075 Simulation of Grinding Circuit at Silver Plant, Turkey

7/26/2019 075 Simulation of Grinding Circuit at Silver Plant, Turkey

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Simulation of grinding circuit

at

G j m j ~ k 6 y

silver plant, Turkey

S.

Ersaym

B.

Sonmez $.

L.

ErgOn,

B.

Aksanl and I.

F.

Erkal

Synopsis

Since

start-up

the comminution circuit

at

Etibank's

i i m i i ~ o y

silver plant, Kiitahya, Turkey,

has

experienced

severe problems ansmg from the

unexpected

amount and moisture content of clays

n

the

feed.

Three

alternatives were

proposed in an

attempt to solve the

problem: first,

replacement of the

existing crushing circuit by tooth roll-type crushers;

second, the use of washing drums;

and,

third, the

introduction of primary

autogenous/semi-autogenous

milling

in

place

of

crushing.

The effects of

the first two

alternatives

on the

capacity

of

the

grinding circuit

were

investigated through computer

simulations.

A

series of laboratory

and

industrial-scale

ball-mill

batch grinding tests was conducted. The

parameters

of

a

cumulative-based

kinetic model were determined.

The

ball-mill circuit was sampled and

operating

model

parameters

were calculated. The

ability

of

batch

grinding

parameters to represent the operating.

conditions was

then

examined.

.

Laboratory

tests failed

to

provide acceptable

model

parameters.

However,

parameters

that were obtained

in

work with an industrial-size batch mill were found to

be suitable for

simulation

studies. The simulations

indicated that a

continuous feed

of

-12

rom provided

by tooth-roll crushers

would

maintain the desired

capacity. Increasing the

feed

size to -25 rom

would

cause

a

reduction

in

capacity of 8-15 .

The

use of

washing

drums

in

the

existing

crushing circuit

could

also solve the capacity problem,

but

small fluctuations

in

water flow

rates might

result

in problems

of excess

water in the grinding circuit.

ackground

The

G i i m i i ~ k O y silver plant, which has a processing capacity

of 1 000 000

t/year run-of-mine (ROM) ore, was designed to

recover silver by

the

cyanidation process after grinding

of the

ore

to

80

-74

1ffi.

The

original comminution circuit

of the

plant comprises three stages of crushing

and

ball-milling

(Fig. 1). After primary crushing to 250 mm the top size is

reduced, to 50 mm by two secondary cone crushers and

further reduced to 12

mm

by tertiary cone crushers, which

are in closed circuit with a screen with an aperture size

of

12

mm. Crushed

ore

is

stockpiled

and then

fed

to

two ball

mills (3.85 m x 5.75 m) running

in

parallel.

The

ball-mills are

in closed circuit with batteries

of

hydrocyclones.

The

circuit

is equipped with belt scales, density gauges, flow meters

and

particle-size monitors. Operating parameters are controlled to

Paper

prepared for presentation

at

the meeting Computer solutions

in mining

and

processing, organized

by

the Institution

in

association

with the

University

of

Leeds

and

held in Leeds

on

23

and

24

September, 1991. Manuscript first received

by

the Institution

of

Mining

and

Metallurgy

on

12 September, 1991; revised manuscript

received

on

27 January, 1993.

Paper

published in

Trans. Instn Min.

Metall.

Sect. C:

Mineral Process. Extr. Metall. ,

102, January-April,

1993.

©

The

Institution of Mining and Metallurgy 1993.

provide the desired

ground product

of

80

-74

1ffi. The

overflow from

the

hydrocyclone is sent

to the

leaching circuit.

The

comminution circuit was designed on the basis

of

the

expected size distribut ion

of ROM

ore. Prior to the design

of

the plant the mineralogy of the ore

had

been thoroughly

investigated and five different ore types

had

been identified in

the

deposit.

Bond

indices

had

been determined for

the

ore

types

and

were found to vary between 7

and

25 kWh/to

It had

been

estimated from these values

that the

average feed

blend

would have a work index

of

15 kWh/to

The

choice

of

ball-mill

size

had

been based on the Bond index data

and

Was aimed at

grinding the ore from 80

-10 mm

to 80 -74 1ffi with a

capacity

of

70 tIhImilL

The

ball-mills were expected to run

with a circulating load

of 250-400 ,

a figure

that

was

used

as

the basis for the selection

of

ancillary equipment.

When

mining

commenced t

was found

that the

actual size

distribution

of

the

ROM

ore showed a large deviation from

that expected. The ore contained

much

larger amounts

of

clay (Fig. 2) and

had

a higher moisture content (16 as

opposed to ·the expected value of 6%). These unexpected

Run of mine

ore - -»

Grizzly

~ _ ~

Screen Oversi ze _,

:

I'

lt £i

one Cone

Undersize I c r ~ e r

L G r : : ~ : r i J

-12 mm I O v e r l l o w ~ To leaching circuit

~ l o Y l o n e m

~ I

.

