Pilot plant trial of the reflux classifier q

Transcription

1 Minerals Engineering 15 (2002) Pilot plant trial of the reflux classifier q K.P. Galvin *, E. Doroodchi, A.M. Callen, N. Lambert, S.J. Pratten Department of Chemical Engineering, University of Newcastle, University Drive, Callaghan, NSW 2308, Australia Received 31 July 2001; accepted 2 October 2001 Abstract The Ludowici LMPE Reflux Classifier is a new device designed for classifying and separating particles on the basis of size or density. This work presents a series of experimental results obtained from the first pilot scale study of the reflux classifier (RC). The main focus of the investigation was to assess the particle gravity separation and throughput performance of the device. In this study, the classifier was used to separate coal and mineral matter less than 2 mm in size. The experimental results were then compared with the performance data on a teetered bed separator (TBS). It was concluded that the classifier could offer an excellent gravity separation at a remarkably high solids throughput of 47 t=m 2 h more than 3 times higher than for a TBS. The separation performance of the RC was also better, with significantly less variation in the D 50 with particle size. A simple theoretical model providing an explanation of the separation performance is also presented. Ó 2002 Elsevier Science Ltd. All rights reserved. Keywords: Coal; Gravity concentration 1. Introduction The Ludowici LMPE Reflux Classifier, developed at the University of Newcastle, is a new device for separating particles on the basis of either size or density. The device is best compared to a hindered settling column or teetered bed separator (TBS) (Galvin et al., 1999), essentially a fluidized bed in which slower settling particles report to the overflow and faster settling particles to the underflow. However, the reflux classifier (RC) contains sets of parallel inclined plates that significantly increase the throughput of the device to a level many times higher than can be achieved in the more conventional system. The purpose of this paper is to present a series of results obtained using a 0:6 m 0:6 m pilot scale RC operated at a coal preparation plant, and to assess its gravity separation and throughput performance using a )2 mm feed coal Basis of the reflux classifier Boycott (1920) was the first to report on the dramatic effect of presenting an inclined surface to a suspension. q Presented at MEGS 01, Falmouth, UK, June * Corresponding author. Tel.: ; fax: address: cgkpg@cc.newcastle.edu.au (K.P. Galvin). Lamella settlers containing parallel inclined plates have since been used as high throughput devices to achieve a solid liquid separation. A feed slurry enters below the set of plates. As the flow passes up through the inclined channels, particles settle onto the plates and slide down, while clear water passes through to the overflow. Consider now the effect of passing fluidization water up through the system. Slower settling particles are unable to settle against the fluidization water, and are forced to separate from the faster settling particles. The higher hydraulic loading then causes the slower settling particles to emerge through the set of plates to the zone above, and hence report to the overflow. Faster settling particles are simply withdrawn through an opening in the base. This mode of operation forms the basis of the RC. With the reject suspension providing an autogenous dense medium, the separation proceeds largely on the basis of density, with the lower density particles adopting the lower settling velocity and reporting to the product overflow, leaving faster settling high density reject particles to report to the underflow. Nguyentranlam and Galvin (2001a,b) have recently investigated the interaction between a set of parallel inclined plates and a fluidized suspension, developing and validating a kinematic model of the system. The fluidization presents to each inclined channel a suspension at a uniform concentration and velocity. Fluidization rates many times greater than the terminal velocity /02/$ - see front matter Ó 2002 Elsevier Science Ltd. All rights reserved. PII: S (01)

2 20 K.P. Galvin et al. / Minerals Engineering 15 (2002) of an isolated particle can be employed before particles emerge through the other side of the inclined channel. The particle trajectory within the inclined channel develops both a motion parallel to the plates and a motion normal to the plates. Thus, the particles need only settle a short distance before they reach the upward facing inclined plate. There the particles join a concentrated sediment that slides downwards at a high rate, returning the particles to the fluidized zone below. A key finding of the second study (Galvin and Nguyentranlam, 2001b) is the potential to achieve a given suspension concentration, using a broad range of possible fluidization velocities. Thus a relatively high suspension density can be developed in the RC, even at high hydraulic loadings, which in turn permits an autogenous gravity separation to occur. Fine high density particles fail to pass through the upper inclined plates, and hence these accumulate in the system. The accumulation of such particles is ideal for effecting a high suspension density, without requiring an excessive volume fraction of