Separation by density difference is a process that is as old as recorded history. Separation of gold by density difference dates back to at least 3000 BC as depicted in writings from ancient Egypt. The principle employed in gravity separation goes back further in time to the formation and weathering of the rocks and the releasing of the minerals they contain and the transport of the mineral grains by heavy rains. It is the driving force for the formation of the alluvial deposits of precious metals and gemstones that have been worked since beyond recorded history as they still are today. Archaeological excavations have discovered mineral concentration activities such as the lead–silver concentrating plant in Attica, Greece, dating from 300 to 400 BC. So gravity separation has a long history as a mineral concentration process.
Not all mineral combinations are amenable to this type of concentration technique. To determine the suitability of gravity separation processes to a particular ore type, a concentration criterion is commonly used. A concentration criterion (CC) can be defined as
where SG = specific gravity (or density), and the fluid is typically water or air.
Some concentration criterion ratios for minerals that are treated by gravity separation are given in Table 1.
A guideline for separability by gravity based on this concentration criterion is given in Table 2. Figure 1 shows these limitations graphically over a separation curve. Separation is possible above the line and impossible for concentration criteria below the line.
Table 2 shows that separation will be easier in a fluid of higher density. The concentration ratio numbers in Figure 1 and Table 2 are a guide only as the ratio is influenced by other factors such as particle shape. Particle shape can be taken into account by including a shape factor defined as the ratio of shape settling factors for the heavy and light minerals.
2. Gravity separation equipment
Stratification in a bed of particles results from the repeated pulsation of a current of fluid up through the bed. The particles in the bed are expanded so that when pulsation ceases, the particles are allowed to consolidate under the influence of gravity. Figure 2 illustrates the expansion and contraction of the bed with the heavier, larger particles falling under hindered settling conditions.
The expansion and contraction of the bed is repeated in a cyclic operation until the heavy and light particles have stratified according to their specific gravity. The frequency of pulsations usually varies from 50 to 300 cycles per minute.
Jigs are commonly used to clean coal but are also used in heavy mineral separations including alluvial gold. When treating coal, the light fraction is the concentrate and in the mineral industry, the heavy fraction is the concentrate. For this reason, gravity separation products will be referred to as light or heavy rather than the concentrate or tailing. The jig is commonly an open tank filled with water, with a horizontal screen near the top. Some early jigs were designed where the screen surface, in the form of a basket, moved up and down in a barrel or tank of water, hence producing the vertical flow of fluid through the bed of particles. This manual operation is reported in the sixteenth century work by Agricola. Modern prospectors may still use this simple manual device in water drums or streams. Some movable screen jigs are still designed today, though most modern jigs employ a stationary screen and pulse the water through it. The differences between the various types of jigs available relate to the methods used to generate the pulsation and the manner in which the heavy fraction is removed from the jig. The screen is there to support the bed of particles and the area underneath the screen is called the hutch. The tank is usually divided into two main sections, one containing the support screen with the bed of ore and another section which generates the fluid pulse.
2.1.2 Operation of Jigs
The control of a jig separation is determined by the water addition, stroke frequency and amplitude, the feed rate and the ragging layer. Water is added to the jig as either top water (water added above the screen) or back water (water added beneath the screen or hutch water). The total water flowing across the top of the jig bed is the cross water. This cross water controls the horizontal flow of particles across the top of the bed. The back water reduces the effect of the suction part of the cycle and hence affects the falling water velocity relative to the rising water velocity during the pulsation part of the stroke. The feed rate must be matched with the discharge rate of the heavy fraction so that a steady state operation can be maintained. If the discharge of the heavy fraction does not keep up with the heavy particles reaching the separated layer, then this layer will build up until ultimately some heavy minerals will be lost to the light fraction. Conversely if the discharge rate of heavies through the ragging or through the discharge gate is greater than the rate of segregation of heavy particles into the separation layer then some light particles will eventually be drawn into the heavy fraction, lowering the grade.
