To start the discussion on grinding circuits I would like to begin by showing you some simple Flowsheet SCHEMATICS of some sample circuits, but before I do there is one thing that I must explain for the schematics to make sense. When the ore is put into the first piece of equipment for grinding, water is added to it to form a SLURRY. This water addition is very important, it makes the rest of the separation process possible as you will see a little later.
To design a circuit there is some information that has to be learned. The HARDNESS OF THE ORE, the TONNAGE that has to be ground per hour and the DEGREE TO WHICH IT MUST BE GROUND. Once these three pieces of information has been discovered the design of the circuit can begin. The four circuits that I am going to show you start with the simplest, the one that is grinding the softest ore to the coarsest grind.
The ore is simply brought from the fine ore storage area and run into one end of the grinding machine called a MILL, the mill grinds it and mixes it into a slurry which is sent to the next stage of processing.
Our second flow sheet, has a couple extra pieces of equipment added. These are a PUMP and a CLASSIFIER, A classifier is a piece of equipment that separates ground material by size, There are different methods that can be employed to accomplish this, but again that is something that we .will take up in greater detail in further sessions.
This is the path that the ore takes in this circuit. The ore comes from the fine ore storage area to be added to the mill. Again it is ground but this time the ground material is pumped to a classifier to be classified by size. The material that has not been ground enough will be sent back to the mill for regrinding. The material that is ready for the next phase of processing will be sent there.
That brings us to our third circuit. Like the second one, the ore is ground in a mill and sent by way of a pump box to a classifier. Now however there has been a second mill added. The rock that requires further grinding is sent to this mill where it is reground. Once the ore has gone through the mill, it is sent by way of the pump back for reclassification.
These circuits have been getting more and more complex. This is because the ore and the required grind has been getting more and more complicated. The ore in the first one was the softest and the grinding wasnt critical. In the second one the ore was probably still very soft, but the grind had become critical, therefore it had to have a positive control on it, hence the classification. The third one however has harder ore with a critical grind.
Now for our fourth and final flow sheet. This is one that has been-designed for exceptionally hard ore and/or a grind that must be very fine. This time the ore is conveyed to the first mill, mixed with water and ground. As it exits the mill it is separated and sent to two different pumps which in turn sends it to be classified. Each of the classifiers will send the material that requires grinding back to its own mill. On this circuit there was one PRIMARY MILL feeding two SECONDARY MILLS each with their own pump and classifier. Again the secondary mill discharge is sent back to its own classifier to be reclassified.
Since several types of flotation circuits can generally be employed in conjunction with the various processes for the flotation of different classes of minerals, an outline of the standard circuits in common use is best given before the processes to which they are applicable are described. The flow sheets illustrating them are diagrammatic, but, in cases where the circuit includes two or more machines, the latter are shown in their approximate relative positions. The diagrams must not, however, be considered to represent exact plant layouts; any standardised arrangement willusually need a certain amount of alteration in minor details to suit a particular ore, not only to meet the requirements of the flotation process but also to conform to the design of the plant as a whole.
is taken off the first few cells and the remainder are run as scavengers, the froth from them being returned to the head of the machine through a middling pipe, as described in the previous chapter in connection with M.S. Sub-aeration and Denver Sub-A Machines. This method can be operated successfully when the gangue is relatively unfloatable, but it requires rather more careful control than when a separate cleaning circuit is used.
Circuit No. 2 (Fig. 49) is, as a rule, preferable to the preceding one, and it is advisable to employ it when a very clean concentrate is desired. The froth from the primary or roughing cells, termed the rougher concentrate, is diluted to a W/S ratio between 6/1 and 10/1 and transferred through a middling pipe to the cleaning section. This consists of a series of similar cells, varying in number from one-half of those in the roughing section for the treatment of a bulky low-grade rougher concentrate to one-quarter of the number when it is required to raise a small quantity of comparatively rich primary concentrate to the highest possible grade. It is usual, though not necessary, for the cleaning cells to be placed at the head of the machine. The tailing from the last cell is then mixed, as it overflows the discharge weir, with the incoming primary feed, so that both enter the roughing section through the transfer passage or pipe.
Circuit No. 2 is easier to regulate than No. 1 because the final concentrate is made in a pulp containing a much larger proportion of mineral than of gangue ; there is therefore less likelihood of gangue entering the froth and lowering the grade of the final concentrate. Moreover, variations in the character and rate of the feed, such as are encounteredperiodically, especially in small plants, may cause considerable alteration in the grade of the rougher concentrate before the change is noticed and the machine readjusted. Variations of this sort are evened out to a large extent in cleaning cells, so that the grade of the finished concentrate is affected very much less than would be the case if circuit No. 1 were in use.
Circuit No. 3 (Fig. 50) is developed from the preceding one. Here the primary cells are divided into roughing and scavenging sections. The concentrate from the first section is re-treated in cleaning cells as before, but that from the latter part of the machine is returned to the first roughing cell through a middling pipe. Thus the cleaning cells receive a comparatively high-grade feed, whilst the scavenging section can be run with an excess of air so as to obtain the maximum recovery of the valuable minerals ; the resulting low-grade froth, being usually little richer in mineral than the normal run of ore, is best returned to the circuit with the primary feed. This circuit, besides being useful for ores that need the maximum amount of aeration at the end of the machine to give a profitable recovery of the valuable minerals, is often employedwhen the gangue is floatable and difficult to separate from the mineral. In the latter case, if one cleaning operation does not yield a concentrate of high enough grade, a re-cleaning section becomes necessary, such as is shown in circuit No. 4 (Fig. 51), in which the cleaner concentrate is diluted and re-treated in a separate series of cells. In this circuit the use of middling return pipes eliminates the necessity for pumping three separate concentrate products.
It should be noted that the diluting water of the cleaning cells in circuits 2 and 3 passes to the roughing cells and dilutes the primary feed, which ought therefore to contain a correspondingly smaller proportionof water as it leaves the grinding section in order that the dilution of the cleaner tailing may bring it to the correct pulp ratio in the roughing machine. Little further dilution of the primary pulp is involved when recleaning is practised, since the water added passes with the recleaner tailing to the cleaning cells where it constitutes the bulk of the diluting water, very little extra being normally required. Thus practically the same amount of water enters the roughing section in the cleaner tailing whether recleaning is practised or not. This also applies to the cleaning operations in the air-lift and pneumatic machines.
