Screening is the passing of material through definite and uniform apertures is the only true and accurate means of grading to a required particle size. Air separation and hydraulic classification depend upon gravity and particle shape, and result in the segregation and retention of material of higher specific gravity and lower surface area irrespective of size.
The use of Screens increases with the education and civilization of a people and with the improving and perfecting of an art. In our advanced civilization practically everything that we eat, wear and use has been in contact with, or dependent upon screens in some phase of its growth, development or processing. In this treatise, we are only concerned with the sorting, grading or sizing as accomplished with a mechanical screening device.
Some materials such as beach sands, clays, native chemicals, etc., occur in nature in a closely graded state resulting from a mechanical water sorting, precipitation or gravity deposition. They require only scalping or some form of treatment for removal of tramp coarse foreign elements. Others such as salt, sugar and various chemicals are crystallized or precipitated in their processing to fairly close limits of size. They require only such sorting or grading as is dictated by market preference and conditions of use.
In mechanical mixtures such as raw cement, finished fertilizers, stock feeds, etc., the ingredients are blended, ground and screened to a definite fineness. This maintains the intimate relationship by preventing segregation of a coarse constituent through automatic sorting. We have all noted how by piling an ungraded material the fines will segregate in the center of the pile and the coarse will automatically run to the outside and bottom. Metallic and non-metallic ores, stone and other aggregates, coal and coke, various furnace products, chemicals, cerealsetcetera, must be crushed, ground, disintegrated or pulverized before they can go on to further processing and ultimate use. In these fields screens are used for sorting into definite grades, top scalping for removal of coarse oversize and foreign material, bottom scalping for elimination of fines and dirt, and to return oversize to a crusher or grinder until it is reduced to a size finer than the opening of the screen. This latter practice is known as closed circuit crushing or grinding.
A nest of standard brass framed screens, with a definite ratio between openings, is used to sort a representative sample into the clean fractions retained on each screen. The tabulated resulting sieve analysis graphically shows the percentages of given sizes present in the sample. (Table I, p. 347). It indicates just what is available for recovery by screening through and over certain openings in a commercial production screening operation and also shows the reduction obtained by passage through crusher or grinding mill.
Another important factor in commercial screening that will be revealed by a sieve analysis is the percentage of near-mesh material present in the screen feed. If, for instance, it is observed that 40 percent of the sample had passed through the 8-mesh testing screen and was retained on 10- mesh and another 40 percent had passed through 10-mesh and was retained on 14-mesh, an efficient productionscreening operation at 10-mesh would require the maximum in screen area, particularly as to length. This preponderance of near-mesh, or go and no-go size of particle, obviously makes a difficult separation condition. In such cases unless the proper care is taken in the selection of the type of screening device and the specification of the wire cloth used on it, the openings may fill up and blind to a point where no separation is obtained.
In addition to the necessary sieve analysis, other factors must be known before a proper and intelligent recommendation can be made on any but the simplest of screening problems. Many cases require a laboratory test, simulating actual operating conditions, before the size and type of the screen can be determined and proper specification of screen cloth selected. The screen doctor must have the answer to the following questions before he can make proper diagnosis and prescribe treatment:
Capacity required in tons or gallons per hour? This should be expressed in both average and maximum, because peak loads, even of short duration, may result in spoiling of products previously graded or may upset subsequent steps in the operation, due to the drop in screening efficiency. Sufficient screen area should be provided to handle the maximum load.
Type of screening, wet or dry? How much water can be added? In the case of wet screening it is necessary to know if a definite density of the through screen product must be maintained and how much spray water can be added to rinse the oversize.
Percentage of moisture present in the feed? The maximum figure should be given here because different materials become unscreenable at varying degrees of moisture. To effect a separationat a given fineness it may be necessary to dry the material or add water and wash it through the screen.
Is material free-screening? An affirmative answer here obviates practically all other questions. Sticky? As clay, some food products, chemicals, etc. This determines if screening is practical and type of wire cloth recommended.
By closed circuit crushing or grinding it is meant that the product from a crusher or grinder is fed to a screen. The material that has been reduced to sufficient fineness passes through the openings and the oversize is returned to the breaker for further reduction. Escape from the circuit can only be through the screen so this product, the undersize, is equal in tonnage to the initial feed to the crusher or mill. The oversize returned for further work is known as the circulating load. It is a most important factor and can be extremely insidious. If the screen is inefficient and rejects finished material or if the crusher will not reduce the oversize fast enough, this load may build up, and rapidly, to a point beyond the capacity of the breaker, the screen or the conveying equipment, whicheverproves to be the neck of the bottle.
For greatest economy and efficiency, fines should be removed by means of a screen just as fast as they are created in each successive stage of crushing or grinding. Most every case must be handled on its own individual merits and proper balance worked out. In some cases a circulating load as high as 1,000 percent is considered economical. Picture how this would affect the requirement in screening capacity with eleven tons of material handled for every ton produced.
The percentage of circulating load can be readily determined from the sieve analyses of the screen feed, the oversize and the undersize (See Table 1). Samples should be taken simultaneously after circulating load has reached its peak. Conditions and analyses will be similar to those set forth in flowsheet at right. The formula can be expressed:
PercentCirculating Load=100 (B-C/A-C -1) A=Percent finer than required sizein the screen feed. B=Percent finer than required sizein the screen undersize. C=Percent finer than required sizein the screen oversize.
In the example, A equals 35.0,B equals 95.0, and C equals5.0. The value of 1 in the formula represents the initial feed to the circuit which is equivalentto the undersize, or product removed through the screen.
Percent Efficiency=100(100 F-D/AF) A=Percent finer than required size in the screen feed. D=Percent coarser than required size in the screen feed. F =Percent coarser than required size in the screen oversize.
There are different schools of thought on this subject and other formulae. Some operators are satisfied to simply use the percentage coarser than the screen opening in the overscreen product as the efficiency figure. This would be F in the above formula and 95 percent instead of 90.22 percent.
Dependent on the nature of the material and type of operation, screening may be accomplished through bars, perforated plate or woven wire screen. The bar screen is used for scalping extremely coarse material where definite sizing is of secondary importance and abrasion is severe. Perforated plate offers a smooth surface upon which heavy oversize will slide very easily, often too easily for good screening. Under some conditions it blinds less readily than woven wire screen. Objections to it are the fact that the openings wear gradually larger and larger, and the percentage of blank area is so high.
For most purposes woven wire screen, or wire cloth, is the best medium. With it the maximum in open area can be obtained. Various weights, metals and alloys, and shapes of openings are available to satisfy conditions of heavy load, abrasion, corrosion, screenability and capacity. Mesh in wire cloth is the number of openings per lineal inch and means nothing unless accompanied by the decimal designation of the wire diameter or the actual opening of the screen. It is best to specify the required screen opening as this can then be obtained in several meshes, dependentupon the weight of wire that is used. Obviously, for a given opening, the greater the mesh count and the finer the wire diameter, the higher will be the percentage of open area in the fabric.
Much as we might like to do so, we cannot have our cake and eat it, too. Therefore, the selection of a screen specification is usually a compromise. Dependent upon conditions, screen life is constantly being sacrificed for screenability and vice versa. For instance, a heavy and abrasive material suggests an extra heavy wire to secure maximum life. It is found, however, that the low percentage of open area restricts capacity and that the large wire diameter promotes blinding and lowers efficiency. A compromise is, therefore, made by easing off on the weight of the wire. Conversely, another material may, for instance, be damp and sticky, dictating the use of an extremely fine diameter of wire to minimize the surface upon which it may build up. Such a screen specification may last only a few hours and capacity and efficiency must be sacrificed in the interest of longer screen life.
