BULK FLOTATION OF COMPLEX COPPER ORE FROM SIOCON, ZAMBOANGA SIBUGAY

BULK FLOTATION OF COMPLEX COPPER ORE FROM SIOCON,

ZAMBOANGA DEL NORTE

________________________________________________________

An Undergraduate Thesis

Presented to the Faculty of the

Metallurgical Engineering Department

College of Engineering

MSU- Iligan Institute of Technology

Iligan City

________________________________________________________

In Partial Fulfillment of

the requirements for the degree of

Bachelor of Science in Metallurgical Engineering

Khmer Lee P. Lugod

May 2010

Mindanao State University ILIGAN INSTITUTE OF TECHNOLOGY

Iligan City, 9200 Philippines _____________________________________________________

COLLEGE OF ENGINEERING

APPROVAL SHEET

The undergraduate thesis attach hereto entitled “Bulk Flotation of Complex Copper Ore From Siocon, Zamboanga del Norte”, prepared and submitted by KHMER LEE P. LUGOD, in partial fulfillment of the requirements for the degree in Bachelor of Science in METALLURGICAL ENGINEERING is hereby recommended for approval.

Prof. MA. TERESA T. IGNACIO

Thesis Adviser

___________________

Date Signed

This undergraduate thesis is approved in partial fulfillment of the requirements for the degree of Bachelor of Science in METALLURGICAL ENGINEERING.

Dr. FELICIANO B. ALAGAO

Dean, College of Engineering

___________________

Date Signed

Prof. JULIUS TORRALBA

Panel Member

___________________

Date Signed

Engr. VANNIE JOY RESABAL

Panel Member

___________________

Date Signed

Prof. ROSALINDA C. BALACUIT

Department Chairman

___________________

Date Signed

You cannot dream yourself into a character: you must hammer and forge yourself into one. Henry D. Thoreau

ABSTRACT

Complex ores contain profitable amounts of more than one valuable mineral. One

of the most common of these ores is the copper-zinc sulfide. In the Philippines such type

of ores can be found in the Bicol and Zamboanga peninsula. The separation of zinc and

copper has posed a challenge to a profitable concentration process. The problem is even

more complicated with the presence of arsenic whose content could be greater than the

smelter’s acceptable limit. This study examined the bulk flotation of copper and zinc

utilizing O-isopropyl ethyl thiocarbamate (NASCOL 446) and di-isobutyl

monothiophosphate (NASCOL 201) as collectors. The behavior of the associated arsenic

was also investigated.

One kilogram of ore, at -200 mesh was used as feed to the bulk flotation that was

conducted for 8 minutes. Two collectors, independently added, were used at 3 different

dosages: 20, 30, and 40 g/ton. The copper and zinc contents were determined as well as

their respective recoveries in the bulk concentrate. Arsenic content and recovery were

also calculated. Flotation using NASCOL 446 yielded 14.95% copper with 91.6 %

recovery using 40 g/ton. The best results for zinc was obtain using 40 g/ton NASCOL

446 which yielded 6.85% zinc with recovery 96.18% Zn. Arsenic lowest yield is 0.22%

arsenic with 96.7% recovery using 40 g/ton NASCOL 446.

ACKNOWLEDGEMENT

First and foremost, praise and thanks goes to my savior Jesus Christ for the many

blessing undeservingly bestowed upon me.

It would not have been possible to write this under-graduate thesis without the

help and support of the kind people around me, to only some of whom it is possible to

give particular mention here. Above all, I would like to thank my parents, brother Josef,

sister Lj and Aunt Lorna, who have given me their unequivocal support throughout, as

always, for which my mere expression of thanks likewise does not suffice.

This thesis would not have been possible without the help, support and patience of

my principal supervisor, Prof. Maria Teresa Ignacio, not to mention her advice and

unsurpassed knowledge of flotation studies.

To Prof. Julius Torralba and Engr Vannie Joy Resabal, my panel, thanks for

the truthful comments and suggestions on my papers.

I would like to acknowledge the academic and technical support of TVI Pacific

INC., and its staff, for providing the reagents and the sulphide ore, also in the analysis

of the bulk feed and bulk tails that provided the necessary financial support for this

research. To NASACO INTERNATIONAL and Engr Enrico Nera, for providing the

flotation reagents which is used in the study.

I am most grateful to Engr. Venice Onog and Engr Arnel Ang-og for providing

me the technical knowledge for the flotation study, which have been a valuable and

reliable method in the whole process. I would also like to thank Engr. Hans Enriquez,

for his kindness and generosity during my stay in TVI. To Engr. Edgardo V. Arellano,

for granting the proposal and etc., without his support, this study would not have been

possible.

I would like to thank Mr. Alexes Jann P. Agudera and Mr.Jonas Karl

Liwanag for their kindness, friendship and support, together with. The generosity and

encouragement of Mr. Belrie Dagasuhan, Jr, for his unending support. To Ms. Shelda

Capiral, who has never forgotten and for her generosity. To Ms. Jenny Suerte and

vi

Shehannie Guliman and Richellie Yuson for the resources, kindness and help on the

study. To Ms. Dulce Libby Garay for lending her thesis softcopy is very will

appreciated. To Ate Marivic, Cecil, and Linda thanks for the endless support and

generosity, for the help and encouragement. To Maam Nannete Abatayo, for her

support, patience, and guidance on the use of laboratory equipments and others.

Amongst my fellow undergraduate students in the Department of Metallurgical

Engineering, the effort made by Ms. Patricia Candari, Aileen Insalada, Jane Flores,

and team MESS for organizing the thesis presentation.

I would also like to thank my colleagues and friends in the College of

Engineering, Education, Arts and Sciences, Science and Mathematics. Last, but by no

means least, I thank my friends in Apexmines (Engr. Anarica Palconet and Ruben

Caballero) , A. Boulton of AB Offices LLC -USA, Val Flint of Reinforce- Australia,

Gee Rod of Webmate-UK, and elsewhere for their support and encouragement

throughout, some of whom have already been named.

For any errors or inadequacies that may remain in this work, of course, the

responsibility is entirely my own.

Khmer Lee P. Lugod May 2010, Iligan, Philippines

LIST OF CONTENTS

CHAPTER DESCRIPTION PAGE

Title Page i Approval Sheet ii Free Page iii Abstract iv Acknowledgement v Table of Contents vii List of Tables ix List of Figures x

I INTRODUCTION

1.1 Background of the Study 1 1.2 Statement of the Problem 3 1.3 Objectives of the Study 3 1.4 Significance of the Study 4 1.5 Theoretical Framework 4 1.6 Scope and Limitations 6

II REVIEW OF RELATED LITERATURE 7

III METHODOLOGY 17

3.1 Sample Preparation 17 3.2 Bulk Flotation 17 3.3 Reagents 18 3.4 Chemical Analysis 19 3.5 Microscopy 19 3.6 Recovery Calculations 20 3.7 Experimental Design and Statistical

Methods of Analysis 21

viii

PAGE

IV RESULTS AND DISCUSSION

22

4.1 Ore Characterization

22

4.2 Flotation Results of Copper, Zinc, and

Arsenic

26

V CONCLUSIONS AND RECOMMENDATIONS 39

5.1 Conclusion

39

5.2 Recommendations

39

BIBLIOGRAPHY

41

APPENDICES

A Calculations

44

B Tables of Obtained and Calculated

Results

45

C Calculated Recoveries

54

D Statistical Analysis

56

E Equipments, Materials, Procedures and

Products

58

LIST OF TABLES

PAGE

Table 3.1 Experimental design layout for bulk flotation

21

Table 4.1 Chemical analysis of the bulk feed using Inductive

Coupled Plasma (ICP).