2.2 Shaking table
2.2.1 Introduction of shaking table
Wet concentrating tables developed from continuous belt concentrators utilized a flowing film of water to effect a separation. The ore moved up an incline slope on an endless belt where the lighter minerals were washed away from the heavy minerals by a film of water flowing down the belt, similar to the Strake tables. The Vanner, a vibrating continuous belt, was developed in the 1860s and bumping tables followed before the modern differential shaking table was developed by Wilfley in 1896.
A bed of particles which experiences a horizontal shaking motion will undergo segregation on the basis of size and density, for example a gold pan and particles of a conveyor belt. If the particles are of the same density, then particles will segregate according to size with the fine particles sinking and the coarse particles rising to the top (Figure 3).
If particles of different density exist in the mixture then particles of higher density will sink to a lower level than similar sized but lighter density particles. To achieve this stratification, the shaking motion must be strong enough to expand the bed to the extent that allows particles to penetrate. The shaking motion, however, must still maintain a particle-to-particle contact. The fact that small particles of light mineral and large particles of heavy mineral segregate to the same position in the bed suggests that density is not the sole-separating force.
2.2.2 Differential Motion Shaking Tables
In the shaking table concentrator, differential motion and a riffled deck with cross flowing water is used to create a particle separation. The shaking motion is asymmetrical, being slow in the forward direction and rapid in the reverse direction. This differential motion imparts a conveying action to the table moving those particles which are in contact with the table deck, through friction, in the direction of the motion.
The table consists of a slightly inclined flat surface or deck with a series of parallel ridges or riffles along the direction of motion (Figure 4). The riffles are tapered towards the opposite end to the reciprocating drive. Feed is introduced at the corner of the table at about 25% solids (by mass) and with the shaking motion, the particles spread out over the table. Wash or dressing water is introduced along the top edge of the deck to assist in segregation and transport of particles on the table. The net effect is that the particles move diagonally across the deck from the feed end.
As the feed material spreads out over the table the particles stratified in layers behind the riffles. The riffles help to transmit the shaking motion to the particles and prevent the particle washing directly off the table. Successive layers of particles are removed from the top of the riffles by the cross-flowing water as they become exposed by the shortening riffle height as the bed moves away from the feed end of the table. When the remaining particles reach the end of the riffles only a thin layer remains on the table surface. If the table has a smooth unrifled end, then the flowing film of water further cleans the heavy particles before discharging them off the end of the table.
2.2.3 Shaking table operating parameters
Factors which affect the operation of the shaking concentrator are particle shape, size and density, the riffle design, deck shape, water and feed flow, stroke and speed of the table and the deck slope. The effects of these variables are summarized in Table 3. The correct operation of the table has the middling fraction discharged at the diagonally opposite corner of the table to the feed. For any feed variation, the operating variables are adjusted to maintain this separation point.
The particle shape is not a major factor in the overall tabling process; however, flat particles do not roll easily across the deck and would tend to be carried along to the heavy mineral discharge end of the table.
Of considerable more importance is the particle size. In both table stratification and hindered settling, the separation of particles becomes more difficult as the range of sizes in the feed increases. If a table feed contains too wide a range of sizes, some size fractions will be separated inefficiently. For efficient table operation, a normal feed size for coal treatment ranges from 0.3 to 9.5 mm. The lower size limit for an effective separation on a table is about 50 mm even if the density difference is high.
For optimum table operation, the feed flow of solids and water onto the table must be uniform and constant. Approximately 90% of the water reports to the light fraction. The dressing water represents approximately 25% of the total water on the table. The table capacity varies according to the size of the feed particles. Tables can handle up to 2 t/h of 1.5 mm sand and 1 t/h of fine sand. Capacities can be as low as 0.5 t/h for a slimes feed. The stroke rate for normal operation is between 250 and 300 strokes per minute. The stroke length required for coal separation can range from 10 to 25 mm. A longer stroke moves the reject (heavy) to the heavy discharge end of the table more rapidly, but requires more water.