An air-lift machine can often be run as a rougher-cleaner cell, as in circuit No. 5 (Fig. 52), by taking a finished concentrate off the first few feet and returning the froth from the rest of the machine to the feed end. This method is also applicable to pneumatic cells, but its use is not common. As with mechanically agitated machines, careful control is needed if the required grade of concentrate is to be consistently maintained, and it is more usual, and generally more satisfactory, to employ a circuit such as No. 6 or 7, in which a separate stage of cleaning is included. Here the primary(rougher) concentrate is diluted as before to a W/S ratio between 6/1 and 10/1 before being sent to the cleaning section, and the cleaner tailing is returned to the head of the primary machine.
Circuit No. 6 (Fig. 53) is suitable when the feed flows by gravity to the rougher ; it is generallyadvisable to adopt it also in cases where the cleaning operation requires the addition of extra reagents, since the pump provides a convenient method of mixing them thoroughly with the pulp, but when it is necessary for the feed to be pumped to the rougher, circuit No. 7 (Fig. 54) has
the advantage that the cleaner tailing can be elevated in the same pump, the primary concentrate flowing by gravity to the cleaning section. When vigorous aeration is needed to obtain a maximum recovery of
the valuable minerals, it is preferable to use separate roughing and scavenging machines, the scavenger froth being returned with the cleaner tailing to the head of the rougher. Of the two usual arrangements, circuit No. 8 (Fig. 55) is the simpler, requiring only one pump to handle both products, but, should the layout of the plant make it advisable to place the cleaning section above the rougher, circuit No. 9 (Fig. 56) withtwo pumps will be necessary. The second method has the advantage that the pumps provide points of agitation ahead of both the roughing and the cleaning sections for the addition of reagents, should they be required. It is usually possible to place the two pumps close together with a third between them so connected that it can be used as a spare for either of the other two. By separating the roughing and scavenging machines in this way a rougher concentrate without an undue proportion of gangue can be sent to the cleaning section.
It will then be correspondingly easy to produce a final concentrate of the highest possible grade, while still maintaining maximum aeration in the scavenging section, the froth from which is returned to the point where the comparatively large amount of gangue in it does the least harm.
Even with separate roughing and scavenging sections it is sometimes found impossible to make a concentrate of high enough grade in the cleaner without allowing too much mineral to return in the tailing to the rougher, with consequent overloading of the latter section. In such a case recleaning is necessary, the cleaner concentrate being diluted with water and re-treated in a separate machine. Circuits 10 and 11 (Fig. 57 and 58) are the two most usual arrangements, but various modifications of the layout are possible to suit special cases. In fact, although the flotation method to be adopted is the primary consideration, the actual positions of the flotation machines in relation to each other and to the pumps handling the products must necessarily depend to a great extent on the requirements of the rest of the plant.
As in the case of mechanically agitated machines, a recleaning circuit involves very little more dilution of the pulp in the rougher than does one cleaning operation, since the diluting water in the second stage passes to the first stage with the recleaner tailing and serves the same purpose there.
It occasionally happens that, although two cleaning operations are necessary to obtain a final concentrate of the required grade, the tailing from neither stage contains much mineral. In snch cases the circuit can be simplified by dispensing with one pump and sending the tailings from both stages to the head of the rougher, as is shown by the dotted line in circuit No. 11 (Fig. 58). The method has the disadvantage, however, that the diluting water from both cleaning sections enters the roughing machine and may make the pulp too thin. It is therefore preferable to pump one or both tailings to the classifiers in the grinding section, especially if they contain imperfectly ground material. Not only is the mineral then given another chance of entering the ball mill and becoming liberated from the gangue, but the water of the two tailings serves todilute the pulp in the classifiers and does not then interfere with the density of the pulp in the roughing machine. The tailing from a single- stage cleaning operation is often returned in the same way to the grinding circuit, if it contains mineral which has not been completely liberated from the gangue.
When a large tonnage is to be handled, the flotation section of the plant is divided up into a number of parallel units, each consisting of one or more lines of roughing and scavenging machines with the appropriate cleaners. Occasionally, if the reagent mixture and control is simple, the grinding section is also designed in separate units, from each of which the pulp is discharged directly to the corresponding flotation unit. It is more usual, however, to send the discharge of the grinding section to a central distributor which splits the pulp up evenly amongst the flotation units ; the number of the latter need not then correspond to the number of grinding units.
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.
S9 Hashboard has 21 voltage domains in total, and each voltage domain contains 3 chips. There is a voltage difference between the voltage domains, if the difference is abnormal, the hash board will not work
Grinding circuits are fed at a controlled rate from the stockpile or bins holding the crusher plant product. There may be a number of grinding circuits in parallel, each circuit taking a definite fraction of the feed. An example is the Highland Valley Cu/Mo plant with five parallel grinding lines (Chapter 12). Parallel mill circuits increase circuit flexibility, since individual units can be shut down or the feed rate can be changed, with a manageable effect on production. Fewer mills are, however, easier to control and capital and installation costs are lower, so the number of mills must be decided at the design stage.
The high unit capacity SAG mill/ball mill circuit is dominant today and has contributed toward substantial savings in capital and operating costs, which has in turn made many low-grade, high-tonnage operations such as copper and gold ores feasible. Future circuits may see increasing use of high pressure grinding rolls (Rosas et al., 2012).
Autogenous grinding or semi-autogenous grinding mills can be operated in open or closed circuit. However, even in open circuit, a coarse classifier such as a trommel attached to the mill, or a vibrating screen can be used. The oversize material is recycled either externally or internally. In internal recycling, the coarse material is conveyed by a reverse spiral or water jet back down the center of the trommel into the mill. External recycling can be continuous, achieved by conveyor belt, or is batch where the material is stockpiled and periodically fed back into the mill by front-end loader.
