Project_NYENDWA_(10th_ed)[1]

ii
THE UNIVERSITY OF ZAMBIA
SCHOOL OF MINES
DEPARTMENT OF METALLURGY AND MINERAL PROCESSING
FINAL YEAR PROJECT
MET 5494
INVESTIGATION OF THE SYNERGIC EFFECTS OF SODIUM ETHYL
XANTHATE (SEX) AND AMONIUM DITHIOPHOSPHATE (ADTP) ON NICKEL
CONCENTRATE GRADES AND RECOVERIES AT MUNALI CONCENTRATOR.
By
NYENDWA JANNY
COMPUTER #: 11016281

i
DECLARATION
I Nyendwa Janny do declare that this project entitled “INVESTIGATION OF THE
SYNERGIC EFFECTS OF SODIUM ETHYL XANTHATE (SEX) AND AMONIUM
DITHIOPHOSPHATE (ADTP) ON NICKEL CONCENTRATE GRADES AND
RECOVERIES AT MUNALI CONCENTRATOR” was written by me and no other part of it
has been written without citation. It is therefore, to my best knowledge that this project has not
been previously presented at this university or any other learning institution for academic
purposes.

ii
APPROVAL
INTERNAL EXAMINER
…………………………………………………….………………………………
DR. L. WITIKA
HEAD OF DEPARTMENT.
………………………………………………………………………………….
DR. E. SIAME
EXTERNAL EXAMINER
……………………………..………………………………….……………….
PROF. A. MAINZA

iii
ACKNOWLEDGEMENT
I firstly thank Jehovah God for granting me an opportunity to be one of the few students to be
trained by the greatest minds of the land to attain my Bachelor’s Degree in Mineral Sciences
at the University of Zambia. Special thanks go to my School Project Supervisor, one of the few
talented intellectuals of our day Dr. L. Witika, the entire School Of Mines staff in the
Department of Metallurgy and Mineral Processing at the University of Zambia for the guidance
rendered to me and the necessary information and constructive criticism I received for my
thesis write-up and presentations. This report would not have come into existence without the
assistance and enthusiasm of my Plant Project Supervisor, Henry Zulu at Munali Nickel Mine,
who was there for me throughout the time the project was being executed. To the general
manager at Munali Nickel Mine Mr Matthew Banda I appreciate the humane opportunity you
rendered to me, to everyone who was there academically you made this possible and this
achievement goes to you.
My sincere gratitude to my parents, Mr and Mrs Nyendwa, for making me who I am today.
Thank you mom and dad, I hope to make you proud. And to my beloved sisters and brothers,
you are awesome. I love you so much. To my classmates for the support during my stay at
campus, thank you very much you are more than family to me.

iv
ABSTRACT
Munali Nickel Mine had average grades and recoveries of 10-13% and 65-78% respectively
when xanthates was used as a sole collector when the plant was in operation. In 2011 Mine
ordered over 100 tonnes of ammonium dithiophosphate (ADTP) to be used as a collecting
reagent despite having Sodium Ethyl Xanthate (SEX) which was being used in the Plant
during operation. Unfortunately, ADPT yielded poor grades and recoveries as a collecting
reagent and efforts to sell the reagent locally and within the Sub-region proved futile. The
aim of the investigation was to assess the synergic effect of the two collectors on the grades
and Recoveries on Munali Ore. One of the major challenges faced by flotation of these ores
is the ever varying grade and mineralogy hence synergies of xanthates and dithiophosphates
have shown to enhance recoveries when combined in the right proportion.The two collectors
were combined in the ratio of SEX:ADTP in the ratio 1:0, 1:1, 2:1 and 3:1 at three different
dosages of 100gpt, 200gpt and 150gpt. Each sample had a mass of 1kg and was subjected to
the same conditions such as collecting time and reagents combination such as frother, lime
and guar. Pure collectors showed lower recoveries than when combined with xanthates. The
grades however showed a gradual decrease with increase in dosages for varying ratios of the
synergies. The increase in recoveries was linked to the synergies and dosages while the low
grades was probably due to the unmonitored operating pH and lost in the selectivity of the
reagents. This project was recommended to see if synergies of the two named collectors
would be enhance recoveries and grades. Synergies had a positive effect on the recoveries
and little impact on the grades, pH on the other side could have contributed to the low grades
obtained.
Key word: synergy, ammonium dithiophosphates, Sodium ethyl xanthates, Munali Mine,
dosage, recoveries, grades, collector.