22

Table 4.2 Calculated grades of copper in the bulk concentrate

27

Table 4.3 Mean grades of copper in the Bulk concentrate

27

Table 4.4 Calculated recoveries of copper in the bulk concentrate

28

Table 4.5 Mean recoveries of copper in the bulk concentrate

28

Table 4.6 Recoveries of zinc in the bulk concentrate utilizing

NASCOL 446 and NASCOL 201

31

Table 4.7 Mean recoveries of zinc in the bulk concentrate using


31

Table 4.8 Calculated grades of zinc in the bulk concentrate

32

Table 4.9 Mean grades of zinc in the Bulk concentrate

32

Table 4.10 Recoveries of arsenic in the bulk concentrate utilizing


35

Table 4.11 Mean recovery of arsenic in the bulk concentrate using


35

Table 4.12 Calculated grades of arsenic in the bulk concentrate

36

Table 4.13 Mean grades of arsenic in the bulk concentrate

36

Table B.1 Specific gravity results of ore

45

Table B.2 Copper analysis using NASCOL 201

46

Table B.3 Copper analysis using NASCOL 446

47

Table B.4 Zinc analysis using NASCOL 201

47

Table B.5 Zinc analysis using NASCOL 446

48

Table B.6 Arsenic analysis using NASCOL 201

48

Table B.7 Arsenic analysis using NASCOL 446

49

Table B.8 Results of chemical analysis 1st batch

50

Table B.9 Results of chemical analysis 2nd batch

51

Table B.10 Results of chemical analysis 3rd batch

52

Table B.11 Bulk analysis results of feed.

53

Table C.1 Calculated recovery of copper (%) in bulk concentrate.

54

Table C.2 Calculated recovery of zinc (%) in bulk concentrate.

54

Table C.3 Calculated recovery of arsenic (%) in bulk concentrate.

55

Table D.1 Analysis of Variance for Percent Recovery at α = 0.05.

56

LIST OF FIGURES

PAGE

Figure 1.1 Principle of froth flotation

5

Figure 3.1 General bulk flotation flowsheet

18

Figure 3.2 Experimental bulk flotation flowsheet

19

Figure 4.1 Pyrite from bulk feed (a) pyrite at -200, +325 mesh; 500x

magnification (b) Reference picture for pyrite

23

Figure 4.2 From bulk feed,(a) sphalerite (b) arsenopyrite (-200, +140

mesh @ 500x magnification)

23

Figure 4.3 Reference picture for (a) sphalerite (b) arsenopyrite

24

Figure 4.4 Bornite from bulk feed (a) bornite -200, +200 mesh @

500x magnification, (b) Reference picture for bornite

24

Figure 4.5 (a) Chalcopyrite at 200x total magnification. (b)

Reference microscopic view of chalcopyrite

25

Figure 4.6 Pieces of crushed bulk feed ore

25

Figure 4.7 Froth formations in the bulk flotation

26

Figure 4.8 Effect of collector dosage on copper recovery in the

flotation concentrate

29

Figure 4.9 Effect of collector dosage on copper grade in the flotation

concentrate

30

Figure 4.10 Effect of collector dosage on zinc recovery in the flotation

concentrate

33

Figure 4.11 Effect of collector dosage on zinc grade in the flotation

concentrate

34

Figure 4.12 Effect of collector dosage on arsenic recovery in the

flotation concentrate

37

Figure 4.13 Effect of collector dosage on arsenic grade in the flotation

concentrate

38

Figure D.1 Interaction plot of the reagents and dosages

57

Figure E.1 Filtered cake from the bulk flotation process

58

Figure E.2 Bulk flotation process

59

CHAPTER I

INTRODUCTION

1.1 Background of the Study

TVI Philippines Inc. is mining a complex ore which is being studied for the

recovery of both zinc and copper. A polymetallic ore could possibly produce more than

one mineral concentrate. However, the ore also contains arsenic. It is one of the elements

that is encountered in copper sulfide concentrates, and its removal from associated

sulphide minerals is recommended in order to reduce contamination of the valuable

concentrate and/or to reduce, arsenic emission to the atmosphere (Kantar, 2002).

The presence of arsenic produces metallurgical problems, which makes metal

extraction difficult, and the recovery of a final product of high purity. It is regarded as a

highly toxic contaminant resulting in environmental problems when released to the due to

the atmosphere and possible water contamination associated to the processing of

arsenic bearing ores and concentrates (Makita, et.al. 2010).

Complex ores contain profitable amounts of more than one valuable mineral.

Metallic minerals are often found in certain associations within which they may occur as

mixtures of a wide range of particle sizes or as single-phase solid solutions or

compounds. Galena and sphalerite, for example, associate themselves commonly, as do

copper sulphide minerals and sphalerite to a lesser extent. Pyrite (FeS2) is very often

associated with these minerals (Wills, 1997).

According to Bulatovic (2007), the complex sulfide copper ores are considered

easy to treat provided that the main copper mineral is chalcopyrite. In case the ore

contains secondary copper minerals, such as chalcocite, bornite and covellite, depression

of pyrite may be a problem because the pyrite can be activated by copper ions generated

during the grinding operation. Some copper sulfide ores can be partially oxidized, also

influencing the selection of a reagent scheme, with the exception being a hypogene

sulfide copper ore.

2

There are two common options practiced in the treatment of these ores. These are

(1) sequential copper flotation from pyrite and other sulfides, the most common practice

in the treatment of sulfides ores and (2) bulk or semi-bulk flotation followed by copper–

pyrite separation after re-grinding of the bulk concentrate. This method is used in the

case where copper is finely disseminated with pyrite or with ore that contains clay

minerals (of acidic nature), which interferes with copper flotation (Bulatovic, 2007).

The reagent schemes used for the treatment of complex sulfide copper ores are

much more diverse and are designed to cope with specific problems associated with

processing the ore. When treating hypogene sulfide copper ores, the reagent scheme is

relatively simple. It uses xanthate as a collector in alkaline pH (11.0–11.5). In some

cases, dithiophosphate is used as a secondary collector when secondary copper minerals

are present in the ore. In the case of stringer ore and copper ores in which the pyrite is

active, the reagent scheme is more complex and involves different depressant

combinations (Bulatovic, 2007).

The choice of collector also depends on the nature and occurrence of copper and

associated sulfides. In most cases, xanthate collectors are used alone or in combination

with dithiophosphates or thionocarbamates. Dithiophosphates and thionocarbamates are

normally used when secondary copper minerals are present in the ore or when the copper

flotation is carried out at lower pH. A mixture of xanthates (i.e. ethyl-butyl, ethyl-

isopropyl) has been successfully used in a number of Russian operations where both

selectivity and recovery were improved when using a mixture of two xanthates

(Bulatovic, 2007).

In the processing of poly-metallic ores, a selective separation process will allow

an economical rejection of arsenic to enable copper concentrates to meet the typical

smelter penalty level of 0.5% As .

The bulk flotation process is essential to pre-clean the mineral of the arsenic

present in its raw stage and prior to sequential flotation of poly-metallic minerals.

3

Flotation is, at the present, the only method that can be used to beneficiate the

complex sulfides such as copper-lead-zinc and copper-zinc. The process involves a

series of flotation known as the differential flotation. In the flotation of copper-zinc

massive sulfide ores, flotation properties of copper and zinc are determined by the nature

and composition of the ore. Copper zinc massive sulfide is even regarded as the most

complex ore and consequently the most difficult to treat. The complexity of the ore leads

to certain ways of treating it by the use of sequential flotation in which copper and zinc

are floated in series of operation that may involve a bulk flotation of copper and zinc

feeding it to another circuit of copper and zinc separation in order to produce both

marketable value of copper and zinc concentrate (Bulatovic, 2007). To optimize a

flotation operation, test works of reagents are conducted.

