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Iranian Journal of Science & Technology, Transaction B, Engineering, Vol. 32, No. B4, pp 415-424 Printed in The Islamic Republic of Iran, 2008 © Shiraz University

EVALUATION OF GRINDING CIRCUIT PERFORMANCE IN ESFORDI PHOSPHATE PROCESSING PLANT*

A. DEHGHANI AHMADABADI1, Z. POURKARIMI2, M. NOAPARAST1**, S. Z. SHAFAEI3, AND E. JORJANI2

1Faculty of Mining Engineering, University of Tehran, Tel: 82084265, Tehran, I. R. of Iran, Email: [email protected]

2Department of Mining Engineering, Islamic Azad University Science and Research Campus, Tehran, I. R. of Iran, 3Department of Mining Engineering, Shahrood University, Shahrood, I. R. of Iran

Abstract– The performance of a grinding circuit in the Esfordi phosphate mineral processing plant, located in Yazd Province, Iran was evaluated. For this purpose, samples from different units including i) rod mill feed and product, ii) first hydrocyclone feed, overflow, underflow, iii) ball mill product, and iv) second hydrocyclone underflow were taken within 7 days for screen analyzing, solid percent and work index tests. From the test results, the work index of the ore was equal to 9.47kwh/t and d80 of the above samples were determined as17686.0, 274.3, 238.7, 100.9, 236.5, 206.4, 94.1 microns, respectively. In addition, the solid weight percent of the samples (except rod mill feed) were measured 55.6, 58.0, 83.6, 76.0, and 54.0 percent. According to the calculations, the reduction ratios of rod and ball mills were 64.65 and 1.15, which show high discrepancies from design values (23 and 7.5 for rod and ball mills). In the next step the diameters of the mill media were estimated. These were conducted at 12.5 and 3.5cm for rod and ball diameters, respectively. Therefore, increasing rod diameter from 8cm (current) to 12.5cm (estimated) could decrease the effective rod surface involved in breakage. This causes a reduction in the number of rod-particle impacts in a mill and results in decreasing the rod mill reduction ratio. In another way, by changing balls from 5cm (current) to 3.5cm (estimated), the breakage in the ball mill is increased and therefore its reduction ratio is higher. In addition, the residence time distribution of particles in rod and ball mills were measured using NaOH as a tracer in mill feeds and measuring pH in mill output in sequential times. The mean residence time of particles in rod and ball mills were equal to 16.68 (maximum 51) and 2.07 (maximum 20) minutes, respectively, which can be considered as one significant reason for the high and low reduction ratios in rod and ball mills.

Keywords– Esfordi, phosphate, reduction ratio, work index, residence time, rod mill, ball mill

1. INTRODUCTION

The Esfordi phosphate processing plant is located in Yazd Province, Iran, 35km far from Bafgh city, with about 16mt of 13.9%P2O5 reserves. The run-off-mine contains about 10-16%P2O5 mineral, which is reported by the processing plant. The major units of this plant are crushing, grinding, magnetic separation, and flotation, which produces concentrate with about a 37.5-39%P2O5.

According to the flowsheet of the Esfordi plant, ores after a grizzly with 60cm*60cm aperture enter the jaw and cone crushers to achieve d80=14mm. It then goes to the grinding section. The feed rate to crushing section is 100t/h, and for the plant after the rod mill is 57.5t/h. The grinding circuit includes two mills, i) Rod mill using media (rods) of 8cm diameter, and ii) Ball mill using 5cm diameter balls. The product of the rod mill (stream no.1) with about d80=600 microns is reported to the first hydrocyclone, ∗Received by the editors November 10, 2007; Accepted April 16, 2008. ∗∗Corresponding author

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having a cut size of 100 microns. The first hydrocyclone underflow (stream no. 5) with d80=750 microns goes to the ball mill for further grinding. The ball mill yields a d80=100 microns product (stream no. 6) which is circulated to the first (previous) hydrocyclone to separate the 100 micron particles. The overflow of the first hydrocyclone (stream no. 4) is deslimed in the second hydrocyclone to remove particles finer than 10 microns as overflow (stream no. 7) which reports to the tailing dam. The underflow of the second hydrocyclone (stream no. 8) is the final product of the grinding circuit and reports for further processing such as flotation, etc.[1]. The flowsheet of the grinding circuit in the Esfordi Phosphate Processing Plant is shown in Fig. 1.