v
TABLE OF CONTENTS:
i. DECLARATION………………………………………………………………....ii
ii. APPROVAL……………………………………………………………………...iii
iii. ACKNOWLEDGEMENT………………………………………………………..iv
iv. ABSTRACT………………………………………………………………………v
v. TABLE OF CONTENT………………………………………………………….vi
vi. LIST OF FIGURES……………………………………………………………...vii
1. INTRODUCTION.........................................................................................................1
1.2. Plant Flowsheet…………………………………………………………………...4
1.3. Problem Statement………………………………………………………………..4
2. LITERATURE REVIEW..............................................................................................5
2.1. Synergies of SEX and ADTP……………………………………………………..5
2.2. Grades and Recoveries…………………………………………………………....6
2.3. Hydrophobicity…………………………………………………………………...7
2.4. Particle/Bubble Contact…………………………………………………………..8
2.5. Froth Layer……………………………………………………………………….9
2.6. Reagents………………………………………………………………………....10
2.7. Collectors………………………………………………………………………..10
2.8. Chemisorption…………...………………………………………………………12
2.9. Physisorption……………..……………………………………………………...12
3. METHODOLOGY......................................................................................................13
3.1.Sample Preparations……………………………………………………………..13
3.2. Chemical Reagents……………………………………………………………...14
3.3.Flotation Tests……………………………………………………………………15
4. RESULTS AND DISCUSSIONS...............................................................................17
5. CONCLUSION AND RECOMMENDATIONS.......................................................21
6. APPENDICES............................................................................................................22
7. BIBLIOGRAPHY.......................................................................................................26

vi
LIST OF FIGURES
Figure 1.1: Flow sheet of Munali Nickel Mine concentrator…………...………….......……...2
Figure 2.1:Typical form of Grades/Recoveries Curves for froth flotation…………………….6
Figure 2.2:Attachment of air bubbles to hydrophobic particles……...………………………..7
Figure 2.3: Principle of three phase contact angle……………...……………………………..8
Figure 2.4: Schematic of a conventional flotation cell………..…………….………………..11
Figure 3.1: Classification of collectors……………………………………………………….11
Figure 3.2: The grindability time at 80% passing -75microns…………,,,……...…………...13
Figure 4.1:Laboratory flow sheet………………………………………….............................16
Figure 4.2:Graph of recovery/grade at 100gpt……....……………………………………….17
Figure 4.3:Graph of recovery/grade at 120gpt……………………………………………….17
Figure 4.4 Graph of recovery/grade at 150gpt……………………………………………….18
Figure 4.5 Synergic Effect of Recoveries……………………………………………………19
Figure 4.6 Synergic effect of Grades……………………………………………...................19
Figure 4.7 General trend in synergic effect on recoveries…………………………………...20

vii
.
LIST OF TABLES
Table 3.1: Synergic dosages (gpt) of collectors………………………..……………….…14
Table 3.2: Synergic laboratory dosages (gpKg) of collectors…………………………….14
Table A.1: Grades and recoveries for synergies at dosage of 100gpt……………..............22
Table A.2:Grades and recoveries for synergy at dosage of 120gpt………………………23
Table A.3 Grades and recoveries for synergy at 150gpt………..………………………....24
Table A.4: Grind of Mesh of Time…………………………….………………………….24

2
CHAPTER ONE.
1. INTRODUCTION
Mabiza Resources Limited (MRL), a 100% subsidiary of Consolidated Nickel Mine (CNM)
limited at the London Exchange listed company, operates Munali Nickel Mine which is located
in the Southern province of Zambia approximately 85km south of Lusaka.
The Mine has a processing plant with a capacity to treat 900,000 tons per annum of ore and a
maximum of 1.2million tons per annum may be achievable when minor modifications are made
to the facility. The initial plant was commissioned in mid-2008, but operations were suspended
and the Plant placed on care and maintenance in March to December 2009 due to the global
economic crunch. The operations resumed in January 2010 after restructuring of the company.
However, in November 2011, the operations were again placed under care and maintenance to
facilitate a restructuring process till date.
Preliminarily results on Munali ore conducted using Ammonium Dithiolphosphate showed
poor grades and recoveries, and from this performance background, various metallurgical test
works were to be conducted to assess the effects of using a combination of various collecting
reagents. It is assumed that any combination of the two collectors should yield better results
than the above obtained. Some of the challenges that were experienced included:
 It was always assumed that the ore body is homogenous while it was clearly observed
that there were changes in grade on both strike and depth.
 The Mine had a shortage of equipment needed for a successful experiment such the
unavailability of a pH meter for pH monitoring at the time of research.
The test works previously undertaken clearly illustrate that with improvements, the recovery
and grades of the concentration can also be improved and achieved >70% recovery. On the
other side Sodium Ethyl Xanthates as a sole collector was successful and yielded an average
grades and recoveries at 10-13% and 65-78% respectively. Combining the two collectors would
therefore enhance metallurgical recoveries and grades when used in the right proportion.
The principal sulphide mineral in nickel ores is Pentlandite (Ni,Fe)9S8 with talc or talcose
(Mg3Si4O10(OH)2) type minerals as the main gangue minerals. Other forms of minerals include
pyrite (FeS2) or pyrrhotite and chalcopyrite (CuFeS2).

3
1.2 Plant Flow Sheet
Figure 1.1 Flow sheet of Munali Nickel Mine concentrator (Munali, 2011).

