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Breakage Characteristics of Coal-Water Slurries Heechan Cho and Mark S. Klima* Mineral Processing Section, 115 Hosler Building, The Pennsylvania State University, University Park, Pennsylvania 16802 Received January 31, 1996X
Grinding tests were conducted under various conditions to characterize the breakage characteristics of coal-water mixtures. Five bituminous coals were ground in a conventional tumbling ball mill for different times at solids concentrations of 40-70% by weight, with and without the use of a dispersant. For a given coal, the grinding rate decreased as the solids concentration increased, whereas grinding at a higher solids concentration produced a higher proportion of fines, giving a flatter distribution. However, even at the highest slurry density, the product size distribution was steeper than the broad size distribution recommended for maximum solids loading and minimum viscosity. Additional grinding tests were performed in a stirred ball mill, aiming at a two-stage grinding method for producing coal-water mixtures. On the basis of the breakage parameters obtained from the laboratory tests, simulations were performed to analyze an industrial scale two-stage grinding circuit.
Introduction The use of coal-water mixtures in combustion requires high-density slurries consistent with slurry handling and storage. The particle size distribution is an important factor for determining the coal loading and subsequent viscosity. It is generally considered that either a bimodal or a broad size distribution is required to produce a highly loaded coal-water slurry with minimum viscosity. Ball mills have been widely used for the preparation of coal-water mixtures because of their availability and capacity. However, the grinding of high-density slurries is not completely understood due to the complexity arising from the dependence of breakage properties on the slurry characteristics such as the solids concentration, the fineness of the size distribution, and the chemical environment.1 It has been shown that the grinding rate changes in a complex manner with the slurry density but generally decreases as the slurry becomes thick and viscous.2 Also, the shape of the particle size distribution is affected by the slurry density, with more fines produced at higher slurry densities. Hence, there is a solids loading that would provide the best tradeoff between the grinding rate and the production of fines during the preparation of coalwater mixtures. However, there is a lack of information available regarding the recommended grinding conditions for coal-water slurry preparation. Furthermore, such conditions are likely to be coal specific. Therefore, a study of laboratory ball mill grinding of coal-water slurries was carried out with five bituminous coals to identify the grinding behavior under different slurry conditions. The effects of a dispersant (Coal Master A-23M) on grinding in dense suspensions were also * Author to whom correspondence should be addressed [telephone (814) 863-7942; fax (814) 865-3248]. X Abstract published in Advance ACS Abstracts, August 15, 1996. (1) Klimpel, R. R. Laboratory Studies of the Grinding and Rheology of Coal-Water Slurries. Powder Technol. 1982, 32, 267-277. (2) Tangsathitkulchai, C.; Austin, L. G. The Effect of Slurry Density on Breakage Parameters of Quartz, Coal and Copper Ore in a Laboratory Ball Mill. Powder Technol. 1985, 42, 287-296.
analyzed. In addition, separate grinding tests were conducted on one of the coals using a stirred ball mill. On the basis of the laboratory results, simulations were conducted to analyze grinding circuits for producing the recommended size distribution for coal-water mixtures. Experimental Section Test Coals. Five test coals were used in this study, which were chosen as part of a project investigating the development of coal-based technologies for Department of Defense facilities.3 Four of the coals (Taggart, Lower Kittanning, Indiana, and Pittsburgh) were precleaned by conventional techniques, while the Upper Freeport seam coal was obtained directly from an auger mining operation. A sample of each coal was passed through a jaw crusher to reduce the top size from approximately 38 to 13 mm and then crushed to -1.2 mm (16 U.S. mesh) using a hammer mill. These samples were used as feed for the ball mill grinding tests. Samples of the Taggart seam coal were ground further to -75 µm (200 mesh) in a high-speed pulverizer for use as feed for the stirred ball mill tests. The proximate and total sulfur analyses, along with the Hardgrove grindability index (HGI) for each coal, are given in Table 1. Test Procedure. The first set of grinding tests was carried out in a batch steel ball mill, which was 190 mm in diameter by 175 mm long. The mill contained six lifters welded at equal distances around the periphery of the mill. The mill was loaded to a bulk volume of 1570 cm3 (30% loading) with 25 mm diameter steel balls. All tests were run at 72 rpm (70% of critical speed) with a constant slurry volume loading of approximately 627 cm3. Variations in the percent solids (4070% by weight) were achieved by changing the water-to-solids ratio to give 627 cm3 of total slurry volume, assuming a coal density of 1.4 g/cm3. Grinding tests were conducted for various times using a new sample for each test. For the Taggart seam coal, additional tests were run to obtain torque measurements from which the power draw was estimated for various grinding conditions. The ground products were wet screened at 38 µm (400 mesh). The (3) Miller, B. G., et al. The Development of Coal-Based Technologies for Department of Defense Facilities. Annual Report for the U.S. Department of Energy, Pittsburgh Energy Technology Center, 1993; Contract DE-FC22-92PC92162.