Rectangular and elongated screen openings assist greatly in increasing capacity and eliminating blinding. The opening in a square mesh screen is shaped similar to a funnel and particles can be wedged into it to bear on all four sides. The rectangular opening limits this contact to three sides and thus minimizes the possibility of wedge blinding. When this slot is further elongated to many times the opening width, a springing of the long wires is possible and permanent blinding is eliminated. Naturally, these long openings can not be used for true sizing of anything but cubical or granular materials. Where flakes and slivers are present and cannot be tolerated in the screen under-size, square mesh cloth must be used at the sacrifice of capacity.
For abrasion resistance, high- carbon spring steel wire is available. Stainless steel and the non- ferrous alloys give a selection where rust and corrosion are a factor. The difference between success and failure of a screening operation may rest with the selection of the proper screen clothspecifications and this subject requires considerable thought and study, plus experience.
Reviewing the foregoing, it is readily understandable that a fixed table of screen capacities would be misleading and dangerous. There are so many variables that two neighbouring plants, working on the same deposit, may have entirely different screening conditions, due, for instance, to a difference in crushing practice. Larger tonnages can be handled on scalping operations, and in some cases with closed circuit crushing, than on close grading into specific fractions. On some materials a scalping deck over the sizing screen increases capacity by breaking and distributing the load and opening- up the mat of material. Washing increases capacity materially over so-called dry screening.
From the grizzly and trommel we have seen the development of screening devices through the shaking, knocking and bumping stages to the high speed vibrating screen of today. This development ran the range of eccentric head motions; knockers; cams; air, cam and electric vibrators; unbalanced shafts and eccentric flywheels; grasshopper motions, etc., up to the present positive-drive, high-speed, circle-throw, eccentric- shaft screen.
In this type the throw and speed must be properly specified and coordinated to secure the best screening action. Bearings should not be under shock and design should not be complicated with compensators and adjustments to eat power and tempt experimentation. The loading of the bearings should be so minimized that the equipment manufacturer evidences his confidence in his design by extending a generous guarantee.
In closing, it is recommended that the screen user select a proved and simple machine that will give uniform, continuous, care-free operation. Your supplier should qualify to consult with you on installation, operation, and selection of proper screen cloth specifications. Do not overlook this important service feature.
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.
Though the gold recovery methods previously discussed usually catch the coarser particles of sulphides in the ore and thus indirectly recover some of the gold associated with these and other heavy minerals, they are not primarily designed for sulphide recovery. Where a high sulphide recovery is demanded, flotation methods are now in general use, but in the days before flotation was known, a large part of the worlds gold was recovered by concentrating the gold-bearing sulphides on tables and smelting or regrinding and amalgamating the product.Though the modern trend is away from the use of tables, because flotation is so much more efficient.
The flotation process, which is today so extensively used for the concentration of base-metal sulphide ores and is finding increased use in many other fields. In1932flotation plants began to be installed for the treatment of gold and silver ores as a substitute for or in conjunction with cyanidation.
The principles involved and the rather elaborate physicochemical theories advanced to account for the selective separations obtained are beyond the scope of this book. Suffice it to say that in general the sulphides are air-filmed and ufloated to be removed as a froth from the surface of the pulp while the nonsulphide gangue remains in suspension, or sinks, as the expression is, for discharge from the side or end of the machine.
For more complete information reference is made to Taggarts Hand book of Mineral Dressing, 1945; Gaudins Flotation and Principles of Mineral Dressing; I. W. Warks Principles of Flotation; and the numerous papers on the subject published by the A.I.M.E. and U.S. Bureau of Mines.
Flotation machines can be classed roughly into mechanical and pneumatic types. The first employ mechanically operated impellers or rotorsfor agitating and aerating the pulps, with or without a supplementary compressed-air supply. Best known of these are the Mineral Separation, the Fagergren, the Agitair, and the Massco-Fahrenwald.
Pneumatic cells use no mechanical agitation (except the Macintosh, now obsolete) and depend on compressed air to supply the bubble structure and tohold the pulp in suspension. Well-known makes include theCallow and MacIntosh (no longer manufactured) the Southwestern, and the Steffensen, the last, as shown in the cross-sectional view in Fig. 47, utilizing the air-lift principle, with the shearing of large bubbles as the air is forced from a central perforated bell through a series of diffuser plates.
The number and size of flotation cells required for any given installation are readily determinedif the problem is looked upon as a matter of retention time for a certain total volume of pulp. The pulp flow in cubic feet per minute is determined from the formula
For ordinary ratios of concentration the effect on cell capacity of concentrate (or froth) removal can be neglected, but where a high proportion of the feed is taken off as concentrates, or where middlings are removed for retreatment in a separate circuit, due allowance should be made for reduced flow and, in consequence, increased detention time toward the tail end of a string of cells. Not less than a series of four cells and preferably six or more cells should be used in any roughing section in order to prevent short-circuiting.
It is not intended here to discuss the subject of flotation reagents in anydetail. The subject is a large one with a comprehensive technical and patent literature. Research leading to the development of new reagents and to our understanding of the mechanism involved has been largely in the hands of academic institutions and the manufacturers of chemical products.
Recent work reported by A. M. Gaudin on the use of Radioactive Tracers in Milling Research described, for instance, the use of a flotation reagents containing radioactive carbon to determine the extent of collector adsorption. The bubble machine devised to measure the angle of contact of air bubbles on collector-treated mineral surfaces has been extensively used for determining the theoretical value of various reagents as flotation collectors, but for the most part the actual reagent combination in use in commercial plants is usually the result of trial-and-error methods.
The following is a brief discussion of the reagents ordinarily used for the flotation of gold and silver ores prepared from notes submitted by S. J. Swainson and N. Hedley of the American Cyanamid Company.
Conditioning agents are commonly used, especially when the ores are partly oxidized. Soda ash is the most widely used regulator of alkalinity. Lime should not be used because it is a depressor of free gold and inhibits pyrite flotation. Sodium sulphide is often helpful in the flotation of partly oxidized sulphides but must be used with caution because of its depressing action on free gold. Copper sulphate is frequently helpful in accelerating the flotation of pyrite and arsenopyrite. In rare instances sulphuric acid may be necessary, but the use of it is limited to ores containing no lime. Ammo-phos, a crude monoammonium phosphate, is sometimes used in the flotation of oxidized gold ores. It has the effect of flocculating iron oxide slime, thus improving the grade of concentrate. Sodium silicate, a dispersing agent, is also useful for overcoming gangue-slime interference.
Promoters or Collectors. The commonly used promoters or collectors are Aerofloat reagents and the xanthates. The most effective promoter of free gold is Aerofloat flotation reagent 208. When auriferous pyrite is present, this reagent and reagent 301 constitute the most effective promoter combination. The latter is a higher xanthate which is a strong and non-selective promoter of all sulphides. Amyl and butyl xanthates are also widely used. Ethyl xanthate is not so commonly used as the higher xanthates for this type of flotation.
The liquid flotation reagents such as Aerofloat 15, 25, and 31 are commonly used in conjunction with the xanthates. These reagents possess both promoter and frother properties. When malachite and azurite are present, reagent 425 is often a useful promoter. This reagent was developed especially for the flotation of oxidized copper ores.
The amount of these promoters varies considerably. If the ore is partly oxidized, it may be necessary to use as much as 0.30 to 0.40 lb. of promoter perton of ore. In the case of clean ores, as little as 0.05 lb. may be enough. The promoter requirement of an average ore will usually approximate 0.20 lb.
The commonly used frothers are steam-distilled pine oil, cresylic acid, and higher alcohols. The third mentioned, known as duPont frothers, have recently come into use. They produce a somewhat more tender and evanescent froth than pine oil or cresylic acid; consequently they have less tendency to float gangue, particularly in circuits alkaline with lime. The duPont frothers are highly active frothing agents; therefore it is rarely necessary to use more than a few hundredths of a pound per ton of ore.
