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sag mill grinding circuit design

sag mill grinding circuit design

AG and SAG mills are now the primary unit operation for the majority of large grinding circuits, and form the basis for a variety of circuit configurations. SAG circuits are common in the industry based on:

Though some trepidation concerning AG or SAG circuits accompanied design studies for some lime, such circuits are now well understood, and there is a substantial body of knowledge on circuit design as well as abundant information that can be used for bench-marking of similar plants in similar applications. Because SAG mills rely both on the ore itself as grinding media (to varying degrees) and on ore-dependent unit power requirements for milling to the transfer size, throughput in SAG circuits are variable. Relative to other comminution machines in the primary role. SAG mill operation is more dynamic, and typically requires a higher degree of process control sophistication. Though more complex in AG/ SAG circuits relative to the crushing plants they have largely replaced, these issues are well understood in contemporary applications.

AG/SAG mills grindore through impact breakage, attrition breakage, and abrasion of the ore serving as media. Autogenous circuits require an ore of suitable competency (or fractions within the ore of suitable competency) to serve as media. SAG circuits may employ low to relatively high ball charges (ranging from 2% to 22%, expressed as volumetric mill filling) to augment autogenous media. Higher ball charges shift the breakage mode away from attrition and abrasionbreakage toward impact breakage; as a result, AG milling produces a finer grind than SAG milling for a given ore and otherwise equal operating conditions. The following circuits are common in the gold industry:

Common convention generally refers to high-aspect ratio mills as SAG mills (with diameter to effective grinding length ratios of 3:1 to 1:1), low-aspect ratio mills (generally, a mill with a significantly longer length than diameter) are also worth noting. Such mills are common in South African operations; mills are sometimes referred to as tube mills or ROM ball mills and are also operated both autogenously and semi-autogenously. Many of these mills operate at higher mill speeds (nominally 90% of critical speed) and often use grid liners to form an autogenous liner surface. These mills typically grind ROM ore in a single stage. A large example of such a mill was converted from a single-stage milling application to a semi autogenous ball-mill-crushing (SABC) circuit, and the application is well described. This refers to high-aspect AG/SAG mills.

With a higher density mill charge. SAG mills have a higher installed power density for a given plant footprint relative to AC mills. With the combination of finer grind and a lower installed power density (based on the lower density of the mill charge), a typical AG mill has a lower throughput, a lower power draw, and produces a finer grind. These factors often translate to a higher unit power input (kWh/t) than an SAG circuit milling the same ore. but at a higher power efficiency (often assessed by the operating work index OWi, which if used most objectively, should be corrected by one of a number of techniques for varying amounts of fines between the two milling operations).

In the presence of suitable ore, an autogenous circuit can provide substantial operating cost savings due to a reduction in grinding media expenditure and liner wear. In broad terms, this makes SAG mills less expensive to build (in terms of unit capital cost per ton of throughput) than AG mills but more expensive to operate (as a result of increased grinding media and liner costs, and in many cases, lower power efficiency). SAG circuits are less susceptible to substantial fluctuations due to feed variation than AG mills and are more stable to operate. AG circuits are more frequently (but not exclusively) installed in circuits with high ore densities. A small steel charge addition to an AG mill can boost throughput, result in more stable operations, typically at the consequence of a coarser grind and higher operating costs. An AG circuit is often designed to accommodate a degree of steel media for circuit flexibility. AG mills (or SAG mills with low ball charges) are often used in single-stage grinding applications.

Based on their higher throughput and coarser grind relative to AG mills, it is more common for SAG mills to he used as the primary stage of grinding, followed by a second stage of milling. AG/SAG circuits producing a fine grind (particularly single-stage grinding applications) are often closed with hydrocyclones. Circuits producing a coarser grinds often classify mill discharge with screens. For circuits classifying mill discharge at a coarse size (coarser than approximately 10 mm), trommels can also be considered to classify mill discharge. Trommels are less favorable in applications requiring high classification efficiencies and can be constrained by available surface area for high-throughput mills. Regardless of classification equipment (hydrocyclone, screen, or trommel), oversize can be returned to the mill, or directed to a separate stage of comminution.

Many large mills around the world (Esperanza with a 12.8 m mill. Cadia and Collahuasi with 12.2-m mills, and Antamina. Escondida #IV. PT Freeport Indonesia, and others with 11.6-m mills) have installed SAG mills of 20 MW. Gearless drives (wrap-around motors) are typically used for large mills, with mills of 25 MW or larger having been designed. Several circuits have single-line design capacities exceeding 100,000 TPD. A large SAG installation (with pebble crusher product combining with SAG discharge and feeding screens) is depicted here below, with the corresponding process flowsheet presented in Figure 17.9.

Adding pebble crushing as a unit operation is the most common variant to closed-circuit AG/SAG milling (instead of direct recycle of oversize material ). The efficiency benefits (both in terms of grinding efficiency and in capital efficiency through incremental throughput) are well recognized. Pebble crushers are effective at reducing the buildup of critical-sized material in the mill load. Critical-sized particles are those where the product of the mill feed-size distribution and the mill breakage rates result in a buildup of a size range of material in the mill load, the accumulation of which limits the ability of the mill to accept new feed. While critical-size could be of any dimension, it is most typically synonymous with pebble-crusher feed, with a size range of 1375 mm. Critical-sized particles can result from a simple failure of a mills breakage rates to exceed the breakage rate of incoming particles, and particles generated when breaking larger particles. Alternatively, a second type of buildup of critical-sized material can result due to a combination of rock types in the feed that have differing breakage properties. In this case, the harder fraction of the mill feed builds up in the mill load, againrestricting throughput. Examples of materials in this category include diorites, chert, and andesite. When buildup of these materials does occur, pebble crushing can improve mill throughput even more dramatically than when the critically sized fraction results purely from a breakage rate deficit alone. For these ore types, a pebble-crushing circuit is tin imperative for efficient circuit operation.

Currently, every AG/SAG flowsheet evaluation is likely to consider the inclusion of a pebble crusher circuit. Flowsheets that do not elect to include pebble crushing at construction and commissioning may include provisions for future retrofitting a pebble-crushing circuit. Important aspects of pebble crusher circuit design include:

The standard destination for crushed pebbles has been to return them to SAG feed. However, open circuiting the SAG mill by feeding crushed pebbles directly to a ball-mill circuit is often considered as a technique to increase SAG throughput. An option to do both can allow balancing the primary and secondary milling sections by having the ability to return crushed pebbles to SAG feed as per a conventional flowsheet, or to the SAG discharge. Such a circuit is depicted here on the right. By combining with SAG discharge and screening on the SAG discharge screens, top size control to the ball-mill circuit feed is maintained while still unloading the SAG circuit (Mosher et al, 2006). A variant of this method is to direct pebble-crushing circuit product to the ball-mill sump for secondary milling: while convenient, this has the disadvantage of not controlling the top size of feed to the ball-mill circuit. There have also been pioneer installations that have installed HPGRs as a second stage of pebble crushing.

The unit power requirement for SAG milling (both individually and as a fraction of the total circuit power) is worthy of comment. It can be very difficult operationally to trade grind for throughput in an SAG circuitonce designed and constructed for a given circuit configuration, an SAG mill circuit has limited flexibility to deliver varying product sizes, and a relatively fixed unit power input for a given ore type is typically required in the SAG mill. This is particularly true for those SAG circuits designed with a coarse closing size. As a result, under-sizing an SAG mill has disastrous results on throughput across the industry, there are numerous examples of the SAG mill emerging as the circuit bottleneck. On the other hand, over-sizing an SAG circuit can be a poor utilization of capital (or an opportunity for future expansion!).

Traditionally, many engineers approached SAG circuit design as a division of the total power between the SAG circuit and ball-mill circuit, often at an arbitrary power split. If done without due consideration to ore characteristics, benchmarks against comparable operating circuits, and other aspects of detailed design (including steady-state tests, simulation, and experience), an arbitrary power split between circuits ignores the critical decision of determining the required unit power in SAG milling. As such, it exposes the circuit to risk in terms of failing to meet throughput targets if insufficient SAG power is installed. Rather than design the SAG circuit with an arbitrary fraction of total circuit power, it is more useful to base the required SAG mill size on the product of the unit power requirement for the ore and the desired throughput. Subsequently, the size of the secondary milling circuit is then sized based on the amount of finish grinding for the SAG circuit product that is required. Restated, the designed SAG mill size and operating conditions typically control circuit throughput, while the ball-mill circuit installed power controls the final grind size.

The effect of feed hardness is the most significant driver for AG/SAG performance: with variations in ore hardness come variations in circuit throughput. The effect of feed size is marked, with both larger and finer feed sizes having a significant effect on throughput. With SAG mills, the response is typically that for coarser ores, throughput declines, and vice versa. However, for AG mills, there are number of case histories where mills failed to consistently meet throughput targets due to a lack of coarse media. Compounding the challenge of feed size is the fact that for many ores, the overall coarseness of the primary crusher product is correlated to feed hardness. Larger, more competent material consumes mill volume and limits throughput.

A number of operations have implemented a secondary crushing circuit prior to the SAG circuit for further comminution of primary crusher product. Such a circuit can counteract the effects of harder ore. coarser ore. decrease the size of SAG mill required, or rectify poor throughput due to an undersized SAG circuit. Notably, harder ore often presents itself to the SAG circuit as coarser than softer oreless comminution is produced in blasting and primary crushing, and therefore the impact on SAG throughput is compounded.

Circuits that have used or do use secondary crushing/SAG pre-crush include Troilus (Canada), Kidston (Australia), Ray (USA), Porgera (PNG). Granny Smith (Australia), Geita Gold (Tanzania), St Ives (Australia), and KCGM (Australia). Occasionally, secondary crushing is included in the original design but is often added as an additional circuit to account for harder ore (either harder than planned or becoming harder as the deposit is developed) or as a capital-efficient mechanism to boost throughput in an existing circuit. Such a flowsheet is not without its drawbacks. Not surprisingly, some of the advantages of SAG milling are reduced in terms of increased liner wear and increased maintenance costs. Also, pre-crush can lead to an increase in mid-sized material, overloading of pebble circuits, and challenges in controlling recycle loads. In certain circuits, the loss of top-size material can lead to decreased throughput. It is now widespread enough to be a standard circuit variant and is often considered as an option in trade-off studies. At the other end of the spectrum is the concept of feeding AG mills with as coarse a primary crusher product as possible.