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Table 1. Proximate Analysis, Total Sulfur, and Hardgrove Grindability Index of the Test Coals coal seam
proximate analysis moisture, % volatile matter, % fixed C, % ash, % total S, % Hardgrove grindability index
Taggart (Wise Co., Virginia)
Lower Kittanning (Jefferson Co., Pennsylvania)
Indiana VII (Sullivan Co., Indiana)
Upper Freeport (Jefferson Co., Pennsylvania)
Pittsburgh (Greene Co., Pennsylvania)
1.7 34.2 62.1 2.0 0.59 47
2.0 33.1 55.5 9.4 0.78 68
12.7 30.3 49.4 7.6 0.41 51
1.3 28.0 59.1 11.5 3.07 76
2.5 34.7 56.1 6.9 1.76 56
Figure 1. Product size distributions after grinding the -16 mesh sample of the Taggart seam coal in the ball mill at 50% solids (by weight) for various times without a dispersant. +400 mesh material was dried and then screened in a RoTap sieve shaker to determine the product size distributions. A 0.6 L volume, horizontal stirred ball mill was also used in this study (Dyno Mill, W. A. Bachofen, Manufacturing Engineers, Basle, Switzerland). The mill consists of a stainless steel chamber and a stirrer shaft fitted with four polyurethane agitator disks. Although the slurry can be recirculated through the mill, all tests were conducted in the batch mode. The mill was loaded to a bulk volume of 480 cm3 (80% loading) with 1 mm diameter steel beads. A shaft speed of 3000 rpm was used for all tests with a constant slurry loading of 252 cm3 (40% solids by weight). The mill was attached to a Watt meter, and the power draw was measured. The coal was ground for various times using a new sample for each test. The products were wet screened at 400 mesh. The +400 mesh size fractions were dried, followed by dry screening. Any -400 mesh material was combined with the wet-screened fines and then analyzed using a Microtrac SRA and SPA.
Results and Discussion Figure 1 shows a typical set of size distribution data for various grinding times in the ball mill. As the grinding proceeded, the curves shifted in a parallel manner to finer sizes. Figure 2 shows the variation of the 80% passing size with grinding time for the Taggart seam coal. It can be seen that for a fixed grinding time, the 80% passing size increased as the solids concentration increased, as indicated by the higher curves. Also, above 60% solids loading, the curve becomes flatter, indicating a slowing of the grinding rates. The addition
Figure 2. Variation of the 80% passing size with time for the Taggart seam coal ground in the ball mill at various solids concentrations.
of 0.5% (by weight) of the dispersant increased the grinding rate, attaining a much smaller 80% passing size. However, a further increase in reagent dosage did not improve the rate of grinding. Similar results were obtained for the other coals. It should be noted that changing the slurry density also changed the shape of the size distribution. Therefore, one characteristic size of the product is not sufficient to describe the size distribution completely. This will be discussed later. The grinding time required to reach a product size of 80% passing 200 mesh was estimated by interpolation or extrapolation of the data (such as given in Figure 2) for each coal and plotted in Figure 3. It is apparent that the required time increased with increasing solids concentration. A rapid increase in the grinding time was observed in the neighborhood of 60% solids for all but the Indiana seam coal, which increased at a much lower solids loading. This increase indicates a slowing of the breakage rate with increasing slurry concentration, resulting from the increase in slurry viscosity. As shown previously, the required grinding time can be significantly reduced by the addition of the reagent. Figure 4 shows a plot of the grinding time required to achieve a product size of 80% passing 200 mesh versus the HGI of each coal. Except for the Indiana seam coal, a more or less straight-line relationship can be seen. For these four coals, which are all from the Eastern coal province, the HGI can be used as a criterion to estimate the time required to produce a specific product size.
