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By Jaime Lomelin

The brines from the sub-surface of the Laguna del Rey dry lake are pumped to Química del Rey, S. A. plant, where they are cooled in adiabatic crystallizers precipitating sodium sulphate decahydrate. This salt is dehydrated in multiple effect evaporators and dried in rotary kilns producing high purity anhydrous sodium sulphate. The spent brines from the sodium sulphate operation are concentrated in solar ponds, precipitating astrakanite and leaving a brine with a higher concentration of magnesium ions. This brine is treated with calcium chloride solution to eliminate the remainder of sulphates, leaving a solution of magnesium chloride ready for the manufacture of refractory grade magnesium oxide or periclase. The magnesium chloride is reacted with dolime to produce magnesium hidroxide which is washed, filtered, and calcined to produce soft burned magnesium oxide, which is briquetted and dead burned to produce the periclase.

Jan 1, 1984

By Kal V. S. Sastry

Modeling and simulation are two major tools available to a mineral process engineer for a consistent and systematic process analysis. The application of these tools to the analysis of specific mineral processing unit operations is well documented in the literature; nevertheless, they are used only to a limited extent in the industry. This is because the fundamental principles of modeling and simulation in the context of mineral processing context are not available. This paper represents the first in a series of papers planned to meet this need. General principles of process modeling and a unified mineral process model development methodology are presented. A simple example of model development using the flotation unit operation is given. A discussion of the practical significance, challenges and opportunities associated with mineral process modeling is included.

Jan 1, 1990

By D. J. Boron, R. Kollrack

Introduction The US has the largest total (364 Gt, or 401 billion st) and recoverable (182 Gt, or 201 billion st) coal reserves in the world (Steam/Its Generation and Use, 1978). However, progressive coal use in this country has been hampered by environmental constraints, such as restrictions on sulfur dioxide, nitrogen oxides, and particulate emissions. Since the promulgation of the Environmental Protection Agency (EPA) New Source Performance Standards (NSPS), it has become apparent that all US coals will require sulfur removal to comply with the new sulfur dioxide emission levels. If more stringent emission standards are imposed, through pending acid rain legislation or amendments to the Clean Air Act, it is likely that the coal industry, and the industrial and utility sectors will look increasingly toward the development of new and advanced coal cleaning technologies to continue and promote the use of coal. Table 1 summarizes the EPA NSPS for sulfur dioxide emissions. Fortunately, the basis for determining sulfur dioxide removal is the Btu and total sulfur content of the run-of-mine coal. Thus, credit is given for any sulfur removed before combustion, e.g., through physical coal cleaning. Unfortunately, physical coal cleaning alone may be unable to produce a compliance coal with respect to the EPA NSPS. Therefore, all industrial and utility coal-fired boilers will require post combustion gas cleanup, e.g., flue gas desulfurization (FGD), to comply with the new regulations. Table 1 - EPA NSPS for SO2 Emission Category SO2 Content of ROM Coal SO2 Reduction Required 1 0-2 Ib/M Btu 70 2 2-6 Ib/M Btu To 0.6 Ib/M Btu (70% to 90%) 3 6-12 Ib/M Btu 90% 4 12 Ib/M Btu To 1.2 Ib/M Btu Alternatives to post combustion gas cleanup, such as staged combustion, lime and coal injection, fluidized-bed combustion, and chemical coal cleaning, are currently under development. Of these, chemical coal cleaning offers the advantages of providing (1) high quality fuel for combustion (low sulfur and ash, high Btu, and attenuated product variability); (2) a solid fuel product for use in retrofitted gas and oil boilers; and (3) a solid fuel product with multiple end uses, e.g., internal combustion engines, carbon electrodes, coke, or chemical feedstock. Gravimelt process In the Gravimelt process, 1.4 mm x 0 (14 mesh x 0), physically beneficiated coal of about 5% moisture is fed into a 3401 to 3901 C (644 ° to 734 ° F) molten bath of sodium hydroxide and potassium hydroxide. Organic and pyritic sulfur and most of the mineral matter in coal react with and are dissolved in a molten mixture of potassium hydroxide and sodium hydroxide. The reacted coal sulfur is converted to alkali sulfides and polysulfides, whereas the mineral matter is apparently converted to alkali-aluminum silicates and other byproducts that are water and acid soluble. Coal residence time is 60 to 240 minutes. The sodium hydroxide to potassium hydroxide ratio is about 9:1. The total hydroxide-to-coal ratio is no less than 10:1. The specific gravity of the caustic melt is about 1.8 g/cm3 . Thus, coal is buoyant in the bath, and insoluble mineral matter with density greater than 1.8 g/cm3 would sink. Reacted coal is first water-washed with an equal weight volume of water in a countercurrent mode of operation. The coal is filtered and the filtrate sent for regeneration. The coal is then acid-washed in 10% sulfuric acid. The acid-water/coal mixture is filtered to recover the final coal product, and the filtrate treated with lime to precipitate dissolved mineral matter. The precipitate is filtered from the mixture and discarded. The water is recycled to the plant. Filtrate from water-washing is treated to remove sodium sulfide. The sodium sulfide is chemically treated to produce hydrogen sulfide, calcium car¬bonate, sodium hydroxide, and other products. The hydrogen sulfide is treated to produce sulfuric acid for the acid-washing step. The calcium carbonate is treated further or discarded, and the sodium hydroxide is concentrated by evaporation and recycled to the reactor. The Gravimelt process is currently at the bench-scale stage of development. A 0.9-kg (2-lb) continuous reactor is in operation with a continuous water- and acid-washing section adjacent. The regeneration section is not developed to a meaningful extent, although key reactions have been tested in the laboratory. Process conditions, product recovery efficiencies, etc., however, have not been determined adequately and continuous regeneration operation has not been attempted. Work is now in progress to develop each section to a continuous bench-scale operation in processing coal at 9.1 kg/h (20 lb per hr). Microwave process The carbon and hydrogen in coal are relatively transparent to microwave radiation. However, water, caustic, pyrite, and other mineral matter components are strong conductors of microwave radiation energy at the proper frequency. As a result, selective heating of the noncarbonaceous materials occurs that enables the chemical reaction of sulfur and ash with caustic. Although very high temperatures may result from