Underpow+®

Fig. 1 Original comminut ion fiowsheet

of plant

showing sampling

points (numbered

14

lO°mlfmrmgmmng

Ii

t

]

100 1000

Particlcslze. un

Fig. 2 Expected

and

actual particle-size distributions

C32



characteristics affected processing adversely. The clay

adhered to surfaces, jamming ore bins and chutes, clogging

crushers and blocking screen openings. s a result, since

start-up the plant has never reached its design capacity except

during very short, dry summer seasons. Furthennore, the clay

caused mechanical problems

and

rendered

the

automatic

control

of

grinding

and

subsequent circuits almost

impossible.

The

problems were partly solved by increasing

the closed-side settings of the tertiary crushers and the

aperture

of

the

screen in

the

closed circuit, where clays

had

the greatest adverse effect, to 25 mm. During summer

periods the coarse fractions containing smaller amounts of

clay were separated so that they could be fed

to

the plant in

rainy winter seasons.

It

appears that the management considered three possible

solutions to the problem: first, replacement

of

the entire

crushing circuit by tooth-roll crushers (which, it is claimed,

are suitable for clayey types

of

ore); second, the use

of

washing drums after the primary jaw crusher to separate clays

prior to secondary and tertijIry crushing; and, third,

replacement of the crushing circuit by primary autogenous

semi-autogenous grinding.

The

management was, however,

not

sure whether the

design capacity

70

tIhImill) could be reached even i f the ore

were crushed down to

-12 mm

since the circuit has never

been fully operational with a feed

of

this size. Such wotries

were further substantiated by the fact that

under

the

prevailing conditions the capacity

of

each mill was only about

30-40 tIh.

In the present study

of

the three alternatives the effect

of

the first two on the grinding circuit was examined by carrying

out

computer

simulations of the proposed alternative circuits.

The

aim was to furnish the management with information on

the effect

of

each alternative

on

the capacity

of

the circuit to

provide a basis for decision-making.

The

task

had to be

completed within a very short time.

Modelling

of comminution

circuit

The

focal

point of

the study was the modelling

of

ball-milling.

Of

the types

of

models

that

are used to define size reduction

in

ball-mills, a kinetic model based

on

the rate

of

breakage of

cumulative proportions of sizes in the feed was chosen for its

simplicity and ease of use. It has been shown that this type

of

model gives reasonably good simulation results by comparison

with more complex kinetic models.

2

The

model

contains only

two parameters, simplifYing interpretation of the results. t has

also

been

reported

that

parameters

detennined in

laboratory

batch experiments may

be

employed directly in the simulation

of

comminution

circuits with

no

scale-up.

3

To define retention time

in

the mills plug flow of material

was assumed. Furuya

and

co-workers

4

have shown that the

plug-flow assumption is reasonable for high circulating loads,

i.e. >150%.

In the cumulative-based kinetic model the kinetic

parameter

is

defined as the rate

of

disappearance of oversize

material for a given sieve size.

For

batch grinding and

continuous grinding with the plug-flow assumption the

following equation is used to describe the comminution

process:

W(x,t) =

W(x,O)

exp(-kt)

1)

where

W(x,t)

is

cumulative percentage

of

oversize material for

sieve size

x

at

time

t W(x, 0) is

cumulative percentage

of

oversize material

of

size

x

in the feed; k

is

the breakage rate

constant, min-I; and t

is

time, min.

For

a continuous mill and with a plug-flow assumption the

retention time, t is calculated from

Instantaneous pulp volume of mill, m

3

=

(2)

Volumetric feed

rate

offeed, m

3

/min

The

relationship between the breakage rate function

and

particle size

is

described

by

the equation

3)

where C and

n

are constants

that

are dependent on material

and mill characteristics. f C

and n

are known, the size

distribution of the product for a given feed can

be

calculated

as

4)

W(x,t)

For

the closed-circuit simulation the Plitt modelS and

measured partition coefficients were used to describe the

hydrocyclone unit.

In

this model the relationship between the

corrected part ition coefficient and particle size is described by

5)

where C

is

the corrected partition coefficient, d

j

is

mean

particle size

of

the

ith

fraction,

d

s

is

the corrected

d

s

value

. c

and

m

IS a model parameter.

