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Studies on Differential Flotation Characteristics of
Arsenopyrite Pyrite Concentrate
R K TUT J J QING U
J T C SIEFKEN
2
AND V N MISRA
2
STR CT
Investigation of the differential flotation characteristics between
arsenopyrite and pyrite in a ore concentrate have been carried out on a
laboratory Leeds flotation cell. Experimental results indicate that some
limited separation between arsenopyrite and pyrite
is
possible by
differential flotation under the conditions
of
high pH and low
pulp
redox
potential. The best perfonnance in this testwonc was achieved by using an
oxidizing agent at pH 10.7. Of the oxidizing agents tested, NaOCl gave
better oxidizing effect than
KMn 4
I t has also
been
observed that the
pulp redox potential depends on the
pulp
pH value.
At
high pH values a
hydroxide layer is fonned on the arsenopyrite surface, which has the
effect
of
a depressant on the arsenopyrite.
INTRODUCTION
Arsenopyrite is an arsenic mineral. It is usually associated with
precious metal ores, and the minerals galena, sphalerite and
pyrite. Its presence
in
an ore deposit can
be
of vital economic
significance. Arsenopyrite may carry significant fraction of the
gold present in certain ores. Such gold m ay b e present as separate
grains between arsenopyrite crystals and may
be
extracted by
direct cynidation Heinen
et
l 1980 . Gold can also be found in
solid solution or small inclusions in arsenopyrite Clark, 1960
which necessitates the us e
of
more unusual treatment techniques
Addison, 1980 . It may be useful to separate arsenopyrite and
pyrite so that they ca n be subsequently processed by different
methods to recover gold.
When arsenopyrite in an ore i s no t associated with gold values,
it is considered to
be
a nuisance impurity and its selective
depression is beneficial.
Th e
presence of arsenopyrite in sulphide
concentrate ca n cause severe health hazards and generate arsine
during pyrometallurgical and hydrometallurgical processing and
refining Habashi and Ismail, 1975 .
Rotation is the
only
cost effective method for separation of
arsenic bearing concentrate from pyrite.
In
the past few decades
researchers
h av e m ad e
attempts to separate arsenopyrite and
pyrite by flotation. In the early-1960s, several Russians scientists
studied the separation of arsenopyrite and pyrite in the following
way Glembotski, Klassen and Plaskin, 1963 :
flotation of pyrite by reducing the dissolved oxygen;
2. depression of arsenopyrite by using lime;
3. depression
of both
minerals followed
by
activation
of
arsenopyrite using copper sulfate;
4. depression of both minerals by using lime, followed by
activation of pyrite by ammonium chloride; and
5. depression of both minerals by using sodium sulfide
which was removed subsequently
by
dewatering,
followed by oxidation with oxidising agents pyrolusite .
Beattie and Poling 1988 conducted laboratory flotation tests
on
several ores and bulk concentrates to evaluate the
1. Western Australian School
of
Mines, PO
Box
597, Kalgoorlie
WA6430.
2. Kalgoorlie Metallurgical Laboratory, Chemistry Centre WA ,
Department
of
Minerals and Energy,
PO
Bo x 881, Kalgoorlie
WA6430.
effectiveness of chemical oxidising agents as selective
depressants for arsenopyrite. They reported that the maximum
flotation recovery
of
arsenopyrite occured at pH values less than
approximately 7.0. Increasing
pH
resulted in decreasing
flotability of arsenopyrite only under oxidising condition. The
selective depression of arsenopyrite from bulk pyrite-arsenopyrite
concentrate was achieved through the use of an appropriate
oxidising agents such as hydrogen peroxide or sodium
hypochlorite.
Li
and Zhan 1989 showed that arsenopyrite could be
depressed heavily in alkaline media. However, the presence of
heavy metal ion, such as Cu
2
+,
made separation difficult
O Conner
and Bradshaw 1990 recovered 74.8
per
cent
arsenopyrite and only 8.4 per
cent
pyrite
by
two-stage differential
flotation. They used dithiophosphate in the first s tage at pH
and copper sulfa te and di th iocarbonate in the second stage of
flotation. Iwasaki et l 1989 found that flotabil ity of
arsenopyrite was improved in a nitrogen atmosphere and
decreased markedly by increasing pH above 7.