Jan 1, 1987

By R. D. Rein

The Henderson mine is a large continuous panel caving operation in central Colorado. The orebody is roughly 760 m by 910 m and will be extracted from two production levels. Each level is divided into three north - south panels 244 m to 317 m wide. The general arrangement of a production level and its associated undercut drifts is shown in Figure 1. The production drifts (3.6 m by 3.7 m) are spaced 24 m apart center to center and have undercut drifts (3.4 m by 3.8 m) 17 m directly above them. The production drifts are interconnected with drawpoint crosscuts (4 m by 4.3 m) driven on 12 m centers. Drawpoint crosscuts receive substantial brow steel and concrete support prior to undercutting. The brow steel is a 15 cm by 46 cm wide flange beam with a 2.5 cm wear plate to box in the 'cave' side of the beam. The concrete is a minimum of 0.3 m thick and the batch design results in 27 500 kPa strength. The finished drawpoint opening is 3.7 m by 3.2 m with a choke area of 3.7 m by 2.4 m. (Figure la) Undercutting and bell formation is accomplished by blasting rings of 7.6 cm holes drilled from the undercut drifts. The rings have a burden of 2.0 m and six rings develop one complete bell.

Jan 1, 1985

By R. S. Mateer

Linear programming on desktop computers is commonly used to optimize coal blending proportions. The objective function is usually the cost of the blend. The constraints consist of materials and heat balances as expressed as linear functions of the blend. This is illustrated in Example 1. Inclusion of a sulfur compliance constraint is more complex because the maximum permitted sulfur percentage is not a pre-specified constant. Rather, it is a function of the calorific power of the blend itself. The current limit of 1.2 pounds of sulfur dioxide per million Btu may be expressed as: Sulfur Max % = 0.6 • H/10,000 with H in Btu/lb.