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Figure 3. Time required to reach a product size of 80% passing 200 mesh for the five coals.
Figure 4. Relationship between the time required to reach a product size of 80% passing 200 mesh and the HGI of each coal.
The power draws for different solids concentrations are shown in Figure 5. For a solids concentration of 40%, the power draw remained relatively constant, indicating that the tumbling action of the balls inside the mill was stable. At 60% solids, the overall power draw increased because of the higher solids loading. However, a drop in power draw occurred after 20 min of grinding as the slurry became more viscous due to the increased production of fines. Above 60% solids, the overall power draw decreased. Apparently, the rheological properties of the slurry had changed such that the balls no longer tumbled properly. At high solids concentrations, the balls begin to stick to the mill shell, reducing the effectiveness of grinding as indicated by the lower power draw.4 Further grinding resulted in a continuous decrease in power draw. Also, the “noise” (4) Tangsathitkulchai, C.; Austin, L. G. Slurry Density Effects on Ball Milling in a Laboratory Ball Mill. Powder Technol. 1989, 59, 285293.
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Figure 5. Power measurements for the Taggart seam coal at various grinding conditions in the ball mill.
Figure 6. Product size distributions after grinding the -200 mesh sample of the Taggart seam coal in a stirred ball mill at 40% solids (by weight) for various times without a dispersant.
level in the torque measurement at this solids concentration became very high after 15 min of grinding, indicating that the rotation of the mill was unstable. It can be seen that the addition of the dispersant stabilized the milling action, resulting in a constant power draw over the entire grinding period. However, the overall power draw was lower than at 40% solids because of the higher slurry viscosity, which still affected the tumbling action of the balls. This observation correlates well with the grinding data shown in Figure 2. Figure 6 shows the size distribution of the products after the Taggart seam coal at 40% solids was ground in the stirred ball mill for various times. It can be seen that the size distribution becomes progressively finer in a parallel manner, a trend very similar to the typical batch grinding data in a ball mill. The power measurement showed that except for the initial period of grinding, the power draw remained fairly constant with an average value of 0.132 kW compared to 0.005 kW for the ball milling.
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Figure 7. Variation of the median size with specific grinding energy for the ball mill and stirred ball mill.
Figure 7 relates median particle size with energy input for both mills. In this log-log plot, the size reduction is correlated linearly with the energy input by a single line, known as the Charles energy-size relationship.5 In this case, a size reduction ratio of 10 requires 80 times more energy. Analysis of the Product Size Distribution. In many cases, the energy-size relationship can be used to control the fineness of the products, but it does not provide information about the size distribution, which may be required in subsequent processes. A better understanding of coal breakage characteristics under various conditions is achieved through the kinetic grinding model.6 This model is also useful in the design and optimization of grinding circuits. The kinetic grinding model is based on the definition of two functions: the breakage rate function and the breakage distribution function.6 The breakage rate function, Sj, is the specific breakage rate of material in the jth size interval. The breakage distribution function, bij, is the fraction of material broken out of the larger jth size interval and into the ith size interval. For batch grinding, the above definitions of S and b lead to the size-mass balance equation
Figure 8. Experimental and simulated size distributions for the Taggart seam coal ground at 50% solids in the ball mill. Table 2. Breakage Parameters for the Taggart Seam Coal at Various Solids Concentrations concn, %
a
R
φ
γ
β
40 50 60 65a 70a
3.37 3.40 1.88 1.54 0.850
1.37 1.37 1.37 1.37 1.37
0.43 0.32 0.39 0.47 0.52
0.67 0.58 0.51 0.50 0.44
5.0 5.0 5.0 5.0 5.0
a
With 0.5% dispersant.
