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Solid-Liquid Separation and CCD washing

MTRL 358

Solid-Liquid Separation

2012

Introduction

This is a critically important operation in both mineral processing and at various stages in hydrometallurgical extraction. It is often quite costly, requiring large equipment and facilities. Poor liquid-solid separation characteristics can undo what would otherwise be a successful process. This requires careful attention.

Water may need to be reclaimed for a couple of reasons. First to adjust process stream characteristics so that they are suitable for subsequent treatment (applies to both tailings and other process streams). Second to recover or recycle water, which reduces consumption (very important in dry areas) and to prevent contamination of fresh water sources. On the other hand solids may need to be removed from a solution in order to obtain a pure metal product, or to further process the solids themselves.

Slurries, suspensions and tailings may contain from 10% to almost 100% water by weight. Dewatering is accomplished by four basic methods:

(a) Screening (larger particles retained on screens as water passes through)

(b) Sedimentation (decantation or thickening - relies on the same principles as

sedimentation classification)

(c) Filtration

(d) Thermal drying (may be costly since the heat capacity of water is high)

The effectiveness of water removal as a function of particle size (on a log scale) is shown in the figure below. The vertical axis lists volume % water retained in the dewatered product. There are numerous types of screens. Screening a slurry can very simply remove excess water from a desired size fraction. The smaller the particle size, the greater the total surface area per unit weight and the more water that is retained by the solids. A sieve bend is a type of screen that uses a curved surface. Slurry flows over the concave surface and in doing so experiences some centrifugal force due to the curvature. They are therefore somewhat more efficient than flat screens.

There are numerous types of classifiers. Classifiers have use in separating small particles from large particles or particles on the basis of density differences. They may also be used for dewatering. In this case the slurry may be fed into a long trough. Solids settle and are dragged up an incline by spirals, rakes etc. A disadvantage is that the aqueous phase may be significantly diluted. Very fine particles may also be deliberately removed from the pulp, with the aqueous phase (desliming). A thickener is a type of classifier. According to the figure a thickener can achieve about 75-80% water content by volume in the discharged underflow slurry. This seems rather large, but it should be remembered that minerals are much more dense than water. In many instances thickeners can achieve 50% water retention by weight. Relatively new paste thickeners use much deeper, narrower vessels and have solids (mud) zones that are several meters thick; much more than in conventional thickeners. This produces a much denser slurry (paste) that contains less water and exhibits non-Newtonian flow properties (rheology). An important type of classifier not indicated in the diagram is a hydroclone. These use centrifugal force to accentuate density differences. They are used for size separations, and for dewatering as well. Centrifuges are accelerated sedimentation devices that spin rapidly, also accentuating density differences. In a hydroclone, the machine stays fixed and the slurry is introduced tangentially into the cylindrical housing.

Filtration is commonly practiced in hydrometallurgy. Common types of filters are drum filters (explained later), belt filters and filter presses. Drum and belt filters employ reduced pressure on the underside of a porous membrane to draw off liquid. Filter presses use positive pressure on the side where the slurry contacts the filter cloth. Vacuum filters, for instance, are good at removing fine material and lower water content to as low as 10% by volume.

It is apparent from the figure that no matter what the particle size range, no mechanical dewatering system can removal all water. Further, as particle size decreases mechanical methods become less effective, due to the increased surface area per unit mass. Reduction in water content to low levels ultimately requires thermal drying. This is increasingly expensive as the cost of energy increases. Feeds to pyrometallurgical smelters may required thermal drying. In some instances this is also required for hydrometallurgical processes, but not commonly.

Figure 1. Overview of dewatering methods [1].

In sedimentation the settling rates of very small particles (a few microns in diameter) may be very slow under gravity alone. Flocculants may be added which act to agglomerate fine particles into more massive aggregates, which settle more rapidly. A wide variety of flocculants are available. Many are polymers with charged sites that attract oppositely surface-charged particles. This is illustrated in the figure below. This is an important technique in sedimentation. The objective now is not to separate one type of solid from another, but to separate all the solids from the water or solution. Thickeners are the most important type of sedimentation unit. They are similar in design and size to sedimentation classifiers. Diameters on the order of ~100 m exist. They are cylindrical and have conical bottoms. A variety of designs and types exist. They operate continuously. The overflow is the substantially clarified water/solution. The underflow slurry is higher in solids concentration than the feed slurry is, but is still quite high in water or solution content. A radially aligned rake turns, moving along the bottom to direct the solids to the discharge.