The overall circuit configuration can guide selection of die classification method of primary circuit product. Screening is more successful than trommel classification for circuits with pebble crushing, particularly for those with larger mills. Single-stage AG/SAG circuits are most often closed with hydrocyclones.

To a more significant degree than in other comminution devices, liner design and configuration can have a substantial effect on mill performance. In general terms, lifter spacing and angle, grate open area and aperture size, and pulp lifter design and capacity must be considered. Each of these topics has had a considerable amount of research, and numerous case studies of evolutionary liner design have been published. Based on experience, mill-liner designs have moved toward more open-shell lifter spacing, increased pulp lifter volumetric capacity, and a grate design to facilitate maximizing both pebble-crushing circuit utilization and SAG mill capacity. As a guideline, mill throughput is maximized with shell lifters between ratios of 2.5:1 and 5.0:1. This ratio range is stated without reference to face angle; in general terms, and at equivalent spacing-to-height ratios, lifters with greater face-angle relief will have less packing problems when new, but experience higher wear rates than those with a steeper face angle. Pulp-lifter design can be a significant consideration for SAG mills, particularly for large mills. As mill sizes increases, the required volumetric capacity of the pulp lifters grows proportionally to mill volume. Since AG/SAG mill volume is roughly proportional to the mill radius cubed (at typical mill lengths) while the available cross-sectional area grows only as the radius squared, pulp lifters must become more efficient at transferring slurry in larger mills. Mills with pebble-crushing circuits will require grates with larger apertures to feed the circuit.

No discussion of SAG milling would be complete without mention of refining. Unlike a concentrator with multiple grinding lines, conducting SAG mill maintenance shuts down an entire concentrator, so there is a tremendous focus on minimizing required maintenance time; the reline timeline often represents the critical path of a shutdown (but typically does not dominate a shutdown in terms of total maintenance effort).

Reline times are a function of the number of pieces to be changed and the time required per piece. Advances in casting and development of progressively larger lining machines have allowed larger and larger individual liner pieces.

While improvements in this area will continue, the physical size limit of the feed trunnion and the ability to maneuver parts are increasingly limiting factors, particularly in large mills. The other portion of the equation for reline times is time per piece, and performance in this area is a function of planning, training/skill level, and equipment.

Abroad range of AG/SAG circuit configurations are in operation. Very large line plants have been designed, constructed, and operated. The circuits have demonstrated reliability, high overall availabilities, streamlined maintenance shutdowns, and efficient operation. AG/SAG circuits can handle a broad range of feed sizes, as well as sticky, clayey ores (which challenge other circuit configurations). Relative to crushing plants, wear media use is reduced, and plants run at higher availabilities. Circuits, however, are more sensitive to variations in circuit feed characteristics of hardness and size distribution; unlike crushing plants for which throughput is largely volumetrically controlled. AG/SAG throughput is defined by the unit power required to grindthe ore to the closing size attained in the circuit. Very hard ores can severely constrain AG/SAG mill throughput. In such cases, the circuits can become capital inefficient (in terms of the size and number of primary milling units required) and can require more total power input relative to alternative comminution flowsheets. A higher degree of operator skill is typically required of AG/SAG circuit operation, and more advanced process control is required to maintain steady-state operation, with different operator/advanced process control regimens required based on different ore types.

Many mills have been built based on data from inadequate sampling or from insufficient tests. With the cost of many mills exceeding several hundred million dollars, it is mandatory that geologists, mining engineers and metallurgists work together to prepare representative samples for testing. Simple repeatable work index tests are usually sufficient for rod mill and ball mill tests but pilot plant tests on 50-100 tons of ore are frequently necessary for autogenous or semiautogenous mills.

Preparation and selection of the test sample is of utmost importance. Procedures for autogenous and semiautogenous mill pilot plant tests are relatively simple for those experienced in running them. Reliable and repeatable results can be obtained if simple fundamental procedures are followed.

The design of large mills has become increasingly more complicated as the size has increased and there is little doubt that without sophisticated design procedures such as the use of the Finite Element method the required factors of safety would make large mills prohibitively expensive.

In the past the design of small mills, up to +/- 2,5 metres diameter, was carried out using empirical formulae with relatively large factors of safety. As the diameter and length of mills increased several critical problem areas were identified. One of the most important was the severe stressing which took place at the connection of the mill shell and the trunnion bearing end plates, which is further aggravated by the considerable distortion of the shell and the bearing journals due to the dynamic load effect of the rotating mill with a heavy mass of ore and pulp being lifted and dropped as the grinding process took place. Incidentally the design calculation of the deformations of journal and mill shell is based on static conditions, the influence of the rotating mass being of less importance. An indication of shell and journal distortion is shown in Figure 1.

Investigations carried out by Polysius/Aerofall revealed that practical manufacturing considerations dictated some aspects of trunnion end design. Whereas the thickness of the trunnion in the case of small diameter mills was dictated by foundry practice which required a minimum thickness of metal the opposite was the case in the design of large diameter mills where the emphasis was not to exceed a maximum thickness both from the mass/casting temperature point of view and the cost aspect.

While the deformation of shell and end plates was acceptable in the case of small mills due in some extent to the over stiff construction, the deformation in the large, more flexible, mills is relatively high. The ratio of the trunnion thickness to trunnion diameter in a mill of 2,134 m diameter is almost twice that of a mill of 5,8 m diameter, i.e. a ratio (T/D) of 0,116 to 0,069 for the large mill.

The use of large memory high speed computers coupled with finite element methods provides the means of performing stress calculations with a high degree of accuracy even for the complex structures of large mills. The precision with which the stress values can be predicted makes the use of safety factors based on empirical formulae generally unnecessary.

In the case of large diameter trunnion bearing mills the distortion which takes place is further compounded by the fact that the deformation varies across the width of the bearing journal due to the fact that the end of the journal attached to the mill end plate is less liable to distortion than the outlet free end of the journal. This raises serious complications as far as the development of the hydrodynamic fluid oil film of the bearing is concerned since the minimum oil gap may be only 0,05 mm.

Obviously a thinner oil film is adequate where the deformation of the journal is less while at the unsecured end of the journal widely varying oil film thickness is necessary to maintain the correct oil pressure to support the mill. A solution to this problem has been the advent of the hydrostatic bearing with a supply of high pressure oil pumped continuously into the bearings.

Incorporating the mill bearing journals as part of the mill shell reduced the magnitude of the problem of distortion although there is always out of round deformation of the shell. The variation across the width of the journal surface is less pronounced than is the case with the trunnion bearing.

The replacement of a single bearing with a number of individual self adjusting bearing pads which together support the mill has lessened the undesirable effects of deformation while improving the efficiency of the bearing.

The ability of each individual bearing-pad to adjust automatically to a more localised area of the shell journal gives rise to improved contact of the oil film with both the bearing surface and the journal and in the case of hydrodynamic oil systems makes it unnecessary to supply oil at constant high pressure once the oil film has been established. A cross-section of a slipper pad bearing is shown in Figure 3.

Kidstons orebody consists of 44.2 million tonnes graded at 1.79 g/t gold and 2.22 g/t silver. Production commenced in January, 1985, and despite a number of control, mechanical and electrical problems, each month has seen a steady improvement in plant performance to a current level of over ninety percent rated capacity.

The grinding circuit comprises one 8530 mm diameter x 3650 mm semi-autogenous mill driven by a 3954 kW variable speed dc motor, and one 5030 mm diameter x 8340 mm secondary ball mill driven by a 3730 kW synchronous motor. Four 1067 x 2400 mm vibrating feeders under the coarse ore stockpile feed the SAG mill via a 1067 mm feed belt equipped with a belt scale. Feed rate was initially controlled by the SAG mill power draw with bearing pressure as override.

Integral with the grinding circuit is a 1500 cubic meter capacity agitated surge tank equipped with level sensors and variable speed pumps. This acts as a buffer between the grinding circuit and the flow rate sensitive cycloning and thickening sections.

The Kidston plant was designed to process 7500 tpd fresh ore of average hardness; but to optimise profit during the first two years of operation when softer oxide ore will be treated, the process equipment was sized to handle a throughput of up to 14 000 tpd. Some of the equipment, therefore, will become standby units at the normal throughputs of 7 000 to 8 000 tpd, or additional milling capacity may be installed.

The SAG mill incorporates a design which allowed expedient manufacturing to high quality specifications, achieved by selecting a shell to head to trunnion configuration of solid elements bolted together. This eliminates difficult to fabricate and inspect areas such as a fabricated head welded to shell plate, fabricated ribbed heads, plate or casting welded to the head in the knuckle area and transition between the head and trunnion.

Considerable variation in ore hardness, the late commissioning of much of the instrumentation and an eagerness to maximise mill throughput led to frequent mill overloading during the first four months of operation. The natural operator over-reaction to overloads resulted numerous mill grindouts, about sixteen hours in total, which in turn were largely responsible for grate failure and severe liner peening. First evidence of grate failure occurred at 678 000 tonnes throughput, and at 850 000 tonnes, after three grates had been replaced on separate occasions, the remaining 25 were renewed. The cylinder liners were so badly peened at this stage that no liner edge could be discerned except under very close scrutiny and grate apertures had closed to 48 percent of their original open area.

The original SAG mill control loop, a mill motor power draw set point of 5200 Amperes controlling the coarse ore feeder speeds, was soon found to give excessive variation in the mill ore charge volume and somewhat less than optimal power draw.