solids. Conversely, conventional liquid fluidized beds, including TBS, have a one to one correspondence between the suspension concentration and the fluidization rate, and hence are forced to operate at the lowest possible fluidization rate in order to effect a satisfactory gravity separation. A schematic representation of the RC is shown in Fig. 1. The pilot scale unit referred to in this study had a square cross-section 0:6 m 0:6 m. Water flowed up through a distributor plate at the base of the vessel, suspending particles within the vessel. The feed slurry was delivered through the side of the vessel. The reject plates near the base of the unit were 0.6 m long and 100 mm apart, and were inclined to the horizontal at an angle of 60. Above the feed entry position a set of middling plates, 0.6 m long and 50 mm apart, was located. A third set of plates 1.2 m long and 30 mm apart was located near the top of the unit. Fluidized mixing zones were located between each set of plates. The three sets of plates and the corresponding mixing zones permitted reject, middlings, and overflow streams, however, in this study the middlings option was not utilized Gravity separation In gravity separation, the objective is to cause particles of a density higher than a specified set point to report to one stream, and particles of lower density to a second stream. In coal preparation, the lower density particles form a low ash, coal product, and the higher density particles a high ash, mineral matter reject. The separation condition is controlled by the suspension media, which in turn can be formed autogenously or by adding a foreign high density particulate species of suitable size. In practice, the separation is governed by the settling rates of the particles, which depend on the particle diameter, density, and the suspension medium. Therefore, the particles of a given size within a feed will separate at a different efficiency and/or density. Here we are interested in describing the particle size dependence of the separation density in a fluidized bed system. Partition curves describe the probability of a particle of given density reporting to the product stream. A typical curve is shown in Fig. 2, with the Ep efficiency index defined as ðd 25 D 75 Þ=2. At the separation density, defined by the D 50, particles have an equal probability of reporting to either output stream. Ideally, different partition curves should be reported for specific, narrow particle, size ranges so that the variation in the D 50 with particle size can be determined. All particles that have an equal probability of reporting to either output stream, i.e. D 50 particles, must exhibit identical Fig. 1. Schematic representation of the Reflux Classifier. Drawing is not to scale. Fig. 2. A typical partition curve describing the variation of partition number with relative density.

3 K.P. Galvin et al. / Minerals Engineering 15 (2002) slip velocities. In effect, they must all settle at the same velocity relative to the liquid. Moreover, those particles will form a mixed suspension, displaying essentially a zero velocity relative to the vessel, and thus settle at a common slip velocity. The upward interstitial liquid velocity, as measured relative to the vessel, must therefore be equal in magnitude to this slip velocity. Galvin et al. (1999) presented a model for describing the gravity separation produced in a conventional fluidized bed system, referred to as a TBS. One way to consider the model described in 1999 is to imagine a batch system containing particles of different size and density, all of which happen to be D 50 particles. The objective of the model is to determine the size and density combination that will yield a mixed suspension. The only model inputs are the suspension density, q m, total volume fraction of solids, U and the superficial fluidization rate, W. The model predicts a particle combination consisting of relatively small denser-particles, and relatively large lighter-particles, etc. All of the particles in the system have the one slip velocity, identical to the interstitial fluid velocity relative to the vessel, W =ð1 UÞ. All of the particles in the system are equally capable, therefore, of reporting to the overflow or the underflow, under continuous conditions. A continuous feed effectively creates the need for an underflow consisting of the faster settling feed particles, and an overflow consisting of the slower settling feed particles. The terminal velocities, U t of particles of different size and density, isolated in a liquid, were determined using the equations of Zigrang and Sylvester (1981). The Reynolds number of the particle at its terminal velocity, U t, is, Re ¼½ð14:51 þðgðq s qþqþ 0:5 1:83d 1:5 =lþ 0:5 3:81Š 2 ; ð1þ where g is the acceleration due to gravity ðm=s 2 Þ; q s is the density of the particle ðkg=m 3 Þ, q the density of the fluid ðkg=m 3 Þ, d the particle diameter (m), and l the viscosity of the fluid ðns=m 2 Þ. In turn, the particle Reynolds number, Re, is defined as Re ¼ qu td l : ð2þ The influence of the surrounding suspension medium on the slip velocity, V, was determined using the following equation, n 1 q V ¼ U s q m t ; ð3þ q s q where q m is the density of the suspension medium. The equation of Garside and Al-Dibouni (1979) provided the dependence of n on the Reynolds number. In the explicit form, n