In Figure 7.35 shows the SAG mill closed with a crusher (recycle or pebble crusher). In SAG mill operation, the grinding rate passes through a minimum at a critical size (Chapter 5), which represents material too large to be broken by the steel grinding media, but has a low self-breakage rate. If the critical size material, typically 2550mm, is accumulated the mill energy efficiency will deteriorate, and the mill feed rate decreases. As a solution, additional large holes, or pebble ports (e.g., 40100mm), are cut into the mill grate, allowing coarse material to exit the mill. The crusher in closed circuit is then used to reduce the size of the critical size material and return it to the mill. As the pebble ports also allow steel balls to exit, a steel removal system (such as a guard magnet, Chapters 2 and 13Chapter 2Chapter 13) must be installed to prevent them from entering the crusher. (Because of this requirement, closing a SAG mill with a crusher is not used in magnetic iron ore grinding circuits.) This circuit configuration is common as it usually produces a significant increase in throughput and energy efficiency due to the removal of the critical size material.
An example SABC-A circuit is the Cadia Hill Gold Mine, New South Wales, Australia (Dunne et al., 2001). The project economics study indicated a single grinding line. The circuit comprises a SAG mill, 12m diameter by 6.1m length (belly inside liners, the effective grinding volume), two pebble crushers, and two ball mills in parallel closed with cyclones. The SAG mill is fitted with a 20MW gearless drive motor with bi-directional rotational capacity. (Reversing direction evens out wear on liners with symmetrical profile and prolongs operating time.) The SAG mill was designed to treat 2,065t h1 of ore at a ball charge of 8% volume, total filling of 25% volume, and an operating mill speed of 74% of critical. The mill is fitted with 80mm grates with total grate open area of 7.66m2 (Hart et al., 2001). A 4.5m diameter by 5.2m long trommel screens the discharge product at a cut size of ca. 12mm. Material less than 12mm falls into a cyclone feed sump, where it is combined with discharge from the ball mills. Oversize pebbles from the trommel are conveyed to a surge bin of 735t capacity, adjacent to the pebble crushers. Two cone crushers with a closed side set of 1216mm are used to crush the pebbles with a designed product P80 of 12mm and an expected total recycle pebble rate of 725t h1. The crushed pebbles fall directly onto the SAG mill feed belt and return to the SAG mill.
SAG mill product feeds two parallel ball mills of 6.6m11.1m (internal diameterlength), each with a 9.7MW twin pinion drive. The ball mills are operated at a ball charge volume of 3032% and 78.5% critical speed. The SAG mill trommel undersize is combined with the ball mills discharge and pumped to two parallel packs (clusters) of twelve 660mm diameter cyclones. The cyclone underflow from each line reports to a ball mill, while the cyclone overflow is directed to the flotation circuit. The designed ball milling circuit product is 80% passing 150m.
Several large tonnage copper porphyry plants in Chile use an open-circuit SAG configuration where the pebble crusher product is directed to the ball mills (SABC-B circuit). The original grinding circuit at Los Bronces is an example: the pebbles generated in the two SAG mills are crushed in a satellite pebble crushing plant, and then are conveyed to the three ball mills (Mogla and Grunwald, 2008).
Hydrocyclones have come to dominate classification when dealing with fine particle sizes in closed grinding circuits (<200m). However, recent developments in screen technology (Chapter 8) have renewed interest in using screens in grinding circuits. Screens separate on the basis of size and are not directly influenced by the density spread in the feed minerals. This can be an advantage. Screens also do not have a bypass fraction, and as Example 9.2 has shown, bypass can be quite large (over 30% in that case). Figure 9.8 shows an example of the difference in partition curve for cyclones and screens. The data is from the El Brocal concentrator in Peru with evaluations before and after the hydrocyclones were replaced with a Derrick Stack Sizer (see Chapter 8) in the grinding circuit (Dndar et al., 2014). Consistent with expectation, compared to the cyclone the screen had a sharper separation (slope of curve is higher) and little bypass. An increase in grinding circuit capacity was reported due to higher breakage rates after implementing the screen. This was attributed to the elimination of the bypass, reducing the amount of fine material sent back to the grinding mills which tends to cushion particleparticle impacts.
Circulation of material occurs in several parts of a mineral processing flowsheet, in grinding and flotation circuits, for example, as well as the crushing stage. In the present context, the circulating load (C) is the mass of coarse material returned from the screen to the crusher relative to the circuit final product (or fresh feed to the circuit), often quoted as a percentage. Figure 8.2 shows two closed circuit arrangements. Circuit (a) was considered in Chapter 6 (Example 6.1), and circuit (b) is an alternative. The symbols have the same meaning as before. The relationship of circulating load to screen efficiency for circuit (a) was derived in Example 6.1, namely (where all factors are as fractions):
The circulating load as a function of screen efficiency for the two circuits is shown in Figure 8.3. The circulating load increases with decreasing screen efficiency and as crusher product coarsens (f or r decreases), which is related to the crusher set (specifically the closed side setting, c.s.s.). For circuit (a) C also increases as the fresh feed coarsens (n decreases), which is likely coming from another crusher. In this manner, the circulating load can be related to crusher settings.
In industrial grinding process, in addition to goal of productivity maximization, other purposes of deterministic grinding circuit optimization have to satisfy the upper bound constraints on the control variables. We know that there lies a tradeoff between the throughput (TP) and the percent passing of midsize classes (MS) from the previous work of Mitra and Gopinath,2004. In deterministic optimization formulation, there are certain parameters which we will assume them as constant. But, in real life that may not be case. There are such six parameters in our industrial grinding process which are R, B, R, B are the grindability indices and grindability exponents for the rod mill (RMGI) and the ball mill (BMGI); and P, S are the sharpness indices for the primary (PCSI) and secondary cyclones (SCSI). These parameters are treated as constant in deterministic formulation. As they are going to be treated as uncertain parameters in the OUU formulation. These parameters are assumed uncertain because most of them are obtained from the regression of experimental data and thus are subject to uncertainty due to experimental and regression errors. In the next part of the section, we consider them as fuzzy numbers and solve the OUU problem by FEVM. In FEVM formulation, the uncertain parameters are considered as fuzzy numbers and the uncertain formulation is transformed into the deterministic formulation by expectation calculations for both objective function and constraints. So, the converted deterministic multi-objective optimization problem is expressed as:
Another spinning batch concentrator (Figure 10.27), it is designed principally for the recovery of free gold in grinding circuit classifier underflows where, again, a very small (<1%) mass pull to concentrate is required. The feed first flows up the sides of a cone-shaped bowl, where it stratifies according to particle density before passing over a concentrate bed fluidized from behind by back-pressure (process) water. The bed retains dense particles such as gold, and lighter gangue particles are washed over the top. Periodically the feed is stopped, the bed rinsed to remove any remaining lights and is then flushed out as the heavy product. Rinsing/flushing frequency, which is under automatic control, is determined from grade and recovery requirements.