When coarse sulphides and moderately coarse gold (65 mesh) must be floated, froth modifiers such as Barrett Nos. 4 and 634, of hardwood creosote, are usually necessary. The function of these so-called froth modifiers is to give more stable froth having greater carrying power.
The conditioning agents used for silver ores are the same as those for gold ores. Soda ash is a commonly used pH regulator. It aids the flotation of galena and silver sulphides. When the silver and lead minerals are in the oxidized state, sodium sulphide is helpful, but it should not be added until after the sulphide minerals have been floated, because sodium sulphide inhibits flotation of the silver sulphide minerals.
Aerofloat 25 and 31 are effective promoters for silver sulphides, sulphantimonites, and sulpharsenites, as well as for native silver. When galena is present, No. 31 is preferable to No. 25 because it is a more powerful galena promoter. Higher xanthates, such as American Cyanamid reagent 301 and amyl and butyl xanthates, are beneficial when pyrite must be recovered. When the ore contains oxidized lead minerals, such as angle-site and cerussite, sodium sulphide and one of the higher xanthates may be used. In some instances reagent 404 effects high recovery of these minerals without the use of a sulphidizing agent.Silver ores require the same frothers as gold oresviz., pine oil, cresylic acid or duPont frothers.
Aero, Ammo-phos, and Aerofloat are registered trade-marks applied to products manufactured by this company. The Great Western Electro-Chemical Company, California, makes amyl xanthate, butyl xanthate, potassium xanthate, and sodium xanthate. In the United States these reagents are used on the gold ores of California and Colorado and in Canada on the gold ores and sulphides of Ontario and Quebec.
Flotation reagents of the Naval Stores Division of the Hercules Powder Company are as follows: Yarmor F pine oil, a frother for floating simple and complex ores; Risor pine oil, for recovering sulphides by bulk flotation; Tarol a toughener of froth, generally used in small amount with Yarmor F, but with some semioxidized ores where high recovery is essential yet the grade of concentrate not so important, Tarol does good work; Tarol a frother for floating certain oxide minerals, but it can be used in selective flotation of sulphide minerals and in bulk flotation where tough frothis desirable; Solvenol, for the floating of graphite in conjunction with Yarmor F.
The statement has come to the attention of the American Cyanamid Company that organic flotation reagents, such as xanthates, even in the small amounts used in flotation, cause reprecipitation of gold from pregnant cyanide solutions. The ore-dressing laboratory of this company is studying the question, and preliminary results indicate that this statement is unfounded. The addition of xanthate, in the amount usually found in flotation circuits, does not precipitate gold from a pregnant cyanide solution containing the normal amount of cyanide and lime.
Valueless slime, in addition to its detrimental effect in coating gold-bearing sulphide, thereby limiting or preventing its flotation, also becomes mixed with the flotation concentrate and lowers its value. Sometimes the problem in flotation is that, although the gold is floatable, the concentrate product is of too low grade. Talc, slate, clay, oxides of iron, and manganese or carbonaceousmatter in ores early form slime in a mill, without fine crushing. Such primary slime, according to E. S. Leaver and J. A. Woolf of the U.S. Bureau of Mines, interferes with the proper selectivity of the associated minerals and causes slime interference. The tendency of primary slime is to float readily or to remain in suspension and be carried over into the concentrate. Preliminary removal and washing of this primary slime before fine crushing is one method of dealing with it. At the Idaho-Maryland mill, Grass Valley, Calif., starch is regularly used as a depressant during flotation. Flotation tests using starch were made on a quartz ore containing carbonaceous schist from the Argonaut mine, Jackson, Calif.; a talcose ore from the Idaho-Maryland mine mentioned; a talcose-clayey ore from Gold Range, Nev.; a siliceous, iron and manganese oxide ore from the Baboquivari district, Nevada; carbonaceous and aluminous slime from the Mother Lode and some synthetic ores. The conclusions from the foregoing tests were in part as follows:
It acts first on the slime; then, if a sufficient excess of starch is present, it will cause some depression of sulphides and metallic gold, either by wetting out or by producing an extremely brittle froth. Therefore, care must be taken in regulating the amount of starch added to obtain the maximum depression of the slime commensurate with high recovery of the gold. In this, as in all other phases of flotation, each ore presents an individual problem and must be so studied.
It wasdescribe by the use of 600 series of flotation reagents which were developed primarily for the purpose of depressing carbonaceous and siliceous slimes in the flotation of gold ores. Carbonaceous material not only greatly increases the bulk and moisture content of a flotation concentrate, but its presence makes cyanidation of the concentrate difficult or impossible owing to reprecipitation of the gold during treatment.
In the treatment of an auriferous sulphide ore associated with carbonaceous shale from South Africa, up to 77 per cent of the carbon was eliminated by the use of 1 lb. per ton of reagent 637 with a 90.5 per cent gold recovery at 20.4:1 ratio of concentration.
A gold carbonaceous sulphide ore from California carrying free gold yielded a 93 per cent recovery into a concentrate at 14.4:1 to ratio of concentration after conditioning with 0.50 lb. per ton of reagent 645.
In each case the ore was ground to about 70 per cent minus 200 mesh and conditioned at 22 per cent solids with the reagents as indicated. Flotation reagents included reagents 301 and 208 and pine oil. In the second case some soda ash and copper sulphate where also used.
It is obvious that the most suitable treatment for ores carrying gold and silver associated with pyrite and other iron sulphides, arsenopyrite or stibnite, will depend on the type of association. Cyanidation is usually the most suitable process, but it often necessitates grinding ore to a fine size to release the gold and silver. Where it is possible to obtain a good recovery by flotation in a concentrate carrying most of the pyrite or other sulphides, it is often more economical to adopt this method, regrinding only the comparatively small bulk of concentrate prior to the leaching operation.
That the trend over the last 10 years has been in this direction will be noted from the numerous examples of such flow sheets in Canada and Australia (see Chap. XV). A number of plants formerly using all-cyanidation have converted to the combined process.
The suitability of the method involving fine grinding and flotation with treatment of the concentrate and rejection of the remainder should receive careful study in the laboratory and in a pilot plant. Mclntyre-Porcupine ran a 150-ton plant for a year before deciding to build its 2400-ton mill. Comparative figures given by J. J. Denny in E. and M. J., November, 1933, on the results obtained by the all-sliming, C.C.D. process formerly used and the later combination of flotation and concentrate treatment showed a saving of 12.1 cents per ton in treatment cost and a decrease of 15 cents per ton in the residue, a total of 27.1 cents per ton in favor of the new treatment.
Flotation may also prove to be the more economical process for the ore containing such minerals as stibnite, copper-bearing sulphides, tellurides,and others which require roasting before cyanidation, because this reduces the tonnage passing through the furnace.
Even when recovery of gold and silver from such ores by flotation is low, it may be advantageous still to float off the minerals that interfere with cyanidation, roasting, and leaching or possibly to smelt the concentrate for extraction of its precious metals. Cyanidation of the flotation tailing follows, this being simpler and cheaper because of prior removal of the cyanicides.
It is a good practice to recover as much of the gold and silver as possible in the grinding circuit by amalgamation, corduroy strakes, or other gravity means to prevent their accumulation in the classifier; otherwise gold that is too coarse to float may escape from the grinding section into the flotation circuit where it will pass into the tailing and be lost.