Figure 2. Schematic illustration of how a flocculant gathers small particles into a floc and makes a denser, more easily settled agglomerate.

CCD Washing

When a solid-liquid separation must be performed it may be preferable to use a thickener. Often solids settle readily under gravity, but filtration and washing turn out to be uneconomically slow. Additionally, as the tonnage of slurry to be separated increases, the cost of the thickener per tonne of material decreases. But, to filter and wash a large tonnage of slurry will require several filters in parallel. (Filters have relatively low capacities.) The capital cost of filters is proportional to the throughput. A single series of thickeners will often suffice to handle a large tonnage throughput. Thickeners can be made very large.

The solution in a slurry of leached ore often needs to be separated from the solids. Passing the slurry through a single thickener would still leave a substantial fraction of the solution (an hence valuable solute) going out the bottom of the thickener with the solids. This residual solution can be recovered by washing the solids. This is achieved by counter-current decantation. In this technique a series of thickeners are used. This is illustrated in the diagram below. Overflow from a thickener becomes the wash solution to the preceding thickener. It is preferable to have the underflow contain as high percent solids as possible. This minimizes the fraction of valuable metal in the underflow solution exiting with the solids. The final underflow slurry still contains some solute in solution. This is minimized when the underflow slurry has the highest possible solids content (less solution, and hence less solute), and with increasing number of wash thickeners.

In the schematic diagram below the total number of stages here is just three. In practice many more stages may be used, depending on the application. The number of wash stages in this example above is two, after the first , or lead (lead as in front, not as in Pb) thickener, to which feed slurry is added and from which the final clarified solution overflows. (The number of wash stages is always 1 less than the total number of thickeners in the circuit.) Wash

Figure 3. Schematic illustration of a simple CCD circuit.

water (or spent solution, low in the solute of interest) enters the last thickener and moves up to preceding thickeners. Underflow solids move from the first thickener to be combined with the wash solution as shown. The solids move in one direction and the solution overflow in the other. Hence the term counter-current decantation, or CCD. )The term decantation refers to the idea of pouring off a solution from a settled mass of solids.) It is important that the concentration of solution be uniform in the thickeners. This is achieved by ensuring good mixing of the underflow with the wash stream, as indicated by the confluence of streams in the diagram. If the flow of solutions and slurry is turbulent, a good rule of thumb is that the combined streams should flow through a distance equal to the radius of the thickener. Then mixing of the streams is sufficient to provide a uniform composition solution in the slurry. It is this mixing that results in the overflow solution and underflow solution form a thickener having the same composition! If mixing is inadequate there will be some loss of efficiency. It is often necessary to add flocculating agents to aggregate small particles in order to facilitate settling. Too strongly flocculated particles may trap some solution phase in the floc and this can result in excessive loss of solute to the final underflow.

Test work needs to be done for new processes in order to determine the settling rates of solids, for example after leaching of an ore. Another critical variable is the efficiency of the circuit as indicated by the recovery. Recovery is the ratio of the mass of solute in the final overflow (the pregnant solution) to the total mass of solute in the feed solution (the feed solution, again, being associated with a slurry). The remainder is lost to the final underflow solution exiting with the solids slurry leaving the final thickener. A key parameter for evaluating efficiency is the concentration of solids in the underflow. And the higher the solids content, the less the amount of solution, and therefore solute, that exits with the underflow slurry from each thickener. Then a lower number of stages may be required to effect the same degree of recovery. Washing results in a substantial dilution of the final recovered solution. The final PLS solute concentration is typically still >50% of the starting value. With enough thickeners virtually complete recovery can be obtained. The theoretical limit is 100%, albeit with a very large number of thickeners. In practice, since thickeners are capital intensive, there is an economic limit. A tradeoff between technical feasibility and economic feasibility must be made. Each additional stage recovers a smaller percentage of the solute; the law of diminishing returns.

To calculate recovery and efficiency a mass balance approach may be used. Consider a thickener into which a leach slurry enters. The overflow is the pregnant solution. In this example there are four stages, three of which are wash stages. In practice there may be several more stages. This is a variable. A diagram illustrating the solution flows is shown below.

Figure 4. Schematic illustration of solution flows and concentrations in CCD streams.