4
On the other hand one of the main problem encountered is selectivity of Pyrrhotite and also in
some cases chalcopyrite during flotation which affects the grades of the concentrate. A clean
and satisfactory separation of Pentlandite from pyrrhotite by flotation is difficult in practice
since pyrrhotite typically contains inter-grown inclusions of pentlandite as well as nickel in
solid solution. In fact pyrrhotite often contains 0.5-1% Ni that cannot be separated by physical
methods. The common occurrence of both monoclinic (magnetic) and the hexagonal
(nonmagnetic) forms of pyrrhotite in association with pentlandite also poses problems (Rao,
2000). Another type of alteration which adversely effects flotation recoveries is that tochilinite
has flotation properties similar to that of pyrrhotite. As a consequence it either reports to the
flotation tailings, thereby decreasing the nickel recovery, or, if it is effectively floated, a
significant amount of pyrrhotite accompanies it, diluting the nickel grade in the concentrate.
There is a distinct difference in silicate mineralogy between types of host rock, which have
their own problems with respect to rejection of gangue by flotation. Talc and other naturally
hydrophobic magnesia-bearing minerals have a tendency to float with sulphides, resulting in a
concentrate exceptionally high
in magnesia. The presence of magnesia causes viscosity problems in the slag during smelting.
Magnesia also promotes conditions favourable to hetero-coagulation of minerals, especially
fine sulphides with coarse gangue minerals, thus leading to nickel loss (Heiskanen et al, 1991).
Nickel Sulphide minerals such as pentlandite can, in general, be separated from their gangue
by Flotation using a thiol group of collectors like xanthates and alkyl dithiophosphates in the
presence of variety of activators, depressants and dispersants. Since nickel contain other
sulphides such as pyrrhotite, pentlandite and chalcopyrite, the enrichment of nickel is generally
carried out by two methods (Wills, 2006):
1. Production of bulk concentrate containing all sulphides together as smelter feed;
2. Production of bulk chalcopyrite pentlandite concentrate by preferentially depressing
the pyrrhotite followed by selective flotation of chalcopyrite and pentlandite.
Although bulk flotation of all sulphides is relatively simple the presence of pyrrhotite,
most of the Sulphur contained in the Flotation concentrate is emitted from pyrrhotite, the
rejection of pyrrhotite is important. Pyrrhotite is known to float poorly in alkaline
media; therefore the general practice is to selectively float pentlandite from pyrrhotite, by
maintaininga highly alkaline pH with lime and guar as depressant and using thiols like

5
xanthates and dithiophosphates as collectors. Although it is possible to reject significant
amounts of pyrrhotite in this way, the concomitant pentlandite losses into flotation tailings are
highly unsatisfactory (Wills, 2006).
1.3 Objective
 To investigate the synergic effect of Sodium Ethyl Xanthate (SEX) and Ammonium
Dithiophosphate (ADTP) on the Nickel recoveries and grades on at the Munali
concentrator.
1.3.1 Specific objectives
 To investigate the effects of dosages on grades and recoveries.
 To determine the grind of mesh of the ore.
1.4 Problem Statement
Concentrator department ordered more than 100 tonnes of ammonium Dithiophosphate
(ADTP) without any prior test work on the plant or laboratory to ascertain suitability of ADTP
as a flotation reagent for Munali ore. ADTP was found ineffective as a flotation reagent for the
Munali ore as it’s use resulted in poor grades and recoveries in flotation process.

6
CHAPTER TWO.
2.0 LITERATURE REVIEW.
2.1 SYNERGIES OF SODIUM ETHYL XANTHATES AND AMMONIUM
DITHIOPHOSPHATE.
It is well documented that ADTP when used with other collectors in the correct proportion or
synergy enhances recovery and improves grades (CYTEC, 2010). ADTP gives an improved
selectivity in the flotation of sulphides especially when used in synergy with Xanthates. ADTP
is known to work very well in low pH conditions less than 9.2. Recent trends in flotation
practice have shown that, in many cases, a combination of two or more different collectors
provides better flotation responses than when a single collector is used. This is not surprising
when one considers that, even in such a simple case as copper ores, there may be a variety of
copper minerals present (eg. chalcopyrite, chalcocite,covellite, bornite, native copper,
tetrahedrite, and oxidized or tarnished copper minerals) each of which responds differently to
different collector chemistries. Most minerals exhibit an optimum pH range for a given
collector. While some minerals can often be floated at the natural pH, in most cases the pH has
to be adjusted for maximum recovery
and selectivity. The most commonly used reagents for alkaline circuits are lime and soda ash.
For acid circuit flotation, the most commonly used reagent is sulphuric acid. These three
modifiers are generally the most cost effective. Other pH modifiers are also used occasionally
when difficult separations are involved (Wills, 2006).
Many collectors and frothers are in use in the flotation treatment of sulphide and metallic ores
containing such metals as copper, nickel, cobalt, molybdenum, iron, precious metals (including
platinum-group metals) and such penalty elements as arsenic, antimony and bismuth. The
principal factors affecting the choice of collectors are the mineral forms (sulphide, oxidized
and/or metallic species) and their associations with each other and the gangue minerals.
Dithiophosphates have been used for decades and most commonly used collector combinations
with xanthate ( Wills 2006). The xanthates are the most important for sulphide mineral
flotation. They are prepared by reacting an alkali hydroxide, an alcohol and carbon disulphide:
ROH + CS2 + KOH = RO.CS.SK + H2O.