If

the water recovery

is

known, equation 6 can be used to

calculate the uncorrected partition coefficient, P;:

(6)

The

main function

of

the washing

drum is to

disintegrate

clay mechanically

so that

it can be separated by a simple

screening process. t

is

reasonable to assume that the size

reduction within the

drum,

if

any, would

be

negligible.

t

may

therefore be suggested that the separation behaviour of a

washing drum could be represented by a screen model. The

screen

model

developed by Whiten

6

was used for the

definition

of both

washing

and

screening processes.

This

model is based

on

probability considerations

and

the details

can be found elsewhere.

6

The size distribution of the product from cone crushing

was also defined by another model developed by Whiten that

is based

on

a matrix approach to size reduction.

6

The

model

comprises breakage, B and classification, C, functions.

For

a

known feed-size distribution,

j

the

product,

p is

expressed

in

matrix form as

p

=

l -C) . l -B .C) - I . j

(7)

where I is the

unit

matrix.

Since no

model

was available to represent tooth-roll

crushers, it was assumed that they would produce a size

distribution similar to

that of

the existing stockpiles for a

setting

of

25 mm. A graphical estimation

of

the size distri

bution

of

a product that has been crushed to --12

mm

by a

tooth-roll crusher was obtained by drawing curves parallel to

the size-distribution curves of the stockpiles.

xperimental studies and determination

of

model

parameters

Laboratory- and industrial-scale

batch

tests were conducted

to

determine model parameters for ball-mill grinding

of

the

ore. A sampling survey was also catried out

on

the existing

grinding circuit for comparison and model verification,

C33



although the

mill was

not

running to design specifications

during the

sampling period.

Laboratory

batch grinding tests

The study commenced with laboratory batch grinding tests to

determine suitable model parameters to represent

the

size

reduction behaviour of the ball-mills

at

the plant. Three

different types of ball-mill feed, code-named ASl, AS2 and

SC, had been

recognized by the plant engineers. These

had

been

crushed

to

-25

mm and

blended

and stockpiled

separately for processing

at

a suitable time

of

year. Samples

were taken by

the

engineers and sent to

the

laboratory.

The

three samples from each stockpile were treated separately. To

disintegrate the clay the samples were

kept

in

water

for

24

h,

after which their size distributions were determined by wet

sieving. The results revealed

that

the main difference

between

them

was

the

clay

content

(Fig. 3). The clays were separated

by

several stages of decantation. The coarse fractions were

dried and

then

crushed to 3.14

mm.

Batches from the

crushed samples were used in the tests.

The

reason for the

exclusion of clays from

the

grinding tests was to provide more

efficient grinding in the mill, thereby giving a better

simulation of closed-circuit grinding.

Batch

grinding

conditions were

kept

as close as possible to those of

the plant

(Table 1). A mix

of

balls with a

maximum

size

of

50

mm

was

used.

100

0

c

-

0

10

10 100

Y

-

+ASI

sc

Sampling of existing circuit

Initially, the sampling of

the

existing circuit was planned with

the primary aims of evaluating the ability of model

parameters determined

from

both the

laboratory

and indus

trial batch mills to

predict the

size-reduction

behaviour

of

the

mill

at the plant and, if

required, of deriving a scale-up

relationship

between the

mills.

t

was also considered

that the

data would be useful for model verification and for defining

the size-separation behaviour of the existing hydro cyclones in

terms

of

partition

coefficients, which

would be required

for

the modelling of the unit during the simulation of the whole

circuit.

1000

10000

100000

Paniciesize,

11m

Fig. 3 Particle-size distributions of samples taken from different

fine-ore stockpiles

Table

1

Grinding

conditions for laboratory

and

industrial

size batch mills

Laboratory mill

Industrial

mill

Dimensions,

m

0.20

x

0.35 1.09

x

1.13

Rotation rate, rev min-

54

30

Void volume,

45

40

Specific gravity of balls

7.8

7.8

Bulk density of ore, g

cm-

3

1.64

1.70

Ball charge,

40

33

Solids, wt%

60

75

Weight of sample, kg

Dry grinding

3.780

245

Wet

grinding

4.440

290

Both

dry-

and

wet-grinding tests were

performed on each

type

of

ore.

The

sample weights were chosen

SO

as to fill

100%

of the

void volume

between

balls

in

dry grinding

and

120%

in

wet grinding. After grinding for 15, 30,

45,

75, 120

and

240

s

the

size distributions of

the

ground samples were

determined. When

they

had

been sieved

and

weighed

the

fractions were combined

and

returned to the mill. The

same

sample was used for the entire duration of

each

grinding test.

Sieving was carried

out

dry down to 1.18 mm and wet

between

1.18 mm and 53

run

atch grinding tests in industrial batch

mill

n the plant there was a 1.09 m x 1.13 m ball-mill that was

used

for batch grinding of smelter slags. Initially, it was

thought

that in

case the laboratory parameters failed to

simulate

the

plant operation

the data

from a medium-size mill

would be

useful for scale-up purposes.

