flota arsenopyrite sep1

8/12/2019 flota arsenopyrite sep1

1/14

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


2/14

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


3/14

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

219


4/14

22

Kalgoorlie

8

OCtober 1992

Extractive Metallurgy of Gold and Base Metals


5/14

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


6/14

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


7/14

RECOVERY AND

GRADE

OF COMPLEX LEAD-ZINC ORES BY FLOTAnON

A

5IJm

I

CMS41

CMS

>

CMS42

CMS41 NaEJ

flota arsenopyrite sep1

Documents