7
Where R is the hydrocarbon group and contains normally one to six carbon atoms, the most
widely used xanthates being ethyl, isopropyl, isobutyl, amyl, and hexyl (Wills 2006).
2.2 Grades and Recoveries
While each of these single calculated values for recoveries and grades are useful for comparing
flotation performance for different conditions, it is most useful to consider both the grade and
the recovery simultaneously, using a “Grade/Recovery Curve”. This is a graph of the recovery
of the valuable metal achieved versus the product grade at that recovery, and is particularly
useful for comparing separations where both the grade and the recovery are varying. A set of
grade/recovery curves is shown in figure 2.1 below. If 100% of the feed is recovered to the
product, then the product will obviously have the same composition as the feed, and so the
curve starts at the feed composition with 100% recovery. Similarly, if the purest mineral grain
that contains the metal of interest is removed, this will be the maximum grade that can be
produced by a physical separation, and so the 0% recovery end of the curve terminates at an
assay less than or equal to the assay of the purest grains available in the ore. In the graphs
shown in Figure 2.1 points that are higher and to the right show better performance than points
that are lower and to the left.

8
Figure. 2.1 Typical form of Grades/Recoveries Curves for froth flotation (Klasen, 1963).
2.3 Hydrophobicity/hydrophilicity.
The basis of froth flotation is the difference in wettability of different minerals. Particles range
from those that are easily wettable by water (hydrophilic) to those that are water-repellent
(hydrophobic). If a mixture of hydrophobic and hydrophilic particles are suspended in water,
and air is bubbled through the suspension, then the hydrophobic particles will tend to attach to
the air bubbles and float to the surface, as shown in Figure 2.2. The froth layer that forms on
the surface will then be heavily loaded with they hydrophobic mineral, and can be removed as
a separated product. The hydrophilic particles will have much less tendency to attach to air
bubbles, and so it will remain in suspension and be flushed away (Whelan and Brown, 1956).
Particles can either be naturally hydrophobic, or the hydrophobicity can be induced by
chemical treatments. Naturally hydrophobic materials include hydrocarbons, and non-polar
solids such as elemental sulphur. Chemical treatments to render a surface hydrophobic are
essentially methods for selectively coating a particle surface with a monolayer of non-polar oil
(Kawatra and Eisele, 1992).

9
Figure 2.2 Selective attachment of air bubbles to hydrophobic particles (Klasen, 1963).
The attachment of the bubbles to the surface is determined by the interfacial energies between
the solid, liquid, and gas phases. This is determined by the Young/Dupre Equation,
γlvcosθ = (γsv – γsl) where γlv is the surface energy of the liquid/vapour interface, γsv is the
surface energy of the solid/vapour interface, γsl is the surface energy of the solid/liquid
interface, and θ is the “contact angle”, the angle formed at the junction between vapour, solid,
and liquid phases, as shown in figure 2.3 below. If the contact angle is very small, then the
bubble does not attach to the surface, while a very large contact angle results in very strong
bubble attachment. A contact angle near 90° is sufficient for effective froth flotation in most
cases (Wills, 2006).
Figure 2.3: Contact angle between and air bubble and a solid surface immersed in liquid
(Wills’ 2006).
2.4 Particle/Bubble Contact.

10
Once the particles are rendered hydrophobic, they must be brought in contact with gas bubbles
so that the bubbles can attach to the surface. If the bubbles and surfaces never come in contact,
then no flotation can occur. Contact between particles and bubbles can be accomplished in a
flotation cell such as the one shown schematically in Figure 2.4.
Figure 2.4: Simplified schematic of a conventional flotation cell. (Klasen, 1963)
2.5 Collection in the Froth Layer.
Once a particle and bubble have come in contact, the bubble must be large enough for its
buoyancy to lift the particle to the surface. This is obviously easier if the particles are low
density (as is the case for coal) than if they are high-density (such as lead sulphide). The particle
and bubble must remain attached while they move up into the froth layer at the top of the cell.
The froth layer must persist long enough to either flow over the discharge lip of the cell by
Stator Rotor Slurry Froth Air Froth Overflow gravity, or to be removed by mechanical froth
scrapers. If the froth is insufficiently stable, the bubbles will break and drop the hydrophobic
particles back into the slurry prematurely. However, the froth should not be so stable as to
become persistent foam, as a foam is difficult to convey and pump through the plant. The
surface area of the bubbles in the froth is also important. Since particles are carried into the
froth by attachment to bubble surfaces, increasing amounts of bubble surface area allows a

11
more rapid flotation rate of particles. At the same time, increased surface area also carries more
water into the froth as the film between the bubbles. Since fine particles that are not attached
to air bubbles will be unselectively carried into the froth along with the water (entrainment),
excessive amounts of water in the froth can result in significant contamination of the product
with gangue minerals (Boutin and Wheeler, 1967).
2.6 Reagents
The properties of raw mineral mixtures suspended in plain water are rarely suitable for froth
flotation. Chemicals are needed both to control the relative hydrophobicity of the particles, and
to maintain the proper froth characteristics. There are therefore many different reagents
involved in the froth flotation process, with the selection of reagents depending on the specific
mineral mixtures being treated (Boutin and Wheeler, 1967).
2.7 Collectors.
Collectors are reagents that are used to selectively adsorb onto the surfaces of particles. They
form a monolayer on the particle surface that essentially makes a thin film of non-polar
hydrophobic hydrocarbons. The collectors greatly increase the contact angle so that bubbles
will adhere to the surface. Selection of the correct collector is critical for an effective separation
by froth flotation. Collectors can be generally classed depending on their ionic charge: they can
be non-ionic, anionic, or cationic, as shown in Figure 2.4.