Apart from the maximum particle size in the feed, which

was 8

mm, the

sample-preparation procedure for the indus

trial batch mill was the

same

as that for the laboratory mill.

Grinding conditions for the mill are given in Table 1. The

maximum ball size was 75

mm. Dry-

and wet-grinding model

parameters

were

obtained

by grinding

the same

sample. The

sample was first

dry-ground

for 1, 2 and 4 min. After

each

grinding cycle

the

mill was

stopped

for sampling. The

samples were analysed by screening as described above. After

sieving the size fractions were combined

and

returned to

the

mill. On completion of the dry test a further 20% of

un

ground ore was

added

to the mill to raise

the

ore volume/void

volume ratio to

120% and the sample

was

wet-ground

for

3 min.

The

mill was

then

stopped

and

emptied and the

sample for sieve analysis was

taken

after

the

ground ore

had

dried.

Since

completion

of the

work

was a

matter

of

some

urgency,

the

sampling

had

to be carried

out during

an

unfortunate

period-just before

the plant

was

shut down

for

maintenance during the

winter

season-when

operating

conditions were

quite

different from

the

design values. The

liners were

worn,

capacity was very low

and,

most

important

of all, the ball charge was only about

20%

rather than the

design value

of

33%. Under the circumstances it was realized

that the

data could not' be used for the primary aiins.

Nevertheless, sampling was undertaken to provide data for

model

verification

and

modelling of

the hydro

cyclones.

Sampling

was

begun

after it

had

been

confirmed

that

flow

rates around

the

circuit were fairly steady by checking values

at the

central control-room. Samples were

taken

from

the

plant feed, mill discharge and hydrocyclone underflow and

overflow (Fig. 1) every 15 min during the sampling period

of 2 h 45 min. Flow rates and operating parameters were

recorded

while the samples were being collected. The size

distribution of

the

samples was

determined

by wet-sieving

in

labora tory facilities

at the

plant.

Detennination of model parameters

Breakage-rate constants were calculated from the laboratory

data

for cumulative oversize versus

time

by a

non-linear

regression technique. Breakage-rate constants were calcu

lated

in

the same

way from

the

dry-grinding

data

for

the

industrial batch mill. However, for wet grinding in

the

latter

C34



Table 2

Breakage-rate constants

and

model

parameters calculated from laboratory

grinding data

Particle size

r e a k a g e ~ r a t e

constant

k

J UTl

ASl AS2 SC

Wet

Dry

Wet

Dry

Wet

Dry

2360

0.295

0.296

1700

0.215 0.230

1180

0.175 0.166

850

0.122

0.103

600

0.092 0.077

425

0.074

0.064

300

0.059 0.042

212

0.046 0.036

150 0.035 0.028

106

0.030 0.022

75

0.025

0.018

53

0.Ql8

0.014

C

0.00103 0.00056

n

0.71667 0.78599

only two data points were available, so breakage-rate

constants were determined directly from

the

slope

of

the line

joining them.

The measured flow rates and size distributions obtained

from the plant were first adjusted by using a mass-balance

program and the rate constants were then calculated directly

from

the

balanced data.

The

calculation was

based on the

m

3

estimated hold-up volume

of

11.1

and the

plug-flow

assumption.

The

adjusted values were

not

found to deviate

markedly from the measured data.

To

calculate C

and n

equation 3 was linearized by taking

the logarithm

of both

sides and applying linear regression.

This method provided a better fit at the fine size range, for

which

the

breakage-rate constants were considered to

be

more reliable.

The

rate constants

and

model parameters

that

were

calculated from the laboratory

and

industrial batch mills

and

from the plant data are presented

in

Tables 2 and 3,

respectively.

For

the construction of a model

of

the existing grinding

Table 3 r e a k a g e ~ r a t e constants and model parameters

calculated from industrial batch mill

and

plant data

Particle size

r e a k a g e ~ r a t e constant k

J UTl

Industrial batch mil l

Plant

Wet Dry Wet

4000

0.791

0.730

0.256 0.266 0.293 0.272

0.172

0.180 0.203

0.197

0.137 0.138 0.151

0.143

0.097 0.097 0.105 0.099

0.081 0.078 0.081 0.068

0.066 0.060

0.064

0.056

0.053

0.047

0.052

0.043

0.044

0.037 0.041

0.034

0.034 0.029

0.034

0.027

0.029 0.024 0.028

0.023

0.021 0.021 0.023

0.020

0.018 0.017 0.020 0.018

0.00114 0.00086 0.00103

0.00072

0.67875 0.71586 0.70373 0.74228

circuits parameters representing the behaviour

of

the

hydrocyclone unit were also required. By using the adjusted

data the partition coefficients for each size fraction were

calculated (Table 4). n attempt was then made

to

use these

to estimate the parameters

of

the Plitt model, m d 5 ~

and

R

f

However, the non-linear optimization technique that was

employed did

not

provide a reasonable fit

to

meaningful

model parameters.