In this investigation an attempt was made to separate
arsenopyrite and pyrite by differential flotation using a laboratory
Leeds flotation cell. Rotation tests were carried ou t to observe
the control that could be exerted over the flotation of arsenopyrite
through the us e of oxidis ing agent. All f lotation tests were
performed with an arsenopyrite-pyrite concentrate obtained from
a Western Australian GoldMine.
EXPERIMENT L
Materials used were as follows:
MIRC frother , PAX collector , NaOH pH modifier , NaOCI
oxidant ,
Ca OClh
oxidant and
KMn04
oxidant .
The
ore
sample studied was flotation concentrate 80 pe r cent passing 200
mesh 74 microns produced from a flotation circuit. The major
elements were determined
by
chemical analysis. Th e average
contents are as follows: Au = 60 g/t,
Fe
=
20
per cent, S = 20 per
cent, and As = ten per cent. Mineralogical examination using
XRD, light microscopy and scanning electron microscopy
showed that the concentrate was mainly composed of pyrite and
arsenopyrite, with talc and minor quartz. The grains of these
minerals ranged between 100 m and sub-micron sizes. The
concentrate
also
contained minor amounts of iron-nickel
sulpharsenide, slightly manganoan magnetite, galena and silver
bearing gold with silver content estimated to
be
three to five
per
cent.
Experimental procedure
The flotation tests were conducted in a three litre Leeds flotation
cell. The ore sample, unless otherwise stated, was tap water
washed using a pressure ftlter; this was then used as the flotation
feed. The pulp density was adjusted by addition oftap water to 25
pe r cent by weight . The pH level was adjusted with NaOH.
Unless otherwise stated, the dosage
of
PAX was
50
g/t and MIRC
38 g/t
of
the flotation feed. The pulp redox potentia l Eh was
adjusted either by NaOCI or by KMn04 and measured using a
redox potential meter. The redox potential readings obtained from
the meter were values relative to a Ag/AgCIIKCI 1.0 M
reference at 25C. Th e feed was conditioned for tenminutes.
Extractive Metallurgy of Gold
and Base Metals
Kalgoorlie 8 October
99
217
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R K TUTEJA, QING LID, T C SIEFKEN and V N MISRA
TABLE 1
lotation results
the
ee s
cell
Test No. Test Conditions
Gr:lde )
Recovery )
Con. )
Note
pH Eh Oxidant
FeAsS
FeS2
FeAsS FeS2
1
5.44
121
-
16.0 28.9
14.9 22.7
21.1
2. 8.44 100
-
13.5 35.4
15.1 32.5
24.9
pH
3.
10.68 19
-
11.3
47.8
8.6 26.8
16 3
4.
8.08 140
NaOC1
16.5 23.4
22.2
28.1
30.7
5.
8.24 175 NaOCI
17.8 12.2 23.0
14.8
29.0
6. 8.10
25 0
NaOCI
17.4 10.0
18.1 8.5
22.7 NaOCI
7.
4.80
2 20
Ca OClh 18.7 20.3 21.3 18.9
26.1
8.
8.36
-9 0
KMn4
16.5 28.6 21.0 29.5
27.0
9.
4.30
10 0
KMn4
21.7
37.4
61.6
793
59.6
KMn4
10.
5.60
16 0
KMn04
17.0
35.9
30.9
53.6 41.4
11.
53 0
36 0
KMn04
17.0 12.4 21.6
27.8 30.8
RESULTS AND DISCUSSION
In the present investigation, the following methods were tested:
Flotation
of
pyrite by depressing arsenopyrite at high pH
level;
2. Flotation
of
pyrite by depressing arsenopyrite using
oxidants; and
3 The combination
of
the above two.
The recoveries and grades of arsenopyrite and pyrite minerals
under various test conditions are listed in Table 1
:0
I
30
t
20
0
0
_
---
_ _ . 2
...n
10 11
Effect o f p H level
From tests 1 to 3 as listed in Table 1 and plotted in Figure
I,
it
can be seen that pH level has a small favourable effect on
differential flotation between arsenopyrite and pyrite. The
recovery
of
arsenopyrite decreases and recovery
of
pyrite
increases, with increase in pH level. The highest recovery
of
pyrite 32.5 per cent) oeeured at pH 8.4 and the lowest recovery
of
arsenopyrite 8.6 per cent) oeeured atpH 10.7.