12
Figure 2.4: Classification of collectors (Wills, 2006).
The non-ionic collectors are simple hydrocarbon oils, while the anionic and cationic collectors
consist of a polar part that selectively attaches to the mineral surfaces, and a non-polar part that
projects out into the solution and makes the surface hydrophobic. Collectors can either
chemically bond to the mineral surface (chemisorption), or be held on the surface by physical
forces or physical adsorption (Wills, 2006).
Flotation is a physio-chemical process which involves both physical and chemical reactions.
To achieve this process, chemical factors include the interfacial chemical reactions which
results in the formation of certain surface chemical species and the physical processes which
leads to the formation of the three contact phases that exist in flotation, viz. Solid-liquid, gas-
liquid and solid-gas phases. In order to understand the chemistry involved in this process,
important roles of various flotation reagents – such as collectors, depressants, frothers,
activators, and pH modifiers – used in the process, have to be understood, as well as water
chemistry and the surface chemistry of the minerals involved. Physical properties such as ,
physical-mechanical and operational factors comprise equipment components such as cell
design, hydrodynamics, cell configuration, aeration rates residence times, feed rates,
mineralogy, particle size distribution, pulp density etc., need to be closely monitored. Thus
flotation is an extremely complex process involving many scientific, technological and
engineering phenomena .In most flotation systems, physical and chemical factors are not
independent i.e. there are significant interactions among the many variables. In theory, when

13
all physical factors are optimized, a change in a chemical factor should clearly record a
measurable change in flotation efficiency (either recovery or grade or both), and vice versa. In
practice, however, this may not be immediately obvious because of certain operational
restrictions, and metallurgists have to revert to statistical tools to demonstrate significant
changes (Glembotskii, Klassen and Plaksin, 1972).
2.8 Chemisorption
In chemisorption, ions or molecules from solution undergo a chemical reaction with the
surface, becoming irreversibly bonded. This permanently changes the nature of the surface.
Chemisorption of collectors is highly selective, as the chemical bonds are specific to particular
atoms (Glembotskii, Klassen, and Plaksin, 1972).
2.9 Physisorption
In physisorption, ions or molecules from solution become reversibly associated with the
surface, attaching due to electrostatic attraction or van der Waals bonding. The physisorbed
substances can be desorbed from the surface if conditions such as pH or composition of the
solution changes. Physisorption is much less selective than chemisorption, as collectors will
adsorb on any surface that has the correct electrical charge or degree of natural hydrophobicity
(Klassen, and Mokrousov, 1963).

14
CHAPTER THREE
3.0 MATERIALS AND METHODOLOGY
3.1 Sample Preparation
The ore was received from underground before being subjected to crushing. This was then
crushed in a laboratory crusher until a product of minus 2.5mm was obtained.
The -2.5mm sample was subjected to grind tests to establish the 80% passing 75µm as per
Munali floatation feed standard (Zulu, 2011). The bulk ore sample was split into 1 kg
samples manually by weighing on a balance. The ore was milled in a Bond Index rod mill
to establish the milling curves in order to simulate the in-plant particle size distribution.
The 1 kg of each crushed samples were milled at different times in 5 minutes intervals i.e.
5, 10, 15, 20, to 35 minutes and subjected to wet screen analyses to determine 80% passing
75µm and results are shown Figure 3.1 below.
s Figure 3.1: Graph showing the grindability time at 80% passing 75µm at 25minutes.
At 80%-75µm, the size is suitable for flotation as has been indicated by previous test work.
0
10
20
30
40
50
60
70
80
90
100
0 5 10 15 20 25 30 35
%PASSING75µ
TIME (MIN)

15
3.2 CHEMICAL REAGENTS
The standard reagents for the process plant were applied in this research. These reagents
were collected from reagents stocks as follows:
 Sodium Ethyl Xanthate (SEX) used as a standard collector for nickel and other
valuable minerals.
 Ammonium Dithiophosphate (ADTP) was used as an auxiliary collector.
 Guar gum depressant used as a standard depressant mainly for MgO and other
gangue minerals.
 Lime for pH adjustment.
 Frother used for froth stabilization was ore prep F-549.
Table 3.1: The actual dosages and stages of dosing at industrial level.
Reagent Dosage(g/t) Addition stage
SEX 150 Rougher 1
Guar gum 500 Rougher 1
Frother As required to maintain Rougher 1
Lime 1000 Rougher 1
Collector dosages was varied at 100gpt, 120gpt and 150gpt according to Plant parameters.
Since the project was done in a laboratory there was need to reduce the dosages to laboratory
level i.e. 100gpt was equivalent to 0.1g/Kg of the ore sample to be floated. Since this was done
for different synergies in varying ratios the table below illustrates different ratios and input for
the synergy.