Jan 1, 1994

By W. Holding

Since the 1975 International Mine Ventilation Congress in Johannesburg there has been a trend in South Africa towards making the mine ventilation departments responsible for monitoring the total environment. This includes the taking of noise and illumination measurements and there is in fact a debate taking place as to whether the departmental title should be changed to 'Environmental Control Department'. The organisers of the 2nd international Mine Ventilation Congress have decided that noise and illumination will not be included as Congress subjects and 'Total Environment' includes only those facets which are open to debate at the Congress. The choice of mine is significant in that while most Congress papers from South Africa will be devoted either to coal mining or to the somewhat unique deep level gold mines, Prieska Copper Mine is a mine which can easily be identified with similar mines in many other parts of the world. The mine is situated some 750 km south west of Johannesburg and about 600 km north east of Cape Town. Current production (ore and waste) is about 340 000 tons per month and the total underground labour force is just over 2 000. The ore body is a massive sulphide, the main minerals being copper and zinc and the method of mining is sub-level open stoping. Daily average surface temperature conditions range from 85°F (29.5°C) dry bulb and 69.8°F (21°C) wet bulb in mid-summer to 42°F (5.5°C) dry bulb and 36.5°F (2.5°C) wet bulb in mid winter. The information presented in-this paper is limited to what can be gathered by mine ventilation staff as part of an on-going exercise and is obviously not as comprehensive as would be the case if a fully equipped research establishment were to become involved. The paper will be divided into three sections 1) Gases 2) Dust 3) Heat In each section statutory and/or other standards will be set down followed by a discussion of assessment and in some cases control methods.

Jan 1, 1980

By Carl Erickson, James Hislop

INTRODUCTION As practiced today, longwall mining is a mechanical version of an established mining system. The equipment for longwall mining consists of four basic units: roof supports, face conveyor, winning or mineral-getting ma¬chine, and stage-loader conveyor. The system mines relatively thin slices of mineral across a predetermined face length in a virtually continuous sequence. As the winning machine cuts, the mineral is loaded automatically onto the face conveyor. In turn, the face conveyor is pushed over into the newly cut track by hydraulic rams attached to the conveyor and the roof supports. The roof supports then are lowered individu¬ally from the roof, advanced to the repositioned con¬veyor, and reset to the roof. At the end of the pass, the longwall cutter is pushed over, and a slice is taken in the opposite direction, so the system repeats itself. Fig. I illustrates a typical longwall mining system. The principle behind the longwall mining system is that the roof is supported only immediately adjacent to the worked seam, keeping the blocks of immediate roof clamped securely from the face to the rear of the sup¬ports. At the rear of the supports, the roof is allowed to fall and break up, with the swell of the broken material eventually establishing an equilibrium in the strata be¬hind the mined area. Fig. 2 illustrates the principle of longwall support. There are two distinct types of longwall mining. The longwall retreating system is the most common in the United States, South Africa, and Australia. The longwall advancing system is the most common in the re¬maining industrialized countries such as Britain, Ger¬many, Poland, and the Soviet Union. The longwall retreating system comprises a mining panel, where the roadways that form the head entry (end at which mineral is delivered from the face con¬veyor to a stage-loader conveyor) and the tail entry are predriven before the longwall units starts production. In the longwall advancing system, the longwall unit makes the head and tail entries as the face advances, utilizing various techniques to prevent the entries from being destroyed as they are left in a consolidating gob area. Fig. 3 illustrates the comparative approaches of the longwall retreating and advancing operations. As with most other underground mining systems, there are many variations on these two basic systems, using combinations of the systems that are dictated by