The armature, weighing 19 tonnes, together with the top half magnet frame, were trucked two thousand kilometers to Brisbane for rewinding and repairs. The mill was turning again on January 24 after a total elapsed downtime of 14 days. After a twelve day stoppage due to a statewide power dispute in February, the mill settled down to a fairly normal operation, apart from some minor problems with alarm monitoring causing a few spurious trips. One cause of the mysterious stoppages was tracked down to the cubicle door interlocks which stuttered whenever the mining department fired a bigger than usual blast.

The open trunnion bearings are sealed with a rubber ring which proved ineffective in preventing ingress of water, and occasionally solids, from feed chute chokes and spillages. Contamination and emulsification of the oil with subsequent filter choking has been responsible for nearly eighteen percent of SAG mill circuit shutdowns. Despite the very high levels of contamination, no damage has been sustained by the bearings which has at least proved the effectiveness of the filters and other protection devices.

Design changes to date have, predictably, mostly concentrated on improving liner life and minimising discharge grate damage. Four discharge grates with thickened ends have performed satisfactorily and a Mk3 version with separate lifters and 20 mm apertures is currently being cast by Minneapolis Electric.

Cylinder liners will continue to be replaced with high profile lifters only on a complete reline basis. While there is the problem of reduced milling capacity with reduced lifter height towards the end of liner life, it is hoped to largely offset this by operating at higher mill speeds.

Mill feed chute liner life continues to be a problem. The original chrome-moly liners lasted some three months and a subsequent trial with 75 mm thick clamped Linhard (rubber) liners turned in a rather dismal life of three weeks.

sag mill slurry rheology - grinding & classification circuits - metallurgist & mineral processing engineer

sag mill slurry rheology - grinding & classification circuits - metallurgist & mineral processing engineer

Rheology can play an important part in residence time in a mill. Trying to classify a thixotropic slurry is most difficult since the particles are no longer acting as 'individuals' but rather as a mass that entraps the fine particles, preventing them from moving on to the next stage.

Why not re-phrase your question a bit more clearly. Are you trying to push the density to such a limit that you are grinding a paste which essentially means that you are no longer grinding! It would seem obvious that grinding in this condition is no longer grinding. Maybe your question is more how far can you push the density and still achieve acceptable performance. Note that this viscosity that you refer to may in fact change from day to day as a function of feed material such as clays.

You have a good point in re-phrasing your question. Certainly, as the ore mineralogy changes, so will the viscosity. Adding more water can help if the viscosity is high but then your mill capacity will decrease. There is always a price to pay when changes are made to a process circuit.

Ore characterisation is core to my work and one of the 'great unknowns' is what role mineralogy plays and what to extent, it is all part of the scope so thanks. You make a good point with regard to viscosity/density, I do need to address the question of limits to which densities may be pushed especially when dealing with ores that have a tendency to assume a paste-like consistency when ground.

I have seen a ball mill circuit where a cyclone cluster (instead of a thickener) was employed for dewatering of the mill product stream upstream from gold leaching, and the cyclone overflow stream (i.e. water with a low % ultra-fines) was used as mill dilution. Needless to say, a few months down the line the mill started showing serious throughput capacity problems, and this was due to a build-up of ultra-fines in the mill circuit over time i.e. the pulp rheology had gone toxic. The pulp inside the mill was so viscous, I could walk on it (my feet did not make much of an impression on the pulp surface).

So yes, maintaining proper pulp rheology (the combination of pulp SG and viscosity) in the mill is extremely important. However, if the ore characteristics dictate otherwise (such as a high clay content in the ore), it may be best to scrub & classify the ore upstream from passing the coarse size fraction to the mill.

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overflow vs grate discharge ball or mill - why retrofit

overflow vs grate discharge ball or mill - why retrofit

Allow metocompare:Ball mills can be of the overflow or of the grate discharge type. Overflow discharge mills are used when a product with high specific surface is wanted, without any respect to the particle size distribution curve. Overflow discharge mills give a final product in an open circuit.

Grate discharge mills are used when the grinding energy shall be concentrated to the coarse particles without production of slimes. In order to get a steep particle size distribution curve, the mill is used in closed circuit with some kind of classifier and the coarse particles-known as classifier underflow-are recycled.

In the past, I had worked with +10% as an expected increase created by the conversion. How much capacity you gain by using grate discharge over overflow discharge on a mill =The +20% data comes from an old paper.

BallMills have a very large discharge opening or area and smaller area for incoming feed. The gradient between the incoming feed opening and the discharge near the periphery of the shell provides a faster migration of the fines than the oversize particles. In deep pulp level mills commonly known as overflow mills this migration can not occur since material enters and leaves at the same level by displacement only. Independent tests have shown that regardless of mill shape or design, the discharge product of an overflow mill will be the same no matter at which end the feed enters.

Grate Discharge Ball mills withlow pulp levels benefit from the full impact of the grinding media acting on the ore particles, as it falls into the shallow pulp. With a deep pulp level the grinding media is cushioned in the pulp, thus losing its energy and reducing its grinding ability. Grate Discharge Ball Mills have shown 25% to 45% more tonnage ground and a substantial reduction in power consumed per ton of material compared size for size with overflow mills.

A general statement can be made that the closer the discharge is to the periphery of the shell, the quicker the material will pass through and less overgrinding will take place. This is important in both rod mill and ball mill grinding. First, regardless of how fine a grind is required, overgrinding is costly and undesirable. The ideal condition is to remove the particles as soon as they have reached the optimum size. Secondly, in grinding applications where a minimum amount of fine material is preferred, again a rapid flow through the grinding mill is required. These can be accomplished with the grate for ball mill operations, or the various Grate Discharge discharge arrangements for the rod mill.

The discharge end of the conventional Open End Rod Mill is virtually open as the name implies. As a means of controlling splash and to prevent unruly rods from moving out of the mill a discharge plug or plug door arrangement is furnished. The use of this construction permits pulp to discharge freely around the annular opening between the plug and the discharge trunnion liner. By simple removal of the plug the full large area of the discharge end may be used for re-rodding, inspection of the mill when in operation, and an easy access to the mill interior for relining. This large opening does away with the necessity of manholes for mill entrance as commonly employed in the overflow type mill. The plug door arrangement is a great time saver during re-rodding and re-lining operations.

On smaller diameter Grate Discharge Rod Mills a discharge plug is furnished mounted on the trunnion liner and extending through to line up with the discharge head liners. The larger diameter Grate Discharge Mills are furnished with a discharge housing arrangement independent of the mill. A hinged door is mounted in this housing and easily swings in or out of the discharge trunnion liner. These housings are also used to control the direction of discharge pulp flow leaving the mill. Such flow may be directed to the left, right, or directly below the mill centerline.

The discharge housing is of very heavy construction for strength and rigidly. Maintenance of this housing is kept at a minimum, the only wearing part being the replaceable Manganese Steel plug door liner.

The discharge end of a Grate Discharge Ball Mill is fitted with grate sections approximately 3 thick, made of special heat treated alloy steel developed for this particular application. The grate sections have tapered openings between and 7/8 dependent upon the specific grinding application. These are selected to provide the greatest efficiency for any particular job. The grate sections are held in place by tapered Manganese Steel side clamp bars, a center discharge liner, and end clamp bars. The discharge grates are very simple to install and require no attention during operation. The overall life of the discharge end parts generally is greater than that of the feed head liners or shell liners. The discharge end of the Grate Discharge Ball Mill has at least ten times the discharge opening area, through the grate slots, compared to the common trunnion overflow type mill. The discharge grates are designed to run clean and free of any blinding or choking. The pulp level in the mill may be varied by merely changing the pulp dilution. There is no complicated mechanical arrangement to compensate for pulp level changes. The side clamp bars and center discharge liner besides holding the grate sections in place, act as a means of stirring up the ball charge and reduces the amount of wear on the grate sections. The pulp discharges through the grate slots into a lifter compartment in the discharge head, lined with replaceable wearing parts. This lifting compartment elevates the discharge pulp up to the level of the discharge trunnion liner opening and spills this against a deflecting cone which directs it out through the trunnion liner.

The above is a Grate Discharge Mill head with discharge grates, side clamp bars, end damp bars, and center discharge liners in place. The grates and side clamp bars are accurately ground to fit the machined surfaces of the discharge head lifters.

We have already discussed grinding in a general way and have referred numerous times to the grate dischargeprinciple of grinding. To illustrate roughly this principle, take a certain weight of crushed ore and grind it with a mortar and pestle until all of the ore particles will pass through a 65-mesh screen. Then take a similar sample but this time grind for a few minutes and screen at 65-mesh removing the finished material, then return the oversize particles and grind for another short period of time and repeat the screening operation. You would find that the actual net grinding time required for the second sample is about half the time required under the first condition. This same process takes place in the Grate Discharge Ball Mill. It must be borne in mind that it is the classifier which determines the size of the finished product, not the grinding mill itself. The Grate Discharge Mill permits a quick discharge of the finished material into the classifier which makes the desired mesh size separation and returns the oversize particles to the mill for another pass.Contrary to the usual belief, material does not discharge through the grates at the bottom. In fact it is carried up in the ball load so that the greater portion passes out from the ball load on the upturning side of the mill, in the grate area from about half way below the centerline of the mill, on up to the point where the balls start to leave the shell on their downward paths. This indicates then that the thick pulp carried in the mill is well within the ball mass where the actual grinding is taking place. The discharge grates are not to control the size of particle discharged, but merely to retain the grinding balls within the mill, provide the full discharge area required, and form the steep gradient between the feed entrance and product discharge.

To illustrate the comparison of the grate discharge Ball Mill to an overflow type of mill we are showing on page 31 several actual case histories of installations where the performance of grate discharge mills versus overflow mills have been proven. In each such test, run for long periods of time, the ore characteristics and size of feed were maintained identical so that the tests could be compared under like conditions. It will be noted that in each case the grate discharge Mill provided a high increase in tonnage with a lesser increase in power consumption so that the actual KWH per ton consumed was reduced. From these field examples you can verify the previous statement that an overflow type of mill has somewhere near 70% the capacity of the grate mill. These tests were conducted independently by the actual operating companies involved.