16
Table 3.2: Illustrating laboratory equivalency for plant parameters for synergies of SEX
and ADTP at 100gpt.
Ratio SEX (mg) ADTP (mg) Total (mg)
1:0 100 0 100
1:1 50 50 100
2:1 67 33 100
3:1 75 25 100
3.3 FLOTATION TESTS
Using the standard laboratory flow sheet Figure 3.2, batch flotation tests was carried out
using a laboratory flotation cell with volumetric capacity of 2.5 litres. This was done by
adding the slurry from the laboratory rod mill to the flotation cell. One kilogramme of the
sample ore was milled at the determined 25 minutes milled time to get the slurry. The slurry
density was adjusted to 38% solids in line with plant requirement. The air flowrate was at 3
l/min and a froth build-up of 120 seconds was allowed for before the first concentrate was
collected.
Four different rougher concentrates were collected at three (3) minutes intervals leaving a final
tailing at the end as shown in Figure 3.2. A synergy of collectors, guar depressant, lime and
frother were added to the slurry during conditioning.
The amount of metal (Ni) recovered was calculated using equation (1) which is based on
assays alone of the feed (f), tailings (t) and concentrate (c).
𝐹
𝐶
=
𝑐(𝑐−𝑓)
𝑓(𝑓−𝑡)
𝑥100……………………………………………………………………………...(1)

17
Figure:3.2 showing Laboratory flow sheet

18
CHAPTER FOUR.
4.0 RESULTS AND DISCUSSIONS.
Figure 4.1: Results of grades against recoveries for the variant synergic ratio of SEX and
ADTP.
0
20
40
60
80
100
0 0.5 1 1.5 2 2.5 3 3.5 4
Recoveries(%)
Grades (%)
RECOVERIES/GRADES
Series1 Series2 Series4

19
Figure 4.4.1: Synergic effect of SEX and ADTP at dosage of 100gpt on Recoveries.
0
10
20
30
40
50
60
70
80
90
Ratio 1:0 Ratio 1:1 Ratio 2:1 Ratio 3:1
Recoveries(%)
Synergic Effect of SEX:ADTP
0
10
20
30
40
50
60
70
80
90
100
Ratio 1:0 Ratio1:1 Ratio 2:1 Ratio 3:1
Recovery(%)
Synergic effect of SEX:ADTP

20
Figure 4.5.1: Effects of Grades of SEX and ADTP at dosage of 100gpt.
0
10
20
30
40
50
60
70
80
90
100
Recoveries(%)
Synergic effect of SEX:ADTP
150gpt
3.4
3.45
3.5
3.55
3.6
3.65
3.7
3.75
Grades(%)
Effect of synergies of SEX:ADTPon Grades
100gpt

21
Figure 4.5.2: Effects of Grades of SEX and ADTP at dosage of 120gpt.
Figure 4.5.3: Effects of Grades of SEX and ADTP at dosage of 150gpt
3.2
3.25
3.3
3.35
3.4
3.45
3.5
3.55
3.6
3.65
Grades
Synergic Effect of SEX:ADTP
3.15
3.2
3.25
3.3
3.35
3.4
3.45
3.5
3.55
Grades(%)
Synergic Effects of SEX:ADTPat 150gpt

22
Figure 4.6: Average synergic Effects of SEX:ADTP on Recoveries.
Figure 4.7. General Synergic Effect of SEX: ADTP on Grades.
4.1 DISCUSSIONS
Increasing the dosages with respect to synergies showed an increase in metal recovery but
showed a general decrease in the grades of the concentrate. At the ratio of 1:0 i.e. SEX:ADTP
results showed a lower recoveries compared to recoveries when synergies were made. The
highest recovery was achieved at 150gpt as 93.48% with an average recovery and grade of
64.7
71.4
87.4
0
10
20
30
40
50
60
70
80
90
100
100gpt 120gpt 150gpt
Recoveries
Dosages
3.2
3.25
3.3
3.35
3.4
3.45
3.5
3.55
3.6
3.65
100gpt 120gpt 150gpt
Grades
Dosages

23
87.47% and 3.36%. The average recovery and grade at 120gppt was 71.4% and 3.46%
respectively. The lowest average recovery and highest grade was at 100gpt i.e. average
recovery of 64.7% and 3.62%.
The plant had an average grades of 10-13% and recoveries of 65-78%. After combining two
collectors i.e. SEX and ADTP, there was a general increase in the average recoveries obtained.
The Plant average grades however were not achieved in this research. Looking at Figure 6.5
there was a variation in the grades obtained despite the increase in the dosage of the synergies
of SEX and ADTP. The variation could have been due to the pyrite reporting to the
concentration. This happens when the operating pH condition is below the optimum operating
pH of pentlandite i.e. pH of 9.4. Since the addition of lime was constant (usually added to
moderate pH) it could be pyrite was not depressed enough and interfered with the grades of the
concentrate. The grades are also affected by the condition of the reagents, if xanthantes have
absorbed moisture it affects the selectivity properties of the reagents (Wills, 2006).