The units come in several designs, the Semi-Batch (SB), Ultrafine (UF), and i-Con, designed for small scale and artisanal miners. The first installation was at the Blackdome Gold Mine, British Columbia, Canada, in 1986 (Nesset, 2011).
These two batch centrifugal concentrators have been widely applied in the recovery of gold, platinum, silver, mercury, and native copper; continuous versions are also operational, the Knelson Continuous Variable Discharge (CVD) and the Falcon Continuous (C) (Klein et al., 2010; Nesset, 2011).
To liberate minerals from sparsely distributed and depleting the ore bodies finer grinding than generally obtained by the conventional Rod Mill Ball Mill grinding circuits is needed. Longer grinding periods in the conventional milling processes prove too expensive mainly due to large power consumption. Stirrer mills have been tried in mineral industry with considerable success and have therefore been increasingly used. In this chapter, the theories involved in the design and operation of these mills, as established till now, are explained. Further theoretical studies and designs of the mills are still in progress for a better understanding and improved operation. Presently, the mills have been proved to be economically viable and the mineral of interest conducive to improved recovery and grade.
IMP Technologies Pty. Ltd. has recently tested a pilot-scale super fine crusher that operates on dry ore and is envisaged as a possible alternative to fine or ultra-fine grinding circuits (Kelsey and Kelly, 2014). The unit includes a rotating compression chamber and an internal gyrating mandrel (Figure 6.13). Material is fed into the compression chamber and builds until the gyratory motion of the mandrel is engaged. Axial displacement of the compression chamber and the gyratory motion of the mandrel result in fine grinding of the feed material. In one example, a feed F80 of 300m was reduced to P80 of 8m, estimated to be the equivalent to two stages of grinding. This development is the latest in a resurgence in crushing technology resulting from the competition of AG/SAG milling and the demands for increased comminution energy efficiency.
The iron oxide crystal grains in most iron ores are not evenly distributed in size. Spiral separators can therefore be used to take out the coarser iron oxide grains in the primary grinding circuit to save grinding energy and help achieve a higher iron recovery. Figure 9.14 presents a typical flow sheet for processing an oxidized ore containing about 30% Fe using a combination of spiral and SLon magnetite separators and reverse flotation. This ore is mainly composed of hematite, magnetite, and quartz, and the iron oxide crystals range in size from 0.005 to 1.0mm with an average size of about 0.05mm. The average size of the quartz crystals is approximately 0.085mm.
In the primary grinding stage of the flow sheet in Figure 9.14, the ore is first ground down to about 60% -75m and then classified into two size fractions, a coarse size fraction and a fine size fraction. The coarse size fraction is treated with spiral separators to recover part of the final iron ore concentrate. Then, drum LIMS and SLon magnetic separators are used to reject some of the coarse gangue minerals as final tailings. The magnetic products from the LIMS and SLon are sent back to the secondary ball mill for regrinding, and the milled product returns to the primary cyclone classifier.
The fine size fraction is about 90% -75m and is processed using drum LIMS separators and SLon magnetic separators in series to take out the magnetite and hematite, respectively. The magnetic products from the magnetic separators are mixed to generate the feed for reverse flotation to produce another component of the final iron ore concentrate.
The key advantage of this flow sheet lies in the fact that the spirals and SLon magnetic separators take out about 20% of the mass of the final iron concentrate and about 20% of the mass of the final tailings, respectively, from the coarse size fraction. This greatly reduces the masses being fed to the secondary ball mill and reverse flotation, thereby greatly reducing the total processing cost. From the plant results for this flow sheet, an iron concentrate containing 67.5% Fe could be produced from a run-of-mine ore containing 30.1% Fe, at a mass yield to the iron concentrate of 34.9%, an iron recovery of 78.0%, and a tailings grade of 10.2% Fe.
The first step of physical beneficiation is crushing and grinding the iron ore to its liberation size, the maximum size where individual particles of gangue are separated from the iron minerals. A flow sheet of a typical iron ore crushing and grinding circuit is shown in Figure 1.2.2 (based on Ref. ). This type of flow sheet is usually followed when the crude ore contains below 30% iron. The number of steps involved in crushing and grinding depends on various factors such as the hardness of the ore and the level of impurities present .
Jaw and gyratory crushers are used for initial size reduction to convert big rocks into small stones. This is generally followed by a cone crusher. A combination of rod mill and ball mills are then used if the ore must be ground below 325 mesh (45m). Instead of grinding the ore dry, slurry is used as feed for rod or ball mills, to avoid dusting. Oversize and undersize materials are separated using a screen; oversize material goes back for further grinding.
Typically, silica is the main gangue mineral that needs to be separated. Iron ore with high-silica content (more than about 2%) is not considered an acceptable feed for most DR processes. This is due to limitations not in the DR process itself, but the usual customer, an EAF steelmaking shop. EAFs are not designed to handle the large amounts of slag that result from using low-grade iron ores, which makes the BF a better choice in this situation. Besides silica, phosphorus, sulfur, and manganese are other impurities that are not desirable in the product and are removed from the crude ore, if economically and technically feasible.
While used sometimes on final concentrates, such as Fe concentrates, to determine the Blaine number (average particle size deduced from surface area), and on tailings for control of paste thickeners, for example, the prime application is on cyclone overflow for grinding circuit control (Kongas and Saloheimo, 2009). Control of the grinding circuit to produce the target particle size distribution for flotation (or other mineral separation process) at target throughput maximizes efficient use of the installed power.
Continuous measurement of particle size in slurries has been available since 1971, the PSM (particle size monitor) system produced then by Armco Autometrics (subsequently by Svedala and now by Thermo Gamma-Metrics) having been installed in a number of mineral processing plants (Hathaway and Guthnals, 1976).