Fig. 1. The flowsheet of grinding circuit in Esfordi phosphate processing plant

The main problem of the Esfordi grinding circuit is the low reduction ratio of its ball mill, which is just less than 2 instead of 7.5 (750/100) based on the design documents. Thus it could, accordingly, be postulated to remove the ball mill from the Esfordi comminution circuit, so it was attempted to verify the rod and ball mills performance. To do that, a certain sampling campaign from different units was performed within seven days. The samples were taken from i) rod mill feed and product, ii) feed, underflow and overflow of the first hydrocyclone, iii) ball mill product and iv) underflow of the second hydrocyclone. Different tests were then carried out on these samples including screen analysis, work index, and solid percent of pulps in order to calculate the circulating load and some other effective parameters. The residency time distributions of the material in the rod and ball mills were individually measured to control the situation of the ground particles. In this paper the results of this research work, including methods of trouble shooting and/or omitting the Esfordi grinding circuit difficulties are presented.

2. SIZE DISTRIBUTION

Size distribution tests using different samples of seven days were performed. According to the size distribution of the samples, the d80 of each test was estimated and are presented in Table 1. The values of d80 were then implemented to calculate of reduction ratios of rod and ball mills (d80, feed/d80, product) (Fig. 2). The average values of reduction ratios for rod and ball mills were equal to 64.65 and 1.15, with standard deviation of 3.91 and 0.03, respectively. However the reduction ratios with 95% confidence level and under a normal distribution were equal to 66.765.64 ± for the rod mill, and 06.015.1 ± for the ball mill, respectively. It is, therefore, concluded that the reduction ratio of the rod mill (64.65) is extremely higher than that of the design value (23), indicating a distinct discrepancy from normal values for rod mills as well. In the case of the ball mill, the estimated value (1.15) is remarkably lower than the design value (7.5).

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Evaluation of grinding circuit performance in…

August 2008 Iranian Journal of Science & Technology, Volume 32, Number B4

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3. SOLID WEIGHT MEASUREMENT

To measure solid weight percent in different pulps, samples of pulp streams including rod and ball mills products, feed, overflow, underflow of the first hydrocyclone, and underflow of the second hydrocyclone, were collected over a period of 7 days. The specific gravity of the samples were first measured, and then using the obtained results and specific gravity of the solid, the solid weight percent for each sample were determined[2]. The specific gravity values of the samples from the above streams were equal to 3.4, 3.8, 3.2, 3.8, 3.8, 3.2g/cm3, respectively.

The solid weight percent results related to various streams during the seven sampling days are presented in Table 2. The average solid weight percent for rod and ball mills products, first hydrocyclone feed, overflow, underflow and second hydrocyclone underflow were estimated 54.75, 75.93, 57.67, 83.64, 16.00, and 53.20 percent, respectively. According to the documents of the Esfordi plant at the design stage, the rod and ball mills solid percent should be equal to 60 and those of the first hydrocyclone feed, overflow and underflow equal to 45, 25, and 75 percent, respectively.

4. CIRCULATING LOAD MEASUREMENT

The circulating load during the 7 days were estimated based on the following equation [3], using the obtained solid percent amount (Table 2) shown in Fig. 2.

100100100

100100

% ×

−−

−

−−

−

=

f

f

u

u

o

o

f

f

CL

where

f: solid weight percent of hydrocyclone feed, o: solid weight percent of hydrocyclone overflow, u: solid weight percent of hydrocyclone underflow, CL: circulating load percent.

Figure 2 shows the results of the circulating load on different sampling days, in which the variation is

considerable. The circulating load amount should be equal to 150%, according to the initial design reports, and the average value during the sampling days was 839.14% (Fig. 3), which shows a quite significant discrepancy.