The above tables list some of the most common Grate Discharge Ball Mill sizes. Capacities are based on medium hard ore with mill operating in closed circuit under wet grinding conditions at speeds indicated. For dry grinding, speeds are reduced and capacities drop between 30% to 50% .

The above dimensions are approximate and for preliminary use only. Right hand mills are shown. For left hand mills put drive on opposite side. Drive may also be located at feed end. but clearance of scoop must be considered.

The above dimensions are approximate and for preliminary use only. Right hand mills are shown. For left hand mills put drive on opposite side. Drive may also be located at feed end, but clearance of scoop must be considered.

mill charge - an overview | sciencedirect topics

mill charge - an overview | sciencedirect topics

Rod mill charges usually occupy about 45% of the internal volume of the mill. A closely packed charge of single sized rods will have a porosity of 9.3%. With a mixed charge of small and large diameter rods, the porosity of a static load could be reduced even further. However, close packing of the charge rarely occurs and an operating bed porosity of 40% is common. Overcharging results in poor grinding and losses due to abrasion of rods and liners. Undercharging also promotes more abrasion of the rods. The height (or depth) of charge is measured in the same manner as for ball mill. The size of feed particles to a rod mill is coarser than for a ball mill. The usual feed size ranges from 6 to 25mm.

For the efficient use of rods it is necessary that they operate parallel to the central axis and the body of the mill. This is not always possible as in practice, parallel alignment is usually hampered by the accumulation of ore at the feed end where the charge tends to swell. Abrasion of rods occurs more in this area resulting in rods becoming pointed at one end. With this continuous change in shape of the grinding charge, the grinding characteristics are impaired.

The bulk density of a new rod charge is about 6.25t/m3. With time due to wear the bulk density drops. The larger the mill diameter the greater is the lowering of the bulk density. For example, the bulk density of worn rods after a specific time of grinding would be 5.8t/m3 for a 0.91m diameter mill. Under the same conditions of operation, the bulk density would be 5.4t/m3 for a 4.6m diameter mill.

During normal operation the mill speed tends to vary with mill charge. According to available literature, the operating speeds of AG mills are much higher than conventional tumbling mills and are in the range of 8085% of the critical speed. SAG mills of comparable size but containing say 10% ball charge (in addition to the rocks), normally, operate between 70 and 75% of the critical speed. Dry Aerofall mills are run at about 85% of the critical speed.

The breakage of particles depends on the speed of rotation. Working with a 7.32m diameter and 3.66m long mill, Napier-Munn etal. [4] observed that the breakage rate for the finer size fractions of ore (say 0.1mm) at lower speeds (e.g., 55% of the critical speed) was higher than that observed at higher speeds (e.g., 70% of the critical speed). For larger sizes of ore (in excess of 10mm), the breakage rate was lower for mills rotating at 55% of the critical speed than for mills running at 70% of the critical speed. For a particular intermediate particle sizerange, indications are that the breakage rate was independent of speed. The breakage ratesize relation at two different speeds is reproduced in Figure9.7.

The blending of different ore types is a common practice to provide a consistent feed to a process in terms of uniform hardness or assay. When several different ore deposits of varying grindabilities are blended prior to closed circuit grinding, the work index of the ore is not an average or even a weighted average of the work indices of the components. The reason for this is that the circulating load will consist predominantly of the harder component and if the circulating load is high then the mill charge will also consist of mostly the harder components. Thus, the work index of the blend will be weighted towards the harder components [39]. Figure3.16 shows the Bond work index of a blend of hard and soft ores as a function of the volume fraction of the softer ore in the blend. The dotted line between the two extremes indicates the weighted average work index based on volume fraction. The work index values of the Magdalinovic method agree with this average Bond work index because the method does not simulate the recycling of harder components into the mill charge. On the other hand, the work index obtained using the standard Bond test shows the weighting of the work index towards the harder component as a result of the circulating load.

Yan and Eaton [39] also measured the breakage rates and breakage distribution functions of the different ore blends in order to predict the work index of the blend by simulation of the Bond batch grinding test. Qualitative analysis of the breakage properties suggests that there is an interaction between the components of the blend that affect their individual breakage rates. The breakage properties of the harder material appear to have a greater influence on the overall breakage properties and the Bond work index of the blend than the softer material.

Whereas most of the ball-milled systems usually prepared with using ball-to-powder weight ratio (Wb:Wp) in the range between 10:1 and 20:1, the effect Wb:Wp on the amorphization reaction of Al50Ta50 alloy powders in a low-energy ball mill was investigated in 1991 by El-Eskandarany etal.[42] They have used 90, 30, 20, 10, and 3g of powders to obtain Wb:Wp ratios of 12:1, 36:1, 54:1, 108:1, and 324:1, respectively.

The XRD patterns of mechanically alloyed Al50Ta50 powders as ball-milled for 1440ks (400h) as a function of the Wb:Wp ratio is presented in Fig.4.32. Single phase of amorphous alloys is obtained when ratios 36:1 and 108:1 were used. The Bragg peaks of elemental Al and Ta crystals still appear when the Wb:Wp ratio is 12:1, indicating that the amorphization reaction is not completed. In contrast, when the Wb:Wp ratio is 324:1, the amorphous phase coexists with the crystalline phases of AlTa, AlTa2, and AlTaFe.

Based on their results,[42] it is concluded that the rate of amorphization depends strongly on the kinetic energy of the ball mill charge and this depends on the number of opportunities for the powder particles to be reacted and interdiffused. Increasing the Wb:Wp ratio accelerates the rate of amorphization, which is explained by the increase in the kinetic energy of the ball mill charge per unit mass of powders. It has been shown in this study that the volume fraction of the amorphous phase in the mechanically alloyed ball-milled powders increases during the early stage of milling, 86173ks (48h) with increasing Wb:Wp ratio. It is noted that further increasing this weight ratio leads to the formation of crystalline phases and this might be related to the high kinetic energy of the ball mill charge which is transformed into heat. When the Wb:Wp ratio was reduced to 12:1, however, the amorphization reaction was not completed. This indicates that the kinetic energy of the mill charge is insufficient for complete transition from the crystalline to the amorphous phase.

It is worth noting that powder particles reached the minimum of extreme fineness when using a high Wb:Wp ratio. One disadvantage of using such a high weight ratio is being the high concentration of iron contamination which is introduced to the milled powders during the MA process, as presented in Fig.4.33.

Romankova etal.[43] have applied the vibration ball milling for coating of stainless steel balls during milling of TiAl powders. They examined metallographically the development of the TiAl coating structure after milling for 60min as a function of the ball-to-powder weight ratio for 6mm balls (Fig.4.34).

The results showed that the milling energy increased with increasing the number of balls. When the weight ratio was 3:1, the substrate could be covered with a thin Al layer (Fig.4.34A). For this case, only small Ti particles were embedded into the Al matrix. It should be noted that the substrate underwent plastic deformation under the ball impacts and its surface became slightly bent. When the weight ratio was increased to 4:1, the energy was sufficient to embed larger Ti particles in the Al layer than at ratio 3:1 (Fig.4.34B). Al bound these Ti particles to the substrate. They notified that, at the 4:1 ratio, the growth for the TiAl coating across the substrate was clustered; this resulted in a hillock-like morphology and increased the surface roughness. Upon further increasing the ball-to-powder weight ratio from 6:1 to 14:1, the coating roughness gradually decreased. They also reported that the lamellar structure was refined when the ball-to-powder weight ratio was 14:1, as presented in Fig.4.34E.

More recently, Waje etal.[44] have studied the effect of the ball-to-powder weight ratio (BPR) on the crystallite size of ball-milled CoFe2O4 nanoparticles, using XRD (Fig.4.35). From their results it can be seen that the particle size decreases linearly from 15.3 to 11.4nm when used BPR of 8:1 and 30:1, respectively.

The mass-size balance models as written above are in the time-domain. To be more practical they need to be converted to the energy-domain. One way is by arguing that the specific rate of breakage parameter is proportional to the net specific power input to the mill charge (Herbst and Fuerstenau, 1980; King, 2012). For a batch mill this becomes:

where SiE is the energy-specific rate of breakage parameter, P the net power drawn by the mill, and M the mass of charge in the mill excluding grinding media (i.e., just the ore). The energy-specific breakage rate is commonly given in t kWh1. For a continuous mill, the relationship is:

where is the mean retention time, and F the solids mass flow rate through the mill. Assuming plug flow, Eq. (5.17) can be substituted into Eq. (5.15) to apply to a grinding mill in closed circuit (where t=).

The distinctive feature of tumbling mills is the use of loose crushing bodies, which are large, hard, and heavy in relation to the ore particles, but small in relation to the volume of the mill, and which occupy (including voids) slightly less than half the volume of the mill.

Due to the rotation and friction of the mill shell, the grinding medium is lifted along the rising side of the mill until a position of dynamic equilibrium is reached (the shoulder), when the bodies cascade and cataract down the free surface of the other bodies, about a dead zone where little movement occurs, down to the toe of the mill charge (Figure 7.3).

The driving force of the mill is transmitted via the liner to the charge. The speed at which a mill is run and the liner design governs the motion and thus nature of the product and the amount of wear on the shell liners. For instance, a practical knowledge of the trajectories followed by the steel balls in a mill determines the speed at which it must be run in order that the descending balls shall fall on to the toe of the charge, and not on to the liner, which could lead to liner damage. Simulation of charge motion can be used to identify such potential problems (Powell et al., 2011), and acoustic monitoring can give indication of where ball impact is occurring (Pax, 2012).

At relatively low speeds, or with smooth liners, the medium tends to roll down to the toe of the mill and essentially abrasive comminution occurs. This cascading leads to finer grinding and increased liner wear. At higher speeds the medium is projected clear of the charge to describe a series of parabolas before landing on the toe of the charge. This cataracting leads to comminution by impact and a coarser end product with reduced liner wear. At the critical speed of the mill centrifuging occurs and the medium is carried around in an essentially fixed position against the shell.