The feed enters the lead thickener (number 0) with a solution flow rate of F in convenient units. The concentration of a solute is C*. The pregnant solution overflow exits at a flow rate of V (> F) and a lower concentration C0 (because of dilution by wash water). Wash water enters the final thickener (number 3) at a concentration of solute = C4. If there is none of the solute in the water then C4 = 0. The wash water flow rate is designated O. This is also the overflow flow rate from each thickener, except number 0. (If this were not so, then solution would accumulate or deplete in the circuit.) It is assumed that the solute concentration in each thickener is uniform. The concentration in solution associated with thickener 2 is designated C2 and so on. Underflow from each thickener exits at a flow rate U of solution phase. The solids flow rate out the bottom adds to the total mass flow, but the mass balance is being carried out on the basis of solution. The concentration of solute in solution exiting thickener 2 is C2 for both the overflow and the underflow.

Now mass balance equations can be written for the solute. The mass balance properly requires knowledge of the mass flow rates and concentrations expressed in terms of mass of solute per unit mass of solution. However, mass flow rates may not always be readily available; pumps deliver known volumes per unit time. If the solution and wash densities are close to 1 g/mL, which in many instances is valid, the concentrations in mass/volume (e.g. Kg/m3) and flow rates in volume/time may be used. If the density of the leach solution, for example, is significantly higher than one, which may occur with concentrated solutions, then volumetric concentrations and flow rates will introduce a degree of error. Incorporating solution densities into the calculations then will overcome this error, but make the equations more complicated. For the sake of simplicity, in this discussion we will assume solution densities are 1 g/mL. Then 1 Kg of water or solution = 1 L volume.

A mass balance equation may be set up for each stage. At steady state, what enters must equal what leaves:

For 0FC* + OC1 = VC0 + UC0

(1)

For 1UC0 + OC2 = OC1 + U C1

(2)

For 2UC1 + OC3 = OC2 + UC2

(3)

For 3 UC2 + OC4 = OC3 + UC3

(4)In general for n wash stages,

UCn-1 + OCn+1 = OCn + UCn

(5)

Note the pattern of concentrations on the left and right. This symmetry leads to convenient simplification of the math later. The equations may be rearranged as follows:

FC*/U + (O/U)C1 = (V/U + 1)C0

(6)

C0 + (O/U)C2 = (O/U + 1)C1

(7)

C1 + (O/U)C3 = (O/U + 1)C2

(8)

C2 + (O/U)C4 = (O/U + 1)C3

(9)

The ratio O/U is called the overflow-underflow ratio, also called the wash ratio. This very important variable is the volumetric ratio of the overflow volume from each wash thickener to the underflow solution volume from each thickener. Let this be x. If C4 is zero, for wash water,

C3 = C2

(10)

x + 1

Now we have C3 in terms of C2. Next we express C1 in terms of C2, and so on. In other words, start with the nth thickener and work back. Substitute (10) into (8):

C1 + xC2 = (x + 1)C2

(11)

x+1

Rearranging gives:

C2 = C1

(12)

x + 1 - x

x + 1Similarly, substituting (12) into (7) yields:

(13)

Finally, substituting (13) into (6) leads to:

(14)

Note the repeating pattern in x terms! There is one set of these per wash thickener. Recovery of solute from the 0th thickener is expressed as:

Recovery = VC0/FC*

(15)

Multiply this by 100 and we have the recovery in %. Note also that the pregnant solution overflow rate must obey the equation:

F + O - U = V

(16)

assuming there are no changes in volume upon mixing. This is the mass balance for solution flows. Recovery then is:

(17)

The concentration C* disappears, which makes sense; fractional recovery cannot depend on concentration. It is just a ratio of masses. F disappears in (17) too, but reappears later when V is converted into F + O - U as per (16). Substituting (16) into (17) and multiplying by (1/U)/(1/U) to convert O terms into O/U = x terms leads to:

(18)It turns out that the function,

(19)

is precisely equivalent to:

xn+1 - x

(20)

xn+1 - 1

where n = the number of wash stages (the total number of thickeners less one). Recovery then is:

R = F/U + x - 1

(21)

F/U + x - xn+1 - x

xn+1 - 1

When terms are collected for equation (19) hairy polynomials result, one in the numerator and one in the denominator. Multiplying these by (x-1)/(x-1) results in equation (20). Hence the n+1 powers. It turns out that these numerators and denominators are always factors of xn+1-x and xn+1-1, respectively; x-1 is the other factor.

Equation (20) is called the wash efficiency. It indicates the effectiveness of the total of all the washing thickeners. Note that as the number of stages increases the wash efficiency gets closer to 1, i.e. 100% recovery:

(xn+1 - x) ( 1

(22)

(xn+1 - 1)as n gets large, and if x > 1. Note too that as O/U increases the wash efficiency improves. The higher O/U is, the less the amount of solution present with the solids in the underflow. In other words, the higher the solids content of the underflow, the better the wash efficiency and recovery. This is reasonable; the less of the solution that is going out the bottom with the solids, the more of it is being recovered with the overflow from each thickener. Hence the O/U ratio is absolutely critical to performance of thickeners.