The PSM system uses ultrasound to determine particle size. This system consists of three sections: the air eliminator, the sensor section, and the electronics section. The air eliminator draws a sample from the process stream and removes entrained air bubbles (which otherwise act as particles in the measurement). The de-aerated pulp then passes between the sensors. Measurement depends on the varying absorption of ultrasonic waves in suspensions of different particle sizes. Since solids concentration also affects the absorption, two pairs of transmitters and receivers, operating at different frequencies, are employed to measure particle size and solids concentration of the pulp, the processing of this information being performed by the electronics. The Thermo GammaMetrics PSM-400MPX (Figure 4.18) handles slurries up to 60% w/w solids and outputs five size fractions simultaneously.
Other measurement principles are now in commercial form for slurries. Direct mechanical measurement of particle size between a moving and fixed ceramic tip, and laser diffraction systems are described by Kongas and Saloheimo (2009). Two recent additions are the CYCLONEtrac systems from CiDRA Minerals Processing (Maron et al., 2014), and the OPUS ultrasonic extinction system from Sympatec (Smith et al., 2010).
CiDRAs CYCLONEtrac PST (particle size tracking) system comprises a hardened probe that penetrates into the cyclone overflow pipe to contact the stream and effectively listens to the impacts of individual particles. The output is % above (or below) a given size and has been shown to compare well with sieve sizing (Maron et al., 2014). The OPUS ultrasonic extinction system (USE) transmits ultrasonic waves through a slurry that interact with the suspended particles. The detected signal is converted into a particle size distribution, the number of frequencies used giving the number of size classes measured. Applications on ores can cover a size range from 1 to 1,000m (Smith et al., 2010).
In addition to particles size, recent developments have included sensors to detect malfunctioning cyclones. Westendorf et al. (2015) describe the use of sensors (from Portage Technologies) on cyclone overflow and underflow piping. CiDRAs CYCLONEtrac OSM (oversize monitor) is attached to the outside of the cyclone overflow pipe and detects the acoustic signal as oversize particles (rocks) hit the pipe (Cirulis and Russell, 2011). The systems are readily installed on individual cyclones thus permitting poorly operating units to be identified and changed while allowing the cyclone battery to remain in operation. Figure 4.19 shows an installation of both CiDRA systems (PST, OSM) on the overflow pipe from a cyclone.
This study is conducted with the aim of investigating the efficiency of open and closed-circuit molybdenite ore comminution processes (primary and secondary mill, respectively), through mineralogical study of mills feed and product. For this purpose, particle size distribution, minerals distribution, degree of liberation and interlocking of minerals in mills feed and product were studied. According to the results, chalcopyrite, molybdenite, pyrite and covellite constitute the major part of the mineral composition of open-circuit mill feed. Minerals at the mill product, in the order of abundance include liberated molybdenite particles, liberated chalcopyrite and interlocked chalcopyrite with pyrite, liberated and interlocked pyrite particles, and associated silicate gangues. The d50 values of the feed and product particles of the open-circuit mill are equal to 13.80 and 13.40 microns, respectively. Degree of liberation of molybdenite for the feed and product of this mill is almost the same and is equal to 98.0%. Closed-circuit mill feed includes, in order of is abundance, liberated molybdenite particles in the form of blades and irregular polygonal shapes, liberated and interlocked chalcopyrite, and liberated and interlocked pyrite particles with gangue minerals. Molybdenite particles in the mill product are almost completely liberated, and the degree of liberation values of chalcopyrite and pyrite are 84.40% and 91.40%, respectively. According to particles size distribution of the feed (d50 equal to 25.03 microns) and the product (d50 equal to 24.24 microns) of closed-circuit mill, it can be stated that comminution is not well-operated in closed-circuit mill due to the low solid percentage of closed-circuit mill feed and the inefficiency of hydrocyclone. Examination of Mo, Cu, and Fe grade variations for 10days in both off and on modes of mill shows that closed-circuit mill does not have an impact on comminution process. It can even be concluded that the mill has a destructive effect the flotation process by producing slimes.
Liberation of valuable minerals from associated gangue minerals is an important and fundamental step in separating an ore mineral from gangue during physical or physicochemical separation processes. Crushing and grinding processes are typically used by crushers and mills to liberate minerals, which are energy-intensive processes (especially fine grinding by mill). Meanwhile, ball mills are known for their lowest energy efficiency. The efficiency of ball mills is about 1.0% and, in some cases, less than 1.0% based on energy consumption.
Mineral comminution theories are often based on the relationship between the size of the primary feed particles entering the mill and the energy consumed (Eq.1); in most of these relationships, it has been assumed that the ground material is brittle1. In other words, the grindability indices of the minerals in a deposit are typically used for designing a comminution circuit for an ore. These indices are proposed with the assumption of homogenic breakage and continuity of the ore, regardless of the textural properties of the minerals on a micro scale2.
The main missing factor in most comminution theories is the relationship between ore grinding and its textural and mineralogical characteristics. Each ore in the mine has different geomechanical characteristics for various reasons, such as the effect of faults, the presence of dikes, as well as the type of the deposit in different zones of the mine (such as oxidized, supergene and hypogene zones). Variations in geomechanical characteristics will cause different comminution behaviors for ore during blasting and comminution operations by crushers and mills. In other words, due to the diversity of minerals and the difference in their grindability indices, the feed of processing plants includes a distribution of mineral types and particle sizes. Therefore, the study of mineralogy and the textural characteristics of an ore and its host rock will provide valuable information from the perspective of ore comminution behavior and minerals content and thus the design or optimization of their comminution processes.
Ore textural parameters including hardness, minerals liberation degree, particle size, particle size distribution, minerals abundance, type of minerals, interlocking between minerals and mineralogical structures are important in the issue of ore grindability1,3,4, and can be considered as optimizing parameters for comminution processes. Many researchers1,5,6,7,8 have extensively investigated the impact of textural parameters of ores in mineral processing. It should be noted that the combination of these parameters with the operating conditions of concentration processes, especially comminution, is very complex due to the randomized mechanism of breakage in mills. In general, the process mineralogy of the feed and product of comminution process (mill) will lead to presenting solutions to optimize the operation of the existing circuit and corrective suggestions in the flowsheet. In other words, mineralogy of the comminution process leads to the determination of the optimal conditions of the process by providing practical information on the liberation degree of valuable minerals in each size fraction, particle size distribution and how the minerals are interlocked9.