Jan 1, 1982

By P. E. Sperry, R. E. Heuer

The United States Bureau of Reclamation is constructing the $200 million Navajo Indian Irrigation Project to irrigate 110,000 acres of land on the Navajo Indian Reservation in northwestern New Mexico. Water from the San Juan River is impounded by Navajo Dam and will be conducted for 26 miles along a series of tunnels and canals to the area to be irrigated. Tunnel No. 3 is the third in a series of six tunnels along the route. Bids for construction of Tunnel No. 3 were received by the Bureau of Reclamation in September, 1970. Fluor Utah was low bidder at $8.7 million. Ten of the 12 bids were based on excavation by tunnel boring machine. The consulting firm of A. A. Mathews, Inc., was retained by the contractor to advise on geologic and rock mechanics conditions after support problems developed in the tunnel. Tunnel No. 3 is 15,264 feet long, has an 18 foot lined diameter, rises approximately one-half inch per 100 feet in the direction being excavated, and is 6,000 feet above sea level. There is one short curve in the alignment. Minimum concrete thickness is 12 inches with eight inches from 'A' to 'B' line making a pay diameter of 21 feet, 4 inches. The tunnel is supported with rock bolts and steel pans. Mole excavation began on May 13, 1971 from a blasted starting chamber. After 46 feet were excavated the mole was shut down and the trailing muck handling equipment was pulled into place. After advancing 1,700 feet in 34 days, unanticipated ground water conditions and ground support problems were encountered slowing moling so that the next 1,000 feet of advance consumed 97 days until December

Jan 1, 1997

By Paul L. Bearden

The accumulation, handling, and utilization of sludges formed from wet process phosphoric acid is one of the major problems in the industry. The purpose of this paper is to explain some of the considerations and techniques that can be employed to provide some measure of control over the problem. An acid made from a Central Florida phosphate rock is used to illustrate the effects of one process design approach used to provide one type of control. Other approaches or techniques are discussed on a general basis. In every case, the discussion is purposely limited to conventional dihydrate processing in which phosphoric acid is produced from the filter at a strength of approximately 30% P205 and is subsequently concentrated to higher strengths such as 44 and 54% P205.

Jan 1, 1966

By Dennis P. Bryan

The southwestern United States has an abundance of natural lightweight aggregate (NLA) deposits of volcanic origin that are being used in the construction industry within the region. New Mexico, Arizona, Nevada and California utilize NLA in portland cement concrete and concrete masonry units for weight reduction and related economic reasons. The geological environment of the western United States is amenable to the occurrence of natural lightweight rocks because it contains Tertiary and Quaternary volcanic products that include low-density glass and vesicular material. There are two major classical types of natural lightweight materials of igneous origin; pumice and cinder. These account for the vast majority of volcanic rocks that are presently utilized as NLA. Pumice is an ejecta product generally found as part of tuffaceous units associated with nearby siliceous volcanic centers. Cinder, also an ejecta product, is associated with mafic volcanics and is most often found as a cinder cone. Both are highly vesicular. A third type of natural lightweight material, though also associated with siliceous volcanics, is less vesicular than pumice. It is best described as being a slightly pumiceous rhyolite or rhyolite-tuff breccia that occurs as a siliceous dome or tuft unit that is most likely to be glassy. Uses of these natural lightweight rocks center on the construction industry where they are utilized in lightweight portland cement concrete and lightweight concrete masonry units. The major reason for usage in portland cement concrete is cost saving effected by weight reduction in high-rise construction. The major usage in concrete masonry units, or "block", is for weight reduction that enables a mason to lay more lock.