Evaluation

of

model parameters and model

verification

Since the available mineral processing packages did

not

include the model

that had been

chosen for ball-mill grinding

and

were

not

able

to

use the measured partition coefficient

for hydrocycloning, programs to model

the

grinding circuit

were written by the authors.

A model

that

incorporated a ball-mill with a hold-up

volume

of

11.1 m

3

and

a hydro cyclone unit in closed circuit

was capable

of

predicting mass flow rates and the size

distribution reasonably well. The performance of the hydro

cyclone unit was initially characterized by the measured

partition coefficients.

The

closed circuit was balanced by

iterative calculations.

During

iteration

the

partition coeffi

cients and (when used) model parameters were kept

Table

4 Measured partition coefficients

Nominal

size J UTl

Partition

coefficient

3360

2360

1700

1000

850

500

355

250

180

125

90

63

40

C

n

0.743

0.635

0.522

0.333

0.289

0.194

0.154

0.126

0.103

0.084

0.071

0.055

0.044

0.002670

0.718900

0.755

0.748

0.713

0.548

0.450

0.240

0.172

0.124

0.098

0.079

0.063

0.053

0.040

0.002890

0.697200

0.623

0.519

0.300

0.199

0.127

0.081

0.049

0.037

0.028

0.021

0.00039

1.05100

10000 1.000

6000 1.000

3180

1.000

1680

1.000

925 1.000

675 1.000

427.5 0.899

302.5

0.949

215 0.993

152.5 0.877

107.5

0.920

76.5

0.829

51.5

0.748

36

0.665

16

0.540

C35



constant. The

model

was first

tested

with

the adjusted plant

data.

When

the rate constants

that had been measured in the

plant

were

used

directly

in the model an

almost per fect fit was

obtained. However, when the rate constants were charac

terized instead by the model parameters C

and

n the size

distribution showed

some deviations-particularly in

the fine

size range. As expected, the deviations

corresponded

to the

points where

the

actual rate

constant

differed

from the

function defined by C

and

n (Fig. 4). Nevertheless, it was

concluded that the model

was suitable for further studies.

100

I

3

6

G

10

10

100

1000

V

,/

I*-

I

I ~ ' , l , ( ~ )

GmeasurW

A calculated C andn)

Panicle

size,

11m

Fig. 4 Comparison of measured and calculated size distributions of

hydrocycione overflow

When the model parameters

that were

obtained

from

the

laboratory

data

were

examined

it was

found

that

the

three ore

types exhibited similar grinding characteristics.

It

also

appeared that there were some differences between the model

parameters obtained under wet- and dry-grinding conditions

(Table 2). To test their ability to predict plant performance

and the

significance of

the

differences

between them

the

capacity of

it

closed-circuit grinding

arrangement

was

calculated

by using each

set

of

model

parameters.

The

circulating load ratio was

the

capacity-limiting constraint and

was

taken

to be

the same

as

the

design value for

the

plant-i.e.

400%. In this part of the work Plitt's models was

used to describe hydrocyclone behaviour. Initially, it was

thought

that

the model parameters

could

be estimated from

the

design and operating variables, but the calculated

parameters (especially d

soc

and R

f

showed considerable

deviation from expected values. This was probably due to the

fact

that

the hydrocyclones

had vacuum-controlled Linatex

underflow pockets. As a result,

the

actual apex

diameter

was

narrowed and

varied with operating conditions.

Therefore,

the

values chosen for

the model parameters m

d

soc

and

R

f

were 2.5, 741JIIl

and

0.4, respectively, so as

to represent

expected hydrocyclone behaviour (details

of

this aspect

of

the

work have been given elsewhere

7

. As can be seen from

Table

Table

5 Capacities of simulated closed-circuit grinding,

tlh

AS AS2

SC Industrial

batch l i l l

Wet

21.30

20.00

20.15 56.00

Dry

15.95 17.75

16.80 55.00

Difference, % 33.50 12.70 19.90

1.80

5, the calculated capacities confirmed the earlier observations

that the three

types

of

ore

had

similar grinding characteristics

and

that dry and wet conditions

had some

influence

on

the

model parameters

obtained

in

laboratory batch tests.

However,

the

most

important

conclusion was that

the model

C36

parameters

derived from

the

laboratory

study could not

be

used

directly

to predict plant

results.