Equation (14) may be rearranged with the same considerations as above to read,

FC*

C0 = U

(23)

F + x - xn+1 - x

U xn+1 - 1

A CCD wash circuit is designed around experimental data obtained during testwork to engineer a process. If it turns out that the % solids in the underflow is significantly less than the testwork suggested, then wash efficiency and total recovery will be correspondingly poor. This results in loss of metal values and it can be economically fatal. In conclusion, recovery and wash efficiency are improved by increasing the number of wash stages and increasing the % solids in the underflow. Each additional thickener improves recovery, but to an ever decreasing extent, as noted previously. With the equations above the effect can be quantified. Sometimes a thickener malfunctions (the walls fall off). While it is being repaired it must be bypassed. This lowers the number of stages and lowers recovery. If this happens too often, overall performance will be decreased. When solute is also present in the wash solution

The equations developed above assumed that the wash water contained none of the solute of interest. Then C4 in equation (9) was zero, or more generally, Cn+1 = 0. If this is not the case then the algerbra is a bit more complicated, but the equations can still be readily solved. The same method is used. The value of the solute concentration in the wash solution must be known. All the same assumptions as previously apply. The solute concentration in the PLS from the zeroth thickener is then found to be,

(24)Note that there are two terms:

FC* xn+1(x - 1) Cn+1

(25)

U + xn+1 - 1

F + x - xn+1 - x F + x - xn+1 - x

U xn+1 - 1 U xn+1 - 1

The first is identical to equation (23). The second is entirely due to the solute in the wash solution. The two contributions are completely separable.

Recovery becomes a matter of how you want to define it. Ignoring the contribution from the wash solution, the recovery of solute that is fed into the lead thickener is the same as it was before and is expressed by equation (21). The wash efficiency is also still the same as equation (20). If this were not so, then we could increase recovery of solute from the feed just by adding some of it to the wash water. That would violate the principle that you cannot get something for nothing, thermodynamically speaking. We can write and equation for recovery of solute that incorporates the solute added both from the feed and the wash solution. Then,

R = VC0

(26) FC* + OCn+1Substituting equation (24) for C0 and collecting terms finally yields the equation below,

(27)For Cn+1 = 0, i.e. wash water, the ratio on the right goes to unity and the familiar recovery from equation (21) results. The ratio on the right is the factor due to solute being present in the wash solution. In the limit of large n the ratio on the right tends toward,

FC* + Cn+1 x - Cn+1

(28)

U

FC + Cn+1 x

U

Thus the ratio on the right in equation (27) is always 1 is used, but W >4 would require too much time. The void volume within the filter cake is very small compared to the filtrate volume. Hence even with W = 4 the dilution of the filtrate is very small. With CCD wash circuits the extent of dilution is substantially greater.Economic Aspects of Solid-Liquid Separation

Some aspects of this have already been mentioned. These are typically rather slow processes. Capital costs are therefore high. Rates of removal of water tend to decrease rapidly as the water content decreases (law of diminishing returns again). Going from 1% to 0.1% water may take as long as going from 50% to 1% did. Likewise the same idea applies to removing suspended solids from a solution. The more clear the liquid needs to be, the greater the time and effort required. Sedimentation processes have moderate electricity requirements. Flocculating reagents may be quite costly, but they tend to be added at low to moderate concentrations. Drying costs may be quite high since high amounts of energy are required to evaporate water. This is not commonly employed in hydrometallurgy.

The highest solid-liquid separation costs are encountered just after leaching of fine solids in a reactor (tank or autoclave etc.). After initial separation to obtain a leach solution, the solids must be washed and then dewatered again to recover valuable metals in solution. The leach solution itself must be clarified to a high degree (very low suspended solids content) for it to be suitable for metal recovery. This may involve considerable capital cost. Good lab testing is always required. Generally the cost of solid-liquid separation increases with decreasing particle size. It is preferable to avoid having to conduct solid-liquid separations on very fine particles. References[1] Hayes, P.C., Process Principles in Minerals and Materials Production," Hayes Publ. Co., 1983, p. 117.

[2] Retrieved May 17/12 at:(a) http://www.komline.com/docs/rotary_drum_vacuum_filter.html(b) http://www.solidliquid-separation.com/VacuumFilters/Drum/drum.htm PAGE 18

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