24
CHAPTER FIVE
5.0 CONCLUSIONS AND RECOMMENDATIONS
An increase in the synergic dosage resulted in an increase in recoveries as seen from results.
There was a general increase in recoveries because synergies enhances collecting and selecting
properties for ores of complex mineralogy. pH has the major contribution in controlling the
flotation of the base metals (Ni, Cu and Co) from the nickel-copper sulphide ore. Depressant
dosage also plays a significant effect in controlling the flotability of nickel but since the dosage
for the depressant was standard to plant parameters, the unknown pH could have contributed
to poor grades.
The maximum levels of parameters for recovery of nickel was achieved at synergic dosage of
150gpt at the ratio 3:1. And the optimum levels for nickel recovery was at 120gpt at the ratio
of 1:1.
Optimal levels of parameter for concentrate grade of nickel was achieved at the synergic
dosage of 100gpt at the ratio 3:1.
The plant had average grades of 10-13% of nickel and average metal recoveries of 65-78%.
Hence using a ratio of 1:0 for SEX: ADTP a similar result was expected. Though recoveries of
60-78 were achieved, the grades were too low to satisfy the findings and consolidate the
research. On the other hand there was no general trend in the grades for the three dosages
insinuating that dosage on synergies had little effect on grades.
In conclusion the unmonitored pH could have led to poor grades as operating in acidic medium
causes pyrite, charcopyrite and other minerals to report to the concentrate and hence yielding
poor grades. Though the addition of lime was arbitrary while acidic conditions may vary this
could have brought the variation in grades. The recommended pH for nickel recovery at munali
is 9.4 from past experiments and any slight change below the recommended would yield poor
grades.
The chemicals have stayed more than five years on site and they could have lost some chemical
properties such as selectivity. Even though the recoveries were satisfied, the grades weren’t
satisfactory enough to ascertain the findings hence the project should be revisited but strictly
monitoring the operating pH by use of a pH meter and working in the recommended medium
in order to ascertain the viability of the synergy of these collectors on Munali Ore. The tailings
should be investigated for the mineralogy leading to the poor grades.

25
APPENDIX: A
The resulting samples from the flotation cell were dried, weighed and assayed. Assayed results
are tabulated below.
Table A.1 Showing grades and recoveries for the two synergies of SEX and ADTP at 100gpt.
Ratio Item Wt (g) Assay (%) Wt (metal g) Recovery
(%)
Feed 739 2.56 18.84 100
1:0 Concentrate 200 3.5775 7.16 42.14
Tailings 539 2.12 5.39 57.9
Feed 754 2.56 7.54 100
1:1 Concentrate 211 3.515 7.42 80.60
Tailings 543 1.20 6.52 19.4
Feed 687 2.56 17.59 100
2:1 Concentrate 190 3.612 6.86 62.67
Tailings 497 1.72 8.55 37.33
Feed 734 2.56 18.79 100
3:1 Concentrate 228 3.675 8.38 77.77
Tailings 506 1.24 6.27 20.23

26
Table A.2 Showing grades and recoveries for synergies of SEX and ADTP at120gpt
Ratio Item Wt (g) Assay (%) Wt.(metal g) Recoveries
(%)
Feed 982 2.56 25.14 100
1:0 Concentrate 198 3.55 7.03 61.55
Tailings 784 1.77 13.88 38.45
Feed 974 2.56 24.93 100
1:1 Concentrate 373 3.59 13.39 68.37
Tailings 601 1.58 9.50 31.63
Feed 995 2.56 25.47 100
2:1 Concentrate 285 3.345 9.53 80.74
Tailings 710 1.40 9.94 22.16
Feed 930 2.56 23.8 100
3:1 Concentrate 263 3.4775 9.14 91.81
Tailings 667 0.65 4.34 8.19
A.1.1 Sample Calculation:
Data
c = 3.55%, f = 2.56%, t = 1.77 and plugging in the equation above.
𝐹
𝐶
=
3.55(2.56−1.77)
2.56(3.55−1.77)
𝑥100=61.54%

27
Table A.3: Showing grades and recoveries for synergies of SEX and ADTP at150gpt.
Ratio Item Wt.(g) Assay % Wt.(metalg) Recovery
Feed 997 2.56 25.52 100
1:0 Concentrate 313 3.5 9.61 73.86
Tailings 684 1.75 11.97 26.14
Feed 926 2.56 23.71 100
1:1 Concentrate 325 3.425 11.13 89.83
Tailings 601 0.79 4.75 10.17
Feed 970 2.56 24.83 100
2:1 Concentrate 360 3.45 12.42 92.38
Tailings 710 0.62 4.40 7.62
Feed 916 2.56 23.44 100
3:1 Concentrate 249 3.295 820.455 93.83
Tailings 667 0.58 3.87 6.16
Table A.4: Grind of mesh of Time.
Time (minutes) 80% Passing 75µ
5 24.5
10 48
15 64.5
20 75.5
25 80
30 91