The amplitude and stroke frequency are interdependent. That is, an increase in stroke length requires a decrease in the stroke frequency to maintain the same transportation speed of the heavy fraction to the discharge point. A fine feed will generally require a higher speed and shorter stroke than a coarse feed. For difficult separations, where the density difference between the two fractions is small, a short stroke length must be used. A general guide to table operation is given in Table 4. Early tables were generally covered with linoleum with wooden riffles. Modern tables use either rubber riffles cemented to a rubber covered deck or the whole deck is moulded in fibreglass.
2.2.4 Types of Tables
A number of different types of tables are available for different applications, and these vary mainly in relation to the type of head motion used.
A sand table is used for treatment of coarse particles, greater than 100 mm. Some types of sand tables, like the Diester table, are used extensively in the coal preparation industry; it is capable of a longer stroke than a standard Wilfley table, which is required for the concentration of the coal particles. The head motion of the Holman or James table is applied to the corner of the table rather than to the centre as in a normal Wilfley table. In cases where floor space is at a premium, tables can be mounted in vertical stacks of two or three high.
The treatment of slime particles (<100 μm) on any gravity separation device is difficult. The separations achieved are not efficient but before the introduction of flotation or centrifugal devices, slime tables were used. The basic principles of slimes tabling are
-the deck area required varies inversely as the feed size,
-the finer the feed, the gentler and slower the table action must be,
-the feed size distribution must be even and channeling avoided,
-the flocculation characteristics of the feed affect its response to gentle flowing action.
The concentration criterion for a mixture of quartz and cassiterite is 3.5 and at less than 50 μm particle size; this is not sufficient to give sharp separation. The concentration criterion for a quartz–gold mixture is nearly 9, and good separation is possible at this size.
Gemeni gold table
The Gemeni table has grooves on the table deck instead of riffles. It thus behaves like a mechanized gold pan in that the heavy gold is trapped in the grooves and the light gangue is washed over the groove and off the table. The gold migrates from groove to groove working its way along the table by the action of the shaking mechanism. The wash water is introduced along the centre of the table, dividing the table into two sloping surfaces.
2.3 Spiral Concentrator
The spiral concentrator first appeared as a production unit in 1943 in the form of the Humphrey Spiral for the separation of chrome-bearing sands in Oregon. By the 1950s, spirals were the standard primary wet gravity separation unit in the Australian mineral sands industry.
In the spiral concentrator the length of the sluicing surface required to bring about segregation of light from heavy minerals is compressed into a smaller floor space by taking a curved trough and forming into a spiral about a vertical axis. The slurry is fed into the trough at the top of the spiral and allowed to flow down under gravity. The spiraling flow of pulp down the unit introduces a mild centrifugal force to the particles and fluid. This creates a flow of pulp from the centre of the spiral outwards to the edge. The heaviest and coarsest particles remain near the centre on the flattest part of the cross-section, while the lightest and finest material is washed outwards and up the sides of the launder (Figure 5).
This separation may be assisted by the introduction of additional water flowing out from the centre of the spiral either continuously or at various locations down the length of the spiral.
This wash water may be distributed through tubes or by deflection from a water channel that runs down the centre of the spiral. Some present designs have overcome the need for this wash water. Once the particle stream has separated into the various fractions, the heavy fraction can be separated by means of splitters at appropriate positions down the spiral. A concentrate, middlings and tailing fraction can be recovered.
In practice spirals are arranged in stacks or modules of roughers, scavengers and cleaners, where the initial concentrate is retreated to upgrade the fraction to its final grade. Spiral length is usually five or more turns for roughing duty and three turns in some cleaning units. For coal concentration, six turns providing a gentler slope with longer residence time for the more difficult separation.
The performance of spirals is dependent on a number of operating parameters, summarized in Table 5
Spirals generally achieve an upgrade ratio of 3:1 (heavy fraction:feed grade) and generally multiple treatments are required. The presence of slimes adversely affects the spiral performance. More than 5% of −45 mm slimes will affect the separation efficiency.