The effect of change in pH on the grades of arsenopyrite and
pyrite was similar
to
that
of
their recovery. The observation that
the higher pH results in the poorer flotability
of
arsenopyrite
agrees with that reported by Huang and Wang 1985). Further,
according to Beattie and Po ling 1987) the oxidation
of
arsenopyrite above pH 7.0 resulted in a ferric hydroxide layer
forming on the surface, which inhibited the oxidation of xanthate
to dixanthogen. Although minor differential flotation between
pyrite and arsenopyrite takes place, their grades in tailings do not
show any significant change at different pH values.
The pulp redox potential Eh was found to be significantly
affected by changes in pH. As can be seen from Table
I,
the
redox potential drops from
121 to
19
mv
as the pH increases from
5.44 to 10.68.
Therefore, it can be concluded that
by
controlling the pulp pH,
a small amount of differential flotation between arsenopyrite and
pyrite is possible.
FIG
I - E ffec t of
pH
level.
Effect of oxidising
agents
Sodium hypochlorite and calcium hypochlorite as
oxidising agents
As shown in Figure 2 corresponding to tests 4 5 and 6) with
increase in redox potential, the recoveries and grades
of
pyrite
decrease more than those
of
arsenopyrite. Therefore it is the
pyrite that is being depressed at higher redox potentials. This was
also observed in test 7 where calcium hypochlorite was applied as
an oxidising agent at pH 4.8.
Potassium permanganate as an oxidising agent
In test 8, where the original sample as received from the plant
was used, much NaOH was added
to
raise the pH
to
8.36.
Consequently, in spite
of
the small addition
of
KMn04, a
negative redox potential was observed. In this test, both at a
higher pH and in the oxidising environment, better performance
was not achieved when compared to
that
of
test
2
2 8
Kalgoorlie 26 28 October 992
Extractive Metallurgy
of
Gold and Base Metals
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Fio 2 - Effect of Eh using NaOCl.
- _ r - - . . . . . . - - . . . . - _ _ : ~ - _ : : : - ~ : : .
110 140 10 111 200
.,.....---------------1
I
t
o ~
......
~ ~ 1
~ . . . . . . . . - . a
FLOTATION CHARAcrERISTICS OF ARSENOPYRITE/PYRITE
arsenopyrite. Above a redox potential
of
150 mv a lower
recovery
of
pyrite compared to arsenopyrite resulted. The
arsenopyrite recovery, however, did not change significantly. The
greater depressing effect
of
KMn04
on arsenopyrite
and
pyrite is
shown at higher redox potentials.
3. The best conditions for differential flotation between
arsenopyrite and pyrite
as
determined by
this
work.
occurs when using an oxidising agent at a high pH.
4. A high pH alone will not usually result in an effective
differential flotation.
Finally, from these series
of
flotation tests, the best differential
flotation conditions for this type
of
sample are: maintaining a pH
of
10.7
and addition
of
a small amount
of
NaOCI. However, these
conditions alone will not result in an effective pyrite and
arsenopyrite separation.
KNOWLEDGEMENT
tests 9,10 and 11 lower pH values were maintained and
KMn04
was used
10
adjust the redox potential. The best
separation, as shown in Figure 3, occured at a redox potential
of
160mv.
~
I
litIloIIII 11
i
, ,- -,.n
=s g
21
10
.
so
1
Ih mvl
FIO 3 - Effec tof Eh using KMn04.
ON LUSIONS
1. In
the differential flotation
of
arsenopyrite and pyrite pH
is
an important factor. The optimum pH suitable for
depressing the arsenopyrite is 10.7.
2. The pulp redox potential was found to
be
dependent on
the pH. Even
the redox potential is negative, flotation
still occurs as long as the pH is suitably maintained.
In
addition, in terms of the surface oxidation, oxidant NaOCl
shows much higher depressing behaviour than KMn04. The
greater depressing effect
of
NaOCl on arsenopyrite at lower
redox potentials resulted in higher recovery
of
pyrite compared
to
The authors wish to thank Or John Hosking, Director, Chemistry
Centre (W.A.) for permission
10
present
this
paper and are also
indebted 10 Or Tony Bagshaw and Professor David Spottiswood
for their comments. This investigation was funded by a MERIWA
grant, for which the authors are grateful.
REFEREN ES
Addison, R, 1980. Gold and silver ext ract ion from sulphide ores. Min
Congr
Oct, pp 47-54.
Beauie, M J V and Poling, G W 1987. A study of the surface oxidation of
arsenopyrite using cyclic voltarnetry, InJernational
Mineral
Proassing pp 87-108.