sented in cumulative form that fits the empirical function6
Bij ) φ(xi-1/xj)γ + (1 - φ)(xi-1/xj)β
(3)
where Si is the specific rate of breakage for particles in size interval i (xi), x0 is a standard (reference) particle size, and a and R are constants for a given material and set of grinding conditions. The b values can be repre-
where β, γ, and φ are constants that depend on the material being ground. To provide the best estimate of the experimental parameters for the S and B functions, the one-sizefraction technique should be used.6 However, reasonable estimates can be obtained by back calculation using a nonlinear optimization scheme.7 Figure 8 compares the experimental size distributions for the Taggart seam coal ground in the ball mill at 50% solids and those computed using the back-calculated values: R ) 1.37, φ ) 0.32, γ ) 0.58, and β ) 5.0. It can be seen that there is good agreement between the experimental and computed results. The breakage parameters for the other solids concentrations are shown in Table 2. It can be seen that the breakage rate (a value) decreased at higher solids concentrations. However, the changes in the parameters for the breakage distribution function (lower γ values, higher φ values) indicated that more fines were produced at higher solids concentrations. This is more evident in Figure 9, where the breakage distribution is flatter for higher density slurries, which will affect the shape of the product size distribution accordingly.
(5) Charles, R. J. Energy-Size Reduction Relationships. Trans. AIME 1957, 208, 80-88. (6) Austin, L. G.; Klimpel, R. R.; Luckie, P. T. Process Engineering of Size Reduction: Ball Milling; SME: New York, 1984.
(7) Klimpel, R. R.; Austin, L. G. The Back-Calculation for Specific Rates of Breakage and Non-Normalized Breakage Distribution Parameters from Batch Grinding Data. Int. J. Miner. Process. 1977, 4, 7-32.
i-1
dwi(t)/dt ) -Siwi(t) +
bijSjwj(t) ∑ j)1
(1)
where wi(t) is the weight fraction of material in the ith size interval at time t. The specific rates of breakage vary with the particle size as a power function relationship given by
Si ) a(xi/x0)R
(2)
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Figure 9. Back-calculated cumulative B values for various solids concentrations of the Taggart seam coal.
Figure 10. Simulated size distributions for the Taggart seam coal to give 80% passing 200 mesh.
To compare the shape of the product size distributions at various solids concentrations for the Taggart seam coal, simulations were conducted using the backcalculated breakage parameters. Grinding times were selected that gave a product size of 80% passing 200 mesh, and the results are plotted in Figure 10. It can be seen that the size distributions fall into three distinct groups: one for slurry densities of 50% and less, one for slurry densities of 60-65%, and one for the slurry density of 70%. The higher density slurries produce comparatively more fines. The variation in the shape of the size distribution indicates that the mechanism of fracture changes when going from low to high pulp densities. These results agree with the findings noted previously.2 Similar trends were observed for the other coals. Figure 11 shows the size distributions for the five coals ground at 65% solids with 0.5% dispersant. Typically, the softer (higher HGI) coals produce a flatter size distribution, containing a bigger portion of fines. However, this trend is not observed for this high solids
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Figure 11. Size distributions for all coals after grinding at 65% solids with 0.5% dispersant to obtain 80% passing 200 mesh, along with the recommended size distribution based on the formula given by Farris.8
loading. It seems that in addition to the solids content, the coal surface properties play an important role in determining the viscosity of the coal slurry, which in turn affects the shape of the size distribution. The solid line represents the recommended size distribution for maximum solids loading given by Farris8 with a maximum particle size of 200 µm and a minimum particle size of 0.1 µm. This equation was applied simply by dividing a continuous size distribution into discrete sizes. It can be seen that the size distributions for all coals do not follow the recommended size distribution. Generally, a natural breakdown of the particles does not produce the recommended size distribution under a simple grinding process. Therefore, a circuit arrangement involving staged grinding will be needed. The simplest way to obtain the desired size distribution is by blending products from two mills. For this purpose, stirred ball milling can be used to produce the finer stream, which can be mixed with the ball mill product to obtain the required size distribution. Figure 12 shows the product size distributions after 16 min in the stirred ball mill and the ball mill product of 70% passing 200 mesh. Since the recommended size distribution of 80% passing 200 mesh lies between the two mill products, blending of the products should produce the desired size distribution. Table 3 shows the proportions required to produce the calculated size distribution. It can be seen that the lower the slurry density, the higher the proportion of the stirred ball mill products needed to compensate for the steeper size distribution of the ball mill products. Table 3 also gives the specific energy requirement for each grinding scenario. The overall specific energy decreased as the slurry density increased because the portion of the energy intensive stirred ball milling product was lower. Therefore, to obtain a flatter size distribution at the minimum energy use, the ball mill should be operated at higher slurry densities to reduce the amount of stirred ball milling required. (8) Farris, R. J. Prediction of the Viscosity of Multimodal Suspensions from Unimodal Viscosity Data. Trans. Soc. Rheol. 1968, 12, 281301.