is n ¼ 5:1 þ 0:2Re0:9 : ð4þ 1:0 þ 0:1Re0:9 Fig. 3. Slip velocities calculated as a function of the density and particle size for a suspension of density 1178 kg=m 3. Eq. (3) converts to the Richardson and Zaki equation for identical particles, and to the Lockett and Al-Habbooby (1973, 1974) equation, applicable when particles of the one density, but different size are present. Galvin et al. (1999) examined the gravity separation of 2 þ 0:25 mm coal in a TBS in the presence of an autogenous mineral matter media, and other media including magnetite, sand, and barite. The coal/mineral matter feed system is examined here. The computed slip velocities for particles of different size and density are shown in Fig. 3, based on the maximum suspension density near the base of the device of 1178 kg=m 3. The upward interstitial velocity of the liquid, 28 mm/s, is represented by the horizontal line. The product and reject streams consist of particles having size and density values located below and above the line, respectively. Size and density conditions that coincide with the horizontal line provide the predicted dependence of the D 50 on the particle size. This predicted dependence is shown in Fig. 4, together with the actual experimental data values obtained from the partition curves. It is evident that the model provides a useful prediction of the variation in the D 50 with the particle size. 2. Experimental This paper describes the first pilot scale trial of the RC at a coal preparation plant, in which the device is used to separate coal and mineral matter less than 2 mm in size. Feed, product, and reject ash data were obtained for a campaign over a single day. The coal slurry was fed to the RC at an average pulp density of 47% solids. The feed contained approximately 1% +2 mm particles, 83% 2 þ 0:125 mm particles, and 16% )0.125 mm particles. A series of 5 steady state sets of samples was obtained. A preliminary analysis was conducted on the 2 þ 0:125 mm fraction of each of the sample sets. A full

4 22 K.P. Galvin et al. / Minerals Engineering 15 (2002) Fig. 4. Variation in the D 50 separation density with particle size for a Teetered Bed Separator (Galvin et al., 1999). The symbols represent the experimental data and the continuous curve corresponds to the slip velocity model prediction. washability analysis was then conducted on a single steady state set of samples over the size range 2 þ 0:250 mm. The washability of the size fraction 0:250 þ 0:125 mm was not examined. 3. Results and discussion The hydraulic loading of the feed slurry to the RC was 75 m 3 =m 2 h, and the solids loading ranged from 33 to 47 t=m 2 h, thus averaging 42 t=m 2 h throughout the campaign. The feed rate was set to the maximum possible at the time. The nominal solids loading of a TBS processing a similar coal is potentially 14 t=m 2 h (Honaker and Mondal, 1999). Clearly, the trial provides a basis for assessing the high throughput capability of the RC device relative to a TBS. The samples taken during the campaign were analysed to obtain the pulp density, mass percent in the size ranges +2, 2 þ 0:125, and )0.125 mm and corresponding ash values in those size fractions. These samples were found to conform to material balance requirements, with little adjustment needed to produce a balance. The +2 mm material was neglected because this fraction was present at less than 1%. The 2 þ 0:125 mm ash values obtained from all of the samples taken during the campaign are reported in Fig. 5. Despite some variation in the feed ash, a consistent product ash of 10.4% on average was obtained, and a reject ash of 73.2% on average obtained, corresponding to an average product recovery of 90%. The pulp densities of the overflow samples were remarkably high, ranging from 35% to 41%. In principle, the fluidization hydraulic loading only needs to be equivalent to that used in a TBS, given that this superficial velocity is sufficient to just fluidize the bed and in turn maintain a high bed concentration. Given that the feed Fig. 5. Feed, product and reject ash values for the 2 þ 0:125 mm fraction obtained during a one day campaign using the pilot scale Reflux Classifier operated on average at 42 t=m 2 h. solids loading was several times higher than in a typical TBS, the quantity of fluidizing water required relative to the quantity of feed was exceedingly low, at about 30% of that needed in a TBS. More recently, we have aimed to achieve a similar product pulp density to that produced by a TBS by adding additional water with the feed. Ideally, it will be possible to run the RC with low feed pulp densities, and hence overcome the need to install cyclones as a pre-concentration step, as required for the TBS Washability analysis A full washability analysis was conducted on the 2 þ 0:250 mm fractions of the feed, product, and reject streams of the fourth separation reported in Fig. 5. The feed washability curve, with the cumulative yield plotted as a function of the cumulative ash, is shown in Fig. 6. Fig. 6. The maximum possible yield obtainable for a given ash product, generally referred to as the feed washability curve. The curve was based on the full mass balance on the 2 þ 0:25 mm fraction. The cross on the plot denotes the process product yield of 87.1% and ash of 9.0% determined by a full mass balance on the 2 þ 0:25 mm fraction.