There

was

no

significant difference

between the model

parameters for dry and wet grinding obtained from the

industrial batch mill. The simulation results confirmed this

finding (Table 5), the difference

between the

capacities

predicted

by simulations being only 1.8 . The capacities

were also reasonable approximations

of the

mill

at the plant,

which

had

a design capacity

of 70 t/h

for a feed of

-12 mm.

To check the validity of the model parameters that had

been obtained in the tests with the industrial batch mill

Bond

work indices were also determined for

the

different

ore

types.

Work index values of 13.79, 15.99 and 15.97 kWhlt were

obtained

for AS1, AS2 and SC, respectively. The conven

tional

method

for estimating

the

capacity

of

a given mill also

predicted

similar capacities for these

work

indices.

It

was

therefore

concluded that the model parameters obtained

from

the industrial batch mill could be used in simulating the size

reduction

behaviour

of the existing grinding circuit.

If

any

scale-up were required, these would,

at

least, provide a

conservative estimate of

the

actual situation.

As

can

be seen

from Table

1,

there

were

some

differences

between

conditions for

the

laboratory and industrial batch

mills.

This

was

due to

initial misinformation

about the

design

conditions

of

the

existing mill.

Tests

with

the

industrial

batch

mill were performed according to actual design conditions.

To test the possibility that differences between the results

obtained with

the

two mills

might

reflect

the

incorrect

conditions used with

the

laboratory mill one set of laboratory

tests was

repeated under the

design conditions.

This did not,

however, produce any significant differences

in terms

of

model parameters and simulated

capacities.

The

model parameters that were obtained from the plant

showed no resemblance to those given by the batch mills. t

appeared, however,

that

there was some similarity between

the

parameters

obtained

from

the latter. The value of n was

similar, although C showed significant differences.

This

gave

rise

to

a parallel type

of

relationship

between the

log

k

versus

log size curves for

the

two

mills-indicating

that n was

dependent

on

ore characteristics, whereas C varied with mill

parameters

(i.e. diameter).

As mentioned earlier, the partition coefficients that were

measured from the existing

hydro

cyclone unit did not provide

a good fit to Plitt's model.

t may

be suggested

that

this could

have

been

due to experimental

error

and/or the heterogeneity

of

the ore, which contains minerals with specific gravities

that

range

from 2.6

to 4.5. The unusual shape of

the partition

curve

could

be attributed to

the

latter feature. The fact

that

56%

of

the

finest fraction (-36 1JIIl) was going to

the

under

flow

indicated that the hydro

cyclone was

not

operating

properly.

t

was believed, however,

that

it

would be wrong

to

draw

firm conclusions from only

one

set

of

data,

but

in

the

limited time available it was not possible to obtain additional

data. A decision was

made

to use

the partition

coefficient

directly in the model to represent

the

prevailing hydro

cycloning behaviour and to use Plitt's model to illustrate

the

effect of better classification

on

the circuit.

iscussion of

simulations

The

aim of

the

simulation studies was to examine the first

two alternative solutions

in terms of

capacity

and

possible

bottlenecks. As explained earlier, .the capacity-limiting

criterion was a recycle load of 400%,

and

breakage

parameters that

were

obtained

from wet grinding

in

the

industrial

batch

mill-0.00276

and 0.7189

for C

and

n

respectively-were used in

all the simulation studies.

Two

simulations were

performed

for

each

flowsheet configuration.



One, which was referred

to

as existing conditions , involved

the use

of

the measured parti tion coefficient

and

a 64 solids

ratio

in

the milL

The

other simulation, referred to as design

conditions ,

used-as

its

name implies-a more

efficient

classification defined by Plitt s model with values for the

parameters m

d

50c

and

f

of

2.5, 74 lID

and

0.40,

respectively, and

75

solids in

the

milL

For both

existing and

design conditions the ball charge was assumed to

be 33 .

Circuit

alternative

1

In

simulations

that

included tooth-roll crushing two- and

three-stage crushing were considered (Fig. 5), Although

three-stage crushing and, hence, a product

of

-12

mm,

were

desired, it would be difficult to guarantee reliable operat ion

of

the crusher at this size range, so two-stage crushing was also

studied. As has already been mentioned,

the

product-size

distribution

of

two- and three-stage tooth-roll crushing was

estimated from the size distribution

of

the existing stockpiles.

These assumed size distributions were used to represent the

feed to the closed-circuit grinding simulator. The simulated

capacities are given

in

Table 6. Other conclusions that were

reached

on

the basis

of

the simulation

of

the first circuit

alternative are set

out

below.