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Figure 13. Two-stage grinding circuit for producing coalwater mixtures with a bimodal size distribution.
Figure 12. Size distributions for the following conditions: (a) after 16 min in a stirred ball mill; (b) after 30 min in a conventional ball mill; (c) 3-to-1 mixture of the conventional and stirred ball mill products; and (d) the recommended size distribution as given by Farris.8 Table 3. Specific Energy Requirements To Obtain the Limiting Size Distribution Given by Farris8 for Different Ratios of the Conventional (BM) and Stirred Ball Mill (SBM) Ground Taggart Seam Coal solids concn, % 40 50 60 65a 70a aWith
specific grinding energy, kWh/t BM SBM 5.8 4.1 7.6 4.5 7.2
308.8
mixing ratio (BM/SBM)
overall grinding energy, kWh/t
65/35 65/35 70/30 70/30 75/25
111.8 110.8 98.0 95.8 82.6
0.5% dispersant.
However, this distribution is far from ideal since the Farris equation requires discrete particle sizes with a size ratio greater than 10:1. As experimentally shown by Farris,8 the viscosity becomes increasingly higher as the size ratio of particles decreases. Therefore, a suspension with a broad size distribution will be more viscous than a binary mixture of particles of size ratio greater than 10:1. Practically, a true binary system of two discrete sizes cannot be produced by grinding, but it can be closely imitated with a size distribution having two modes via a circuit arrangement involving staged grinding. One such circuit arrangement is shown in Figure 13. In the first stage, crushed coal (e.g., -12 mesh), water, and dispersant are introduced into a ball mill, where it is ground in closed circuit. The product from the firststage grinding product is then classified. The fine stream is further ground in a closed stirred ball mill circuit to produce an ultrafine fraction. This product is then combined with the coarser size fraction, producing a bimodal size distribution. The ideal mixing ratio of coarse and fine particles for maximum packing is 60:40. Therefore, the production rate of the coarse and fine size fraction must be controlled accordingly. At the same time, the maximum size of the coarse fraction must be controlled for proper combustion while the size ratio of the coarse and the
fine fractions must be kept sufficiently large (at least 10) for maximum solids loading and low viscosity. Circuit Simulation. The prediction of the performance for such a complete grinding circuit requires the combination of the mill model and a classifier model via the appropriate mass balance equations. One such simulator is the general two-mill grinding circuit simulator developed at The Pennsylvania State University.9 In its original form, the circuit consists of two ball mills in normal closed circuit, with a circuit preclassifier and a preclassifier before the second mill. Additionally, the circuit has three dividers so that the streams of the first and second mills can be connected. Assigning the appropriate values to the classifiers and dividers allows the general two-stage circuit to be reduced to any desired circuit varying from a simple single-stage open circuit to a very complicated two-stage circuit. However, before the simulator can be used to evaluate the circuit shown in Figure 13, several modifications are needed. The scale-up procedure involving the mill sizes and operating conditions for ball milling is well established. However, the design procedure for stirred ball milling has not been completely established. It is often assumed that the specific energy is constant between the test and large mill for the same reduction ratios. The power draw of a stirred ball mill is a function of mill diameter and operating conditions. If operating conditions are selected for large mills to consume the same specific energy as the test mill, the simulation can be performed using the same breakage parameters as those determined by the laboratory milling tests. The make-up feed size distribution (i.e., crushed product) was chosen as 80% passing 700 µm with a Gaudin-Schuhmann slope of 0.9. The ball mill was assumed to have a residence time distribution equivalent to one-large/two-small fully mixed tanks-in-series with relative hold-up ratios of 0.5, 0.25, 0.25, whereas the stirred ball mill was assumed to be fully mixed. The simulations were conducted for a 1 m diameter by 1.5 m long wet-overflow ball mill operating at 70 wt % solids, a fractional ball load of 0.35, and a rotational speed 70% of critical, with a make-up ball size of 31.75 mm (1.25 in.). The relative size of the stirred ball mill to the ball mill was 1:10, operating at 40 wt % solids, 0.6 fractional ball loading, and a total charge volume (balls and slurry) of 0.9. The classifiers were assumed to be hydrocyclones. The characteristic parameters for the selectivity values at various cut sizes were determined by laboratory testing and plant sampling and are shown in Table 4. (9) Austin, L. G.; Luckie, P. T.; Yildirim, K. A Useful General TwoMill Circuit Model. Int. J. Miner. Process. 1987, 21, 205-215.