To prevent this, several companies including the Mclntyre-Porcupine at Timmins, Ontario, have inserted a combination of flotation cell and hydraulic cone in their tube-mill classifier circuits. At the Mclntyre- Porcupine, according to J. J. Denny in E. and M. J., November, 1933, this cell is a 500 Sub-A type. The total pulp discharged from each tube mill passes through 4-meshscreens which are attached to the end of the mills. The undersize goes to the flotation cell, and the oversize to the classifiers. Tailing from the cell flows to the classifiers, and the flotation concentrate joins the concentrate stream from .the main flotation circuit. The purpose of the hydraulic attachment is to remove gold that is too coarse to float, thus avoiding an accumulation in the tube-mill circuit. The cones have increased recovery from 60 to 75 per cent. Every 24 hr. the tube-mill discharge is diverted to the classifiers. Water is added for 15 min. to separate the gangue in the cells from the high-grade concentrate, after which a product consisting of sulphides and coarse gold is removed through a 4-in. plug valve equipped with a locking device. Each day approximately 400 lb. of material worth $2000 to $3000 is recovered. This is transferred to a tube mill in the cyanide circuit,with no evident increase in the value of the cyanide residue. The object of this arrangement is, of course, primarily to deplete the circulating load of an accumulation of free gold and heavy sulphides.
Flotation is used to recover residual gold-bearing sulphides and tellurides. The Lake Shore mill retreatment plant is an interesting example of this technique. The problem here was, of course, to overcome by chemical treatment the depressing action of the alkaline cyanide circuit on the sulphides. A full discussion of this and of the somewhat controversial subject as to whether flotation should in such an instance be carried out before, or after cyanidation will be found in J. E. Williamsons paper Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant elsewhere referred to. Summing up the specific considerations governing the choice oftreatment, the author says:
Incidental matters that influenced the choice of treatment scheme included the realization that preliminary flotation would have involved two separate treatment circuits with additional steps of thickening and filtration following the flotation. Furthermore, in the conditioning method evolved, as much as 60 per cent of the dissolved values in the cyanide tailings were precipitated and recovered.
There are, however, cases where flotation equipment was put in for the purpose of recovering the gold in a concentrate and rejecting the tailing only to find that the tailing was too valuable to waste and had finally to be cyanided before discarding.
It is generally true that cyanidation is capable of producing a tailing of lower gold content than flotation. At a price of $35 per ounce for gold this fact is of much greater importance than when gold was valued at $20.67 per ounce. The possible gold loss in the residue to be discarded will influence the choice of a method of treatment.
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A flotation plant is being erected at the Falcon mine, Rhodesia, to treat ore containing gold and copper. With the exception of the Mt. Morgan, the Etheridge, and the Great Fitzroy mines, Queensland, I have not heard of the flotation process being used successfully to treat ore containing an appreciable amount of gold. The Elmore thick oil process was installed at the Lake View Consols gold mine, Kalgoorlie, several years ago, but was not successful, as the ore was not suitable, and unsuccessful experiments were made by Minerals Separation, Ltd., on orefrom the Lancefield mine, Western Australia, which contains mispickel. The Elmore vacuum process was installed at the Cobar gold mines, New South Wales, and at the New Ravens- wood gold mines, Queensland. Both these mines contain copper in the form of sulphide, as well as gold, but the plants only ran a few weeks. I was informed that the plant at the former mine (where the ore contains about $8 gold and 1.5% copper) gave a fair recovery of copper, but left too much gold in the tailing or left enough copper in the tailing to prevent profitable cyanidation of the gold.
To return to the Mt. Morgan mine, the laboratory apparatus had a capacity of one pound of ore at a time, and the results now being obtained in the experimental mill approximate closely those obtained in the laboratory. The object of concentration was, of course, to obtain a concentrate containing as much gold, copper, and iron, and as little silica as possible, commensurate with a good extraction of the gold, because it was found that the less silica the concentrate contained the poorer was the extraction of gold. It costs 13 cents to flux one unit of silica, and it was necessary to steer a middle course. Experiments made with Sonstadt solution on ore from one part of the mine showed that clean quartz (after separation by specific gravity from all mineral) contained not less than $1.50 gold per ton. In practice, of course, it is impossible to float all the mineral and sink all the gangue.
The agitator in the laboratory plant was at first run at 1100 r.p.m., but was afterward reduced to 800. Tests were made with pulps of different proportions, each separate pulp being agitated for the same length of time, that is, 6 minutes, and it was found that there was not much difference, in the extraction of gold and copper, between a pulp containing three parts solution to one of ore, and a pulp containing seven parts solution to one of ore. A pulp of 1 to 1 was too thick and gave poor results. In practice, the thinner the pulp the smaller the capacity of the flotation machine. Tests were also made to ascertain the effect of agitating for different lengths of time. Two tests were made in the laboratory of which I have a note: one for 10 minutes and one for 15 minutes. The ore contained $6.50 gold and 2% copper; 12% of this sample would remain on a 60-mesh screen. The first one gave a concentrate containing $22.70, 9.4% copper, and 18% insoluble, with an extraction of 51% of the gold and 84.5% of copper. The second gave a concentrate containing $20.20 gold, 7.8% copper, and 27% insoluble, with an extraction of 64.5% of gold and 91.8% copper. The gold left in the tailing was probably in the gangue, as the extraction was poorer than usual. As a rule, the longer agitation and separation are continued, the more silicious the concentrate is. In practice, the length of treatment is regulated by the thickness of pulp and the number of boxes in the flotation machine. Tests made to ascertain to what degree fine crushing was necessary showed emphatically that the ore must all pass through a screen of 60 holes to thelinear inch if a good extraction is to be obtained, and that the finer it was crushed, at any rate down to 120-mesh, the better the extraction was. Tests showed that when using eucalyptus oil there was no advantage in using an acid solution, but that, on the other hand, slight acidity did no harm. Much of the copper pyrite in the ore readily floats on water without any previous agitation. On treating ore containing $25 gold direct by agitation and flotation, without amalgamating or concentrating on tables, it was proved that fine free gold can be floated by using eucalyptus oil.
A few years ago some experiments were made by crushing in ball-mills and concentrating on Wilfley tables, but they were not successful. Last year it was decided to make a thorough trial of the Minerals Separation process, and a small testing plant was erected in the laboratory. At the same time a full-sized experimental unit, capable of treating 300 to 400 tons per 24 hours, was erected in one of the abandoned chlorination plants. Both sets of experiments were carried out by the metallurgical staff of the Company. After they were finished, a representative of the Australian branch of Minerals Separation, Ltd., paid a visit to the mine and conducted a few tests, which confirmed the results obtained by the mine staff.
As mentioned in the Companys annual report, these flotation experiments were successful, the extraction being higher and the costs lower than expected. The company is now building the first unit of a plant to treat 1000 tons per 24 hours. The ore will be crushed by rock-breakers, Symons disc crushers, rolls, and tube-mills. It will then be concentrated on Wilfley tables, after which it will go through a second set of tube-mills, thence to the flotation machines. It is presumed that no royalty will be payable on the Wilfley concentrate. This concentrate will either be briquetted or sintered in a Dwight-Lloyd machine, and smelted in blast-furnaces along with the copper ore and ironstone and limestone fluxes. The Company has no reverberating furnaces.
Many oils were tested, and, generally speaking, it was found that only essential oils gave a coherent froth and good extraction, other oils like petroleum, oleic acid, and lubricating oils tending to form granules which sank. The. best results were obtained from eucalyptus, closely followed by Essential C and Pinus lam us vulgaris. Oleic acid, which was used for years at Broken Hill on zinc ore with hot solution, and gave good results when tried on this ore with neutral and acid solutions, gave an enormous froth and floated most of the silica. A mixture containing 95% of eucalyptus and only 5% of oleic acid gave a concentrate containing 47% silica, showing the power of the oleic to float silica. Experiments were afterward made with a mixture of oils, and one combination (known as Mt. Morgan mixture) was found to give a better extraction of both gold and copper than any of the individual oils, and at less expense. When the sample was all crushed to pass 80 mesh, an extraction of 80% of the gold and 90% of the copper could be obtained every time, with a concentrate containing about 25% insoluble, which can be reduced to 10% by re-treatment. Hot solutions and a solution containing 1% of common salt were found to be detrimental to good recoveries.