In traveling around inside the mill, the medium (and the large ore pieces) follows a path which has two parts: the lifting section near to the shell liners, which is circular, and the drop back to the toe of the mill charge, which is parabolic (Figure 7.4(a)).

Consider a ball (or rod) of radius r meters, which is lifted up the shell of a mill of radius R meters, revolving at N rev min1. The ball abandons its circular path for a parabolic path at point P (Figure 7.4(b)), when the weight of the ball is just balanced by the centrifugal force, that is when:

Mills are driven, in practice, at speeds of 5090% of critical speed. The speed of rotation of the mill influences the power draw through two effects: the value of N and the shift in the center of gravity with speed. The center of gravity first starts to shift away from the center of the mill (to the right in Figure 7.4(a)) as the speed of rotation increases, causing the torque exerted by the charge to increase and draw more power (see Section 7.2.2). But, as critical speed is reached, the center of gravity moves toward the center of the mill as more and more of the material is held against the shell throughout the cycle, causing power draw to decrease. Since grinding effort is related to grinding energy, there is little increase in efficiency (i.e., delivered kWh t1) above about 4050% of the critical speed. It is also essential that the cataracting medium should fall well inside the mill charge and not directly onto the liner, thus excessively increasing steel consumption.

At the toe of the load the descending liner continuously underruns the churning mass, and moves some of it into the main mill charge. The medium and ore particles in contact with the liners are held with more firmness than the rest of the charge due to the extra weight bearing down on them. The larger the ore particle, rod, or ball, the less likely it is to be carried to the breakaway point by the liners. The cataracting effect should thus be applied in terms of the medium of largest diameter.

As already discussed, this control loop is provided to maintain the PA header pressure before the mixing of hot and cold PA duly controlled for temperature. FigureVIII/4-2 is also applicable for this type of mill when the PA is common to all the mills. The control loop is of course different for individual PA fan systems, as the above is applicable for the common PA system only. For control loop description, see Section 4.3.2.3 of this chapter. Common PA fans are provided with suction normally from the atmosphere or it may be from the FD discharge header. Header pressure control is performed through various types of final control elements.

As the fuel/load control is solely done by position adjustments to the PA damper near the mill, this control loop assists smooth and bumpless control of the fuel flow transported by the PA flow to the mill as the upstream PA header pressure control takes responsibility for providing an adequate quantity of air at any environmental condition without sacrificing the required downstream pressure,

FigureVIII/5.3-3 later in the chapter depicts the simple control loop. Any of the mill DP transmitters or level (sound-detector) transmitters is selected and the selected signal is connected to the controller as the process or measured variables against a fixed-level set point. Sufficient redundancy in measurement may vary according to the plants operating philosophy. The controller output is utilized for adjustment of feeder speed with the help of a VFD or SCR control for the gravimetric feeder/feeder speed variator.

At the higher load the charge level inside the drum decreases and the feeder speed should increase accordingly to replenish the material. For a decreasing load, the reverse action takes place. To take care of the sudden load change, the deviation between characterized PA flow and DP acrossthe mill is used to modify the controller output to achieve the desired mill charge level.

Mill load or fuel flow control follows the fuel demand from the boiler master demand control signal and is achieved by regulating the quantity of PA that is transporting agent only. Figures VIII/5.2-4 and VIII/5.2-4 depict the functioning of the control loop, which is similar to that of other mill types. For other mills the fuel demand signal from the boiler master demand is first taken care of by the mill-wise PA flow control system if the demand is less than the prevailing air flow control system. The characterized PA flow is then construed as the feeder speed demand. The ball-and-tube mill control system, on the other hand, uses feeder speed control for maintaining mill level control only and so the fuel flow control is achieved through control of the feeder-wise PA flow to mill itself.

However, the feeder-wise PA flow as measured after redundant transmitters voting selection and density compensated through temperature correction is again determined to get equivalent fuel flow. The total fuel flow is then computed by summing all the fuel (PF) flow of the running feeders and the supporting fuel (oil or gas) if any are being utilized at that time with proper weightage, taking consideration of their thermal or calorific value. The higher selection of this total equivalent fuel flow signal and the air flow demand signal from the boiler master demand (FigureVIII/2.1) is then taken as actual air demand just as in other type of mills.

As already discussed in Section 5.2.1, there is another feeder-wise control system associated with fuel flow control known as a bypass damper control. This feeder-wise damper is provided for each mill end for preheating the raw feed, which is an essential requirement during startup. No process measurement signal is utilized in this subloop. The same fuel demand from the boiler master demand (FigureVIII/ 2-1) is taken as the set point for the position demand of the bypass damper after due characterization, as shown (refer to Figures VIII/5.2-4 and VIII/5.2-5) in the control strategy and the graphical representation of approximate positions of the two final control elements. The previously mentioned two-position demands operate in opposite directions. After being in a fully open position for a certain load, ensuring elimination of initial moisture, this bypass damper begins to close gradually as the load increases.

There are two main types of fuel flow controls achieved through the proportionate PA flow only: (1) common PA fans with individual PA dampers and (2) individual PA fans with vane or speed control. There is also one known as a mill-wise PA flow control that is common to both sides.

FigureVIII/5.2-4 may be referred to for this type of control along with FigureVIII/5.2-2. Here the mill PA flow and bypass PA flows are combined to form the total mill-wise PA flow to the furnace. The boiler master demand acts as a set point here, where the mill-wise PA flow is the measured value as this air flow is only responsible for transporting the fuel to the furnace. The controller output is the demand signal for the individual PA damper. For bypass dampers, the boiler master demand generates the set point while the actual position of this damper acts as the measured value for the controller output, which is the demand signal for the bypass damper.

For any load change, the two flows readjust their positions to deliver the required PA flow. For higher load the bypass damper tends to close to allow less flow for preheating of raw feed and the PA damper to the mill opens more to take care of the load demand.

FigureVIII/5.2-5 may be referred to for this type of control along with FigureVIII/5.2-2(a). Here bypass PA flows need to be subtracted from the total mill-wise PA flow for the fuel flow control, and the total mill-wise PA flow to the furnace is required for air flow control. The reason for this is that the final control element and the flow element are both located in the common primary air path to the individual mill. The boiler master demand acts as a set point here, where the PA flow to the mill is the measured value. The controller output is the demand signal for the individual PA vane or variable speed drive as the case may be.

This type of mill design vis--vis operation is somewhat different from other types, as discussed earlier. FigureVIII/5.2-2(b), which is mainly followed by manufacturers, such as the Foster Wheeler Energy AG corporation, may be referred to for information. Here the boiler combustion control signal regulates the output of the mill by PA flow control dampers placed in the common line to both the ends or sides. The predrying of coal feed is done at the entry of each side before it enters the drum, unlike what is done by the bypass PA damper in many types of tube mills.

Another significant difference is the provision of an auxiliary air and purge air supply line taken from the cold PA for each side of the mill drum. The same is designed to the required minimum velocities of the PA/fuel mixture for maintaining proper flow inside the coal duct and to prevent fuel settling during startup or in extreme low-load operation. This feature extends the individual mill load range without encountering drifting or pulsating fuel flow to the burners. The other purpose is to purge the coal air line automatically when burners are taken out of service.

The feed level control in the drum, classifier outlet temperature control, and seal air DP control are very much similar to those in the other type of mills with the exception of the source of the seal air. Here the seal air supply is taken from the cold PA without any provision of a seal air fan.

Selecting dispersion equipment for a specific application is a complex task. Dispersion of the mixture must be complete and the process and equipment must meet economic constraints. But much more is involved. In practice, such simple criteria are complicated by a variety of parameters related to fillers and to the materials in which they are dispersed. These parameters complicate the problem to the degree that it is not easy to formulate general guidelines. In this discussion we will consider the available equipment types most frequently used for filler dispersion and illustrate their applicability with some examples.

A ball mill is an effective means of dispersing solid materials in solids or liquids.8,9 Ball mills have several advantages which include versatility, low cost of labor and maintenance, the possibility of unsupervised running, no loss of volatiles, and a clean process. The disadvantages are related to discharging viscous and thixotropic mixtures, and considerably lower efficiency when compared with other mixing equipment. The millbase viscosity is usually restricted to about 15-20 Poise, and therefore ball mills are most frequently found in production applications such as paints, flexographic, publication gravure, and letterpress news inks, and carbon paper inks which are dispersed at elevated temperatures.

The mill should rotate at 50-65% of the theoretical centrifugal speed in order to allow balls to cascade, since the cascading balls grind most effectively and do not cause an excessive loss of ball material

Viscosity, the order of filler addition, and the quantity of material should be chosen so as not to cause a viscosity increase above the specified range, since the milling efficiency drastically decreases at that point

The degree of dispersion and jetness achieved when grinding carbon black depends on the wetting properties of the dispersing material and to some degree on the filler form. For instance, pelletized carbon black is easier to disperse than a fluffy type

The sandmill has some drawbacks. It is a two stage process (premixing followed by milling). Milling develops high temperatures in the mixture which causes loss of volatiles and requires cooling. If the millbase is high in viscosity or dilatant, the sandmill process may not work at all. Agglomerated or extremely hard pigments are difficult or impossible to disperse

Both ball and sand mills operate based on a viscous shear principle, thus the viscosity of the millbase is a critical factor in achieving dispersion. The size of filler particles is critical, especially in sandmills. It was found that the shearing force is inversely proportional to the square of the linear size of filler agglomerate. An agglomerate of diameter of 7 m attains 100 times the shear stress of an agglomerate of 70 m diameter. The difference between the ball mill and the sand mill is in the size and density of the grinding media, which is reflected in their performance. Sandmilling uses small particles of low density, and therefore, there is no noticeable reduction in the size of the sand particle, whereas the balls in ballmills are very much larger and may have a high density (steel), which results in a more complex mechanism of grinding including shattering and impacting which cause this mill to be more effective in disintegrating hard particles and agglomerates containing sintered particles.