5 K.P. Galvin et al. / Minerals Engineering 15 (2002) Fig. 7. The cumulative yield as a function of the particle relative density. The cross on the plot denotes the process product yield of 87.1% and ash of 9.0% determined by a full mass balance on the 2 þ 0:25 mm fraction. Fig. 8. Partition curves obtained from the washability analysis of the feed, product and reject streams, for size fractions of 2 þ 1:4, 1:4 þ 1:0, 1:0 þ 0:7, 0:7 þ 0:5 and 0:5 þ 0:25 mm. Fig. 9. The separation efficiency, Ep, as a function of the particle size. This curve, derived from the balanced data set, provides the maximum possible yield for a given product ash. The cumulative yield is also shown as a function of the particle density in Fig. 7. The cross on each of the plots denotes the product yield of 87.1% and ash of 9.0%, as determined by a full mass balance of all the 2 þ 0:25 mm data. It is evident that the product yield obtained using the RC was 0.9% less than the theoretical maximum of 88.0% at 9.0% ash. The loss of yield, 0.9%, is considered accurate, because the value is derived from the balanced data, and very little adjustment in the experimental data was required. The combustible yield was 96.4%, which was also close to the theoretical limit of 97.2%. The data were self-consistent. For example, the actual 2 þ 0:25 mm feed, product, and reject ash values of 17.8, 9.0, and 75.1 correspond to a yield of 86.7%, which is close to the mass balance value of 87.1%. The partition curves obtained from the washability analysis of the feed, product and reject streams are shown in Fig. 8. Size fractions of 2 þ 1:4, 1:4 þ 1:0, 1:0 þ 0:7, 0:7 þ 0:5, and 0:5 þ 0:25 mm were used as the basis for the five partition curves. It is evident that there is a shift in the partition curves towards higher densities as the particle size decreases. The Ep values, which denote the separation efficiency of a given partition curve, increase from 0.03 to 0.08 as the particle size decreases, as shown in Fig. 9. These Ep values are in part a consequence of the finite range of particle sizes in a given size fraction, and hence the inherent difference in the settling velocities of particles in a given size range. The Ep values are all relatively low, and hence the Ep values should have relatively little effect on the overall partitioning of the feed. The D 50 values, however, range from about 1.49 to 1.95 relative density units. Thus, it is the variation in the D 50 with the particle size that really matters, as this will tend to govern the overall partitioning of the feed. The overall partition curve for the size range, 2 þ 0:25 mm, presented in Fig. 10, shows an overall D 50 value of 1.81 and an overall Ep value of Variation in the D 50 with particle size The variation in the D 50 values with the particle sizes for the TBS and the RC are shown in Fig. 11. The data points for the TBS are the same as those reported in Fig. 2 (Galvin et al., 1999), and are consistent with those reported by Hyde et al. (1988) and Honaker and Mondal (1999). The data for the RC correspond to the partition curves presented in Fig. 8. The continuous curves are based on the slip velocity model presented earlier. In the case of the TBS, the curve is based on the measured suspension density and critical slip velocity (interstitial liquid velocity). This work was performed carefully under laboratory conditions and hence it was possible to readily obtain a full compliment of data values.

6 24 K.P. Galvin et al. / Minerals Engineering 15 (2002) Fig. 10. The overall partition curve for the size range of 2 þ 0:25 mm. Fig. 12. Slip velocities calculated as a function of the particle density and particle size for a suspension density of 1370 kg=m 3. Fig. 11. Variation of D 50 with particle size. A comparison between the separations produced using the Teetered Bed Separator and the Reflux Classifier. The symbols represent the experimental data and the continuous curves correspond to the slip velocity model prediction. For the Reflux Classifier, the model was adjusted to fit the experimental data. In the case of the RC, however, it was more difficult to obtain a full compliment of data given this work was conducted at a much larger scale in a coal preparation plant. Hence, in this case, the suspension density and critical slip velocity values were obtained by adjusting the model to fit the D 50 versus particle size data. The required suspension density was found to be 1370 kg=m 3, and the required critical slip velocity was found to be 8.9 mm/s. The high suspension density is consistent with the value of 1375 kg=m 3 derived from a pressure transducer reading of h 1 ¼ 75 mm of water, over a height of h 2 ¼ 0:2 m near the base of the vessel. The transducer pressure reading was equivalent to, DP ¼ qgh 1 ¼ : ¼ 735 Pa. The suspension pressure over the vessel height of 0.2 m, relative to the column of water connecting the transducer to the RC, was therefore DP ¼ðq m qþgh 2 ¼ ð Þ9:8 0:2 ¼ 735 Pa. Assuming the suspension particles have a density of 2000 