The second best combination (20µl of S8989 and 33µ of fuel oil) is with a medium
level of S8989 and a high dosage of fuel oil. The trade-off is only 1% recovery decrease, thus
showing that increasing fuel oil dosage and decreasing S8989 is detrimental for a good
Molybdenum recovery. If this trend is followed for the copper recovery, the trade-off will be
only 0.5% decrease in Cu recovery.
Figure 23: ANOVA plots illustrating the main effects on molybdenum recovery (left) and the interaction of
effects on molybdenum recovery (right) after 5 minutes
32
5.3. TEN MINUTES FLOTATION ANALYSIS
The analysis of the interaction plots after 10 minutes showed that increasing the levels
of S8989 and fuel oil was beneficial for the recovery of all minerals. However, when analyzed
individually, it was observed that the reagent dosage followed the same trend as the 5 minutes
analysis. The recoveries observed with the increasing levels of S8989 show that S8989 plays a
significant role in the recovery of the valuable minerals. Although it is not linearly related to
fuel oil, S8989 is statistically significant for the recovery of the minerals and changes on the
dosage may significantly affect the process.
Figure 24: ANOVA plots illustrating the main effects on copper recovery (left) and the interaction of effects on
copper recovery (right) after 10 minutes.
If the combination of 20µl of S8989 and 33µl of burner oil is analyzed, the recovery of
copper is not the best but the difference between the highest recovery and the recovery under
these conditions is only 1.5%. The recovery is very good recovery and the losses are minimal.
The same occurs with the recovery of molybdenum. Overall, a good combination would be
the use of 20µl of S8989 and 33µl of fuel oil in the process.
Figure 25: ANOVA plots illustrating the main effects on molybdenum recovery (left) and the interaction of
effects on molybdenum recovery (right) after 10 minutes
33
5.4. KINETICS ANALYSIS
Flotation kinetics analysis was performed using the command solve of Microsoft
Excel for the mathematical expression: Rx = Rmax * (1-e-kt
), where Rx is the cumulative
recovery in the selected time, Rmax is the maximum recovery of the reaction, k is the kinetic
constant, and t is the selected time. To calculate the k value of each test, 2 rules for the
mathematical expression were created: (1) Rmax < 1; (2) Rmax >R10min.
The Figure 26 shows the k value and the maximum calculated recovery for each test of
this experiment.
Figure 26: K values and maximum recovery calculated for each test.
Faster kinetics was usually a good indicator of good recovery performance and vice
versa. The best recovery was found with high S8989 and high fuel oil, but for this good
recovery a long residence time is necessary what cannot be industrially viable.
The Figure 27 shows the kinetic curve for copper of the experiment.
Figure 27: Flotation recovery for copper graphic of the tests.
34
6. CONCLUSIONS
Evaluating the results obtained in this work, the following conclusions can be
affirmed:
Collector is the most important component for recovery, but big amounts of
collector not mean one greater recovery rate within 5 minutes flotation, 20 µl is
the ideal amount.
Fuel oil as previously investigated in literature, has an important role to
molybdenum recovery, large quantities of fuel oil with small amounts of
collector results in better recovery rates for molybdenum.
Kinetics constant values showed that small amounts of reagents resulting in a
faster kinetics and higher recovery values, however to achieve these high
recovery levels a long residence time is necessary, which makes it
uneconomical.
So the main objective of this work, which was maximize the copper and molybdenum
recovery of QMPSW ore is possible with 20 µl of collector and 23 µl of fuel oil, however
more tests with different percent of solids, different pH, and different frother is necessary
before use of this ore in the plant.
35
7. REFERENCES
[1] Kodali, P. .: Pretreatament of copper prior to heap leaching , Master thesis(2010);
[2] http:// en.wikipedia.org/wiki/Froth_flotation, accessed in Feb/06/2011;
[3] http://en.wikipedia.org/wiki/Copper_extraction_techniques, accessed in Feb/06/2011;
[4] Martinek, K. .: High precision U-Pb (zircon) dating of the mineralization history of
Bingham Canyon porphyry Cu-Mo-Au deposit, Utah, Master thesis (2009);
[5] Inan, E.: Metasomatic processes in contact aureoles of porphyry Cu deposits – A case
study of Bingham District, Utah, Master thesis (2003);
[6] Wills, B.A.: Mineral Processing Technology, Seventh Edition, BH (2006);
[7] Davenport, W.G.: Extractive Metallurgy of Copper, Fourth Edition, Pergamon (2000);
[8] Handbook of Flotation Reagents: Elsevier (2010);
[9] Cytec: Mining Chemicals Handbook (2010);
36
8. APPENDIX A
Figure 28: Results of all tests.
2 min 5 min 10 min 2 min 5 min 10 min
Test #1 88.65 93.91 95.81 79.15 87.53 90.78
Test #2 85.14 93.62 95.53 74.97 86.80 90.68
Test #3 81.78 92.65 94.75 69.67 84.54 88.81
Test #4 87.89 93.56 95.40 76.60 86.72 90.71
Test #5 88.95 94.28 95.44 77.40 86.34 89.04
Test #6 87.82 92.94 95.21 77.92 86.04 90.36
Test #7 89.44 94.71 96.05 75.88 87.53 90.63
Test #8 90.68 94.01 95.20 79.82 86.43 88.63
Test #9 86.96 94.31 96.05 78.32 89.06 92.31
Test #10 88.65 93.91 95.81 79.15 87.53 90.78
Test #11 86.22 94.65 94.91 72.90 84.65 89.56
Test #12 83.64 92.74 94.85 70.50 84.91 89.49
Test #13 81.45 92.55 95.06 64.98 83.22 89.11
Test #14 87.59 92.97 94.95 76.44 86.07 89.82
Test #15 89.51 94.20 95.80 77.90 87.67 90.96
Test #16 85.68 92.17 93.46 64.60 77.00 80.13
Test #17 83.04 91.17 93.87 73.78 84.65 89.01
Test #18 77.19 86.58 95.93 72.34 83.33 92.05
Copper Recovery (%) Molybdenum Recovery (%)