There is another mill type called an attritor, which is similar to both the ball mill and the sandmill. In construction, it is similar to a sandmill. It also has a vertical shaft, but in the attritor the agitator bars replace the milling discs of the sandmill. It is also similar to a ball mill because it uses balls, usually ceramic ones having 5-15 mm in diameter. Because the motion of the balls is independent of gravity, an attritor can handle thixotropic materials and slightly higher viscosity of millbases, but the principle of action and type of forces operating are similar to those of the ball mill. An attritor applied to pigment dispersion gives several advantages. These include rapid dispersion, the possibility of either a continuous or batch process, low power consumption, small floor space, and easy cleaning and maintenance. Their main disadvantage is high heat generation. Attritors are equipped with a cooling water jacket which can control the heat flow to some extent, but conditions are often too severe for some resins, which may degrade during the process.

Three-roll, one-roll, and stone mills constitute a more mature dispersion technology still in use with medium viscosity millbases. A three-roll mill consists of the feed, center, and apron rolls. In roll mill operation:

The speeds of feed and apron rolls are adjustable, and each roll rotates with a different speed in order to induce shear in the material at the nip and facilitate the material transfer from one roll to the other

For mechanical reasons the gap between rolls cannot be less than 10 m and it usually ranges from 40 to 50 m.7 Small particles will not be affected as they pass through the nip, but agglomerates smaller than the distance between rolls will be disintegrated due to the shear stress imposed on the material

The one-roll mill works on a similar principle but the nip is regulated by a pressure bar. Shearing takes place between the roller and the shearing bar. Stone mills have similar principles of operation. The rotor turns on a stator to achieve shearing

With current raw materials, both the primary particles and agglomerates are very small, and if any positive action can be achieved during the milling process, it can only be done by affecting these small particles. It is thus necessary to operate these machines at very tight gaps which causes abrasion of the mechanical elements, rapid deterioration of equipment, and contamination of the product by the abraded material. This affects the properties of the millbase and the color of the product

The high-speed impeller or shear mixer is the most common equipment to prepare dispersions of solids in liquid. High speed shear mills and kinetic shear mills have retained their usefulness because of their ability to deagglomerate material that is not adequately dispersed in the premixing step. A high-speed shear mill is composed of two elements a container and an impeller. These factors are important in the design:

In the first stage, the viscosity changes from low to high as fillers are incorporated; in the second stage, viscosity remains constantly high because of the disintegration of particles which occurs during the application of the highest shear stress

Long mixing increases temperature and decreases viscosity. This does not provide the conditions for the best filler dispersion. By extending mixing over, for example, a 15 min period, the degree of dispersion is not improved, but the resin may actually be degraded

If the quality of dispersion is not satisfactory, the parameters of mixing should be changed. If the expected result cannot be attained, the range of conditions available is not adequate in this particular mill

In the third stage, the viscosity changes from high to low due to the addition of diluent. The viscosity range which can be handled by high speed mixers is similar to the range of a three-roll mill, i.e., up to about 200 Poise

The range of shear rates available in high-speed mixers is not broad. The flow rate of fluid in motion decreases as viscosity increases and is inversely proportional to the width of the flow passage which, in this case, is the distance between the disperser and the container which is very large in a high speed mixer. It is not so much due to an improvement in mixing equipment that high-speed mixers have become so popular, it is mostly because of the high quality raw materials (pigments, fillers) which are available now. High structure carbon blacks can be more easily dispersed. But with the increased structure, the size of the primary particles decreases, inhibiting dispersion. Because of the interrelation between both parameters, only the medium structure, coarser particles of carbon blacks can be dispersed by high-speed mixers. Other carbon black types demand further treatment. It should be noted that this is only true of a few fillers which are known to possess strongly bonded, small sized particles. In most cases, fillers can be successfully dispersed in high-speed mixers. However, care should be taken that the filler is selected with an appropriate particle size.

High-speed mixers have several important advantages over other existing equipment including the possibility of processing a batch in the same vessel, easy cleaning, and flexibility in color changes. The main disadvantage is that the final dispersion depends greatly on the chosen composition and technology, and these are sometimes limiting factors. Frequently, the proper conditions for quality dispersion cannot be achieved at all.

The basic construction of a high-speed mixer can easily be modified to one's special requirements. For example, a change from impeller to turbine rotor changes both the principle of dispersion and the range of application. The tangential velocities of filler particles can be as high as 500 m/sec. Such particles have a very high kinetic energy, sufficient to cause size reduction. Size reduction is due to particle-particle or particle-wall collisions, and this in turn, is related in efficiency to the relative velocities at the moment of collision. Relative velocity can be increased by decreasing the viscosity of the millbase. The upper limit of millbase viscosity is somewhere around 3 to 4 Poise. It is not viscosity alone which is important but the entire rheological character of the millbase. The best results are obtained when the millbase is nearly Newtonian. For this reason, the dispersion process is best performed in a diluted millbase. As is the case with high-speed mixers, a proper dispersion should be achieved in a matter of 10-20 min. If such is not the case, the conditions of processing should be modified. Once dispersion has been achieved, it should be stabilized, with the mixer continuously running, by the addition of more resin to increase the viscosity in order to prevent sedimentation or flocculation of the pigment.

The other possible modification to such a mixer can be achieved by a substantial lowering of the speed and a change in the motion of the mixing element to planetary. This configuration can process material of a much higher viscosity, up to several thousand Poise. The high speed mixer can be modified in various ways to match its capabilities to the process requirements. Stationary baffles may be added to increase the shear rate. The distance between the rotating and stationary elements can be decreased again increasing the shear rate. The mixer may be designed to work under both pressure and vacuum and with inert gas blanketing which permits deaeration and processing of volatile or moisture sensitive materials.

The other group includes heavy-duty mixers, such as the Banbury mixer and double-arm kneading mixers. The Banbury mixer with a power input of up to 6000 kW/m3 is the strongest and the most powerful mixing unit used by industry. Nearly solid materials are mixed by a rotor which is a heavy shaft with stubby blades rotating at up to 40 rpm. The clearance between the walls and rotor is very small, which induces a very high shear in the material. The high shear generates a great amount of heat which melts the polymer rapidly and allows for quick incorporation of filler. After the filler is incorporated, the dispersion process begins, with rapid distributive mixing along and between two rotors and between the chamber walls and rotor tips. Within 2-3 min, mixing is normally completed and the compound discharged into a pelletizing extruder or a two-roll mill which converts it to a sheet form.8 Carbon black, which is most frequently processed in a Banbury mixer, is usually placed between two layers of polymeric material in order to reduce dusting.

Double-arm kneading mixers are very popular in some industries. They consist of two counter-rotating blades in a rectangular trough carved at the bottom to form two longitudinal half cylinders and a saddle section. A variety of blade shapes are used, with a clearance between them and the blades and the side walls of up to 1 mm. The most popular blade shapes include: sigma, dispersion, multiwiping overlapping, single-curve, and double-naben blades. It is important for filler dispersion in this mixer that the viscosity of the millbase be kept high enough to create the required shearing force to disperse the material. The strong construction of the mixer and its high power allow one to work with concentrated compositions of pigments which could not be processed by any other method.

High volume production is done by mixing in an extruder.11 This method offers several advantages such as a continuous process, material uniformity, a clean environment, high output, and low labor. The biggest disadvantage of this method is a high investment cost. The twin-screw extruder is the most flexible type of extruder and most appropriate for compounding. Their screw designs can be varied as can the method of dosing and the output rate. The abrasiveness of the filler may affect the life-span of the equipment, and particle size and its distribution may influence the quality of filler dispersion and material uniformity. But in general, there is adequate machinery available for almost all requirements. For instance, glass-fiber reinforced materials can be produced by this technique with little change to the initial structure and dimensions of the glass fibers, which shows the versatility of the technology. The production rate of this method is comparable to the Banbury mixer, and an additional advantage comes from the fact that the material can be completely processed in one pass through the machinery.

The importance of the proper dispersion of fillers and the complexity of techniques for measuring the degree of dispersion are reflected in numerous publications. Further information on the mixing of fillers is included in Chapter 18.

The renewable power sources are being explored due to possibility of lack in availability of conventional resources in future. The major drawback of Renewable energy resources are dependency on geographical locations and environmental conditions however, the high initial cost, increased maintenance cost, and different rates of depreciation are the main challenges associated with these hybrid systems[18]. The irregular pattern of natural resources necessitates developing a hybrid system which can generate maximum conceivable energy for continuous and reliable operations [17]. The design of hybrid system is influenced by various factors such as condition of sites, energy availability, efficiency of energy sources as well as technical and social limitation In this specific situation, a combination of optimal sizing method is an indispensable factor to accomplish higher reliability quality with least expense [21,79,87,149]. The fundamental parts of the hybrid energy systems are renewable power source, nonrenewable generators, control unit, storage system, load or grid some times, sources and load may be AC/DC [102].

An arrangement of the renewable power generation with appropriate storage and feasible amalgamation with conventional generation system is considered as hybrid energy system or some time referred as a micro grid [155]. This system may be any probable combination of Photovoltaic, wind, micro turbines, micro hydro, conventional diesel generation, battery storage, hydrogen storage and Fuel Cell in grid-connected or off grid arrangement,

An assembly of interconnected loads, conventional distributed energy resources like distributed generators (DG), renewable resources and energy storage systems in a specified boundary as a controllable single entity referred as micro grid. It may be eternally connected to grid, or isolated by grid. There are worldwide numerous remote communities those are not directly connected to grid, and fulfill electricity demand from distributed generators based on fossil fuel in isolated Microgrids[97,165]. In this paper a assimilated arrangement of solar PV and wind renewable energy resources is discussed which is slightly different from the concept of microgrid.