Run of mine ore

o

leaching

Fig. 5 Flowsheet of alternative crushing circuit with two- or three

stage tooth-roll

crushing

Table

6 Simulated capacity

of

circuit for each type

of

ore

under

existing operating conditions

Ore type

Size, mm

Capacity,

t h

SC

SC

-25

-12

50.4

57.0

AS1

ASl

-25

-12

58.3

62.7

AS2

AS2

-25

-12

50.0

55.5

1) If

the mill were operated

under

eXlstmg operating

conditions and with a feed

of

-25

mm,

the expected capacity

would be between 50

and

58

tIh,

depending

on the

fineness

ofthe

feed.

(2) Use of a finer feed

of

-12

mm

would increase capacity by

8-15 .

(3) Although it is known that mill rheology has some effect

on

the model parameters,

8

on

the assumption

that

this would

not

be significant increasing the pulp solids ratio in the mill

from

64

to

75

(i.e. to

the

design value) would give a

further capacity increase

of 10-12 .

(4)

More

efficient hydrocyc1oning would provide

an

addi

tionaI1O-15

increase in capacity.

(5)

Under

the design conditions and with a -12-mm feed the

mill could achieve its

design,

capacity.

Under the

same

conditions a coarser feed size

of -25

mm

would reduce

capacity

to 58-62 tIh.

Circuit alternative

2

The

circuit

that

was considered

in

the second alternative is

shown in Fig, 6. After primary crushing the ore would be

subjected

to

scrubbing, followed

by

screening

of

the

-3-mm

fraction. Fines would

be

sent directly

to the pump sump and

the

coarse fraction would go through secondary crushing

(-50

mm) and closed-circuit tertiary crushing. Crushed ore

would

be

stockpiled

or

fed directly to the mill.

Fig. 6 Flowsheet

of

alternative crushing circuit with washing drum

This

alternative was examined by considering three

variants.

h e s ~ e

were: first, the fiowsheet as described above;

second, feeding only the crushed ore to the mill; and, third,

feeding only the

-3-mm

fraction to the

pump

sump.

The

last

two variants were studied because it was thought

that

the

plant might process the fines

(-3 mm)

during the

summer

season

and

the coarse fractions in winter. All three variants

were investigated by applying the existing and design

operating parameters described above.

The

measured

product-size distribution from

the

primary jaw crusher was

use to represent the feed to the washing and crushing circuit

simulator, which consisted

of

the models

of

the washing

drum,

the secondary

and

tertiary cone crushers

and

c1osed

circuit screening.

The

products

of

this circuit

then

constitu ted the feeds to the grinding-circuit simulator.

When

the simulations were

run

the capacity

of

the circuit

was found to

be much

lower

than

expected, especially for the

circuits treating - 3-mm fractions. Examination

of

the size

distribution data for the underflow revealed that this was due

to the fact that the use

of both

existing

and

design partition

coefficients in the simulations resulted in a

much

finer

product than desired. The value

of

d

soc

was, therefore,

increased to

90 lID

and the desired product size was

obtained.

The

results are presented in

Table

7. As expected,

it

was found

that if

the ore could be crushed

down

to 12

mm,

the design capacity would be reached. When only the fines

(-3

mm)

were fed the capacity exceeded the design capacity.

Since the circuit is

not

capable

of

handling more

than

70

tIh,

the

capacity would

then

have to

be

reduced

by

decreasing

the

circulating load ratio.

f

only the crushed ore wcre fed, the

capacity would

be

around

30-54 tIh. It

also appeared that as

the mill

product

became finer the importance

of

efficient

hydrocyc1oning increased.

C37



Table

7 Simulated capacity

of

circuit for washed ore

Feed

and

operating conditions Capacity, tJh

Exist ing conditions

Combined

feed 48.0

Fines only (-3

mm) 78.0

Crushed ore only 30.0

Design conditions

Combined feed 74.5

Fines only

(-3

mm)

115.0

Crushed

ore only

54.0

However, the simulation studies also indicated some

potential problems,

an

example being that diluted pulp from

the washing

drum

couId make it difficult to maintain the

desired water balance around the circuit. Small fluctuations

in the water content could result in considerable disturbances

to

normal operation. Although two-stage hydrocydoning

of

this stream could provide a solution, such an alternative was

not thoroughly investigated. Another potential source

of

problems was control

of

the

product

size.

If

different modes

of operation-Leo

using only fines

or

crushed ore as

feed-are

to

be implemented, operating parameters will have

to be

adjusted

in

accordance with the size distribution

of

the feed.

Conclusions

Several conclusions were drawn from the simulation study

of

the grinding circuit

at

the

G i i m i i ~ k o y

silver plant.

1)

Although the kinetic parameters obtained from the

laboratory investigation were good indicators

of

the grind

ability

of

the ore types, they failed to provide reasonable

parameters for use in full-scale simulation

of

the grinding

circuit.