Molybdenite is a sulfide mineral and is often found in copper-molybdenum porphyry deposits, which is processed as a by-product of copper during the flotation process10. Molybdenite mineral is one of the most stable members of Transition Metal Dichalcogenides (TDMs), which are present in the hexagonal system as a 2H poly as well as a 3R. Figure1 shows a schematic image of the MoS2 layered structure along with the XRD pattern crystallographic plates. Sulfur atoms at higher and lower surfaces surround smaller molybdenum atoms in the form of sandwiches11. Molybdenum and sulfur atoms inside the layers are bonded together by strong covalent bonds, but the successive layers of sulfur atoms are joined bonded by weak Van der Waals bonds12.
Due to the crystalline structure mentioned, molybdenite has anisotropic properties, which has led to its different behavior in different faces14, including the fact that the anisotropic property leads to the preferred orientation of molybdenite mineral during grinding; and as the particle size decreases, this orientation increases. In this way, molybdenite is broken under grinding at two different surfaces. Surfaces created by breaking SS bonds (non-polar surfaces) and surfaces resulting from the breakage of strong Mo-S bonds (polar edges)15. It is worth mentioning that in the structure of molybdenite, the bond between SS and Mo-Mo is of Van der Waals type and SMoS is of covalent type. Due to the fact that Van der Waals forces/bounds are relatively weak compared to covalent bonds, it is more likely to break at the edges. On the other hand, the behavior of molybdenite in other concentration processes, such as flotation, is greatly influenced by how it breaks and its liberation degree during milling. The natural floatability of molybdenite is related to its textural characteristics such as flatness, roundness of particles, longitudinal elongation ratio and smooth surfaces. Due to the preferred cleavages along weak SS and MoMo bonds during the grinding process, plate-like fragments are produced from larger particles. Flat and long particles cause poor performance of particle-bubble attachment and thus reduce recovery. Therefore, molybdenite flotation behavior is the result of a combination of the property of natural floatability and particle morphology.
The study of the flotation of copper and molybdenite ores indicates that the recovery of molybdenite and copper flotation is reduced in the coarse, fine and very fine size ranges of these minerals. The highest recovery of molybdenite and copper occurs at sizes 2755 microns16, therefore the optimal grinding of these minerals is of great importance. Due to the anisotropic behavior of molybdenite and its association with other sulfide minerals (and other associated gangue minerals), performing process mineralogy studies can lead to results for proper design or optimization of its comminution circuit. In the present study, the type and behavior of copper sulfide, molybdenite and associated gangue minerals, especially pyrite, have been identified through a process mineralogy approach toward molybdenite comminution circuit (Sungun copper-molybdenum processing complex located in northwestern of Iran). For this purpose, mineralogical studies have been performed on mill feed and product. As a result of these studies, the liberation degree and the particle size, distribution and the interlocking mode of the minerals have been determined. Analysing and combining this information with the operating conditions of the plant led to solutions for optimizing the current comminution circuit. In other words, according to the mineralogy of feed and product of mills, the most optimal operating conditions are determined and implemented in order to improve the efficiency of the circuit.
The present study was performed on the grinding efficiency of molybdenum flotation circuit mills of Sungun coppermolybdenum processing complex. Sungun copper mine and complex with geographical coordinates of 46 43 east and 38 42 north is located in northwest of Iran. In the molybdenum processing plant of Sungun complex (flowsheet is shown in Fig.2) uses two ball mills to perform grinding operations. The primary ball mill operates in an open circuit (after the middle thickener) and the secondary ball mill operates in a closed circuit with a hydrocyclone. The underflow of the middle thickener with a solid percentage of about 55.0% enters the primary ball mill, and the mill product enters the cleaner 3 flotation cells after dilution. The length and diameter of the ball mill used in this department are 2.44 and 1.52m, respectively, which has a slurry mass capacity of 4.13 t/h. Cleaner 4 concentrate is introduced into hydrocyclone clusters (two hydrocyclone clusters, each of which consists of 3, 6-in. cyclone devices) and is separated into a fine overflow fraction with particles size smaller than 38.0 microns and an overflow fraction with particles size larger than 38.0 microns. The overflow of each cluster is transferred to 58 cleaner cells and its underflow enters a regrinding ball mill, which is in a closed circuit with these hydrocyclones. The goal of the regrinding stage is to achieve the maximum liberation degree of molybdenite and copper minerals and their liberation from each other. The length and diameter of the closed-circuit ball mill are 1.83 and 1.22m, respectively, which has a slurry mass capacity of 3.6 t/h and a solid percentage of 5060% (designed for the plant).
In order to study the mineralogy of milling process in the molybdenum processing circuit, samples of feed (middle thickener underflow) and product of open-circuit ball mill, as well as overflow and underflow of hydrocyclone (feed) and product of closed-circuit ball mill were prepared. Sampling points are marked in red in Fig.2. It is worth mentioning that 3 samples were collected using a sampling spoon from each location and at 30-min intervals, (to investigate the effect of plant feed variations). After filtering and drying, 30g of the sample was prepared using a riffle sample splitter to check size distribution, preparation of polished sections and performing microscopic studies.
In order to perform the process mineralogy studies on the samples, optical microscopic studies were performed on polished sections after ehaviour the size of the feed and product particles using Laser Particle Size Analyzer (SLS: mastersizer 2000/Malvern Panalytical technology). The results of particle size analysis of feed and product samples from grinding circuits are shown in Fig.3, and comparison of their d10, d50 and d90 values are performed in Table 1. According to the results, grinding by open-circuit ball mill has caused the particles size to decrease from 90.0% smaller than 45.0 microns to about 99.0% smaller the 45.0 microns. Grinding by closed-circuit ball mill has also reduced d90 value of particles from 67.96 (mill feed) to 65.13 microns (mill product).
Microscopic study of polished sections is the most common method of studying the mineralogical properties and textural association between minerals in mineral samples. In order to investigate the grinding ehaviour of the minerals in the molybdenum processing circuit, microscopic studies were conducted on polished sections prepared from the collected samples. Microscopic studies were performed using the Leitz polarizing microscope of model SM-LUX-POL equipped with a digital imaging camera at the college of Mining Engineering, University of Tehran. Based on the results of mineralogical studies, open-circuit mill feed contains chalcopyrite, molybdenite, pyrite and covellite. Besides, molybdenite, chalcopyrite, and pyrite are the major minerals that make up the feed of closed-circuit mill (hydrocyclone underflow). Table 2 shows the important physical/chemical properties of various minerals in the feed of open and closed-circuit mill of the molybdenite processing circuit.