Beauie, M J V and Poling, G W, 1988. Flotation depression of
arsenopyrite through use of oxidising agents, Trans IMM 97(C), pp
15-20 (The Institution of Mining and Metallurgy: London).
Clark, L A 1960. The Fe-As-S system: phase relations and applications,
Econ Geol 55,
pp
1631- 1652.
Glembolski,
V
A, Klassen,
V
I and Plaksin, I N, 1963. Flotation,
MonumenJ press New- York, USA, p
540.
Habashi, F and smail, M I, 1975. Health hazards and pollution in the
metallurgical industry due to phosphine and arsine, CIM Bull
August, pp 99-104.
Heinen, H J, McClel1and,
G
E and Lindstrom, R E, 1980. Recovery
of
gold from tusenopyrite concentrates by cyanidation-carbon
adsorption, USBM,
RI
8458, pp 1-40.
Huang, K and Wong, D 1985. A study
of
selective flotation
of
antimonite
and arsenopyrite, Nonfe ous Metals 37:(2), pp 22-29.
Iwasaki, I Malicsi, AS
Li X
and Weiblen,
P
W, 1989. Insights into
beneficiation losses
of
platinum group metals from gabboric rocks,
Challenges in Mineral Processing Society of Mining
Enginurs
(Ed:
P Somasundran), pp 433-447.
Li G and Zhang, H 1989. Effect of alkaline oxidants and cupric ions on
arsenopyrite flOlation,
Nonfe ous
Met Chin Sac Met, 41:(4), pp 27
32.
O Conner , C T Bradshaw, D J and Upton, A E, 1990. The use of
dithiophosphates and dithiocarbonates for flotation of arsenopyrite,
Miner Eng 3:(5), pp 447-459.
Extractive Metallurgy of Gold and Base Metals Kalgoorlie, 26 28October 1992
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Kalgoorlie
8
OCtober 1992
Extractive Metallurgy of Gold and Base Metals
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Factors Affecting the Recovery and Grade of
Complex Lead Zinc Ores by Flotation
U NAY WIN
AND D S Y
2
STR CT
The size-by-size batch flotation behaviour and flotation rate constant
of
galena and sphalerite from three lead-zinc ores were determined. The
experimental data were fitted
to
the Klimpel model
o
first order flotation
kinetics.
t
was found that the flotation behaviour of the coarse +63
J IlI1
and
fine -10
lUD)
size ranges of galena was significantly poorer than that of
the intermediate -45+10
J IlI1
size range, and the fine size range of
sphalerite was p oore r tha n the othe r size ranges. The flotation rate
constant was found to be a maximum at some intermediate particle size.
The contact angle
o
galena and sphalerite was measured under the
same reagent condition as the flotation tests
by
using a simple method of
bubble-particle attachmenL
t
was found that the ov er all flotation
behaviour varied according to the trends in the contact angle.
Bubble size was measured for the different types
o
frother used in the
flotation tests.
t
was found that stronger frothers produced the smallest
bubbles and gave high recovery and high flotation rates, while weaker
frothers produced larger bubbles and gave higher grades.
t was
also found that the more complex mi;neral assemblages resulted in
p oo re r fl ota tion b eh avi ou r t ha n t he o re c onta ining relative ly simple
mineral assemblages.
Differentiation between true flotation and the entrainment of mineral
pa:ticles during the flotation process was determined. The results show
that fine galena was entrained in the froths at
short
flotation times, and the
true flotation rate constants were higher
than
the overall flotation rate
constants for all size fractions
o
sphalerite and for all fractions greater
than 10 J IlI1 for galena.
INTRODUCTION
Production o lead and zinc concentrates from complex lead-zinc
ores
by
using froth flotation is
an
important part o the production
o lead and zinc metals. the past, ores were high grade, there
was only o ne metal o interest, and it was relatively simple
to
extract. However, this is no longer the case. Nearly every
mineralisation has some problem either in the mining or the
extraction o the metal.
Most o the complex lea4-zinc ores contain lead-zinc minerals
in finely disseminated form. Although flotation is by far the most
important unit operation o mineral concentration, the recovery
achieved by using flotation for these fine ores
is
often poor. This
is because
o
the relationship among the various physical and
chemical properties
o
fine particles, and their behaviour in
flotation.
