tcc_gustavo_pereira

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



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



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 -



/ 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


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


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


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]

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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

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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,

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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.

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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.

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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.

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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

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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

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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.



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

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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.



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

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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.

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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.

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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);

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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 (%)

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