Solar Photovoltaic /Wind based Hybrid Energy System shows its adequacy to provide the essential electrical demand for off grid utilization. The at most imperative feature of a Solar Photovoltaic (PV) and Wind based Hybrid Energy System is that it uses at least two sustainable power sources which enhances reliability, efficiency and financial restrictions emerges from single energy resources of renewable nature [18,89,133]. Solar Photovoltaic and Wind based Hybrid Energy System is considered as amalgamation of solar PV panel, Wind mills, charge controller, storage system, power conditioning units, diesel based generator set and load [19]. The assessment of performance of Hybrid system can be done by recreating their models at Simulink platform for the accessible insulation, speed of wind, electrical load and various components [20].

The essential objective for evaluation of Hybrid System are building up the suitable models for various components and their simulation in a sequential manner as firstly availability of speed wind, accessibility of sunlight and the demand of load models are simulated after that model of battery storage and diesel generator can be Simulated. Last strides in the entire procedure of assessment is deciding the coveted criteria and exploring the optimum structure of system. [21]. The optimal hybrid system arrangement should satisfy and compromise the objectives of power reliability and cost of system. The load demand frequently considered as limitation of the optimization issue and ought to be totally satisfied [22]. The solar PV/wind hybrid system is mostly reliant on execution of individual segments. To estimate the performance of solar PV/wind hybrid system, individual components are modeled initially after that entire system evaluated to meet the demand [23]. In general key aspects to analyze a hybrid system are hybrid system configuration with respect to the available resources, the optimization of the available renewable resources exploitation and the optimization of the output power quality [24].

Solar energy and wind energy are analogous to each other in nature and both are well appropriate to develop a hybrid system [26]. Availability of solar radiations are relatively greater in summer, winds are more accessible in the evening times of winters. This hybrid renewable energy systems give a more reliable output throughout the year can be planned to fulfill craved qualities on more decreased possible cost [27]. The constraints of Photo voltaic system, the assessed energy of wind energy system and the battery storage are the majorly considered parameters for evaluation of solar and wind based hybrid energy system. In addition, the precise angular attitude of Photo voltaic panels and the tower height of wind turbines are considered for achieving the minimum levelised cost of energy. Ribeiro [31] proposed multi-criteria based analytical decision scheme abbreviated as MCDA which consider several issues like economic, quality of life, technical and environmental issues of local populations.

Metrological data based on technological, economical, socio-political and environmental factors having major impact for estimation and selection of various components of Solar Photovoltaic and Wind based Hybrid Energy System [32]. Hourly climate information as sun oriented radiation, wind speed and temperature are raw information illustrates the inconstancy of the parameter input. Place to place data is hard to obtain for designing purpose at remote location [3,73]. Statistical metrological climatic information can be delivered by the average of month to month meteorological information. The information of climate can be anticipated from an adjacent site or synthetically can be produced [32]. Simulation for performance of Solar PV/Wind Hybrid Energy System required climate data including solar radiation, speed of wind and temperature which can be find from web sources and also from local meteorological station, it is best to find realistic solution preference should be given to the specified location based weather data [28]. To optimize solar photovoltaic and wind based hybrid energy system are hourly or day by day climate information of solar and wind energy are considered as required significant inputs [29]. Meteorological data determined the receptiveness and amount of sunlight based radiation and wind energy sources at a particular region. An investigation of characteristics of sun based emission and availability of wind at a specific location ought to be concluded before starting [28]. Bianchini A et al. gives stress on the metrological data in the form of solar irradiance and wind distribution and considered hybrid renewable energy system as a amalgamation of PV panel of rated power, horizontal axis wind turbine of rated power, a diesel generator of precise nominal power able to manage peak load and a battery bank of specific storage capacity [33]. Hall et al. [34] proposed the well-known engineered climate information term Typical meteorological year (TMY) utilized in simulation of solar energy model is first time. It is observational technique picking particular months from different years using the Fleckenstein Schafer accurate system [35].

load demand play a very important role in establishment of solar PV/wind hybrid renewable energy system provides more reliable power for off-grid and standalone applications compared to individual systems [21] The most of the reviewed studies are about the alone Solar Photovoltaic /Wind based Hybrid Energy System and few studies are available for grid connected system. The unsatisfied load request is procured from the grid. Along this way the hybrid system became noticeably trustworthy. The stand-alone systems with storage infused surplus energy to the grid at a prime cost. Along these lines, the grid connected system becomes more financially acceptable.

what's the difference between sag mill and ball mill - jxsc machine

what's the difference between sag mill and ball mill - jxsc machine

A mill is a grinder used to grind and blend solid or hard materials into smaller pieces by means of shear, impact and compression methods. Grinding mill machine is an essential part of many industrial processes, there are mainly five types of mills to cover more than 90% materials size-reduction applications.

Do you the difference between the ball mill, rod mills, SAG mill, tube mill, pebble mill? In the previous article, I made a comparison of ball mill and rod mill. Today, we will learn about the difference between SAG mill vs ball mill.

AG/SAG is short for autogenous mill and semi-autogenous mill, it combines with two functions of crushing and grinding, uses the ground material itself as the grinding media, through the mutual impact and grinding action to gradually reduce the material size. SAG mill is usually used to grind large pieces into small pieces, especially for the pre-processing of grinding circuits, thus also known as primary stage grinding machine. Based on the high throughput and coarse grind, AG mills produce coarse grinds often classify mill discharge with screens and trommel. SAG mills grinding media includes some large and hard rocks, filled rate of 9% 20%. SAG mill grind ores through impact, attrition, abrasion forces. In practice, for a given ore and equal processing conditions, the AG milling has a finer grind than SAG mills.

The working principle of the self-grinding machine is basically the same as the ball mill, the biggest difference is that the sag grinding machine uses the crushed material inside the cylinder as the grinding medium, the material constantly impacts and grinding to gradually pulverize. Sometimes, in order to improve the processing capacity of the mill, a small amount of steel balls be added appropriately, usually occupying 2-3% of the volume of the mill (that is semi-autogenous grinding).

High capacity Ability to grind multiple types of ore in various circuit configurations, reduces the complexity of maintenance and coordination. Compared with the traditional tumbling mill, the autogenous mill reduces the consumption of lining plates and grinding media, thus have a lower operation cost. The self-grinding machine can grind the material to 0.074mm in one time, and its content accounts for 20% ~ 50% of the total amount of the product. Grinding ratio can reach 4000 ~ 5000, more than ten times higher than ball, rod mill.

Ball mills are fine grinders, have horizontal ball mill and vertical ball mill, their cylinders are partially filled with steel balls, manganese balls, or ceramic balls. The material is ground to the required fineness by rotating the cylinder causing friction and impact. The internal machinery of the ball mill grinds the material into powder and continues to rotate if extremely high precision and precision is required.

The ball mill can be applied in the cement production plants, mineral processing plants and where the fine grinding of raw material is required. From the volume, the ball mill divide into industrial ball mill and laboratory use the small ball mill, sample grinding test. In addition, these mills also play an important role in cold welding, alloy production, and thermal power plant power production.

The biggest characteristic of the sag mill is that the crushing ratio is large. The particle size of the materials to be ground is 300 ~ 400mm, sometimes even larger, and the minimum particle size of the materials to be discharged can reach 0.1 mm. The calculation shows that the crushing ratio can reach 3000 ~ 4000, while the ball mills crushing ratio is smaller. The feed size is usually between 20-30mm and the product size is 0-3mm.

Both the autogenous grinding mill and the ball mill feed parts are welded with groove and embedded inner wear-resistant lining plate. As the sag mill does not contain grinding medium, the abrasion and impact on the equipment are relatively small.

The feed of the ball mill contains grinding balls. In order to effectively reduce the direct impact of materials on the ball mill feed bushing and improve the service life of the ball mill feed bushing, the feeding point of the groove in the feeding part of the ball mill must be as close to the side of the mill barrel as possible. And because the ball mill feed grain size is larger, ball mill feeding groove must have a larger slope and height, so that feed smooth.

Since the power of the autogenous tumbling mill is relatively small, it is appropriate to choose dynamic and static pressure bearing. The ball bearing liner is made of lead-based bearing alloy, and the back of the bearing is formed with a waist drum to form a contact centering structure, with the advantages of flexible movement. The bearing housing is lubricated by high pressure during start-up and stop-up, and the oil film is formed by static pressure. The journal is lifted up to prevent dry friction on the sliding surface, and the starting energy moment is reduced. The bearing lining is provided with a snake-shaped cooling water pipe, which can supply cooling water when necessary to reduce the temperature of the bearing bush. The cooling water pipe is made of red copper which has certain corrosion resistance.

Ball mill power is relatively large, the appropriate choice of hydrostatic sliding bearing. The main bearing bush is lined with babbitt alloy bush, each bush has two high-pressure oil chambers, high-pressure oil has been supplied to the oil chamber before and during the operation of the mill, the high-pressure oil enters the oil chamber through the shunting motor, and the static pressure oil film is compensated automatically to ensure the same oil film thickness To provide a continuous static pressure oil film for mill operation, to ensure that the journal and the bearing Bush are completely out of contact, thus greatly reducing the mill start-up load, and can reduce the impact on the mill transmission part, but also can avoid the abrasion of the bearing Bush, the service life of the bearing Bush is prolonged. The pressure indication of the high pressure oil circuit can be used to reflect the load of the mill indirectly. When the mill stops running, the high pressure oil will float the Journal, and the Journal will stop gradually in the bush, so that the Bush will not be abraded. Each main bearing is equipped with two temperature probe, dynamic monitoring of the bearing Bush temperature, when the temperature is greater than the specified temperature value, it can automatically alarm and stop grinding. In order to compensate for the change of the mill length due to temperature, there is a gap between the hollow journal at the feeding end and the bearing Bush width, which allows the journal to move axially on the bearing Bush. The two ends of the main bearing are sealed in an annular way and filled with grease through the lubricating oil pipe to prevent the leakage of the lubricating oil and the entry of dust.

The end cover of the autogenous mill is made of steel plate and welded into one body; the structure is simple, but the rigidity and strength are low; the liner of the autogenous mill is made of high manganese steel.