Surface and electrochemical propertie o fine particles tend to
be different from coarse particles of the SaI1)e material. Due to
the small mass and momentum o fine particles, they are carried
into the froth by entrainment, which
is
different from the
mechanism o particle-bubble attachment in flotation.
gangue
minerals are included in such entrained particles, the result is a
reduction in the grade
o
the concentrate.
Finer mineral particles have higher specific surface energies
and this may influence flotation in a number o ways. It may
introduce undesirable impurities into solution, affecting
1. Metallurgical Engineer, No 1 Mining Company, Kanbe Road,
Yankin PO Rangoon, Myanmar.
2. Senior Lecturer and Acting Head, Department ofMinerals
Engineering and Extractive Metallurgy, WA School
o
Mines,
PO Box 597, Kalgoorlie, WA 6430.
collector/mineral interactions. Rapid oxidation may also render
some minerals non-floatable under the conditions used for their
flotation. The high surface energy
o
fine particles also increases
the tendency
o
collectors to adsorb non-specifically. Fine
particles have low collision probabilities because o their small
mass, which results in a low flotation rate
and
low recovery. Fine
particles at the liquid/vapour interface may also stabilise the
froth, causing concentrate handling problems.
Because
o
the extremely fine dissemination and interlocking
o
minerals in complex sulphide ores, the treatment
o
these ores
represents one o the most complicated problems in base metal
flotation. The difficulties
arise
in producing high grade or high
recovery or both in flotation circuits. This problem comes from
incomplete liberation, poor flotation response at fine particle size
and/or interaction with some components
o
the complex ores.
However, due
to
losses
o
mineral and metal values in the fine
size range, considerable interest is growing in developing new
processes and improving old processes for the recovery
o
fine
particles. The objective
o
this work is
to
examine the flotation
behaviour
o
fine particles in terms
o
the flotation kinetics of
different size classes in lead and zinc concentration from bench
scale flotation test work. The variables studied were type
o
frother, type o collector, collector concentration, and ore type.
The principle recovery mechanisms are presumed to be
genuine flotation bubble attachment and levitation) and
entrainment carry-over with water which enters the concentrate
via the froth). The other possible recovery mechanisms ,
including entrapment in the froth, carrier flotation and the
influence
o
slime coatings, froth modification by fines, or
possible size effects associated with the return
o
particles from
froth to pulp, are not considered.
EFFECT
OF P RTI LE
SIZE ON FLOT TION
Particle size is recognised as being a very important flotation
variable, and major problems in flotation arise in many instances
from the relatively poor response
o
coarse and very fine
particles. Recovery falls sharply above 100
lJ Il
but only
gradually below 10
lJ Il
No t all minerals show a maximum recovery in exactly the
same size range, but there is no doubt that recovery
is
best for
particles of an
intermediate size.
The presence
o
gangue minerals in the pulp
~ i g h t
also effect
different particle sizes differently. ~ v r y
o
glmgue durIng the
tlotation
o
lead and zinc at Broken Hill was found to increase
withdecreasing particle size, particularly below 10
lJ Il
Kelsall
l 1974; Lynch and Thome, 1974). Granite gangue was
recovered much better as the pulp became finer when synthetic
mixtures o galena and granite were floated in laboratory batch
cells Gaudin
l
1931). I t was attributed
to
mechanical
carry-over
o
fme gangue. The entrained fine gangue causes a
decrease in both concentrate grade and the flotation rate
o
the
valuable mineral, with decreasing size.
E X P E ~ E N T L P R O E D U R E
Flotation tests
Lead-zinc ores from three deposits, Cadjebut (WA), Woodlawn
NSW) and Bawdwin Myanmar) were used for the flotation
tests.
ExtractiveMetallurgy of Gold
and
Base Metals
Kalgoorlie 6 8 October 99
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UNAYWlNand
OS
YAN
15
00
80
60
QI
>
0
40
QI
l
20
10
FIG 1 -
Cumulative recovery
of
PbS at different times as a function
of
different size fractions, (0.023 kg/t NaCN, 0.05 kg/t NaEX, H407 0.022
kg/t, pH 8.5 in PbS flotation of Cadjebut ore).
20 - 40 60
120 > 240 s _ Rate
20
30
40 50 60
Mean Size (microns)
Flotation tests were conducted in a modified Leeds cell.
Recrystallised sodium ethyl xanthate (NaEX), sodium amyl
xanthate (NaAX), liquid CMS 41 (secondary butyl
dithiophosphate) and CMS 42 (hexyl dithiophosphate) were used
individually or in combination with one another as collectors.