The feed of the primary ball mill or the open-circuit mill is the underflow pulp from the middle thickener (Fig.4). According to Fig.2 circuit, the middle thickener with a diameter of 12m and free settling mechanism (in the molybdenum plant of the Sungun coppermolybdenum Complex) is located after cleaner 2 and before the open-circuit mill. The feed pulp to this thickener has a solid percentage of 13.87, which after settling, the underflow is discharged with a solid percentage of about 60.0% and the overflow weight percent is 0.04%. Based on the grade analysis performed on the underflow of the thickener or open-circuit mill feed, the grade values of Mo, Cu and Fe elements are 23.71, 17.57 and 14.84%, respectively. Due to the 23.71% value of molybdenum grade, this product cannot be supplied as a final concentrate and it is necessary to perform more processing stages (cleaner flotation stages). Therefore, the purpose of grinding at this stage is to achieve more liberation of copper minerals from molybdenite.
According to the PSD diagram shown in Fig.5 as well as Table 1, the grinding process in open-circuit mill produces about 70.0% of the fine product with particle size smaller than 20.0 microns; of this amount, 40.0% is smaller than 10.0 microns in size. Particles with a size smaller than 7.0 microns also have a significant volume and account for approximately 25.0% of the mill product particles. Based on the results, it can be concluded that the highest amount of grinding occurred for particles in the size ranges of d75-d25 of feed. Grinding also occurred for feed particles smaller than d25 (7.0 microns), but grinding did not have a desirable result for particles larger than 23.0 microns. Given that the product of open-circuit mill is the feed to cleaner 3 flotation cells, the size distribution of the mill product particles (in other words, the amount of grinding in the mill) is of great importance. Since with increasing the amount of particles smaller than 10.0 microns, flotation recovery of molybdenite gradually decreases due to reduced probability of collision and attachment to air bubbles 17.
According to optical reflected light microscopic studies, chalcopyrite, molybdenite, pyrite, and covellite (Table 2) make up the major part of the composition of the middle thickener underflow or open-circuit mill feed. Liberation studies of molybdenite in the feed of open-circuit mill (Fig.6A) indicate that this mineral has achieved a proper liberation degree (about 98.0%). Therefore, at this stage, grinding leads to more fine production of molybdenite particles and does not cause a significant change in their liberation degree. The study of PSD of the mill product also confirms this; In other words, at this stage, the particles of molybdenite and copper sulfides have become finer and have turned into slimes.
(A) Molybdenite liberated particle in feed, (B) Distribution of minerals in the product, and (C) Molybdenite liberated particles in the product of open-circuit ball mill (Mo molybdenite, Cpy chalcopyrite, Py pyrite).
As can be seen in Fig.6B, the minerals present in the product of open-circuit mill, in the order of abundance, include molybdenite, which is mostly free and has become fine after grinding in mill, liberated chalcopyrite particles (about 32.0%) and interlocked with pyrite and other associated gangue minerals, and liberated and interlocked particles of pyrite. Liberation studies for chalcopyrite and pyrite in the product of mill shows 85.0% and 85.40% values of liberation degree for these minerals, respectively. It can be stated that these two minerals have the same liberation degree. On the other hand, there was no significant interlocking between copper sulfide minerals and molybdenite (Fig.6C). Coarse particles of molybdenite are observable in some cases, but in general the molybdenite particles are ground to very fine size ranges (slime range).
Due to its anisotropic properties, molybdenite behaves differently from other sulfide minerals. SEMTEM images (Fig.7) show how MoS2 breaks (cleaves) in layers. Because in molybdenite with hexagonal structure, the SMoS layers are connected with the covalent bond by the Van der Waals forces. Once molybdenite is ground, its cleavage occurs more easily within the weak Van der Waals forces. Hence, the outer layers of the mineral surface are removed under the influence of shear forces; While the compressive forces in the grinding environment affect the mineral edges and cause breakage in the direction of the edges. The above-mentioned grinding processes, which reduce the thickness of the mineral, occur in the early stages of grinding. As the grinding time increases and in the final stages of grinding, the particle size of the produced layers decreases at a slower rate when compared to the early stages of grinding13. As the size of the molybdenite particles decreases, the surface-to-edge ratio decreases, resulting in an increase in its hydrophilic properties, which reduces its floatability.
The cleaner 4 concentrate with a molybdenum grade of 39.34% and a copper and iron grade of 4.94% and 10.34%, respectively, must be re-floated in order to achieve a higher grade (Fig.2). In this regard, this concentrate is introduced into the hydrocyclone with a separation limit of 38.0 microns, and the underflow is rejected to the closed-circuit ball mill for regrinding (Fig.8). The goal of the secondary grinding stage is to achieve the maximum liberation degree of molybdenite and copper minerals from each other. Examination of the hydrocyclone underflow size distribution (ball mill feed) and mill product (Fig.9 and Table 1) shows that grinding did not have much effect on reducing particle size. In general, the highest grinding occurred in the size range of feed d25d50 values.
Microscopic studies have been performed on mill feed (hydrocyclone underflow) and product. Figure10A,B shows a picture of the minerals distribution in the studied samples. According to studies, the composition of hydrocyclone underflow (mill feed), in order of abundance, includes liberated molybdenite in the form of blades and polygonal fragments (Fig.11), liberated and interlocked chalcopyrite with gangue minerals and liberated and interlocked pyrite particles. The liberation degree of chalcopyrite and pyrite in mill product are 84.40% and 91.40%, respectively. It is worth mentioning that despite the performed grinding, interlocking between molybdenite and chalcopyrite have rarely been observed. In general, according to mineralogical studies and PSD of feed and product of closed-circuit ball mill, the minerals in feed and product are almost similar in terms of particle liberation degree and only particle size has become finer.