1.2 Statement of the Problem

This study examined the bulk flotation of copper, zinc and arsenic utilizing

NASCOL 201 (O-isopropyl ethyl thiocarbamate) and NASCOL 446 (di-isobutyl

monothiophosphate) as collectors.

1.3 Objective of the Study

This study aimed to:

1. Determine the effect of NASCOL 201 (O-isopropyl ethyl thiocarbamate) and

NASCOL 446 (di-isobutyl monothiophosphate) on the recovery of copper, zinc

and arsenic in the bulk concentrate.

2. Determine the effect of dosages and collector on the recovery of copper, zinc,

and on arsenic removal.

3. Preliminary identify the minerals present in the bulk feed.

4

1.4 Significance of the Study

Early removal of arsenic in the bulk flotation will yield better quality of copper

concentrate. With early removal of arsenic, there will be a better feed in the subsequent

flotation system. With a better copper quality in the bulk feed, there is less cost in the

differential separation of copper and zinc. This will result to higher recoveries of copper

after the bulk concentration. It will also result to a more efficient separation since the

gangue mineral has been pre removed and pose fewer problems in the copper and zinc

flotation.

1.5 Theoretical Framework

Flotation is a physico-chemical separation process that utilizes the difference in

surface properties of the valuable minerals and the unwanted gangue minerals. The ore is

first conditioned with chemicals to make the copper minerals water-repellent (i.e.,

hydrophobic) without affecting the other minerals. Air is then pumped through the

agitated slurry to produce a bubbly froth. The hydrophobic copper minerals are

aerophillic and they are attracted to air bubbles, to which they attach themselves, and then

float to the top of the cell. As they reach the surface, the bubbles form froth which

overflows into a trough for collection. The minerals that sink to the bottom of the cell are

removed for disposal (Wills, 1997).

5

Figure 1 Principle of froth flotation (Source: Wills, 2004).

The purpose of bulk flotation is usually to separate one or more minerals from the

other minerals, and the selective adsorption of collector on floatable minerals is a

prerequisite for a successful separation. The selectivity of collector adsorption is often

affected by regulating agents. They can either enhance or prevent the adsorption of

collector on a particular mineral.

Alkalinity plays an important, though complex, role in flotation. Selectivity in

complex separation is dependent on the delicate balance between reagent concentrations

and pH. Flotation is carried out in an alkaline medium, as most collectors including

xanthate are stable under these conditions. Corrosion of cells, pipeworks and etc. is also

minimized (Wills, 1997).

According to Wills (1997), 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.

6

Modifying reagents are reagents that render either floatability or hydrophobicity

on the mineral particles. In the case of separation of sphalerite and copper sulphide, these

minerals float together and an addition of sphalerite depressant is needed to separate

copper from zinc which is useful on for the next processes (Bulatovic, 2007).

1.6 Scope and Limitations

The complex sulphide concentrate was procured from TVI Philippines in

Zamboanga del Norte. Trials conducted per flotation treatment were limited to two, due

to resources limitations. A portable pH meter to check the alkalinity of the slurry was

used throughout the test.

The chemical analysis of the sample was done in TVI. Analysis was limited to

base metals, and analysis for gold was not conducted due to financial constraints. The

age of the ore was not considered. Liberation size analysis was not taken and the -200

mesh grinding was assumed to be sufficient based on the physical appearance of the ore.

Analysis of the pulp for the presence of ions was also not conducted.

CHAPTER II

REVIEW OF RELATED LITERATURE

General sulfide flotation

Some minerals are naturally hydrophobic like Sphalerite and Chalcopyrite coal

and molybdenite. However natural hydrophobicity of sulphides minerals is the most

debatable issue in the field of sulphides flotation (Mendiratta, 2000).

The native floatabilty of minerals according to Mendiratta (2000) is due to

contamination of surface by hydrocarbon bearing groups or due to contamination of

surface by hydrocarbon bearing groups or due to the intentional modification of the

surface. The surfaces formed due to rupture of Van der Waals bonds are naturally

hydrophobic.

According to Mendiratta (2000), sulfide minerals could be rendered floatable by

the presence of sulfur at their surfaces as unlike oxides, sulfur does not form hydrogen

bonds with water. Mild oxidation was necessary for collectorless flotation of

chalcopyrite, the oxidation of chalcopyrite led to formation of hydrophobic elemental

sulphur, S0, which was responsible for flotation. The presence of oxygen would cause

formation of hydrophilic oxidation products and strip sulphides minerals of their natural

floatabilty

Collectorless flotation of chalcopyrrite in presence of Na2S, which is being

strong reducing agent, is used for cleaning the surfaces of oxidation products. It was

suggested that sulphide ions would displace the hydrophilic sulfoxy, which may then

become naturally hydrophobic. Although Na2S striped the surface of hydrophilic species,

slightly oxidizing conditions were still required for collectorless flotation of chalcopyrite.

The results supported the previous findings that oxidizing conditions were required for

collectorless flotation of chalcopyrite. Based on flotation and surfaces studies, Na2S

played a twofold role. Firstly, it displaced the hydrophilic sulfoxy species, such as SO42-

8

and S2O32-

, and created relatively fresh surface. Secondly, it sulphadized the mineral

sulphur species (Mendiratta, 2002).

Hydrophobicity could be due to formation of Cu2S or CuS species along with

elemental or excess sulphur on the surface. There are three different forms of sulphur on

oxidized gold surface: atomic sulphur, S0, polysuphides, Sx+1

2-, and elemental sulphur

represented by S8 which rendered the gold surface hydrophobic (Mendiratta, 2002).

Depending upon pH, excess sulphides ions on the surface may be oxidized to

elemental sulphur or polysulphides, rendering it more hydrophobic. It is proposed that

there is the formation of hydrophobic metal-deficient sulphur layer upon mild oxidation.

On the formation of iron oxide on chalcopyrite, a metal deficient layer (unexposed metal

surface) can be floated without collector (Mendiratta, 2002).

Based on the work of Mendiratta (2002) sulphides minerals are ranked according

to their natural hydrophibicities in the descending order as follows: Chalcopyrite > galena

> Pyrrhotite > Pentlandite > Covellite > Bornite > Chalcocite > Sphalerite > Pyrite >

Arsenopyrite.

The flotation of complex sulfide copper ores

Complex ore are often characterized by particularly fine intergrowth of the

mineral values. Due to this specific mineralogical characteristic, it is necessary to finely

grind and concentrate ore prior to the solubilization of valuable metals (Makita, et.al.

2010). Apart from high energy consumption in fine grinding, its efficiency deteriorates

rapidly with decreasing particle size below approximately 10 µm. Many investigators

have delineated various reasons for this difficulty, including high reagent consumption,

high rate of surface reactions, slime coating, morphological and surface chemical changes

during fine grinding (Yoon, et.al. 2002).

According to Bulatovic (2007), sulfide copper ores are considered easy to treat

provided that the main copper mineral is chalcopyrite. In case the ore contains secondary

copper minerals, such as chalcocite, bornite and covellite, depression of pyrite may be a

9

problem because the pyrite can be activated by copper ions generated during the grinding

operation. Some copper sulfide ores can be partially oxidized, also influencing the

selection of a reagent scheme, with the exception being a hypogene sulfide copper ore.

Copper sulfide ores are normally finer grained than porphyry copper ores and require

finer grinding (i.e. 70–80% <200 mesh). He also added that copper sulfide ores are

disseminated and in some cases, would require fine re-grinding of the rougher

concentrate. Fine copper minerals have a low rate of flotation, which may result in losses

in recovery. Unlike porphyry copper ore, where the reagent schemes are similar for most

operations, the reagent schemes used for the treatment of sulfide copper ores are much

more diverse and are designed to cope with specific problems associated with processing

the ore.