2) The use of wet and dry conditions in the laboratory batch

grinding was

to

some extent reflected in the

model

para

meters.

Wet

grinding appeared

to

increase the grinding rate.

3)

On the other hand,

simulations

that

were based

on the

model parameters obtained from

an

industrial-size

batch

mill

provided satisfactory results. Consequently, they were used

in

the computer-aided study

of

two alternative circuits.

(4)

The

simulations

of

the two alternative circuits indicated

that both were capable

of

solving the problem

of

under

capacity.

t

must

be emphasized, however, that in the work

described here the alternatives were studied solely from

that

point

of

view-i.e. in

terms

of

their effect

on

capacity.

Following

an

examination

of

the third alternative primary

autogenous/semi-autogenous grinding) a final decision would

also have to take into account design

and

cost parameters,

which have

not been

considered here.

After completion

of

the present study the plant engineers

increased the ball charge to

33

and improved the hydrocy

cloning. Following these modifications another two sampling

surveys were carried

out

and, as predicted, a capacity

of

approximately 60 tIhImill with feeds

of -25 mm

was recorded.

Later the management

ruled

out

the first alternative since the

company

that

produces tooth-roll crushers was unable

to

guar

antee

that

tertiary crushers would operate in the

12-mm

range.

On the

basis

of

an

assumed product-size distribution for semi

autogenous/autogenous

SAG/AG)

milling the effects

of

the

third alternative were also investigated.

It

was predicted

that

the SAG/AG milling alternative would produce

an

ample

capacity

of

20-40 .

Because

of

the expected excess water

problems in the washing drum alternative the management

decided to investigate the SAG/AG miling alternative further

by carrying

out pilot-plant testing.

C38

cknowledgement

The

authors wish to express their gratitude to Etibank,

Turkey, for financial assistance for the study

and

for

permission

to

publish the results. They are also grateful to

Dr. Selahattin Anac for his invaluable views and fruitful

discussions.

References

1.

Ramirez-Castro

J and

Finch

J

Simulation

of

a grinding circuit

change to re duce lead sliming,

IM

Bull.,

73,

1980,132-39.

2.

Laplante A. R Finch

J A.

and

del Villar

R

Simplification

of

grinding equation for plant simulation. Trans. Instn Min. Metall.

Sect. C: Mineral Process Extr. Metall.), 96, 1987, C108-12.

3. Finch J A. and Ramirez-Castro J Modelling mineral size

reduction

in the closed-circuit ball mill

at

the

Pine

Point

Mines

concentrator. Int. J. Min. Process., 8,1981,61-78.

4.

Furuya

M. Nakajima

Y.

and Tanaka

T.

Theoretical analysis of

closed-circuit grinding system based

on comminution

kinetics. Ind.

Eng. Chem.

Process

Design Develop.,

10,1971, no. 4,449-56,

5. Flintoff B. C. Plitt L R

and Turak

A.

A.

Cyclone modelling: a

review

of present

technology, IM Bull., Sept. 1987,

39-50.

6.

Whiten W.

J Simulation of crushing plants with models

developed using multiple spline regression, APG01W

972:

proceedings of the 10th international symposium, Johannesburg

Salamon

M. D.

G.

and

Lancaster F. H.

eds Gohannesburg: South African

Institute

of

Mining and

Metallurgy, 1973),

317-23.

7. Ersayin S. et al.

The

effect of laboratory

wet and

dry grinding

on

the kinetic characterization of grinding circuit,

II.

Symposium on

appli<;ation

of

mathematical methods and computers in geology, mining

and

metallurgy, Belgrade, Yugoslavia, 1991. Preprint.

8. Austin L G. K1impell R R Luckie P. T.

Process engineering

of

size reduction: ball milling AIME, 1984),

chapter

15,385-407.

Authors

S. Ersaym

graduated from

the Technical University of Istanbul,

Turkey, in 1979 before working

at

a copper

mine

in southeast

Turkey.

In

1986 he obtained a

Ph.D.

degree in mineral processing

from the University of Leeds, England,

and

later joined the

Department of Mining

Engineering

at

the University of Hacettepe,

Ankara, Tur key, as a lecturer in mineral processing. His

main

interest

is

in modelling

and

simulation.

Address:


Engineering, University of

Hacettepe,

Beytepe 06532, Ankara, Turkey.

B.

Sonmez,

L

Ergiin,

B. Aksam

and

I F.

Erkal

are graduates

of the


Engineering

at

the University

of

Hacettepe.

They hold M.Sc.

degrees

in

mineral processing

and now

work as research assistants

in the department.

Documents

075 Simulation of Grinding Circuit at Silver Plant, Turkey