A test on a sample, crushed to pass a screen of 120 holes per linear inch, containing $37 gold and 4.8% copper, gave a recovery by flotation alone of 90% of the gold and 98.5% of the copper, but left $8 gold in the tailing. The concentrate carried 44% insoluble matter, which could be reduced by re-treatment. A different oil (eucalyptus) would have given a poorer recovery and a cleaner concentrate.
Tests made on ore containing $9 gold, 3.5% copper, and 45% insoluble, showed that after crushing to pass 60 mesh and treating by direct flotation, an extraction of 82% of the gold and 96% ofthe copper could be obtained, with a concentrate containing only 21% insoluble. No doubt with finer crushing even better recoveries would be had. These results leave tables and vanners far behind. It was found decidedly advantageous to re-use the solutions.
A Wilfley table was erected in the mill, some tests made, and the tailing treated by flotation in the laboratory. Sometimes these tailing samples were dried before flotation, and sometimes they were not. It was invariably found that a better extraction was obtained from those which had not been dried, as no matter how carefully the operation was conducted, some of the iron pyrite got sufficiently oxidized to resist flotation, and it carried some of the gold.
In some of the tests the crushed ore was concentrated by panning in the laboratory, and afterward subjected to flotation. In this case the water in the laboratory was used, which did not come from the same source as the water used in the mill. It was noticed that the longer the sample was allowed to remain in the water after panning, the worse the subsequent flotation was. For example, where flotation took place immediately after vanning, the residue assayed $2.60 gold and 0.30% copper, but where tailing from panning was allowed to remain under water for 6 hours before flotation, the residue assayed $3.10 gold and 0.67% copper. An analysis of this water was made, and this incident shows what might happen in a mill where the ore is in contact with bad water for some hours before reaching the flotation machine, such as the time it is going through rolls, Chilean mills, tube-mills, and classifiers, over tables and through thickening devices, and perhaps through secondary tube-mills. The water in question was neutral, both before and after coming in contact with the ore.
Some tests were made both in mill and laboratory in which air was drawn into the agitation boxes through pipes fixed vertically in the corner with the top open to the air and the bottom ending in a bent pipe terminating under the impeller of the agitator. No improvement was, however, noticeable.
Grading tests were conducted on crude ore and flotation products. They showed that as regards crude ore, after crushing either in mill or laboratory, the finest grade of concentrate or ore was the richest and the coarsest grade of tailing was richest, both in gold and copper. The fact that the finest grade of tailing was the poorest shows that this process will float the finest sulphides successfully.
In the experimental mill the ore is crushed in rock-breakers andKrupp dry-crushing ball-mills without drying. This plant was formerly used to crush oxidized ore for chlorination and, being on the spot, it was naturally utilized in preference to buying new machinery. The crushed ore drops into a bin at the bottom of which are two Challenge feeders. These deliver the ore into a launder where it is met by a stream of water which carries it direct to a six-compartment Minerals Separation machine. Each spindle is driven by a half-crossed belt, thus eliminating the noise and grease incidental to the old Broken Hill method of gearing. The machine is of the Hoover single-level type, by which one man can attend to all the flotation boxes. The concentrate was collected at first in circular wooden vats with filter-bottoms of cocoa matting, and later in shallow rectangular concrete tanks which formed part of the old chlorination works. The whole plant is extremely simple and requires very few men to run it. It has not been found practicable to use a screen finer than 35 mesh on the ball-mills. It is found that the gold, copper, and iron contents are greater in the concentrate overflowing from No. 1 box and that they gradually decrease until No. 6 is reached, while the silica content increases from 10% in the concentrate from No. 1 box to about 50% in that from No. 6. About 56 hp. is required to drive the agitators at 350 revolutions per minute.
As it is intended to use Wilfley tables in the new mill to assist in recovering the iron pyrite in the ore for fluxing and other purposes, two of these machines were placed in the experimental mill and some tests made to find out what results may be expected of them. Taking an average of several tests on ore from different parts of the mine, the grading of the table feed was as follows: 10% remained on 60 mesh, and 19% passed through 60 but remained on 120 mesh. It contained $4.50 gold, 1.8% copper, 9% iron, and 76% insoluble. The concentrate assayed $17 gold, 2.9% copper, 34% iron, and 18% insoluble; the recoveries were 33% of the gold, 13% of the copper, and 38% of the iron. No doubt, had the pulp been classified and the fine material passed over slime tables or vanners, better results would have been obtained, but the Company does not intend to use mechanical concentrators for the slime, preferring to rely on the flotation process, so it was not worth while experimenting with them.
During the flotation experiments with eucalyptus oils some tailing was produced which contained a fair amount of gold, and attempts were made to recover some of this by amalgamating and cyaniding.It was found that no extraction by amalgamation was possible, nor was any extraction by cyaniding possible without either roasting or finer grinding. On unroasted tailing assaying $3 gold and 0.44% copper, after crushing to pass 120 mesh, separating the slime, and leaching the sand for 9 days, an extraction of only 60c. per ton was obtained with aconsumption of 3.6 lb. of cyanide per ton. On a different tailing crushed to pass 80 mesh, which after slime was separated assayed $2.90 gold and 0.30% copper, an extraction of $1 was obtained in 5 days with a consumption of 2 lb. of cyanide.
Samples of slime were treated by agitation and washed by decantation, and gave slightly better extractions, but the consumption of cyanide went up to 6 or 7 lb. The strength of solution used in these tests was 0.10% KCN. It should perhaps be noted that all samples of flotation tailing had been dried before being tested by cyanidation.
Two samples of sand from tailing were roasted and treated by percolation. The value was $3. The roasting reduced the sulphur to 0.5%. Although the copper and iron were oxidized by roasting, the consumption of KCN was less than in treating the unroasted tailing, which was contrary to expectation. With three days treatment, the residue was reduced to $1 per ton, and about one-third pound of copper was dissolved from each ton of tailing by the cyanide. The consumption of cyanide was 1.4 lb. per ton, so that the extraction was higher and the loss of cyanide less than in treating unroasted tailing. Speaking from memory, I think that attempts to regenerate the cyanide in solution by means of sulphuric acid and lime were not very successful. The solution contained 0.05 gram copper per litre.
These cyaniding tests were merely done for information, as it is not expected that the tailing from the new mill will be profitable for cyaniding. The subject of extracting gold from flotation tailing arose a few years ago at the Cobar gold mines, as already mentioned, but in that case the difficulty was overcome by selling the mine, which contained highly silicious ore, to a company which owned a smelter, and had, or thought it had, plenty of basic ore for flux. Unfortunately, the amount had been overestimated and the problem is still unsolvedbut that is another story.
The Froth Flotation Method is means separating minerals according to their different physical and chemical properties. According to classification, the flotability of gold and silver minerals is included in the first category of natural and non-ferrous heavy metal sulfides, characterized by low surface wettability and easy flotation, which can be flotation by xanthate collectors.
The froth flotation method is widely used to treat various veins of gold and silver ores for the following reasons: (1) In most cases, the froth flotation process can enrich gold and silver in sulfide concentrate to the greatest extent and discard a large number of tailings, thus reducing the smelting cost. (2) When the flotation machine is used to treat polymetallic gold and silver ores, concentrates containing gold, silver and non-ferrous heavy metals can be effectively separated, which is conducive to the comprehensive utilization of valuable mineral resources. (3) For refractory gold and silver ores which cannot be treated directly by mercury amalgamation or cyanidation, a combined process including flotation is needed. However, there are some limitations in flotation, such as ores with gold particles larger than 0.2-0.3 mm or pure quartz gold ores without metal sulfides, which are difficult to deal with by flotation separation alone.