In the flotation of copper–zinc massive sulfide ores, the flotation properties of

copper and zinc are determined by the nature and composition of the ore. Selectivity

between chalcopyrite and sphalerite, in principle, is determined by the type of copper

minerals present in the ore. The simplest separation was achieved when only

chalcopyrite is present in the ore. The presence of secondary copper minerals (i.e.

bornite, covellite and digenite) represents a significant problem in the separation of

copper from sphalerite. This is because the secondary copper minerals are soluble and

during grinding, or in situ, they release copper ions, which activate sphalerite. It is quite

common that copper–zinc ores that contain secondary copper minerals have a covellite

layer on the sphalerite surface (Bulatovic, 2007).

Sphalerite is the most important mineral which appears in many lead–zinc and

copper– lead–zinc ores. The composition of sphalerite is highly variable and depends on

the impurities contained in sphalerite. According to Takeuchi and Gondo (1957), the

difference of floatability of zinc ores is due to the difference in kinds or inclusions in it,

such as iron or copper minerals. In zinc fine particles of copper minerals could be found.

Cu++

activates zinc ore, the floatability of zinc ore is controlled by the solubility of

coexisting copper ore, the surface of zinc is activated by Cu++, and then the floatability

of zinc ore increases. These impurities are either replacements of zinc in the crystal

10

structure of sphalerite or the formation of emulsions in the mineral itself, as micron

inclusions or ‗disease‘ in sphalerite. The most common impurities in sphalerite are iron,

cadmium, copper, indium, gallium, tin and other elements. Iron content of sphalerite can

vary 1 from to 25%, cadmium can be as high as 1.5%. The copper can vary from traces

to 20%. These impurities in the sphalerite are critical for determining reagent schemes

suitable for the treatment of copper–zinc ore (Bulatovic, 2007).

Apart from the copper activation of sphalerite, the presence of silver, arsenic or

other ions, which come from sulfosalts, may activate sphalerite and create a problem in

the selective separation of copper and sphalerite. Although literature and textbooks

contain vast references on flotation, activation and de-activation of sphalerite, little to

nothing is known regarding the flotation behavior of sphalerite that contains impurities,

even though it is the mineral that has been studied the most. In actual practice, the

separation of sphalerite from copper can be very difficult on one hand, or flotation of

sphalerite from pyrite and/or pyrrhotite can be relatively easy, on the other hand

(Bulatovic, 2007).

Lime utilization and pH conditions

In flotation, several parameters are to be observed such as the pH of the slurry and

its percent solids. In controlling the pH, lime is usually used in plant. The use of a lime

circuit is practically universal in the flotation of copper ores. Lime alkalinity is generally

maintained in the pH range of 9.5 to 11.5. The higher pH serves to depress the iron

sulfide gangue minerals which are commonly present. The pH can also influence the

froth structure and floatability of the copper minerals. These characteristics are adversely

affected below some minimum pH value which varies from ore to ore, especially when

xanthates and dithiophosphates are used. (Cytec Industries Inc., 2002). The presence of

iron ions in solution depends on the oxidation behaviour of the sulphide ore and on the

chemical conditions of the pulp. A common way for activating the surface of pyrite and

arsenopyrite is through the addition of copper ions, which improve their recoveries

(Monte, 2002).

11

Arsenic and associated copper minerals

Arsenic occurs at varying levels in some copper ore bodies, and is a significant

environmental hazard in the copper smelting process when emissions are released into the

atmosphere. The arsenic in the ore is contained in copper-arsenic sulfide minerals, such

as arsenopyrite and tennanite (Smedley and Kinniburgh, 2002). Arsenopyrite is one of

the comparatively little studied sulphide minerals in the flotation technology (O'Connor,

et al, 1990).

Most often in ores, bornite is represented as a secondary copper mineral, together

with chalcopyrite and chalcosine, mainly in porphyry copper–molybdenum and copper–

gold ores. Bornite is relatively stable and does not oxidize. Its floatability depends very

much on the size. Fine bornite (<20 µm) does not float readily and this may represent a

significant problem during beneficiation of disseminated sulfides where the bornite is

present. Floatability of bornite is also pH related, where at a pH >10 its floatability

improves greatly.

Flotation studies on arsenic removal

Trace elements such as Arsenic, bismuth, cadmium, lead, and etc. may also be

present in varying amounts. Arsenic removal in sulphide flotation has been studied

extensively by Ma

and Bruckard, (2009) with various approaches, including pre-

oxidation of flotation pulp, Eh control during flotation and the use of selective

depressants/collectors. Pre-oxidation of flotation pulp using oxidizing agents or aeration

conditioning represents a simple approach in arsenic removal and was found effective in

many cases. Selective flotation of arsenic minerals through Eh control has made

significant advances in recent years with promising results achieved.

In addition, various depressants and collectors have also been studied in arsenic

removal. O'Connor, et al, (1990) suggested the recovery of arsenopyrite from an

arsenopyrite/pyrite ore is desirable for a number of reasons. This can be optimized using

a two stage flotation process in which a dithiophosphate is added at pH=11 in the first

12

stage and copper sulphate and a dithiocarbamate in the second stage. It was found that

better separations were obtained when aged ore was used. It was possible to simulate this

ageing process by heating. The basis of the separation relies on findings that the lower

limiting pulp potential threshold for tennantite is lower than that for chalcopyrite such

that there is a potential window in the reducing region where tennantite is strongly

floatable but chalcopyrite is not. Little or no selectivity between tennantite and

chalcopyrite was found in the oxidizing pulp potential region for the range examined (Ma

and Bruckard, 2009).

From the composite sample tested by Ma and Bruckard (2009), which had a head

grade of 0.11% As and 1.2% Cu, it was possible to produce a low-arsenic high-copper

concentrate containing 52% of the non-tennantite copper and assaying 2600 ppm As.

Computer simulations have shown that for a feed containing a more typical arsenic and

copper level (200 ppm As and 1% Cu) the efficiency of separation should be sufficient to

concentrate about 61% of the copper in a product assaying less than 2000 ppm As. Aside

from flotation studies to remove contaminants like arsenic in the polymetallic minerals by

flotation , studies done by Curreli et. al.( 2008), on alkaline leaching shows an increase

of arsenic extraction which is enhance by the effect of mechanical activation. Arsenic

leaching provides a better separation of gangue minerals by means of alkaline leaching

with mixtures of sodium sulfide and sodium hydroxide. In their study, the influence of

the most significant process variables like specific area of the solid, temperature, pH, and

reagent concentration of the leach solution has been investigated. Leaching selectively

solubilises the arsenic and some gold but does not affect the copper which transform

entirely in the leach residue as a new species. Increasing the surface area of the

concentrates at temperature of 100 degrees improves the efficiency of the whole process.

The theory and practice of sulfide flotation again state that effectiveness of all

classes of flotation agents, to a large extent on the degree of alkalinity or acidity of the

ore pulp. As a result, modifiers that regulate the pH are of great importance. The most

commonly used pH regulators are lime, soda ash and, to a lesser extent, caustic soda. In

sulfide flotation, however, lime is by far the most extensively used. In copper sulfide

13

flotation, which dominates the sulfide flotation industry, or example, lime is used to

maintain pH values over 10.5, more usually above 11.0 and often as high as 12 or 12.5.

In prior art sulfide flotation processes, preadjustment of the pH of the pulp slurry to 11.0

and above is necessary, not only to depress the notorious gangue sulfide minerals of iron,

such as pyrite and pyrrhotite, but also to improve the performance of a majority of the

conventional sulfide collectors, such as xanthates, di-thiophosphates, trithiocarbonates

and thionocarbamates (Nagaraj and Wang, 1986).