28
REFERENCES
Boutin, P. and Wheeler, D. A. (1967), “Column Flotation Development Using an 18 Inch Pilot Unit”,
Canadian Mining Journal, March 1967, Vol. 88, pp. 94-101.
Davis, V.L., Jr.,2. Bethell, P.J.,Stanley, F. L., and Lutrell, G. H. (1995), “Plant Practicesin Fine Coal
Column Flotation”, Chapter 21, High Efficiency Coal Preparation (Kawatra, ed.) Society for Mining,
Metallurgy, and Exploration, Littleton, CO, pp. 237-246.
Degner, V.R., and Sabey, J. B. (1988), “Wemco/Leeds Flotation Column Develompent”,
ColumnFlotation ’88 – Proceedings of an International Symposium, Chapter 29, pp. 267-280.
Dobby, G. S., Amelunxen, R., Finch, J. A. (1985), “Column Flotation: Some Plant Experience and
Model Development”, International Federation of Automatic Control (IFAC), pp. 259-263.
Eberts, D. H. (1986), “Flotation- Choose the Right Equipment for your Needs”, Canadian
Mining Journal, March, pp. 25-33.
Fuerstenau, M. C., Miller, J. D., and Kuhn, M. C. (1985), Chemistry of Flotation, Society of Mining
Engineers, AIME, New York, 170 pages.
Glembotskii, V. A., Klassen, V. I., and Plaksin, I. N. (1972), Flotation,Primary Sources, New York,
633 pages.
Kawatra, S. K., and Eisele, T. C. (1987), “Column Flotation of Coal”, Chapter 16, Fine Coal
Processing (Mishra and Klimpel, eds.) Noyes Publications, Park Ridge, NJ, pp. 414-429.
Kawatra,S. K., and Eisele, T. C. (2001), Coal Desulfurization: High-Efficiency Preparation Methods,
Taylor and Francis, New York, 360 pages.
Kawatra, S. K., and Eisele, T. C., (1992) “Recovery of Pyrite in Coal Flotation: Entrainment or
Hydrophobicity” Minerals and Metallurgical Processing, Vol. 9, No. 2, pp. 57-61.
Klassen, V. I., and Mokrousov, V. A. (1963), “An Introduction to the Theory of Flotation” (translated
by J. Leja and G. W. Poling), Butterworths, London.
Klimpel, R.R.(1995), “The Influence of Frother Structure on Industrial Coal Flotation”, HighEfficiency
Coal Preparation (Kawatra,ed.),Society for Mining, Metallurgy, and Exploration, Littleton, CO, pp.
141-151.
Munali nickel mine (2011) feasibility report study.
McKay,J.D.,Foot, D.G., and Shirts, M.B.(1988), “Column Flotation and Bubble Generation Studies at
the Bureau of Mines”, Column Flotation ‘88, SME-AIME, Littleton, Colorado pp. 173-186.
Metso (2006) Basics in Minerals Processing, 5th Edition, Section 4 – Separations, Metso Minerals,
http://www.metso.com.
Mining Chemicals HANDBOOK. (2004). CYTEC.

29
Rao T. C.,Govindarajan, B.,and Barnwal, J. P. (1995), “A Simple Model for Industrial Coal Flotation
Operation”, High-Efficiency Coal Preparation (Kawatra, ed.), Society for Mining, Metallurgy, and
Exploration, Littleton, CO, pp. 177-185.
Rubinstein, J. B. (1995), Column Flotation: Processes, Designs, and Practices,Gordon and Breach,
Basel, Switzerland, 300 pages.
Tyurnikova, V. I.,and Naumov, M. E. (1981), Improving the Effectiveness of Flotation, English edition
translated by C.D.Zundorf, Technicopy Ltd., Stonehouse, England (originally published by Izd. Nedra,
Moscow), 229 pages.
Whelan, P. F., and Brown, D. J. (1956), “Particle-Bubble Attachment in Froth Flotation”,
Bulletin of the Institute of Mining and Metallurgy, No. 591, pp. 181-192.
Wills, B. (1988). Mineral Processing Technology. 6th ed. Oxford:Butterworth-Heinemann.
Yoon, R. H., and Luttrell, G. H. (1986), “The Effect of Bubble Size on Fine Coal Flotation”,
Coal Preparation, Vol. 2, pp. 174-192.
Yoon, R.H.,Luttrell, G.H.,and Adel, G.T.(1990), “Advanced Systemsfor Producing Superclean Coal”,
Final Report, DOE/PC/91221-T1 (DE 91004332) August, 1990.

Project_NYENDWA_(10th_ed)[1]

Recommended

Recommended

More Related Content

Similar to Project_NYENDWA_(10th_ed)[1]

Similar to Project_NYENDWA_(10th_ed)[1] (20)

Project_NYENDWA_(10th_ed)[1]