The crushing and screening stage in the industry is mainly composed of three-stage and a closed-circuit process. Gold ores need to go through coarse, medium, and fine crushing processes to be minimized into smaller pieces. The screening equipment is used to sieving the smaller gold ores into the proper size for the next steps.
The grinding operation usually adopts one or two ball mills with types of lattice and overflow. The second stage grinding operation forms a closed circuit with a spiral classifier or a hydro cyclone to ensure the grinding fineness.
Since traditional ball milling equipment appears some shortcomings such as fast wear and large energy consumption, many manufacturers adopt new wear-resisting rubber lining boards, sliding bearing to improve a mill operation efficiency and prolong a machine's service life.
The beneficiation stage is a crucial part of gold extraction during the whole gold ore processing plant. Placer gold mine and rock gold mine are most widely processed to extract gold concentration.
The gold slurry process of the carbon slurry method (CIP and CIL) is to put activated carbon into cyanide ore slurry, adsorb dissolved gold on activated carbon, and finally to extract gold from activated carbon.
Equipment required for carbon slurry gold mining process: Leaching mixing tank, activated carbon screen, Two-layer (three-layer) washing and thickening machine, fast desorption electrolysis system with high-efficiency and low-consumption, high-frequency dewatering screen.
It means that by ion exchange resin, gold also can be extracted from ore pulp. Like carbon, the process makes gold absorbed onto solid spherical polystyrene resin beads instead of activated carbon grains.
According to different physical and chemical properties of different types of gold ores, flotation separation utilizes various reagents to make the gold attached to the bubbles then scraping these gold particles from blades to get the concentrate.
A jigger is one of the main pieces of equipment in the gravity separation process. The jigging process mixes gold ore particles of different specific gravity together, then stratifying these particles. The minerals with small specific gravity will be on the upper layer while the minerals with large specific gravity will be on the lower layer.
A shaking table is used to process gold ores in the horizontal medium flow. The motor drives the surface of the shaker to perform the longitudinal reciprocating motion, as well as the differential motion of the washing stream and the surface of the bed. Gold ore particles are stratified perpendicular to the surface of the bed, then being separated parallel to the surface of the bed in reciprocating motion which allows gold ores with different particle sizes to be discharged from different parts to achieve separation.
It adopts lope water flow to achieve separation. With the effect of the combined force of water flow, mineral gravity, the friction created by the bottom of the tank, and ore particles, the gold ore particles will settle in different areas of the tank. The ore particles with small specific gravity will flow away with the water, while ore particles with larger specific gravity would stay.
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Henan Fote Heavy Machinery Co., Ltd. (FTM) has more than 40-year experience in the design of gold mining equipment processes. Its beneficiation equipment and plants sales to many countries including Tanzania, India, South Africa, the United Kingdom and other regions. According to the actual needs of customers, all machines can be customized here.
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Flotation is the most widely used beneficiation technique in the minerals industry. The kinetics of flotation is governed by the rates of bubble-particle attachment and bubble-pulp separation. Both processes interact significantly in a conventional flotation unit operation, making it virtually impossible to optimize one of the processes without sacrificing the effectiveness of the other. In most cases, the bubble-particle attachment is the slowest step in the flotation process, often requiring 5 to 15 min residence time to complete the mineral flotation. Ores processed in the future will be of lower grade and finer grained. As a result, the bubble-particle attachment will decrease significantly. Larger flotation circuits with much longer residence times will be required for the same production capacity. Therefore, a faster flotation process will be needed to efficiently process larger tonnages at smaller particle sizes.
Recent hydrodynamic flotation research by the Bureau of Mines has shown the importance of turbulent agitation on the recovery of fine particles. This research showed that the agitation caused by microscopic eddies contained within the turbulent fluid flow increased the particle-bubble collisions, produced faster flotation kinetics, and resulted in higher recovery of the fine particles. This research also showed that the most effective agitation within a conventional flotation cell occurred in only a small portion of the cell volume. Consequently, to optimize flotation kinetics, the agitation energy needs to be concentrated on the pulp only long enough to attach the particle to the bubble.
As a part of the Bureau of Mines effort to advance the technology of flotation beneficiation, an improved flotation system was developed made up of discrete unit operations for bubble-particle attachment and bubble-pulp separation. This system features an in-line static mixer for bubble-particle attachment and a shallow-depth unit for bubble-pulp separation. Bubbles of approximately 100 m in size were generated externally using a spinning disk bubble generator. Sebba, showed that at very high rotational speed, a spinning disk shears air and water and forms fine-sized bubbles. Both the ore pulp and the bubble slurry were combined and vigorously mixed as they passed through the static mixer. After passing through the static mixer, the froth concentrate was recovered in the bubble-pulp separator.
The in-line static mixer flotation system was tested with a western porphyry copper ore to demonstrate flotation rate improvements on a common sulfide copper ore containing chalcopyrite. This paper presents the design parameters of the in-line static mixer flotation system and its application for recovering copper from a western porphyry copper ore.
A semi-continuous flotation system was designed to agitate the bubble and mineral pulps as they flowed through an in-line static mixer and to continuously remove the mineral laden bubbles from the ore pulp. A diagram of the in-line static mixer flotation system is shown in figure 1. The bubbles are externally generated by a spinning disk generator. A dual head peristaltic pump was used to meter both the bubble slurry and ore pulp and to maintain a constant ratio of bubble-slurry to ore slurry over the range of feed rates tested. Both streams were combined in a 0.44 cm I.D., T-fitting positioned immediately ahead of the static mixer.
Several static mixers were tested. The basic static mixer was a 0.44 cm diam, stainless steel pipe with a spiral mixer inside. The spiral mixer protruded 0.2 cm inward from the pipe wall at about 0.05 cm width. The spiral completed one revolution every 0.5 cm throughout the length of the pipe. The agitation energy dissipated in the static mixer was equal to the headloss:
where hL is the fluid energy headloss, f is the friction factor, l is the length of the mixer, d is the diam of the mixer, V is the flow velocity, and g is the gravitational constant. In the static mixer the flow resistance was caused by the drag of the spiral mixer as it projects into the flow stream within the pipe. The friction factor of the flow through the static mixer was determined by balancing the fluid energy headloss through the static mixer system. Using water, the friction factor of the static mixer was 0.038. Plotting this friction factor and the expected Reynolds number for the flow in the static mixer on a Stanton diagram indicates that the fluid flow within this static mixer is characterized as transition turbulent flow for all flow rates tested. This type of fluid flow generates dissipation eddies within the fluid that intimately mix the bubbles and particles as they flow through the static mixer.
mixer was colled 1.5 revolutions and the 84 and 204 cm long mixers were coiled 3.5 and 8.5 revolutions, respectively. Smaller diameter static mixers (0.22 cm diam) were constructed, but plugged repeatedly in subsequent testing. Larger (0.88 diam) mixers were considered, but the expected Reynolds number and relative roughness were low.
After passing though the static mixer, the bubble-pulp slurry entered a 20-cm-diam tank, 44 cm deep, where the froth concentrate was recovered. For rapid bubble separation from the pulp only a shallow depth was required. The mixture entered the separator through a central sparger that distributed the ore slurry radially throughout the tank. The sparger was positioned only 0.5 cm from the top of the tank. The bubbles rose quickly to the top where they overflowed at the outer edge of the separator. The effective volume of the bubble separator (315 cm) was considered to be the height above the sparger plus an equal distance below the sparger. The remainder of the tank volume was used to collect the tailings product. An overflow weir pipe was used to discharge the tailings pulp beneath the effective volume of the separator. The froth height could be maintained by adjusting the height of this overflow weir pipe.