Reagent scheme used for flotation of sulfide copper ores

Most of operating plants, treating gold bearing sulphide ores, use various types of

xanthate, as primary collector, in combination with dithiophosphate as secondary

collector. Mercaptobenzothiazole is usually employed for the treatment of oxidised

pyrite containing ores, with little or even no xanthate additions to the scavenger flotation

operation. The thiol collector-mineral adsorption reaction strongly affects the floatability

of sulphide minerals. The rate of collector adsorption can be influenced by many factors.

Pre-treatment conditions, such as: the grinding environment, the dissolved oxygen

concentration, the pulp potential and the pH, are the key factors which determine the

extent and kinetics of this reaction (Monte et al, 2002).

It is well established that the adsorption of xanthate occurs via a mixed potential

mechanism, involving the anodic oxidation of xanthate and the cathodic reduction of

oxygen.The adsorption of xanthate results either in the formation of dixanthogen or metal

xanthate. In the first case, the mineral itself does not participate in the reaction except

offering a passage for the transfer of electron. This would be the case for xanthate

adsorption on pyrite, pyrrhotite, and gold (Richardson, 1976).

The adsorption of xanthate results either in the formation of dixanthogen or metal

xanthate. In the first case, the mineral itself does not participate in the reaction except

offering a passage for the transfer of electron. This would be the case for xanthate

adsorption on pyrite, pyrrhotite, and gold (Dung, 1995)

14

For the case of xanthate adsorbing on some of the other sulfide minerals (e.g., chalcocite

and galena), the mineral itself is participating in the adsorption process resulting in the

formation of metal xanthates. The mechanism may be viewed as a two-step process

involving an initial electrochemical reaction (E), which is the oxidation of the mineral to

release the metal ions, followed by the chemical reaction (C) between the metal ions and

xanthate to form metal xanthate. In organic electrochemistry, such mechanisms are

referred to as coupled electrochemical and chemical reactions of the EC-type (Dung,

1995).

In the EC mechanism, the electrochemical reaction is controlled by the

electrochemical potential (E) of the system, while the chemical step is controlled by its

stability constant (pK), as suggested by the chemical theory of collector adsorption. It

may be stated, therefore, that the adsorption of thiol collectors on sulfides is controlled by

both the E and pK values of the system. The E determines the availability of metal ions,

while the pK of the metal thiol complex determines whether this complex can be formed.

The EC mechanism simplifies the understanding of the adsorption process, specifically

for cases where the mineral itself undergoes oxidation and participate in the adsorption

reactions. The EC mechanism was employed to explain the adsorption of modified thiol-

type collectors, including MTP, on precious metals and that of DTPI on copper and

copper sulfides (Dung, 1995). .

The beneficial effects of the synergy between two or more reagents were realized

long time ago. The purpose of using a mixture of collectors was to increase both the

recovery and selectivity. Two thiol collectors, isopropyl xanthate (SIPX) and di-isobutyl

dithiophosphinate (DTPI), having different chemical and functional properties, were

used. The adsorption behavior of these collectors from their mixtures was investigated at

various SIPX:DTPI ratios and sequence of addition by cyclic voltammetry and adsorption

experiments at pH 9.2. The results revealed that the maximum synergistic effect of using

mixture of SIPX and DTPI was strongly influenced by the ratio of the collectors in the

mixture and particularly sequence of addition (Bagci, et.al. 2007).

15

Unlike porphyry copper ore, where the reagent schemes are similar for most

operations, the reagent schemes used for the treatment of sulfide copper ores are much

more diverse and are designed to cope with specific problems associated with processing

the ore.

When treating hypogene sulfide copper ores, the reagent scheme is relatively simple.

It uses xanthate as a collector in alkaline pH (11.0–11.5). In some cases, dithiophosphate

is used as a secondary collector when secondary copper minerals are present in the ore.

In the case of stringer ore and copper ores in which the pyrite is active, the reagent

scheme is more complex and involves different depressant combination (Bulatovic,

2007).

The choice of collector also depends on the nature and occurrence of copper and

associated sulfides. In most cases, xanthate collectors are used alone or in combination

with dithiophosphates or thionocarbamates. Dithiophosphates and thionocarbamates are

normally used when secondary copper minerals are present in the ore or when the copper

flotation is carried out at lower pH. Good metallurgical results are obtained with

thionocarbamate during the flotation of clay-containing sulfide copper ore (Bulatovic,

2007).

Monothiophosphates are relatively new collectors for sulfide ore flotation.

Monothiophosphates (MTPs) are effective copper collectors at typical pH usage range

from neutral to alkaline. They are highly selective against Iron Sulphides. They also

improved the recovery of valuable metals such as Pb, Cu, Zn, PGM‘s and Ni

(Dithiophosphates, 2010).

O-Isopropyl-N-ethyl thionocarbamate is another excellent collector in flotating

nonferrous metallic sulfides, with less collecting pyrite and higher selectivity. The

collectivity of ethyl thiocarbamate is similar with xanthates and dithiophosphates. Ethyl

thiocarbamate displays strong collective power and rapid flotation, low dosage in

comparison with xanthates and dithiophospates. Because its collectivity for pyrite is

weak, Ethyl thiocarbamates exhibits excellent selectivity in flotation of sulphide ores.

16

Ethyl thiocarbamate also exhibits the better flotation results than xanthates and

dithiophosphates in flotation of copper, lead, zinc, antimony and other poly-metallic

sulfides (Flotation Reagents, 2010)

As each of the collectors had a similar probability of attachment, preferential

adsorption was unlikely. Cuprous ethyl xanthate has a lower solubility product than

cupric ethyl dithiophosphate and cupric ethyl dithiocarbamate; indicating that the

xanthate has the highest attraction for copper ions (Mermillod et. al, 2005).

The processes that affected the hydrophobicity of the particles and consequently

affected copper recovery and grade may have been the selective adsorption of the

different collectors on particular sites or changes in the orientation of the alkyl chains

resulting in superior surface coverage (Hangone, et al., 2005). In addition, for a mixture

of xanthate and dithiophosphate there may have been enhanced co-adsorption of

collectors at low collector concentrations. These phenomena may have played a role in

determining the degree of hydrophobicity of the mineral surface and may therefore have

been responsible for the differences in froth properties (Mermillod et. al, 2005)

CHAPTER III

METHODOLOGY

3.1 Sample Preparation

The raw ore was from TVI Phils. Inc, in Siocon, Zamboanga del Norte is a

complex copper-zinc ore obtained from the mine site‘s high sulphide zone. The samples

were crushed to -20 mesh and was ground using a laboratory ball mill. The feed to the

bulk flotation was -200 mesh. A representative sample was taken for chemical analysis.

The density of the ore was obtained using the Volume Displacement Method

using a 100 ml flash and a 32 gram ore sample. The difference in volume of the water

after filling the flask with the ore specimen was used to solve the density of the ore using

this equation:

D = 𝑚

𝑣

Where m is the mass of the ore and v is displaced water volume.

3.2 Bulk Flotation

A one kilogram ore, passing -200 mesh was fed to the flotation machine for the bulk

flotation of copper and zinc. A 30% solid by weight slurry was used. Collectors were

added separately and independent of each, in the flotation set-up. The collectors are


monothiophosphate ) , in the three different dosages. NASCOL HEL was used as a

frother. Lime was used to modify pH to 11 after collectors and frothers were added.

Flotation was carried out after 5 minutes of conditioning. Flotation time was 8 minutes.

18

Figure 3.1 General bulk flotation flowsheet.

3.3 Reagents


monothiophosphate) were used separately as a collector and NASCOL HEL as frother for

the bulk and copper flotation. The dosage for each collector was set to 20, 30, and 40 g/t

of ore and 10 g/t for NASCOL HEL based on the initial tests conducted.

19

3.4 Chemical Analysis

Chemical analysis of copper, zinc, and arsenic for the bulk feed and tails were

carried out by TVI Pacific Inc. using Inductive Coupled Plasma (ICP) method.