Jan 1, 1987

By W. F. Riggs

INTRODUCTION Flotation equipment is becoming more definitive in design as well as greater in dimension. The last ten years have witnessed the implementation of large flotation cells ranging in size from 28 to 85 cubic meters. However, small plants still operate with cells as small as 3 cubic meters. These changes have increased the operational sensitivity of the flotation system to the chemical reagents used. No scientific study has been undertaken to develop the relationship between the size of equipment and the response of reagents in the flotation system. Hence, this article is an attempt to present several observations from past experience for the purpose of stimulating future research in an area previously ignored and perhaps subsequently not used to advantage. CONVENTIONAL FLOTATION Recent advances in conventional flotation cell design have been made in two areas: a. Flotation cells are becoming larger, and b. Basic design of flotation cells has changed. They can be classified into two types: deep cell and shallow draft cell. Size and design changes of flotation cells have a greater effect on frother performance than on the performance of other chemical reagents. Hence, the discussion in the following sections is restricted to the impact of frothers in the flotation system. Flotation Cell Size The advent of increasing flotation cell size has resulted in the increase of frother consumption. There appears to be a correlation between reduced horsepower (kilowatt) consumption and increased frother consumption per ton of ore processed in the circuit. This is more prevalent when the frother chemistry remains unchanged. Another reason for increased frother consumption is the increased distance between the rotating central shaft and the lip of the cell for larger flotation machines. Increasing this distance results in increased surface area corresponding to the larger cell size. Reduced power input per ton combined with the need for longer, stronger load support of the froth bubble as a result of increasing cell size has resulted in the development of a new series of stronger and more persistent frothers. By the same token, these newer, stronger frothers are difficult to control when applied to the smaller cells. This is true even at lower dosages because the break point between too much frother and not enough frother for optimum metallurgy is normally very narrow. This problem is often exaggerated when operators have established an optimum setting for pulp level and air for the particular frother in use. Generally, the tendency is to operate the flotation cells simply by adjusting frother dosage. The increase in frother strength to accommodate the large cells often results in the realization, after a retrofit, that the existing pump box or sump which is still handling the same volume of concentrate is no longer adequate in size. Flotation Cell Design The two most common flotation cell designs installed during the past ten years are the deep cell and the shallow draft cell. The deep cell provides for complete mixing of air and pulp at the bottom of the cell. The shallow draft cell does all the mixing/pumping in the bottom haft of the cell whereas the mixing of the air is performed further up the shaft and is brought into contact with the pulp midway up the vortex. There are many preferences for one or the other. It is the author's contention that both designs have very valid concepts and must simply be suited to the particular situation as required by the operator. The following comments are subjective observations which should be explored in greater detail. The deep cell mixes both air and pulp together at the bottom of the cell requiring a full depth ascending of the homogenized air to the surface to form the froth layer. This design, in combination with charged air, provides for good concentrate grade control under adverse conditions of selectivity, especially when faced with a fine grind and a slime-laden gangue. A selective frother can be applied without a competitive loss in recovery. One of the conditions for good selectivity in this environment is operating at medium pulp solids concentration in the range of twenty-six to thirty-four per cent (26-34%). When a higher solids concentration is used in combination with a coarse grind, it becomes difficult to maintain the same recovery with a selective frother due to higher pulp viscosity. This condition sometimes results in an unacceptable coalescence of the air bubbles as they ascend the full height of the deep cell. The increased pulp viscosity at higher solids concentration can also retard the flotation kinetics. The combination of decreased surface area of dispersed air and slower kinetics can result in reduced recovery causing coarse particles to continue to drop

Jan 1, 1989

By B. Hines

Grinding efficiency in a vertical stirred mill is affected by a wide range of process and mill configuration variables. Of these, the media fill, size, and intrinsic density have a significant effect on power draw and grinding performance. Media-related factors that influence grinding performance include, e.g.,a higher media filling providing additional particle breakage sites, or a lower bead density decreasing the running torque at a given media filling. It has been estimated that the grinding media can cause up to a 30% change in the specific energy required to grind. FLSmidth has generally recommended that the intrinsic density (specific gravity) of the grinding media be more than triple the slurry density. This density difference helps to ensure that sufficient stress intensity is generated in the mill to provide particle breakage. It also reduces the chance of media ?float? if there is a spike in either slurry density or slurry flow rate. The VXPmills have historically operated at slurry density values between 1.2 and 1.5 kg/L, which translates to nominal bead intrinsic density range between 3.6 and 4.5 kg/L. The current trend towards regrinding bulk flotation concentrates to even finer sizes, to improve recoveries of fine grained minerals, has provided motivation to consider closed-circuit operation of stirred media mills. The slurry density in this application is typically higher than in open-circuit operation due to the higher percent solids in cyclone underflows. An additional effect is that closed-circuit operation increases the potential for media float due to higher flow rates as slurry is recycled to the mill. The normal progression in extending the operating range of existing equipment is to carry out laboratory testing before pilot-scale and plant-scale trials. The results presented in this paper examine the media density to slurry density ratio and its effect on grinding performance.