kg=m 3, the volume fraction of solids in the suspension would have to be 0.37, which is certainly reasonable. The critical slip velocity of 8.9 mm/s would in turn correspond to a superficial fluidization velocity of 8:9ð1 0:37Þ ¼5:6 mm=s. Hence a mixed suspension of D 50 particles is produced at a fluidization rate of 5: :6 0:6 ¼ 2: m 3 =s ¼ 120 l= min. A simple water balance indicated a fluidization rate of 69 l/min, however, a consideration of small errors in the measured pulp densities and flow rates easily accounted for any discrepancy between this value and the value of 120 l/min. In a follow up study, a magnetic flow measurement device was fitted and it was shown that the RC typically operates at fluidization rates of l/min, a result consistent with that obtained above using the model. The slip velocities computed for particles of different size and density, based on a suspension density of 1370 kg=m 3, are shown in Fig. 12. A comparison between Figs. 3 and 12 shows the effect of increasing the suspension density. Clearly, in Fig. 12 the curve corresponding to a relative density of 1.4 has shifted significantly away from those of higher density, and hence particles of this density have no chance of reporting to the reject stream. However, particles with a relative density of 1.5, which are larger than about 2 mm in size, should report to the reject stream. It is interesting to consider the effect of increasing the fluidization rate on the separation. In a TBS an increase in the fluidization rate will generate an immediate and sustained reduction in the suspension density, thus shifting the lower density curves in Fig. 3 towards the higher density curves. Thus, large lower-density particles become even more likely to report to the reject stream. However, in the RC, the inclined plates permit maintenance of the high suspension density at the higher

7 K.P. Galvin et al. / Minerals Engineering 15 (2002) fluidization rate (Galvin and Nguyentranlam, 2001b). Therefore, we can continue to use Fig. 12, and apply a higher critical slip velocity by raising the horizontal line. This will yield a higher recovery of the large low-density particles, but also an increase in the particle size of fine high-density particles reporting to the product. Thus, the RC provides a greater degree of flexibility in its operation, thus overcoming one of the major concerns of the TBS, loss of relatively large low ash coal particles. When relatively fine high-density particles are present in the feed, it is essential to combine these fluidized bed systems with other unit operations. For example, a separation of the product at about 350 lm will generally remove all of the fine high ash particles. The 350 lm product may then be sent to froth flotation to selectively recover the low ash coal particles. 4. Conclusions The Reflux Classifier produced an excellent gravity separation of )2 mm coal at a high feed throughput of 75 m 3 =m 2 h of slurry, and solids loading as high as 47 t=m 2 h. This throughput is significantly higher than achieved in a TBS. The RC features sets of parallel inclined plates that permit very much higher volumetric loadings, and therefore higher throughputs. A full washability analysis provided details on the variation in the D 50 with the particle size. It was apparent from a comparison with data on a conventional hindered settling bed that the RC offers significantly less variation in the separation density with size. The improved performance can be explained through a theoretical analysis showing the importance of operating at a higher suspension density. Acknowledgements The authors thank ACARP for its financial support of this work on the Reflux Classifier. References Boycott, A.E., Sedimentation of blood corpuscles. Nature 104, 532. Galvin, K.P., Nguyentranlam, G., 2001b. Influence of parallel inclined plates in a liquid fluidized bed system. Chemical Engineering Science, in press. Galvin, K.P., Pratten, S.J., Nicol, S.K., Dense medium separation using a teetered bed separator. Minerals Engineering 12 (9), Garside, J., Al-Dibouni, M.R., Particle mixing and classification in liquid fluidised beds. Transactions of the Institution of Chemical Engineers 57, Honaker, R.Q., Mondal, K., Dynamic modelling of fine particle separations in a hindered bed classifier. In: Society for Mining, Metallurgy, and Exploration, Denver, Colorado, March 1 3. Hyde, D.A., Williams, K.P., Morris, A.N., Yexley, P.M., The beneficiation of fine coal using the hydrosizer. In: Mine and Quarry. Lockett, M.J., Al-Habbooby, H.M., Differential settling by size of two particle species in a liquid. Transactions of the Institution of Chemical Engineers 51, Lockett, M.J., Al-Habbooby, H.M., Relative particle velocities in two-species settling. Powder Technology 10, Nguyentranlam, G., Galvin, K.P., 2001a. Particle classification in the reflux classifier. Minerals Engineering 14, Zigrang, D.J., Sylvester, N.D., An explicit equation for particle settling velocities in solid liquid systems. AIChE Journal 27 (6),