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being 16. The proportion of the coarser and the finer size fraction is about 60:40, which is an ideal composition. The size ratio can be increased by reducing the cut size for the second-stage postclassifier. This will produce a more distinct bimodal shape for the final product size distribution. However, this will also result in a decrease in the production rate for the second-stage grinding and, hence, require a larger stirred ball mill. Conclusions
Figure 14. Product size distributions around the two-stage grinding circuit. Table 4. Characteristic Parameters for the Hydrocyclone Tests Used in the Simulations cyclone O/F U/F diam, diam, diam, cut size, sharpness flow rate, mm mm mm µm index bypass L/min 25.4 50.8 254
7.1 9.7 102
2.3 3.3 76.2
17.3 39.8 148.0
0.44 0.51 0.53
0.06 0.08 0.20
16.2 37.7 1128.0
It was assumed that an appropriate operation could be obtained to give suitable underflow and overflow slurry densities and the flow rates for the desired cut sizes. Figure 14 shows the size distributions of the various streams when the Taggart seam coal was ground with a mean residence time of 4 min in the first mill and cut sizes for the first, second, and third cyclones at 150, 40, and 17 µm, respectively. The mean residence time for the second mill was determined as 3.8 min. It can be seen that the final product size distribution has a bimodal shape resulting from combinination of the firstand second-stage products. The bimodal shape, though, is less well-defined than in the batch grinding system due to the imperfect separation of cyclones and the backmixing occurring in the continuous mill. The median sizes for each stage product are 52 and 3.3 µm, respectively, with the size ratio of the coarse and finer sizes
It was found that the specific rate of breakage decreased as the solids concentration increased from 40% to 70%. For a given coal, the grinding rate decreased as the solids concentration increased. The addition of a dispersant improved the breakage rate. The shape of the size distribution also changed as a function of solids concentration, with more fines being produced at higher solids loadings. The result was that the net production rate of fines passed through a maximum. However, single-stage grinding in a conventional ball mill is unlikely to produce the shape of size distribution (i.e., broad) required for a stable coalwater mixture. Stirred ball milling was also unable to produce the shape of size distribution needed for coalwater mixtures, although the size distribution was much finer than that produced in the ball mill. A combination of the ball mill and stirred ball mill products can be used to obtain the desired size distribution. The calculated size distribution of minimum viscosity using the Farris equation, which is applied simply by dividing a continuous size distribution into discrete sizes, produces a slurry with a viscosity greater than that of a simple binary system of size ratio greater than 10:1 because of the crowding effect. Therefore, it is more practical to produce a bimodal size distribution with two modes separated by size ratio greater than 10:1 using a two-stage grinding/classification circuit. The specific energy requirement also favors the production of the bimodal system. Circuit simulations for industrial scale mills demonstrate that bigger size ratios may be required to obtain the advantage of the binary mixtures because of the imperfect size classifications in actual cyclone operations. EF960019G