As mentioned, the cut size of hydrocyclone in closed circuit with the ball mill is 38.0 microns, however, according to the microscopic images of the hydrocyclone underflow sample (Fig.12), small particles are also observed in this sample, which is due to the inefficiency of the hydrocyclone. Figure13 shows the particle size distribution diagram for hydrocyclone feed, overflow, and underflow, which indicates poor classification performance of the hydrocyclone in the circuit, in the separation of fine particles. According to PSD diagram, d90 value of feed, overflow and underflow of hydrocyclone are 62.0, 55.0 and 68.0 microns, respectively. On the other hand, by measuring the solid percentage of feed (12.0%) and underflow (15.0%) of hydrocyclone, used with the aim of dewatering and particle size control, it can be concluded that this device did not have a good dewatering performance 18. Given the above, the low solid percentage of feed and the abundance of fine particles in the feed can be considered factors responsible for the poor performance of the mill.
In order to more accurately examine the performance of closed-circuit mill, the results of the final molybdenum concentrate analysis were studied in two modes of mill-on and mill-off in a 10-day period (each day including three 8-h shifts). Figure14 shows diagrams for grade analysis of molybdenum, copper, and iron in two modes of mill-on and mill-off for closed-circuit ball mill. According to the figure, the plant circuit is operating in optimal conditions when the mill is switched off. As the circuit became out of optimal conditions, the closed-circuit mill switched on, but the start-up of the closed-circuit mill did not improve the conditions of the processing circuit. It is worth mentioning that the optimal condition means the grade of molybdenum, copper and iron in molybdenum concentrate is more than 50.0% and less than 1.0% and 3.0%, respectively (according to the standards of commercial markets). According to the results, the average grade of molybdenum, copper and iron in molybdenum concentrate in mill-on (non-optimal) mode is 51.21%, 1.32% and 3.95%, respectively. While the average values for Mo, Cu and Fe grade in mill-off mode are 53.83%, 0.71% and 2.04%, respectively. As can be seen, in mill-off mode and optimal conditions, grade standards for molybdenum, copper and iron are available in molybdenum concentrate. However, in non-optimal conditions, starting of the mill has no effect on the optimization of these values, and this indicates the inefficiency of the mill.
Grade analysis of the final molybdenum concentrate for elements (A) molybdenum, (B) copper and (C) iron in the off and on mode of closed-circuit ball mill in a period of 10days (every day includes 3, 8-h shifts).
It is important to know the type and properties of the minerals in the ore being processed in order to design and optimize the circuit of a processing plant. In this study the efficiency of grinding was investigated by studying the mineralogical properties of feed and product streams to the grinding circuits in the molybdenum processing plant. Analysis of particle size distribution for open and closed-circuit ball mills feed and product showed that d90 value of feed and product of open-circuit mill is 43.83 and 42.33 microns, respectively, and d90 value of feed and product of closed-circuit mill were 67.95 and 65.13 microns. The performed liberation study also shows that in the feed and product of the open-circuit mill, the liberation degree of molybdenite is almost the same and about 98.0%. Therefore, in the milling stage, the molybdenite particles are only ground and there is no significant change in their liberation degree. On the other hand, because there is no controlling equipment for particle size in open-circuit mill, fine materials turn into slimes and due to the slime coating, entrainment and less efficient collision of the particles with the air bubble, the flotation rate and the grade is reduced. With excessive grinding of materials in the mill, the surface-to-edge ratio, especially in the case of molybdenite, is reduced, and due to reduced hydrophobicity and floatability, fine molybdenite particles are introduced to tailing product or copper concentrate. In the case of closed-circuit mill, the minerals present in the feed and product of the mill are almost identical in terms of particle liberation degree, and only the particle size gets finer. An examination of the molybdenum, copper and iron grade changes over a 10-day period for both mill on and off modes of closed-circuit mill showed that in the mill-off mode, the plant circuit is in optimal conditions (molybdenum, copper and iron grade in the molybdenum concentrate were more than 50.0% and less than 1.0% and 3.0%, respectively), but as the circuit gets out of optimal condition, the start-up of closed-circuit mill has not had an effect on improving the circuit and creating optimal conditions.
Becker, M., Brough, C., Reid, D., Smith, D. & Bradshaw, D. Geometallurgical characterisation of the Merensky Reef at Northam Platinum Mine; comparison of normal, pothole and transitional reef types. in International Congress for Automated Mineralogy. 391399 (2008).
Hunt, J. A., Berry, R., Walters, S., Bonnici, N., Kamenetsky, M., Nguyen, K. & Evans, C. L. A new look at mineral maps and the potential relationships of extracted data to mineral processing behaviours. in Ninth International Congress for Applied Mineralogy. AusIMM. 429432 (2008).
Ansari, A. & Pawlik, M. Floatability of chalcopyrite and molybdenite in the presence of lignosulfonates. Part II. Hallimond tube flotation. Miner. Eng. 20(6), 609616, https://doi.org/10.1016/j.mineng.2006.12.008 (2007).
Shalchian, H., Vahdati Khaki, J., Babakhani, A., Taglieri, G., Michelis, I. D., Daniele, V. and Veglio, F. On the mechanism of molybdenite exfoliation during mechanical milling. J. Ceram. Int. 43(15), 1295712967 (2017).
Kim, Y., Huang, J. L. & Lieber, C. M. Characterization of nanometer scale wear and oxidation of transition metal dichalcogenide lubricants by atomic force microscopy. Appl. Phys. Lett. 59, 34043406 (1991).
Abdollahi, M. The effect of texture and mineralogy on flotation recovery of molybdenite at the Sungun copper complex /concentrator plant. in Master of Science Thesis in Mining EngineeringMineral Processing, Urmia University (in Persian) (2019).
Open Access This article is licensed under a Creative Commons Attribution 4.0 International License, which permits use, sharing, adaptation, distribution and reproduction in any medium or format, as long as you give appropriate credit to the original author(s) and the source, provide a link to the Creative Commons licence, and indicate if changes were made. The images or other third party material in this article are included in the article's Creative Commons licence, unless indicated otherwise in a credit line to the material. If material is not included in the article's Creative Commons licence and your intended use is not permitted by statutory regulation or exceeds the permitted use, you will need to obtain permission directly from the copyright holder. To view a copy of this licence, visit http://creativecommons.org/licenses/by/4.0/.
Bahrami, A., Abdollahi, M., Mirmohammadi, M. et al. A process mineralogy approach to study the efficiency of milling of molybdenite circuit processing. Sci Rep 10, 21211 (2020). https://doi.org/10.1038/s41598-020-78337-8Get in Touch with Mechanic