Liquid Nalflote series frothers (polyoxypropylene glycol ethers)
and Dowfroth frothers (polypropylene glycol ether) were used
individually as frothers. Sodium cyanide (NaCN), or a
combination
of
sodium cyanide and zinc sulphate (NaCN +
ZnS04), were used as depressants for sphalerite in the lead
sulphide flotation. Copper sulphate (CUS04) was used as
activator for sphalerite in the zinc sulphide flotation.
Flotation concentrates and tailing were wet screened to
produce six size fractions, -75+63 llm, -53+45
l ffi,
-45+38 llm,
-38+20
l ffi,
-20+10
l ffi
and -10 llm. These size fractions were
studied, in terms
of
flotation kinetics.
Flotation response is a function of the three factors; chemical,
equipment and operation factors. In this study, the equipment
factors and operation factors (per cent solid, pulp density,
temperature, air flow rate, ete ) were held constant.. Different
types of kinetic models (First-order model, Gamma, Kelsall,
Modified Kelsall and Klimpel model) were used to fit the
experimental data, but the Klimpel model gave the best fit
of
the
data.
FIG
2 - Cumulative recovery
of
Zns at different times as a function of
different size fractions, (0.182 kg/t CUS04, 0.08 kg/t NaEX, H407 0.022
kg/t,
pH
10.5 in Zns fltation of Cadjebut ore.
60
24 _ Rate
20
30
40 50
Mean Size (mlcrona)
10
lOO
15
80
::tl
;
10 n
60
0
i:
>
a
40
:
5
20
Contact angle measurement
The contact angle of galena and sphalerite were estimated using
the bubble-particle attachment method (Hanning and Rutter,
1989). This involves determining the diameter
of
the largest
particle from a population of particles immersed in water that can
be raised against gravity by a captive air bubble.. The contact
angle can be calculated by using the equation
of
Scheludko t l
(1976),
as
given below:
d
max
=2
3ywv 2 p g
1{2 sin
8 2
(1)
where d
max
maximum particle size captured by bubble
Ywv surface tension
of
liquid-vapour
p density difference between the solid and water
g gravitational acceleration, and
8 equilibrium contact angle
Bubble size
measurement
The bubbles generated from the Leeds flotation cell were
captured by a capillary tube at the centre
of
the cell. The bubbles
are sucked up the capillary tube where the number
of
bubbles,
as
well
as
their diameter, were calculated using a Randall bubble
size measurement unit. The unit detects the beginning and end
of
a bubble
in
the capillary as it passes a photo diode. For a
capillary
of
known diameter, the volume
of
each bubble is
measured and hence the bubble diameter is calculated.
RESULTS AND DISCUSSION
The experimental data were fitted to the Klimpel model of first
order flotation kinetics, as represented by equation
2
R = R_ [l-(l/(kt(1-exp(-kt] (2)
where, R the cumulative recovery for time
1
R_ recovery at infinite time
k flotation rate constant
The effect of particle size on recovery and flotation rate
Different particle sizes
of
galena and sphalerite exhibited
different flotation rates.
The results
of
the flotation tests using NaEx and Dowfroth 400
are shown in Figures 1 and 2. Galena recoveries for all tests had
similar characteristic curves. Sphalerile recoveries in other tests
also had characteristic curves similar to
Figure 2.
Significant differences in recoveries and flotation rate constants
were obtained for the different size fractions. The highest
recovery and highest flotation rate were exhibited by the
intermediate size fraction (-38+20
l ffi)
for PbS. With sphalerite
flotation, there is no clear maximum in recovery and flotation rate
over the size range measured, with xanthate as collector. The
minimum flotation rate of sphalerite in Zns flotation is obtained
in the fine size range, and the rate slightly increased with size.
In
the discussion of flotation kinetics here, the flotation rate
constants of sphalerite in PbS flotat ion and the flotation rate
constants
of
galena in
Zns
flotation have not been considered.
The content
of
each mineral in the other mineral s concentrate
may be due to incomplete liberation and the relation between the
flotation rate constants of one mineral in the other mineral s
concentrate and particle size was random. Hence their behaviour
is uncertain.
222
Kalgoorlie, 26 -
October 1992
Extractive Metallurgy of Gold and Base Metals
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RECOVERY AND
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OF COMPLEX LEAD-ZINC ORES BY FLOTAnON
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