Feb 23, 2014

By W. Burger

For the development of deep, hard rock, mines fast access to the orebody is critical. In some cases the mineralized area is more than 1,500 m (5,000 ft) below surface. Excavation of shafts or declines is typically on the critical path of the project schedule. Saving time on those activities can significantly increase the net present value of the mining project. Mechanical excavation methods are a step-change in excavation performance and labor safety compared to drill-and-blast operations. In hard rock conditions, roadheaders cannot be used effectively, so disc cutting is the first choice. For excavation of declines, tunnel boring machines (TBM) can be used in many cases. These provide, in certain rock conditions, considerably higher production rates com-pared to drill-and-blast excavation. Herrenknecht AG has developed a vertical shaft sinking machine (VSM) that is able to excavate shallow shafts in soil and medium soft rock. The ground is excavated by a roadheader boom. The muck is removed by a slurry system or a pneumatic system in combination with hoisting of skips. These machines have been applied successfully on shafts down to 100 m (300 ft). A pre-sink for a deep mine shaft has been excavated by this technology. For deep, hard rock shafts, a new shaft boring system has been developed by Herrenknecht AG in collaboration with Rio Tinto. Based on proven technologies, the system uses a unique arrangement of existing proven technologies. The system integrates excavation, mucking, primary rock support, installation of the final lining and shaft infrastructure. This new system improves the health and safety of shaft construction. Detailed performance estimates indicate significantly higher shaft construction rates are possible compared to conventional shaft sinking methods.

Jun 1, 2010

By Ian McFeat-Smith, Robert J. Fowell

Roadheader machines have a specific place in modern tunnelling, not simply as a compromise between drill and blast techniques and full face machines, but as an economic and versatile form of mechanised tunnelling where the method of excavation is easily integrated with tunnelling operations. In the past, roadheaders have been subjected to misuses by engineers who failed to realise that, as in all mechanised tunnelling, a planned and systematic approach is essential if the full potential of these machines is to be gained. An increasing awareness of this, together with the introduction of the more powerful roadheader carrying their cutting potential into the hard range of rock is changing this situation and now makes the prospect of using roadheaders particularly attractive (1). The advantages gained by employing these machines can be considerable and Table 1 lists some of them. This table illustrates that from both cost, and installation points of view, the roadheader has considerable benefits over the full face machine. However, the principal advantage gained is in broken ground where heavy supports can be installed immediately at the tunnel face. In addition, access for planned maintenance and repairs to most working parts of the machine is at a premium and average machine utilisations of 40 - 60% can be expected. In comparison to drill and blast methods, up to 40% better progress can be regularly achieved, and the subsequent savings in overbreak costs alone would more than pay for the cost of a roadheader used to drive a 3km long tunnel. As with full face machines, ground disturbance is minimal and for short tunnels in urban areas a road-

Jan 1, 1979

By W. Kemp Wait

Electronic ore sorting is a coarse particle beneliciating process which performs a function which in the past was primarily achieved through the use of hand picking belts. Photometric ore sorting is capable of separating waste from mineral bearing rock having a distinct difference in surface light reflectance levels. This type of ore sorter can be applied to gold and silver ores which are contained in quartz veining. Photometric ore sorting has been used at gold mines in South Africa as well as the Black Pinc silver mine in Montana. A new ore sorter has recently been developed with precious metal ores in mind. This ore sorter is capable of sorting ores on both photometric and conductivity sensing parameters. The first production machine has been installed at a gold mine near Beardmore, Ontario and currently is being used to process low grade waste dumps.

