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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
NASCOL 446 and NASCOL 201
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
NASCOL 446 and NASCOL 201
35
Table 4.11 Mean recovery of arsenic in the bulk concentrate using
NASCOL 446 and NASCOL 201
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
NASCOL 201 (O-isopropyl ethyl thiocarbamate) and NASCOL 446 (di-isobutyl
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
NASCOL 201 (O-isopropyl ethyl thiocarbamate) and NASCOL 446 (di-isobutyl
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)
NASCOL 446 NASCOL 201
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)
NASCOL 446 NASCOL 201
Trial 1 Trial 2 Trial 1 Trial 2
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.
Dosage (g/t) NASCOL 446 NASCOL 201
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)
NASCOL 446 NASCOL 201
Trial 1 Trial 2 Trial 1 Trial 2
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.
Dosage (g/t) NASCOL 446 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.
Dosage (g/t) NASCOL 446 NASCOL 201
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)
NASCOL 446 NASCOL 201
Trial 1 Trial 2 Trial 1 Trial 2
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)
NASCOL 446 NASCOL 201
Trial 1 Trial 2 Trial 1 Trial 2
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.
Dosage (g/t) NASCOL 446 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)
NASCOL 201 NASCOL 446
Trial 1 Trial 2 Trial 1 Trial 2
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.
Dosage (g/t) NASCOL 446 NASCOL 201
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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arsenic content in a complex galena concentrate by Acidithiobacillus, BioMed
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Acidithiobacillus ferrooxidans, <www.biomedcentral.com/1472-6750/4/22>
(Accessed :May 7, 2000) .
Mining Chemicals Handbook Revised Edition. 2002. Cytec, West Paterson, NJ
Monte, MB. Lins, FF., Dutra, AJ., Albuquerque, CR., Tondo, LA., 2002. The
Influence of the oxidation state of pyrite and arsenopyrite on the flotation of an
auriferous sulphide ore. Minerals Engineering, 15(12):1113-1120
Nagaraj, DR., Wang, SS. 1986. ―Monothiophosphinates as acid, neutral, or mildly
alkaline circuit sulfide collectors and process for using same ‖Dokl. Akad. Nauk
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
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from Yonaihata Mine, Fukushima Prefecture. The Research Institute of Mineral
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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
R3- cooper concentrate, NASCOL 201 dosage 2 trial 2
R4- bulk tails, NASCOL 201 dosage 3 trial 1
R5- bulk tails, NASCOL 201 dosage 1 trial 2
R6- bulk tails, NASCOL 201 dosage 2 trial 2
R7- bulk tails, NASCOL 201 dosage 3 trial 2
R9- cooper concentrate, NASCOL 201 dosage 1 trial 1
R10- zinc concentrate, NASCOL 201 dosage 2 trial 2
R16- zinc concentrate, NASCOL 201 dosage 1 trial 1
R17- zinc concentrate, NASCOL 446 dosage 1 trial 2
R18- cooper concentrate, NASCOL 201 dosage 3 trial 1
51
Table B.9 Results of chemical analysis 2nd
batch.
R11- bulk tails, NASCOL 201 dosage 2 trial 1
R12- cooper concentrate, NASCOL 446 dosage 3 trial 2
R13- zinc concentrate, NASCOL 201 dosage 3 trial 1
R14- zinc concentrate, NASCOL 201 dosage 3 trial 2
R41- zinc concentrate, NASCOL 446 dosage 1 trial 1
R42- zinc concentrate, NASCOL 446 dosage 2 trial 1
R42-B- zinc concentrate, NASCOL 446 dosage 2 trial 2
R43- zinc concentrate, NASCOL 446 dosage 3 trial 1
R44- bulk tails, NASCOL 446 dosage 1 trial 2
52
Table B.10 Results of chemical analysis 3rd
batch.
R19- zinc concentrate, NASCOL 201 dosage 2 trial 1
R20- cooper concentrate, NASCOL 201 dosage 3 trial 2
R21- cooper concentrate, NASCOL 201 dosage 2 trial 2
R29- bulk tails, NASCOL 446 dosage 1 trial 1
R30- bulk tails, NASCOL 446 dosage 2 trial 1
R31- bulk tails, NASCOL 446 dosage 3 trial 1
R32- cooper concentrate, NASCOL 201 dosage 1 trial 2
R33 bulk tails, NASCOL 446 dosage 2 trial 2
R34- bulk tails, NASCOL 446 dosage 3 trial 2
R35- cooper concentrate, NASCOL 446 dosage 1 trial 1
R36- cooper concentrate, NASCOL 446 dosage 2 trial 1
R37- cooper concentrate, NASCOL 446 dosage 3 trial 1
53
R38- cooper concentrate, NASCOL 446 dosage 1 trial 2
R39- zinc concentrate, NASCOL 446 dosage 3 trial 2
R40- zinc concentrate, NASCOL 201 dosage 1 trial 2
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.
NASCOL 201 NASCOL 446
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.
NASCOL 201 NASCOL 446
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.
NASCOL 201 NASCOL 446
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