Figure 3.2 Experimental bulk flotation flowsheet.

3.5 Microscopy

The concentrate produced in the bulk flotation was analyzed by microscopy using

the optical microscope. Representative samples of bulk concentrates and bulk feed were

mounted separately on glass slides. The mineralogy of the ore was determined using

Nikon Optiphot – 100 Metallurgical Microscope. Pictures were taken at 200x

magnification.

BULK FLOTATION NASCOL HEL

LIME

1000 g Ore

-200 Mesh

30% Solid

BULK

CONCENTRATE

BULK TAILS

Collector A or

Collector B

20

3.6 Recovery Calculations

The recoveries of the concentrate were set up using the material balance. Bulk

feed and bulk tails were weighed after filtration and drying. The weights gathered were

used in the calculation. Results from the chemical analysis (ICP) were used for the

computation of the concentrate‘s assay and percent recoveries using the following

equations:

F = C + T (3.1)

fF = cC + tT (3.2)

% Recovery = cC x 100 (3.3)

fF

where:

F – weight of feed

C – weight of concentrate

T – weight of tails

f – assay of feed

t – assay of tails

c – assay of concentrate

21

3.7 Experimental Design and Statistical Methods of Analysis

A Randomized Complete Block Design was used as an experimental design and

an ANOVA was taken at α = 0.05.

Table 3.1 Experimental design layout for bulk flotation.

Where A and B (NASCOL 201 and NASCOL 446) are collectors used during

bulk flotation with 1, 2, and 3 as dosages. The frother used in the bulk flotation was only

NASCOL HEL.

Dosage

COLLECTOR

A B

NASCOL 201 NASCOL 446

1 A11 A12 B11 B12

2 A21 A22 B2 1 B22

3 A31 A32 B31 B32

CHAPTER IV

RESULTS AND DISCUSSION

4.1 Ore Characterization

The minerals present in the ore were bornite, sphalerite, arsenopyrite, and pyrite

as observed by microscopy. XRD analysis by Garay (2010) confirmed the analysis as

major mineral components of the copper ore. Covellite, cubanite, stibnite, bismuthinite,

hematite, magnetite, and quartz are also present Garay (2010). Finely disseminated

pyrite can be observed on the Fig 4.1. The ore was milled up to 100 percent passing 200

mesh to liberate minerals from pyrite. Table 4.1 below provides the chemical analysis of

the bulk feed which confirmed the existence of a copper-zinc polymetallic system.

Table 4.1 Chemical analysis of the bulk feed using Inductive Coupled Plasma (ICP).

Element Assay, % Element Assay, %

Ag 57.8 ppm Hg 17.94 ppm

As 0.21 Pb 0.07

Bi 71 ppm S 26.61

Cd 303.22 ppm Sb 165 ppm

Cu 9.28 Zn 4.03

Fe 36.51

23

Figure 4.1 Pyrite from bulk feed (a) pyrite at -200, +325 mesh; 500x

magnification (b) Reference picture for pyrite (Chesterton, 2000)

Figure 4.2 From bulk feed,(a) sphalerite (b) arsenopyrite (-200, +140

mesh @ 500x magnification)

24

Figure 4.3 Reference picture for (a) sphalerite (b) arsenopyrite

(Chesterton, 2000)

Figure 4.4 Bornite from bulk feed (a) bornite -200, +200 mesh @ 500x

magnification, (b) Reference picture for bornite (Chesterton, 2000)

25

Figure 4.5 (a) Chalcopyrite at 200x total magnification. (b) Reference

microscopic view of chalcopyrite.

Figure 4.6 Pieces of crushed bulk feed ore.

a b

26

4.2 Flotation Results of Copper, Zinc, and Arsenic

4.2.1 Copper Flotation

Fig. 4.7 shows copper flotation. The colour of the froth during copper flotation

was observed to be greyish. It appeared that zinc could have floated with copper during

the bulk flotation.

Figure 4.7 Froth formations in the bulk flotation.

The Cu grade of the bulk feed was 9.28% Cu. The highest grade of copper in the bulk

concentrate was 14.95% Cu using a 40 g/t NASCOL 446, and 12.85% Cu using 30 g/t

of NASCOL 201. The lowest mean grade was obtained using a 20 g/t of NASCOL 201

at 9.5% Cu.

27

Table 4.2 Calculated grades of copper in the bulk concentrate.

Dosage

(g/t)


Trial 1 Trial 2 Trial 1 Trial 2

20 13.6 10.1 8.9 11.7

30 11.1 15.1 10.6 9.5

40 15 14.9 7.5 8.3

Table 4.3 Mean grades of copper in the bulk concentrate.

Dosage (g/t) NASCOL 446 NASCOL 201

20 11.85 9.5

30 13.1 12.85

40 14.95 11.2

Table 4.4 shows the calculated recoveries of copper for both NASCOL 446 (di-

isobutyl monothiophosphate) and NASCOL 201 (O-isopropyl ethyl thiocarbamate). For

the recovery of copper, the best results were obtained using 30 g/t of NASCOL 446

which is 92.7% as shown on table 4.5. The highest was seen at 30 g/t using NASCOL

446 with 92.70 % recovery as shown in table 4.4. The lowest recovery at 62,22 % Cu

was obtained using a 40 g/t of NASCOL 201.

28

Table 4.4 Calculated recoveries of copper in the bulk concentrate.

Dosage

(g/t)



20 96.73 82.52 63.33 80.49

30 92.68 92.74 84.91 75.50

40 93.87 89.29 54.48 69.97

Table 4.5 Mean recoveries of copper in the bulk concentrate.


20 89.62 71.91

30 92.7 80.2

40 91.6 62.2

Figures 4.8 shows the copper recoveries and that the lowest recoveries were

obtained with NASCOL 201 (O-isopropyl ethyl thiocarbamate). NASCOL 446 showed

higher recoveries and better copper grades in the concentrate compared with NASCOL

201. The highest mean grade, 14.95% was obtained using NASCOL 446 at 40kg/ton

dosage with a mean recovery of 91.58%. Highest recovery at 92.71% was obtained using

a lower dosage of 30 kg/ton but a lower copper grade of 13.1%. What appeared to be a

better flotation of the copper minerals with NASCOL 446 may be due to the added

hydrophobicity which is provided by the additional alkyl group of di-isobutyl

29

monothiophosphate as compared to O-isopropyl ethyl thiocarbamate. However, the

lower results obtained using NASCOL 201 may be due to the differences in the

composition of the functional groups of the collectors (Hangone et al, 2005).

Figure 4.8 Effect of collector dosage on copper recovery in the flotation

concentrate.

Statistical analysis showed that the effect of collector was significant on the

recovery of copper, while variation of dosages was insignificant. The lower grade but

high recoveries could be due to the inclusion of pyrites in the float. Pyrite is activated by

copper ions (Monte, 2002). The presence of bornite which releases copper ions activates

pyrite during the grinding operation (Bulatovic, 2007). In most cases according to Wills

(2007), the activated sphalerite and pyrite in the bulk concentrate are covered with a layer

of collector, and are difficult to depress unless extremely large amounts of reagent are

0

10

20

30

40

50

60

70

80

90

100

0 20 40 60

% R

eco

very

of

Co

pp

er

Dosage (g/ton)

NASCOL 201

NASCOL 446

30

used. The formation of a complete monolayer of monothiophosphate ion on copper ore

surface may have already formed.

Figure 4.9 Effect of collector dosage on copper grade in the flotation

concentrate.

Although the effect of collectors‘ dosage was insignificant grade increased with

increase in concentration up to 40 g/ton, NASCOL 201 exhibited low recovery (62.2%)

and low grade (9.5 %) on the bulk concentrate as shown on table 4.4 and 4.6.