Jan 1, 1998

By Hassan El-Shall, G. L. Zucker, T. S. Jordan

INTRODUCTION This chapter will discuss, both from fundamental and practical points of view, the flotation of non-sulfide min- eral s. The flotation characteristics of silicates, salt-type minerals, oxides and coal will be reviewed. Dup- lication of the thorough and excellent presentations by many authors is not intended and a comprehensive list of references is included. GENERAL PRINCIPLES Collectors for Non-Sulfide Minerals : All minerals, with only a few exceptions are readily wetted by water, that is, they are said to be hydrophilic. The purpose of a collector is to coat the mineral surface and render it hydrophobic, preferring an environment of air over an environment of water. Relative to sulphide minerals, the non- sulfides are more naturally hydrophilic and therefore, require a collector with a greater chain length. This is accomplished by using a collector which has a proportionally longer hydrocarbon chain. Both anionic and cationic flotation collectors are used in non- sulfide flotation. Anionic Collectors. These are collectors which ionize in solution to give a negatively charged radical. They have the general structure R-A, where R is a hydrocarbon having 8 or more carbon atoms and A is a hydrogen or an alkali metal. The most common collectors in this group are fatty acids which could have saturated or unsaturated hydrocarbon chains. Such chains may be straight or branched. The configuration of the chains affects their solubility in water and their adsorption characteristics. In general, as the chain length increases the adsorption and the collecting power increases. However, it should be noted that the solubility decreases. The solubility of unsaturated chains is greater than that of saturated fatty acids because double or triple bonds are polar and thus attract water molecules. Solubility data for collector metal salts, even though important for a proper understanding of flotation are not available for all collector systems. Solubilities and structures of different fatty acids are reported by Fuerstenau (1982). Values for the solubility products of fatty acid salts selected from a paper by Du Rietz (1958) are given in Table 1.

Jan 1, 1986

By J. B. Gwiazda

In regard to the paper by J.B. Gwiazda, it makes a highly technical approach to show that the .u factor used by designers of lemniscate-guided roof supports has never really been confirmed as a maximum and assumes that convergence is vertical. Also, the paper does not appear to take into account deflection of structures, which occurs when the lemniscate and base members are fully loaded to their maximum stress level, nor the front to back line of the support in relation to differential roof to floor movements caused by strata movements under pressure. It is not unusual for differential movements to be slightly diagonal to the line of the support, particularly in faulted areas and on gradient faces. The paper also does not take into account consolidation of fines immediately above and below the support. Generally speaking, any differential movement is from face to waste and under these conditions the .u of 0.3, which appears to be an international standard, has worked in practice. However, if the face end of the support is lower than the waste end, then the µ of 0.3 can be considerably increased, giving rise to the damage mentioned in the paper. The ideal design should aim for a slightly forward bias in the lemniscate guide so that the last increment of setting is toward the face, tending to close any fissures that may have developed during the support advance cycle. The support should also be fitted with positive set valves to ensure that a high setting load density is attained to minimize bed separation. As far as powered supports are concerned, convergence is irresistible and all powered supports converge at their rated yield load. A similar principle can be applied to the differential roof to floor movements to drastically reduce the very high forces that would otherwise be applied to the lemniscate structures and pins and that, in turn, are transferred to the base arrangement and floor loading. Any differential movements are usually catered for by the 0.3 µ factor or deflection of structures in the lemniscate guide arrangement and consolidation of the floor. The floor loading, due to differential movement, is in addition to the support convergence load and requires additional bearing area to avoid possible floor penetration. Some seven years ago, Fletcher Sutcliffe Wild Ltd. (FSW) introduced a lemniscate-guided shield support where the lemniscate linkage is connected to the roof bar through two horizontally converging rams to allow differential movement to take place above a given rated figure. This is a known force and can be guarded against, whereas with rigid connections the forces, as yet, are unconfirmed. By careful design, a horizontal force in excess of 6 MN (60 tons) opposes differential movements for a total ram loading of only 2.5 MN (25 tons), or 1.25 MN (12% tons) each. This principle can considerably reduce the length and weight of the support in comparison with a rigid pin-type structure; also, the yield load rating can be increased without affecting the lemniscate forces. The graph shows the tensile and compressive forces in a lemniscate linkage of a support with and without hydrostore. These forces react into both the roof beam and base members and, as can be seen from the support height to linkage load graph, a considerable reduction in these reactions is gained by the use of the FSW patented hydrostore system. Floor loading is considerably reduced under maximum µ conditions, and by allowing the roof bar to move with the strata, some degree of improvement to strata control is achieved in line with the assumptions in the paper. In practice, these movements have only been in the region of a few millimeters, which, in turn, reflects on the improvements to strata control by the addition of positive set valves.