Table 1. The d80 values (microns) of various streams in 7 sampling days

2nd Hydrocyclone u/f

Ball mill product

1st Hydrocyclone Rod mill

Day

u/f o/f feed product feed

117.2 215.4 258.0 99.7 257.9 280.3 18110 D1

99.1 201.4 231.5 97.5 242.2 264.0 17617 D2

100.4 215.1 249.6 131.5 246.4 270.4 17600 D3

94.8 187.5 210.0 112.5 228.2 268.0 17884 D4

83.2 218.1 250.6 98.4 240.6 291.7 16484 D5

87.6 208.8 239.5 77.4 233.8 287.5 18083 D6

76.6 198.3 216.3 89.6 221.9 257.5 18084 D7

94.1 206.4 236.5 100.9 238.7 274.2 17686 Ave.

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0

10

20

30

40

50

60

70

0 1 2 3 4 5 6 7 8Day

Red

uctio

n R

atio

Rod Mill

Ball Mill

Fig. 2. The reduction ratios values of rod and ball mills in 7 sampling days.

Table 2. The results of solid weight percent of pulp in various streams

2nd Hydrocyclone u/f

1st Hydrocyclone Ball mill product

Rod mill product

Stream

Day o/f u/f feed

68.10 26.20 80.70 59.48 72.86 62.26 D1

48.50 14.40 87.20 58.17 76.94 48.25 D2

68.40 20.10 82.50 63.16 81.41 59.29 D3

47.20 9.50 81.85 50.89 70.15 48.25 D4

66.40 9.50 84.70 51.95 78.43 55.26 D5

35.30 14.40 83.10 57.72 74.03 53.11 D6

38.50 17.90 85.40 62.35 77.68 56.82 D7

53.20 16.00 83.64 57.67 75.93 54.75 Ave.

0

200

400

600

800

1000

1200

0 1 2 3 4 5 6 7 8Day

Per

cen

t Circ

ula

ting

Loa

d

%CL

Fig. 3. The calculated results of circulating load in 7 sampling days

5. ROD MILL FEED WORK INDEX

To investigate the variety of feed entering the Esfordi processing plant, in terms of its grindability, the Bond work index of rod mill feed samples prepared in 7 days were estimated using a standard Bond ball mill[4, 5]. The results of these tests are shown in Fig. 4. The average work index in 7 days was 9.47kwh/t with a 0.56 standard deviation. Therefore the rod mill feed work index was 56.047.9 ± with a 95% confidence level. The work index of Esfordi was considered equal to 7.5kwh/t in design documents [1].

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8

8.5

9

9.5

10

10.5

11

0 1 2 3 4 5 6 7 8Day

Wi (

kwh

/t)

Work Index

Fig. 4. The work index results of rod mill feed samples in 7 days

6. APPROPRIATE MEDIA ESTIMATION

Due to the results obtained from different experiments on prepared samples from the 7 days from various streams, it is now clear that there are some significant discrepancies between the design values and the tests results. For example the reduction ratios for rod and ball mills were estimated 64.65 and 1.15, respectively, but should have been about 23 and 7.5. In addition, the rod mill feed d80 was 17686 microns instead of 14000, and its work index was tested equal to 9.47kwh/t, as it was 7.5kwh/t in the design documents [1]. Therefore, according to these changes in operating conditions, it was attempted to estimate the mills media (rods and balls diameters) based on the experiments results, in order to compare with current media diameters. It should be noted that the diameters of the rods and balls used in rod and ball mills are currently 8cm and 5 cm, respectively.

a) Rod diameter estimation

The following equations are usually applied to estimate the largest diameter of rods, as the initial

charge, and for the make-up charge [2, 6]. The first equation is as follows [6]:

5.0

5.075.0

)281.3(10016.0

×××

×=DC

WSFR

S

ig

where

R: diameter or rod (mm), F: feed size 80% passing (micron), wi: work index (kwh/t), Sg: specific gravity (kg/m3), Cs fraction of critical speed, D diameter inside shell liners (m),

Another equation which has been used by Bond to calculate the appropriate rod diameter of a rod mill

is as follows (where R is maximum diameter of rods that are used as media) [2]:

××

×××=

DC

WSFR

S

ig

1001000

77.2 75.0

where

R: rod diameter (m)

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wi: work index (kwh/t) F: feed size 80% passing (m) Sg: specific gravity of ore (kg/m3) D: mill inside diameter (m) Cs: fraction of critical speed

In this paper the second equation was implemented to determine the rods with convenient diameters

required for the Esfordi rod mill. The Esfordi rod mill with a diameter of 2.4m, has a 19rpm rotation speed. The results of sampling within 7 days and performing various experiments on the samples indicated that the d80 of rod mill feed was equal to 17686 microns, and its bulk density 2.16gr/cm3. Furthermore, the ore work index was estimated to be 9.47kwh/t. Thus applying these data, the diameter of the required rods was estimated as follows:

6960.030.27

1930.27

4.2

3.423.42 ==⇒=== SC CD

N

cmR 58.134.26960.0100

907.047.92160017686.0

1000

77.2 75.0 =

××××××=

Moreover, due to analyzing the operational data of the plant during the previous years, the d80 of the rod mill feed was obtained at 13379 microns. Thus using this data, the diameter of the rods was estimated as follows:

cmR 01.114.26960.0100

907.047.92160013379.0

1000

77.2 75.0 =

××××××=

As it is observed, using two series of data, two diameters of: i) 13.58cm, based on the results of 7

days sampling and performing different experiments, and ii) 11.01cm based on the results from plant operation in previous years were achieved. Therefore, the average of these values yielded rods 12.30cm in diameter. It is worth mentioning that the current diameter of rods is 8cm, which is remarkably smaller than the estimated one (12.5 cm). By changing the current rods (8cm) with larger ones (12.5cm), the surface of the rods decrease and the probability of the rod-particle is accordingly decreased. This prevents overgrinding of particles in the rod mill, and therefore, its reduction ratio decreases.

b) Ball diameter estimation

There are some equations which enable the diameters of balls required in ball mills to be achieved [2,

6]. The equation to estimate the largest diameter of balls as the initial charge and for the make-up charge is as follows [6]:

34.0

5.0

5.0

)281.3(100)(4.25

××

××=

DC

WS

K

FB

S

ig

where

B: diameter of ball (mm), F: feed size 80% passing (microns), wi: work Index (kwh/t), Sg: specific gravity (kg/m3), Cs: fraction of critical speed, D: diameter inside shell liners (m), K: constant, equal to 350 for wet-overflow-closed circuit,

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Bond has proposed the following equation for determining the maximum ball diameters in ball

mills[2], and this equation was implemented for further ball diameter estimation:

3

1

5.0

100

××

×××=

DC

WSFKB

S

ig

where:

B: ball diameter (m) wi: work index (kwh/t) F: feed size 80% passing (microns) Sg: specific gravity (kg/m3) D: mill inside diameter (m) Cs: fraction of critical speed K: constant equal to 0.111 for wet milling

The ball mill in the grinding circuit of the Esfordi processing plant has a 2.4m diameter, and a rotation

speed of 21rpm. The d80 of the ball mill feed is 750microns (according to the design value). Therefore, the ball diameter required would be estimated as follows:

7691.030.27

2130.27

4.2

3.423.42 ==⇒=== SC CD

N

cmB 64.14.27691.0100

907.047.9216000075.0111.0

3

1

5.0 =

××××××=

According to the results of the above calculation, it is possible to apply the 2cm diameter balls. With

regard to the fact that the current balls have a 5cm diameter, using balls with a 2cm diameter could produce more slime, as the calculated balls have a higher surface. However, another equation was used to calculate the ball diameter which is as follows [7]:

×

×

×

××

=m

C

f

sim N

N

DK

WdD

ρρρ 7800

5.05.0

25.0

80

where

d80: feed size 80% passing (microns) Dm: new calculated ball diameter (m) D: extra ball diameter in the mill (m) wi: work index (kwh/t) K: constant equal to 0.46 for ball milling

sρ : specific gravity of solid (kg/m3)

fρ : specific gravity of feed (kg/m3)

mρ : average of balls specific gravity (kg/m3)

N: rotation speed (rpm) Nc: critical rotation speed (rpm)

By implementing data in the new equation:

cmDm 45.44650

7800

21

3.27

3500

2160

05.046.0

907.047.900075.05.05.0

25.0=

×

×

×

×××=

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It is observed that under the new condition, the estimated ball diameter was calculated equal to 4.45cm. As is clear, two different equations yielded two different values of ball diameter, 2cm and 4.45cm. Therefore, the average of these values presented 3.25cm and/or 3.5cm as ball the diameter.