With the steep pitch of a spiral, two or three spirals can be wound around the same common column and these types of spirals have been used in China for more than 20 years. The multistart spirals conserve floor space and launder requirements. These triple-start spirals are built into a 12 spiral module and for these modules, the design of the distributor is critical to ensure that each spiral has a uniform feed.
The splitter blades on these spirals are all adjustable to direct the heavy fraction into pipes or a collecting launder. The current range of spirals available consists of a number of different profiles which have individual separation characteristics. The dimensions of some of the available spirals range from 270 to 406 mm pitch, 590 to 700 mm diameter and 2.1 to 2.4 m high.
The advantages that modern spirals offer are simple construction requiring little maintenance, low capital cost and low operating cost – no reagents required, no dense media losses occur, low operating personnel required.
2.4 Centrifugal Separator
Poor separation of heavy minerals occurs at fine particles sizes. This poor performance can be overcome to some extent by increasing the settling rate of the fine particles by using centrifugal acceleration rather than gravitational acceleration. This led to the development of the centrifugal separator.
The Knelson Concentrator was one of the first of this type of concentrator to find commercial success and is a highly efficient centrifugal separator for free gold or other heavy mineral recovery. The concentrator is a compact high-capacity centrifugal separator with an active fluidized bed to capture heavy minerals. A centrifugal force equal to or exceeding 50 times the gravitational force acts on the particles enhancing the specific gravity difference between the heavy particles and the lighter gangue. The strong centrifugal force traps the heavy mineral in a series of rings located in the spinning drum while the gangue overflows the rings and is flushed out.
The concentrator has the capability to treat 40 t/h of minus 6 mm feed material and will recover in excess of 95% of the precious metal values. Gold particles from 6 mm to minus 30 μm are more efficiently recovered than with other known gravity method at high capacity. Even fine, platy or flour gold particles can be recovered.
The hydraulic section forms a self-cleaning fluidized bed that efficiently entraps high-density minerals while pushing out lighter material. This eliminates problems previously encountered with other centrifugal concentrators, including frequent shutdowns to remove black sands, low throughput and low concentrate ratios in the concentrate. The concentrate grade can be as high as 1000 times the feed grade.
The original design was a batch operation but the latest designs offer semi-continuous and continuous extraction of the heavy fraction. These concentrators are popular units in the grinding circuit of gold plants that contain free gold. The cyclone underflow or a percentage of the cyclone underflow is passed through these concentrators with the concentrate going to the gold room for further upgrading or smelting and the tailings going back to the ball mill. This removes the coarse gold which would remain in the mill circulating load due to the high density of gold. These devices are suitable to capture recirculating gold in the grinding circuit because of their high % solids feed capability allowing a minimum of additional water into the high % solids grinding streams.
3. Gravity Separation Performance
The evaluation of the separation method or performance of a gravity separation device is usually based on a sink-float analysis and washability curves. A great many applications of the washability curves are applied to cleaning operations in the coal preparation field (coal washing) so that most reference is made to coal washability curves; however, it must be remembered that the principles will also apply to separations of heavy minerals.
An ideal separation process would be one in which all particles of density lower than the separating density would be recovered in the light or clean product (coal) and all material denser than the separating density would be rejected as the heavy or refuse fraction. This is not achievable in practice. The type of separation between the light and heavy components that might actually occur is illustrated in Figure 6. Material that is much heavier or lighter than the separating density tend to report to their proper fraction but as the density of the particle approaches the density of separation of the unit, the amount of misplaced material increases rapidly. At the density of separation, the amount of misplaced material peaks at 50% as this is how we have defined the density of separation.
The imperfect separation of materials is characteristic of all gravity separation processes. The shape of the curve is determined largely by the inherent difficulty in stratifying materials of only slight density difference and will depend on the feed particles themselves, the feed rate, the media viscosity and separator characteristics. The more efficient the processes the sharper the separation and the narrower the peak. With low efficiency separation processes, the two arms of the peak are more widely separated. The peak will not necessarily be symmetrical.