4.3 Zinc Flotation

The zinc grade of the bulk feed was 4.03% Zn. The amount of zinc in the bulk

concentrate was 6.85 % for the NASCOL 446 and 5.35% for the NASCOL 201. This

0

2

4

6

8

10

12

14

16

0 20 40 60

% G

rad

e o

f C

op

pe

r

Dosage (g/ton)

NASCOL 201

NASCOL 446

31

observation was also verified by calculating the recovery of zinc in the bulk concentrate

shown on table 4.6.

Table 4.6 Recoveries of zinc in the bulk concentrate utilizing NASCOL 446 and

NASCOL 201.

Dosage (g/t)



20 98.53 82.41 79.63 92.52

30 95.82 96.57 94.03 89.09

40 97.82 94.54 77.64 83.90

Table 4.7 Mean recoveries of zinc in the bulk concentrate using NASCOL 446 and

NASCOL 201.


20 90.47 86.1

30 96.19 91.6

40 96.18 80.8

Table 4.6 and 4.7 show the calculated recoveries of zinc for both NASCOL 446

(di-isobutyl monothiophosphate) and NASCOL 201 (O-isopropyl ethyl thiocarbamate).

For the recovery of zinc, the best results were obtained using 40 g/t NASCOL 446 with

32

96.19 % recovery (see also tables 4.8 and 4.9 for dosage results). The ANOVA showed

that the effect of the dosages of the collectors were not significant on the recoveries of

zinc. However, collector variation was significant.

Table 4.8 Calculated grades of zinc in the bulk concentrate.

Table 4.9 Mean grades of zinc in the Bulk concentrate.


20 5.22 5.35

30 5.92 4.98

40 6.85 4.49

The impurities that may be present in the ore such as iron, copper and cadmium

may result to sphalerite activation. The secondary copper minerals that may be present in

the ore could be soluble and during grinding, or in situ, could have released copper ions,

which activated sphalerite. This is common in copper–zinc ores that contain secondary

Dosage (g/t)



20 6.02 4.42 4.87 5.83

30 5.02 6.83 5.08 4.89

40 6.81 6.89 4.64 4.34

33

copper minerals with a covellite layer on the sphalerite surface. Figures 4.11 and 4.12,

show O-isopropyl ethyl thiocarbamate had the lowest recovery and zinc content in the

bulk concentrate. Galvanic interaction between sphalerite and chalcopyrite-pyrite

mineral mixtures causes dissolution of copper ions from chalcopyrite and thus activation

of sphalerite and pyrite by these copper ions (Ekmekci,et al.,2004). Soluble cations that

may be present could also activate pyrite minerals, increasing collector consumption and

often activate sphalerite. Ores with strongly activated sphalerite, either by lead cations or

by copper, which comes from the secondary copper minerals such as bornite, digenite

and covellite, may contain iron hydroxides, slimes and clay minerals (Bulatovic, 2007).

Cu++

activates zinc ore, the flotability of zinc ore is controlled by the solubility of

coexisting copper ore, then the flotability of zinc ore increases (Takeuchi, et.al, 1957).

Figure 4.10 Effect of collector dosage on zinc recovery in the flotation

concentrate.

34

Figure 4.11 Effect of collector dosage on zinc grade in the flotation

concentrate.

4.4 Recovery of Arsenic

The lowest grade of arsenic was obtained using NASCOL 446 at dosage 40 g/ton

as shown on figure 4.13. The lowest recovery was obtained using 40 g/ton NASCOL 201

as shown on figure 4.11. Arsenic recovery with respect to copper might indicate that

arsenic is in solid solution with copper. The analysis of Garay (2010) showed arsenic at

0.09% As in the run-of-mine ore, and however XRD did not reveal any presence of the

suspected mineral tennantite. Arsenopyrite was also detected on microscopy, but was not

confirmed on XRD analysis. Orpiment (As2S3) was also seen, but XRD only inferred its

presence (Garay, 2010).

Table 4.10, 4.11, 4.12, and 4.13 show the calculated recoveries and grades of

arsenic in copper the bulk concentrate. Arsenic is a penalty element for concentrates

35

which a concentrator wishes to eliminate. Arsenic in the feed was 0.21%, the highest

percentage, 0.36%, was found using NASCOL 446 at 40 g/t. This was below the typical

smelter penalty level of 0.5% As. The best result was obtained using NASCOL 446

which yielded 0.22% Arsenic with recovery of 96.7%.

Table 4.10 Recoveries of arsenic in the bulk concentrate utilizing NASCOL 446

and NASCOL 201.

Dosage (g/t)



20 97.34 85.49 81.10 91.10

30 95.08 96.69 91.84 88.87

40 96.68 96.69 72.20 83.28

Table 4.11 Mean recovery of arsenic in the bulk concentrate using NASCOL 446

and NASCOL 201.


20 91.4 86.1

30 95.9 90.4

40 96.7 77.7

36

Table 4.12 Calculated grades of arsenic in the bulk concentrate.

Dosage (g/t)



20 0.26 0.30 0.31 0.24

30 0.26 0.25 0.26 0.36

40 0.22 0.22 0.35 0.37

Table 4.13 Mean grades of arsenic in the bulk concentrate.


20 0.28 0.30

30 0.25 0.30

40 0.22 0.36

The Analysis of Variance (ANOVA) showed that the effects of the variation of

collectors were significant on the recovery of arsenic. The effects of the different

dosages of the two collectors were not significant on the recoveries of arsenic.

37

Figure 4.12 Effect of collector dosage on arsenic recovery in the

flotation concentrate.

0

20

40

60

80

100

120

0 1 2 3 4

% R

eco

very

Dosage

Dosage vs Recovery of As

NASCOL 201

NASCOL 446

38

Figure 4.13 Effect of collector dosage on arsenic grade in the flotation

concentrate.

CHAPTER V

CONCLUSIONS AND RECOMMENDATIONS

5.1 Conclusions

Based on the results obtained, the following conclusions were drawn:

1. The effect of the collector variation was significant on both recoveries and grades of

copper and zinc. Using 40 g/ton of NASCOL 446, the highest copper percentage in

the bulk concentrate was 14.95% Cu, with a recovery of 91.6 %; while for zinc; best

grade was 6.85% Zn, and a recovery of 96.18 %. For arsenic, lowest grade was at

0.22% As using NASCOL 221.

2. Variation of dosage for both collectors had no significant effects on copper, zinc, and

arsenic recovery in the bulk concentrate, however, the highest was observed at 40

kg/ton of NASCOL 446 with 91.6 % % for copper, 96.18% for zinc. The lowest

recovery, 96.7% for arsenic was at 40 g/ton kg/ton dosage using NASCOL 446.

3. Bulk feed was made of chalcopyrite, bornite, pyrite and sphalerite, and other

associated mineral.

40

5.2 Recommendations

1. The experiment should be further studied at higher collector dosages.

2. More replicate samples are recommended to increase efficiency of the study.

3. Adding variations on the conditioning and flotation time.

4. Conduct flotation using a mixture of reagents.

41

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Tadzh. SSR, 13(4): 26-30

Richardson, P. E.; Maust Jr., E. E. 1976. In Flotation; Fuerstenau, M. C. Ed. pp 78

(Wiley: New York)

Smedley PL & Kinniburgh DG, 2002. A review of the source, behavior and

distribution of arsenic in natural water. Appl Geochem 17:517-568

43

Takeuchi, T.,Gondo, K., 1957. Studies on the treatment of Complex Sulphide Ores

from Yonaihata Mine, Fukushima Prefecture. The Research Institute of Mineral

Dressing and Metallurgy. pp 451-452

Wills, B A, 1997. Mineral Processing Technology, 486 p (Butterworth Heinemann:

London)

Yoon, R.H., 1989. The Effect of Bubble Size on Fine Particle Flotation, Mineral

Processing and Extractive Metallurgy Review, 1(4):101-122

44

APPENDIX A

CALCULATIONS

A. Recovery Formula

Mass balances and material balances were set up to calculate for recoveries.