7. RESIDENCY TIME DISTRIBUTION OF PARTICLES IN ROD AND BALL MILLS

In order to measure the particles residency times in rod and ball mills, it was attempted to add 25kg of NaOH to the feed of the mills. The amount of NaOH was then determined in the mill output by measuring pH within different sequential times. In fact, NaOH was the tracer in the mill to specify the residency times in the rod and ball mills. In the next step, the residency time distribution was obtained using a simulation software (simulation of residence time distribution in open and closed circuits, version 1.1) [8]. Average residence time values of particles in the rod and ball mills were estimated 16.68 (maximum 51) and 2.07 (maximum 20) minutes, respectively [9]. The results obtained are presented in Figs. 5 and 6.

8

9

10

11

12

13

0 5 10 15 20 25 30 35 40 45 50 55

Time (min)

Co

nce

ntr

ate

(P

H)

Output

Fig. 5. The residency time distribution particles in rod mill

7

8

9

10

11

12

13

14

0 5 10 15 20 25

Time (min)

Co

nce

ntr

ate

(P

H)

Input

Output

Fig. 6. The residency time distribution of particles in ball mill (closed circuit)

It is observed that the particles residence time in the rod mill is considerably longer than those of the

ball mill, and is even significantly longer than normal rod milling (usually about 10 minutes). This means that particles in the rod mill have longer residence time, and accordingly, its reduction ratio is remarkably higher than those in the ball mill.

8. RESULTS AND DISCUSSION

To study the grinding circuit of the Esfordi Phosphate Processing Plant, samples were collected over 7 days from different streams including rod mill feed and product, first hydrocyclone feed, overflow and

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underflow, ball mill product and second hydrocyclone underflow. These samples were tested for screen analyzing and solid percent of pulp work index. The results obtained from screen analyzing of the above streams indicated that their d80 for the period of sampling was 17686.0, 274.2, 238.7, 100.9, 236.5, 206.4, 94.1 microns, respectively. In addition, the average reduction ratios in rod and ball mills were 64.65 and 1.15. According to the 95% confidence level, the reduction ratios of rod and ball mills were estimated as 64.65± 7.66 and 1.15± 0.06.

Samples of different streams such as; rod mill and ball mill products, first hydrocyclone feed, overflow and underflow, and second hydrocyclone underflow were tested to measure the solid weight percent of pulp. Their average values were achieved as 54.57, 75.93, 57.67, 83.64, 16.00, and 53.20%. The work index test was carried out using a rod mill feed sample with an average value of9.47kwh/t. According to the 95% confidence level, the work index of ore was estimated 9.47± 1.10kwh/t. It should be noted that based on the design documents, the reduction ratio of rod and ball mills were considered 23, and 7.5, respectively. However there is a high discrepancy between the calculated values from the sampling data (64.65 and 1.15) and those of the design (23 and 7.5). This indicates that the rod mill reduction ratio is much larger than those at normal condition, and ball mill reduction ratio is much smaller than those at normal level. In addition, based on the design values for the solid weight percent of pulp in rod and ball mills products, the first hydrocyclone feed, overflow, and underflow streams should be equal to 60, 60, 45, 25, and 75 percent, however, there were some differences with the measured values. The work index of ore was estimated 9.47kwh/t, and it was 7.5kwh/t in the design stage.

The circulating load in the ball mill was estimated by using the amounts of solid weight percentage. Its average was 839.14%, which describes the increase in the circulating load, and should be decreased by adjusting the hydrocyclone feed solid percent, hydrocyclone pressure, changing the cut size and the feed rate.