A sample of copper ore containing 0.67 pct Cu was obtained from a porphyry copper mine in Arizona. The major sulfide mineral was chalcopyrite with smaller amounts of pyrite. The gangue was mostly plagioclase feldspar with smaller amounts of biotite and magnetite. The chalcopyrite was widely disseminated throughout the ore pieces and the liberation size was around 210 m.
For each flotation test, the ore sample was ground with clear saturated lime water for 10 min in a laboratory rod mill at 70 pct solids and screened at 210 m size. The plus 210 m size fraction was reground for an additional 8 min. The ground ore pulp was diluted to 50 pct solids for flotation with clear saturated lime water (pH 12.5). The ore sample was conditioned with 0.1 g/kg potassium amyl xanthate for 5 min in a stirred tank. Then the conditioned ore suspension was pumped through the in-line static mixer, along with the bubble slurry, to the separator. Saturated lime water was used as the process water to maintain the pH. To stabilize the bubbles, Dowfroth 1012, at a dosage of 25 ppm was used in the process water. Reagent dosages in excess of those required to float the copper were used so that the resulting flotation kinetics would indicate the effectiveness of the flotation system. This scheme, however, did allow the flotation of some gangue materials which lowered some concentrate grades.
After pumping the conditioned pulp through the flotation system, the tailings product was repumped through the system three more times to simulate the effect of flotation staging. The froth concentrate from each of the four stages and the final tailings were weighed and analyzed for copper. The copper distribution after each of the four passes through the system was used to determine the flotation rate constant. The pressure drop across the static mixer system was measured to calculate the headloss and energy consumption of the system. The energy required to generate the bubbles was not included in this calculation.
Conventional laboratory batch flotation tests were conducted with a 500-g DR flotation cell on the Cu ore under the same reagent and pH conditions for comparison with the in-line static mixer flotation tests. Samples of the flotation concentrate were taken after 2 and 5 min to establish the flotation kinetics of the conventional laboratory flotation cell. Torque measurements were made on the impeller shaft to determine the energy requirements for the conventional laboratory cell.
The in-line flotation system was tested at pulp flow rates of 350, 600, and 1320 mL/min. The dual head peristaltic pump maintained the air to ore ratio at 0.44 mL/g, a value well in excess of the theoretical requirement. The bubble slurry contained 35 pct air by volume and the ore pulp was at 50 pct solids by weight. As shown in table 1, the combined flow rates of the
ore pulp and bubble slurry were 670, 1145, and 2520 mL/min. After combining both streams, the in-line static mixer contained 17 vol pct air bubbles, 69 vol pct water, and 14 vol pct ore. This is equivalent to a 35 wt pct solids ore slurry with bubbles.
where Rt is the Cu recovery at time t, R is the Cu recovery at t=, and k is the flotation rate constant. Neither the conventional nor In-line static mixer flotation results exactly fit first order kinetics, however, in order to use a simple basis for comparison, the time required to obtain 90 pct recovery, t90, was determined graphically from the flotation data. The first order flotation rate, k90, was calculated as follows:
While the k90 is not a perfect fit to the data, it allows direct comparison of the flotation results from different flotation tests in an easily understandable fashion. For the conventional cell, 90 pct recovery was obtained after 2.2 min with a flotation rate (k90) of 1.0 min-.
As shown in table 2, the best results were obtained at a pulp flow rate of 1320 mL/min with the 36 cm long static mixer. In four stages through the flotation system, over 99 pct of the copper was recovered in a 4.5-pct Cu concentrate at a flotation rate 7 times faster than conventional flotation. The flotation kinetics at each flow rate, shown as recovery versus time, are given in figure 2. The flotation kinetics at lower flow rates were both slower and less effective than those at the high flow rate. Flow rates above 1320 mL ore pulp/min were not tested, because that was maximum capacity of the present pumping system.
In figure 3, the energy consumption versus Cu recovery shows that, at the lower flow rates, the energy was more efficiently used co recover the Cu. At the 350 mL/min ore pulp feed rate, the first stage flotation recovered 40 pct of the Cu using only 0.0025 kWh/mt. At 1320 mL/min ore pulp feed rate, the first stage flotation used 20 times more energy at 0.049 kWh/mt to recover 53 pct of the Cu. Each data point in figure 3 represents a separate flotation stage. After four flotation stages at 350 mL/min ore pulp feed rate, 74 pct of the Cu was recovered. But at 1320 mL/min ore pulp feedrate, 82 pct of the Cu was recovered in just two flotation stages. Half as many flotation stages were required at the high feed rate to recover more Cu than at the lower feed rate. At 90 pct Cu recovery, figure 4 shows the relationship between the energy consumption and the flotation rates. The high ore pulp feed rates produced the fastest flotation, but at the highest energy consumption. Improved flotation kinetics were obtained only at the expense of higher energy consumption. In general, test results show that recovery was related to the number of bubble-particle collisions and the number of bubble-particle collisions was related to agitation intensity.
A plot of the Reynolds number and friction factor on the Stanton diagram showed transitional turbulent flow characteristics within tho mixer. The headloss through the static mixer system is due to the dissipation of the fluid energy in small eddies that mix the chalcopyrite and air bubbles. Based upon the calculated headloss through the mixer, the calculated mixing intensity in the static mixer at the high feed rate was 0.12 W/cm. At tho 350 and 600 mL ore pulp/rain feed rates, the mixing Intensity was only 0.002 and 0.014 W/cm, respectively. This lower mixing intensity resulted in fewer bubble-particle collisions and slower flotation kinetics. However, at 0.12 W/cm with the 1,320 mL ore pulp/min feed rate, the mixing intensity was sufficient for rapid attachment of the bubbles and chalcopyrite particles.
The effect of the static mixer length is shown in figure 5. Tests were conducted at 1,320 mL ore pulp/min without a static mixer (length = 0) and at each of the three mixer lengths. Without a static mixer some flotation did occur, but it was not very effective. The calculated mixing intensity in each mixer was the same, but the different lengths resulted in longer exposure to the mixing agitation. Both of the longer static mixers showed a significant decline in the flotation response. The froths recovered during the two long static mixer tests contained large slugs of air indicating that coalescence of the bubbles was occurring in those tests. Possibly, as the bubbles coalesced some of the chalcopyrite particles were detached resulting in lower Cu recovery. Based upon the flow rates, the residence time in the 36 cm long, static mixer was only 0.12 s and there were no signs of bubble coalescence. The residence times for the 84 and 204 cm long, static mixers were 0.28 and 0.69 s, respectively, and both of those static mixers did show signs of bubble coalescence.
Although 99 pct of the Cu was recovered in four stages, after three stages, the in-line unit had recovered over 94 pct of the copper in a 5.7 pct copper combined concentrate. A still higher grade combined concentrate was obtained with just two stages which produced a 8.7 pct Cu, combined concentrate and recovered 82 pct of the copper, as shown in table 3. The first two
flotation stages recovered 82 pct of the copper in a high grade 8.7 pct concentrate and the third and fourth stages recovered the remaining 17 pct of the copper in a low grade, 1.4 pct concentrate. This low grade concentrate is best suited as a scavenger concentrate and should be recirculated to the first flotation stage. This demonstrates the flexibility afforded by multiple staging of the in-line flotation system to optimize the flotation circuit for recovery and grade.
Finally, a comparison of the in-line static mixer flotation system and the conventional laboratory flotation cell is shown in figure 6 and table 4. The in-line static mixer flotation system recovered more copper, faster, and at lower energy consumption than the conventional laboratory flotation cell.