F = C + T A.1

fF = cC + tT A.2

% Recovery = cC x 100 A.3

fF

Assay Calculations for Copper

Solving for the weights of the Concentrates

Feed Wt.= 1000 g

C = F – T

C = 1000 g – 340.6

C = 659.4 g

Solving for Cu assay:

fF = cC + tT

c = (fF – tT)/ C

Solving for the % Recovery of Copper:

% Recovery = cC x 100

Ff

45

APPENDIX B

TABLES OF OBTAINED AND CALCULATED RESULTS

A. Specific gravity results.

Table B.1 Specific gravity results of ore.

SPECIFIC GRAVITY OF ORE RESULTS

Trial Sample Wt. 100ml Wt. of flask Gross Weight

Volume Water Sp. Gr.

No. Weight Flask + sample displaced in ml

1 32.00 56.67 88.67 180.94 7.73 4.140

2 32.49 56.67 89.16 180.45 8.71 3.730

3 32.61 56.67 89.28 181.23 8.05 4.051

46

A. Assay Results of Collectors

Table B.2. Copper analysis using NASCOL 201.

Copper Grade

NASCOL 201 Dosage

1 2 3

Trial 1

Bulk Con 8.91 10.57 7.50

Bulk Tails 9.99 5.50 12.97

Bulk Feed 9.28 9.28 9.28

Trial 2

Bulk Con 11.68 9.54 8.33

Bulk Tails 5.02 8.57 12.65

Bulk Feed 9.28 9.28 9.28

47

Table B.3. Copper analysis using NASCOL 446.

Copper Grade

NASCOL 446 Dosage

1 2 3

Trial 1

Bulk Con 13.61 11.18 15.04

Bulk Tails 0.89 2.95 1.35

Bulk Feed 9.28 9.28 9.28

Trial 2

Bulk Con 10.19 13.14 15.02

Bulk Tails 6.52 1.57 2.22

Bulk Feed 9.28 9.28 9.28

Table B.4. Zinc analysis using NASCOL 201.

Zinc Grade

NASCOL 201 Dosage

1 2 3

Trial 1

Bulk Con 4.87 5.08 4.64

Bulk Tails 2.41 0.94 2.77

Bulk Feed 4.03 4.03 4.03

Trial 2

Bulk Con 5.83 4.89 4.34

Bulk Tails 0.84 1.66 2.95

Bulk Feed 4.03 4.03 4.03

48

Table B.5 Zinc analysis using NASCOL 446.

Zinc Grade

NASCOL 446 Dosage

1 2 3

Trial 1

Bulk Con 6.02 5.02 6.81

Bulk Tails 0.17 0.73 0.21

Bulk Feed 4.03 4.03 4.03

Trial 2

Bulk Con 4.42 6.88 6.89

Bulk Tails 2.85 0.32 0.49

Bulk Feed 4.03 4.03 4.03

Table B.6 Arsenic analysis using NASCOL 201.

Arsenic Grade

NASCOL 201 Dosage

1 2 3

Trial 1

Bulk Con 0.26 0.26 0.22

Bulk Tails 0.12 0.07 0.18

Bulk Feed 0.21 0.21 0.21

Trial 2

Bulk Con 0.30 0.25 0.22

Bulk Tails 0.05 0.09 0.16

Bulk Feed 0.21 0.21 0.21

49

Table B.7. Arsenic analysis using NASCOL 446.

Arsenic Grade

NASCOL 446 Dosage

1 2 3

Trial 1

Bulk Con 0.31 0.26 0.35

Bulk Tails 0.02 0.04 0.02

Bulk Feed 0.21 0.21 0.21

Trial 2

Bulk Con 0.24 0.36 0.37

Bulk Tails 0.12 0.02 0.02

Bulk Feed 0.21 0.21 0.21

50

B. ICP Analysis Results Assay Results of Collectors

Table B.8 Results of chemical analysis 1st batch.

R0- ROM or bulk feed

R1- cooper concentrate, NASCOL 446 dosage 2 trial 2

R2- bulk tails, NASCOL 201 dosage 1 trial 1







R10- zinc concentrate, NASCOL 201 dosage 2 trial 2




51

Table B.9 Results of chemical analysis 2nd

batch.







R42-B- zinc concentrate, NASCOL 446 dosage 2 trial 2



52

Table B.10 Results of chemical analysis 3rd

batch.








R33 bulk tails, NASCOL 446 dosage 2 trial 2





53




Table B.11 Bulk analysis results of feed.

Element Assay, %

Ag 57.8 ppm

As 0.21

Bi 71 ppm

Cd 303.22 ppm

Cu 9.28

Fe 36.51

Hg 17.94 ppm

Pb 0.07

S 26.61

Sb 165 ppm

Zn 4.03

54

APPENDIX C

CALCULATED RECOVERIES

Table C.1 Calculated recovery of copper (%) in bulk concentrate.

% Recovery of Copper

Trial

No.


1 2 3 1 2 3

1 63.33 84.91 54.48 96.73 92.68 93.87

2 80.49 75.50 69.97 82.53 92.74 89.29

Mean 71.91 80.21 62.22 89.63 92.71 91.58

Table C.2 Calculated recovery of zinc (%) in bulk concentrate.

% Recovery of Zinc

Trial

No.


1 2 3 1 2 3

1 79.63 94.03 77.64 98.53 95.82 97.82

2 92.52 89.09 83.90 82.4 1 96.57 94.54

Mean 86.08 91.56 80.76 90.47 96.19 96.18

55

Table C.3 Calculated recovery of arsenic (%) in bulk concentrate.

% Recovery of Arsenic

Trial

No.


1 2 3 1 2 3

1 81.10 91.84 72.19 97.34 95.08 96.68

2 91.10 88.87 83.28 85.49 96.69 96.69

Mean 86.10 90.36 77.73 91.412 95.88 96.68

56

APPENDIX D

STATISTICAL ANALYSIS

Table D.1 Analysis of Variance for Percent Recovery taken at α = 0.05.

Source of

Variation

Sum of

Squares

Degrees of

Freedom

Mean

Square F-Com P-Value

Reagent 1,439.17 1.00 1,439.17 24.96 0.00

Dosage 314.47 2.00 157.24 2.73 0.08

Reagent*Dosage 336.23 2.00 168.11 2.92 0.07

Error 1,729.85 30.00 57.66

Total 276,893.61 36.00

57

Figure D.1 Interaction plot of the reagents and dosages.

Estimated Marginal Means of Percent Recovery

Dosage

Dosage 3Dosage 2Dosage 1

Estim

ate

d M

arg

ina

l M

ea

ns

100

90

80

70

Reagent Type

nascol 201

nascol 446

58

APPENDIX E

EQUIPMENTS, MATERIALS, PROCEDURES AND PRODUCTS

Figure E.1 Filtered cake from the bulk flotation process.

59

Figure E.2 Bulk flotation process.

CURRICULUM VITAE

Name: KHMER LEE P. LUGOD

Date of Birth: May 11, 1986

Place of Birth: Ozamiz City

Father‘s Name: Jesus Fuentes Lugod

Mother‘s Name: Teresita Ponce Lugod

Home Address: P-12 Catadman – Manabay, Ozamiz City, Mis. Occ, 7200

Mobile Number: +63-926-357-0095

Email Address: [email protected]

Educational Background

College: Mindanao State University – Iligan Institute of Technology

A. Bonifacio Ave., Iligan City

Bachelor of Science in Metallurgical Engineering 2010

Documents

BULK FLOTATION OF COMPLEX COPPER ORE FROM SIOCON, ZAMBOANGA SIBUGAY