The rods and ball diameters were calculated using some equations which yielded 12.5cm and 3.50cm for rods and balls, respectively. It should be noted that the current values for rods and balls are 8cm and 5cm. Therefore, using rods with larger diameters cause a lower surface for particle breakage, and thus a lower reduction ratio in the rod mill, which means the larger product might be reported to ball mill. It is postulated that changing the rods and balls could also change the reduction ratios to lower value for the rod mill, and higher values for those of the ball mill. Hence, in the first step, the rod mill media diameter can be changed from 8cm to 12.5cm, and then new balls with 3.5cm diameters should be used instead of 5cm.

Measurement of particles residence time in rod and ball mills yielded very surprising results, 16.68 (maximum 51) and 2.07 (maximum 20) minutes for rod and ball mills, respectively. These results indicated that a very significant difference in rod mill residency time with those of ball mill. Thus one reason for the high reduction ratio in the rod mill could stem from its high residency time.

9. CONCLUSION The performance of a grinding circuit in the Esfordi Phosphate Processing Plant was investigated. The main reason for this study was to verify the reasons for the ball mill low reduction ratio, and its high ratio in the rod mill. To perform this, some verification was done such as: a) review of initial design documents and manuals, b) gathering past data (using reports of plant), c) current condition of grinding circuit by sampling in seven days. This study yielded some significant results which are listed as follows: - Average amount of d80 of the rod mill feed and product, input, overflow and underflow of the first hydrocyclone, ball mill product and underflow of second hydrocyclone in 7 days were 17686.0, 274.2, 238.7, 100.9, 236.5, 206.4, 94.1 microns, respectively. These amounts were measured and averaged for

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these streams using archive data gathered from last year which were 13379.0, 290.0, 248.3, 88.9, 253.7, 221.7, 91.4 microns. - Average of reduction ratios for rod and ball mills were measured 64.4 and 1.15, according to sampling results. However these amounts should be 23 and 7.5 according to the design documents. - The work index of the rod mill feed was measured 9.47kwh/t and the design value 7.5kwh/t. - Average amount of the circulating load in the ball mill was obtained at 839.14%, which showed a high discrepancy with the design value (150%). - According to the data gathered from the samples, an appropriate media diameter of 12.5cm and 3.5cm for rod and ball mills was obtained, respectively. The current rod and ball are 8cm and 5cm. This change in rod and ball diameters can cause a decrease in the reduction ratio of the rod mill and increase the ball mill reduction ratio. - Average residence time of particles was calculated 16.68 (maximum 51) and 2.07 (maximum 20) minutes for the rod and ball mill, respectively. The high value for rod mill residence time can be one reason for its high reduction ratio. Acknowledgements- This research work was supported by Yazd-Iran Mineral Research Center. The authors would like to express their appreciation for all their help and efforts conducted in this project.

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2. Weiss, N. L. (1985). SME Mineral Processing Handbook. SME Publishers.

3. Wills, B. A. (2006), Mineral processing technology. 7th edition, Butterworth Heinemann.

4. Barratt, D., Sherman, M. (2002). Factors which influence the selection of comminution circuits. in Book

entitled: Mineral Processing Plant Design, Practice and Control, edited by: Mular, A.L., Doug, N.H., Derek,

J.B., SME Publishers, pp. 539-565.

5. Bond, F. C. (1961). Crushing & grinding calculations: Part I. British Chemical Engineering, Vol. 6, No. 6, pp.

378-385.

6. Rowland, Jr., C. A. (2002). Selection rod mills, ball mills and regrind mills. in Book entitled: Mineral Processing

Plant Design, Practice and Control, edited by: Mular, A.L., Doug, N.H., Derek, J.B., SME Publishers, pp. 710-

754.

7. Kelly, E. G. & Spottisswood, D. J. (1989). Introduction to mineral processing. Mineral Engineering Services.

8. Javadi, F. & Banisi, S. (1997). Simulation of residence time distribiution in open and closed circuits, Version

1.1. University of Shahid Bahonar Kerman, Iran.

9. Gharabaghi, M., Noaparast, M. & Shafaei Tonekaboni, S.Z. (2007). Lar mountain phosphate ore processing

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