In-line static mixer agitation effectively mixed the air, water, and ore to quickly attach tho chalcopyrite particles to the surface of the air bubbles. As the mixture entered the shallow-depth separator, the mineral-laden bubbles were quickly floated away from the ore pulp. The in-line static mixer flotation system
was effective for copper flotation at a flotation rate 7 times faster than conventional laboratory flotation, recovering 99 pct of the copper ore in a rougher concentrate containing 4.5 pct copper. This rougher Cu concentrate, which represents the combined concentrate from four flotation stages, could be increased to commercial grade by recirculating the lower grade concentrates to previous circuits for cleaning and upgrading. Also, the in-line flotation system has a higher capacity per unit volume of flotation cell than conventional laboratory flotation. Energy requirements for the in-line static mixer flotation were also lower than the conventional flotation cell, but only a laboratory scale system has been tested. Additional testing is needed with a larger system to more accurately measure the power consumption of the process.
DOVE Diamonds and Gold mining equipment are configured for different ore type, laterite, heavy clay, gravel and black sand. Wash plants are designed for highest recovery and minimum operator requirements.
DOVE is a major manufacturer of hard rock gold mining equipment, and hard rock mining equipment, and crushing plants for base metals, ferrous metals and light metals, producing Ball Mills, Jaw Crushers, Cone Crushers, Magnetic Separators, Shaking Tables, Gold Concentrators, Rotary Dryers, and Flotation Process.
Each processing plant is designed tailor made according to the ore characteristics and the mineral composition of the ore, and designed for the 100% recovery of gold and other metals production, with no loss.
DESERTMINER is a Dry Mineral Processing Plant developed by DOVE to simultaneously concentrate, separate and recover gold, platinum group metal, base metals, ferrous metals from Alluvial deposits, as well as Hard Rock deposits, that are located in dry areas, where water not available.
DESERTMINER is designed and configured to deliver a high efficiency in mineral processing, with 100% recovery of gold and other metals and minerals, with no loss, similarly to the wet mineral processing plants.
DOVE designs and manufactures high recovery Beneficiation Plants. Beneficiation plants are composed of different types of machines and separators depending on the metals and minerals composition and characteristics, designed for high economic recovery in a customized configuration that will meet the projects specifications. These plants can include Flotation Machines, Dryers, High Intensity Magnetic Separators, High Tension Separators, etc. They can be used for both hard rock (primary) and alluvial (secondary) deposits.
DOVE supplies the state of art inGold Refinery technology, which is designed to produce and refine raw gold production to international standard purity of 999.95. DOVE refining units are supplied in capacity ranges of 6 kg up to 150 kg. The units are based on advanced Aqua Regia process and it is designed to refine gold, silver, copper. The refining units are configured with the latest technology to comply with the highest environmental standards, which includes neutralization towers for fumes, etc.
Every mining operation requires sophisticated gold room in order to ensure the highest recovery of gold production. To this end DOVE supplies and manufactures a complete range of equipment, instruments, tools and accessories, which includes Gold Concentrating Table, Gold Centrifugal Concentrator, melting furnace, crucibles, ingot molds, scales, and assay instruments.
DOVE supplies Flame Atomic Absorption Spectrometer (AAS) for Fire Assay Analysis, Portable Mineral Analyzer and Professional Lab Mineral Analyzer for XRF Test, and equipment for Gravity Separation Test such as Lab Jig Concentrator and Lab Concentrating Tables. Several other types of equipment including Lab Induced Roll Lift Magnetic Separator, Lab Isodynamic Magnetic Separator, Lab High Tension Separator, Sampling Pulverizer, Sieve & Shakers, Bench Drill Machine and High Voltage Rectifier are part of the range of DOVE laboratory equipment.
DOVEsupplies advanced and highly accurate range ofMetal Detectors. It is designed for ease of operation in highly mineralized soil condition. It provides a practical solution during exploration and prospecting different terrain with ease and dependability. DOVE Metal Detectors are designed for various depths and conditions.
DOVE provides a complete range of minerals assay testing services, for both alluvial (placer) and hard rock (primary) deposits. DOVE mineral assay services include Gravity Separation, Fire Assay Analysis and XRF Analysis for identification of gold, platinum, other metals and minerals concentration and simultaneously analyze up to 32 other elements, metals and minerals.
Reliable assay and minerals testing of your mine samples can lead you to the most efficient and best equipment configuration and plant design for the highest recovery of production and return of investment.
Driven by a global team of process engineers and metallurgical specialists, Multotec designs, builds, manufactures, installs and maintains equipment throughout the entire value chain of mineral processing plants across all commodity sectors, from diamonds to coal, gold, iron ore, platinum and phosphates.
Today, Multotec mineral processing equipment is used in over 100 countries on 6 continents, and by the worlds leading mining houses such as Glencore Xstrata, Anglo Coal, BHP Billiton, OceanaGold, QM and Rio Tinto.
Multotec has consistently grown its international footprint in order to serve its customers with greater flexibility, agility and technological support. With operations in almost 90 countries on 6 continents, and a high focus on knowledge sharing and strategic global research and development, Multotec is a world of mineral processing knowledge.
Through our strategically located network of sales and service branches, we provide this knowledge to our customers, wherever their operations may be. By developing local capacity including both skills and infrastructure as close to our customers plants as possible we ensure a quick and effective response to your challenges, with leading metallurgical expertise on your doorstep.
With our rapidly growing support network and proven range of products, Multotec is increasingly assisting customers with operations contracts to take over the maintenance of plant equipment, aligned to service level agreements.
Backed by a world-leading range of specialised mineral processing equipment, Multotec provides complete life-of-plant technical services aimed at increasing metallurgical efficiency to optimise your plant throughput.
We assist our customers not just in meeting their demands for products and equipment, but by optimising the life and efficiency of these products to optimise the entire supply chain for our customers. Through holistic plant evaluation, sampling and testing, optimising flow sheets and looking in detail at any plant-wide processing problems, we are able to offer solutions that offer a direct improvement to the process, with significant tangible impacts on your bottom line.
Reducing downtime saves your plant vast amounts of money. Multotec understands this, and, through our flexible and agile global network, ensures we respond to your requirements with maximum speed and efficiency. In most locations, we can respond to customer requests in under 4 hours.
Some of the challenges minerals processing plants face include the high cost of replacing capital equipment, the labour requirements in changing out heavy equipment, such as a DMS cyclone, and production downtime while staff have to comply with safety regulations while equipment is being replaced. Multotec strives to be a plug-and-play solutions provider, providing maintenance and delivery of high-quality equipment through your local branch.
Our investments in R & D are geared towards 4 key optimisation areas: the need for high recovery and yields, for high levels of product consistency and reliability, for lower lifecycle costs and the ability to develop new products on behalf of our customers.
We prioritise skills transfer and capacity-building across our worldwide branch network, and also train our customers staff in the maintenance of our equipment so that they can help ensure maximum efficiency of their plant.
Multotec is able to provide equipment and services to ensure compliance with safety, health and environmental management standards. This includes the supply of equipment, technical expertise and maintenance.
We help the mining industry improve the efficiency of processing plant performance through our approach to training and education. Multotec has invested in several pilot plants and testing equipment that are used for experiential training, as well as in conducting research and development projects, with the results shared among the relevant stakeholders in the industry for continuous product improvements.
Leading metallurgists and engineers from Multotec deliver training courses for mine managers, plant managers, process equipment operators, metallurgists, project house engineers, original equipment manufacturers and chemical engineers.
Multotecs education and training initiatives include a What is Happening in Industry forum, which brings together industry experts and academics to share best practice, knowledge, ideas and the latest trends in technology.
The forum helps ensure that all industry stakeholders and minerals processing faculties at universities keep abreast of changes in terms of technology to develop relevant and updated curricula, and to facilitate participation in research and development programmes.
Majola has over 20 years of experience in mining and extractive metallurgy. He has held a wide range of positions in the mineral processing industry, including Production Metallurgist, Senior Plant Metallurgist and in sales. Bheka has been with Multotec since 2013.