The end cover and the hollow shaft can be made into an integral or split type according to the actual situation of the project. No matter the integral or split type structure, the end cover and the hollow shaft are all made of Casting After rough machining, the key parts are detected by ultrasonic, and after finishing, the surface is detected by magnetic particle. The surface of the hollow shaft journal is Polished after machining. The end cover and the cylinder body are all connected by high-strength bolts. Strict process measures to control the machining accuracy of the joint surface stop, to ensure reliable connection and the concentricity of the two end journal after final assembly. According to the actual situation of the project, the cylinder can be made as a whole or divided, with a flanged connection and stop positioning. All welds are penetration welds, and all welds are inspected by ultrasonic nondestructive testing After welding, the whole Shell is returned to the furnace for tempering stress relief treatment, and after heat treatment, the shell surface is shot-peened. The lining plate of the ball mill is usually made of alloy material.

The transmission part comprises a gear and a gear, a gear housing, a gear housing and an accessory thereof. The big gear of the transmission part of the self-grinding machine fits on the hollow shaft of the discharge material, which is smaller in size, but the seal of the gear cover is not good, and the ore slurry easily enters the hollow shaft of the discharge material, causing the hollow shaft to wear.

The big gear of the ball mill fits on the mill shell, the size is bigger, the big gear is divided into half structure, the radial and axial run-out of the big gear are controlled within the national standard, the aging treatment is up to the standard, and the stress and deformation after processing are prevented. The big gear seal adopts the radial seal and the reinforced big gear shield. It is welded and manufactured in the workshop. The geometric size is controlled, the deformation is prevented and the sealing effect is ensured. The small gear transmission device adopts the cast iron base, the bearing base and the bearing cap are processed at the same time to reduce the vibration in operation. Large and small gear lubrication: The use of spray lubrication device timing quantitative forced spray lubrication, automatic control, no manual operation. The gear cover is welded by profile steel and high-quality steel plate. In order to enhance the stiffness of the gear cover, the finite element analysis is carried out, and the supporting structure is added in the weak part according to the analysis results.

The self-mill adopts the self-return device to realize the discharge of the mill. The self-returning device is located in the revolving part of the mill, and the material forms a self-circulation in the revolving part of the mill through the self-returning device, discharging the qualified material from the mill, leading the unqualified material back into the revolving part to participate in the grinding operation.

The ball mill adopts a discharge screen similar to the ball mill, and the function of blocking the internal medium of the overflow ball mill is accomplished inside the rotary part of the ball mill. The discharge screen is only responsible for forcing out a small amount of the medium that overflows into the discharge screen through the internal welding reverse spiral, to achieve forced discharge mill.

The slow drive consists of a brake motor, a coupling, a planetary reducer and a claw-type clutch. The device is connected to a pinion shaft and is used for mill maintenance and replacement of liners. In addition, after the mill is shut down for a long time, the slow-speed transmission device before starting the main motor can eliminate the eccentric load of the steel ball, loosen the consolidation of the steel ball and materials, ensure safe start, avoid overloading of the air clutch, and play a protective role. The slow-speed transmission device can realize the point-to-point reverse in the electronic control design. When connecting the main motor drive, the claw-type Clutch automatically disengages, the maintenance personnel should pay attention to the safety.

The slow drive device of the ball mill is provided with a rack and pinion structure, and the operating handle is moved to the side away from the cylinder body The utility model not only reduces the labor intensity but also ensures the safety of the operators.

sag mill pulp dischargers - grinding & classification circuits - metallurgist & mineral processing engineer

sag mill pulp dischargers - grinding & classification circuits - metallurgist & mineral processing engineer

Turbo Pulp Lifters have been shown to be effective at discharging slurry where other designs have failed. I believe the Cortez gold mine had success with this design after struggling many years with slurry pooling in their SAG mill.

If you are able to share some of your operating parameters, we may be able to offer you advice more specific to your issue: grate/pulp lifter designs (drawing or really good description); mill diameter and operating speed (rpm or %C.S.); typical pulp density (total feed/(total feed + water)); a description of the mill behavior that has you interested in the TPL system.

Craig's information above is great and I would just like to add a comment based on our experience as a manufacturer of radial pulp lifters (Turbo Pulp I believe is a trademarked name for radial lifters).

Perhaps have as Craig has suggested provide detailed operating parameters and have DEM (Discrete Element Modelling) and FEA (Finite Element Analysis) done which will show the evacuation as well as expected wear points on the pulp lifters.

Pulp lifters take the form of a radial array of box-section channels mounted inside the discharge end wall of the mill. The number of pulp lifters used in conventional mill design is equal to the number of feet in the measure of the mill diameter.

Ground material flows into the pulp lifters via grates mounted on the outer part of the pulp lifter channels. As these channels rotate with the mill, they collect, lift and direct the captured charge towards the mill-centre and out through the external discharge trunnion, Figure 1.

Denlay et al (1997) reviewed the motion of the outer solid charge along a pulp lifter during a mill rotation cycle. The motive force is the component of the gravitational force acting radially inwards, it is opposed by the centrifugal force acting radially outwards. Friction opposes motion and is a function of the normal components of centrifugal force and gravity acting on the lifter and both static and dynamic friction factors. Discharge motion is initiated at points where the inward gravitational force exceeds opposing forces. The locus of all such points is a circle. For a mill operating at around 76% of critical speed, this circle has a diameter similar to the diameter of the mill. (Mill speed is expressed as percent of the critical speed at which outer layers of rocks in the mill would begin to centrifuge). From initiation of motion to TDC, i.e. the first stage of motion, the inward-discharging force is that of gravity opposed by both friction and centrifugal force. After TDC, the second stage of motion is entirely ballistic i.e. controlled by gravity alone. The third stage of motion follows the re-contact of the charge with the opposite side of the pulp lifter, motion from this point to discharge is promoted by gravity and opposed only by friction. The stages in the motion of the outer solid charge are shown in Figure 2.

This interpretation of solid motion indicates that, for straight radial pulp lifters, a significant part of the solid motion occurs after mill top dead centre (TDC). This contrasts with the more fluid components where (due to initial segregation and displacement of the liquid charge) the motion of liquids to the discharge trunnion can be complete around TDC.

Curved Pulp lifters assist the movement of the charge to the centre by initiating motion of the charge earlier than is the case with straight lifters. This moves the charge towards the centre of the mill where, relative to straight pulp lifters, it can be subject to earlier and stronger application of the inward motion gravity forces. In this action the curved lifter enables solids motion to imitate the levelling action of a fluid charge. The overall motion for curved pulp-lifters is illustrated in Figure 3.

This interpretation of solid charge flow is supported by field observation, including video of the discharge flow into trunnions of operating mills, and by the wear patterns due to slurry and solid flow in pulp lifters. The wear data from both curved and straight radial pulp lifters was reviewed by Royston et al (1998).

DISCLAIMER: Material presented on the 911METALLURGIST.COM FORUMS is intended for information purposes only and does not constitute advice. The 911METALLURGIST.COM and 911METALLURGY CORP tries to provide content that is true and accurate as of the date of writing; however, we give no assurance or warranty regarding the accuracy, timeliness, or applicability of any of the contents. Visitors to the 911METALLURGIST.COM website should not act upon the websites content or information without first seeking appropriate professional advice. 911METALLURGY CORP accepts no responsibility for and excludes all liability in connection with browsing this website, use of information or downloading any materials from it, including but not limited to any liability for errors, inaccuracies, omissions, or misleading statements. The information at this website might include opinions or views which, unless expressly stated otherwise, are not necessarily those of the 911METALLURGIST.COM or 911METALLURGY CORP or any associated company or any person in relation to whom they would have any liability or responsibility.

sag mill liner packing - grinding & classification circuits - metallurgist & mineral processing engineer

sag mill liner packing - grinding & classification circuits - metallurgist & mineral processing engineer

My understanding is that packing is due to the steel balls getting jamming in between the lifter bars, and this could probably happen at the shell and ends. The remedy is to design the lifters so the space in between bars isn't an even multiple of the balls diameter. I recall there being a good paper from the SAG 2006 conference.

You know packing usually happens in WET SAG mills and as you mentioned one of the solutions is increasing space between lifters but I was wondering that does it happen in DRY SAG mills too? We are trying to change head radial lifters in a dry SAG mill.

One is, as suggested above, the jamming of steel balls in between liners, where geometry of the liners facilitates.The other is packing of wet ore in between the liners, which cause the "lifting" function of the liners to be severely compromised. I am not too sure what is causing this to happen, I am more inclined to believe is the nature of the ore and % of water addition. The shape and size of the pocket between the shell lifters will have an influence, not sure how much though.

In my view, it should not matter whether the mill is dry or wet. Packing will result if the gaps between the liners permit it, and if the particle size produced in the mill will fit in these gaps. We see packing in our dry vertical roller mills and also in cement mills. Granted the packing is not as severe as in wet mills, as the presence of water can cause some ores to behave like a self-setting concrete mix. Liner design in terms of lifter spacing should consider the best spacing for charge lift. Quality of liner fit up and casting accuracy will determine the spacing between liners. Proper design and use of rubber wedges will prevent packing to some degree, but packing is not necessarily a bad thing, think of it like an autogenous wear layer.

DISCLAIMER: Material presented on the 911METALLURGIST.COM FORUMS is intended for information purposes only and does not constitute advice. The 911METALLURGIST.COM and 911METALLURGY CORP tries to provide content that is true and accurate as of the date of writing; however, we give no assurance or warranty regarding the accuracy, timeliness, or applicability of any of the contents. Visitors to the 911METALLURGIST.COM website should not act upon the websites content or information without first seeking appropriate professional advice. 911METALLURGY CORP accepts no responsibility for and excludes all liability in connection with browsing this website, use of information or downloading any materials from it, including but not limited to any liability for errors, inaccuracies, omissions, or misleading statements. The information at this website might include opinions or views which, unless expressly stated otherwise, are not necessarily those of the 911METALLURGIST.COM or 911METALLURGY CORP or any associated company or any person in relation to whom they would have any liability or responsibility.

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