Jan 1, 1987

By Wallace Jarman

Introduction As with many gravity concentration processes, pneumatic concen¬tration traces its origin in antiquity. The use of winnowing to separate chaff from grain has long been known, and such procedures were undoubtedly practiced in ancient civilizations for the concentration of ores and the separation of slag from metal whenever specific gravity, size, and shape differences were favorable. Taggart142 has reported on the use of dry rockers and dry panning techniques for use with gold ores in arid regions such as Western Australia. The present day air table was developed in the last century for this purpose. Devices and Processes Classification. All classifiers make a separation on the basis of size, shape, and specific gravity. In addition to specific gravity and size, particle shape factor is often particularly important to pneumatic processes, and advantage is taken of the fact that both flat and fibrous particles settle at velocities substantially lower than that of their equiv¬alent spheres. Another factor of importance in this process is the bulk density of the material since substances such as exfoliated ver¬miculite or partially opened fibrils of asbestos have bulk densities significantly lower than those of the pure in situ mineral. Winnowing takes advantage of all of these physical factors to affect a separation starting with a closely sized feed, and vermiculite has been separated from rock by this process.142 When air classifiers are used to size minerals, the separation will often not be perfect because heavy particles will respond similarly to somewhat coarser light particles while flat, elongated, and low bulk density particles will act as would particles finer than the cut size. An example of a classification process that also results in concentration is the zig-zag classifier"' used to separate paper from glass, metal bottles, and cans during the recycle of municipal solid waste. Application of this process to minerals showing the proper specific gravity, size, and shape differ¬ences is obvious. An important device of this type is an air-aspirated screen used to remove asbestos from its associated rock during the dry processing of that mineral.144 The raw ore is crushed and screened and the asbes¬tos removed by aspiration from the surface of the screen by virtue of its shape and low bulk density. Hindered Settling. Based on the discussion in the section on "Hindered Settling Concentration and Jigging," the use of air devices could be anticipated in desert areas or where moisture may be deleteri¬ous to the product to be separated. Such processes have three major defects: (1) dust may be difficult to contain, (2) fines are difficult to process, and (3) the process is inherently less efficient than are wet processes as may be seen from Eqs. 3 and 4 in the section on "Hindered Settling Concentration and Jigging p. 4-47." Taggart142 has reported on the use of both dry panning and a dry rocker for gold separation in and regions. The dry panning makes use of winnowing, defection on the basis of specific gravity and shape during fall, and hindered settling. Prescreened placer material is poured from one pan to another as air blows across the falling stream. The process is repeated many times with hand picking to produce a rough concentrate. Magnetic minerals are then removed, and the residue, one grain deep in the pan, is air blown. In dry rocking the gravel is sized on a steeply inclined screen with the undersize fed to a riffle box with a porous bottom which is blown from underneath."' This device was the forerunner of the present-day air table. Air Tables and Jigs.145. 146 At least nine types of these devices have been developed, all but two of them for coal. While the devices used for coal may be characterized as pneumatic jigs, pneumatic tables, and pneumatic landers,145 only pneumatic jigs survive today. The same is true for ores.'" In these devices presized particles are fed to the separator which consists of an inclined vibrating conveyor with a porous surface through which air is carefully introduced to form a fluid bed. The lighter particles are lifted by the air out of uphill conveyance and float downhill, while the heavier particles in contact with the surface are conveyed uphill by the vibrator. Typically both transverse and longitudinal slopes are used. A plan view of a Triple S air table and the type of separation made is shown in Fig. 44.147 Because of its external appearance, the device has been called an air table, although it functions essentially as a jig. However, some of the asymetrical acceleration conveying function of a standard shak¬ing table is also performed by the eccentric vibrator incorporated into this device. Low pressure air is admitted below the vibrating table surface consisting of cloth (e.g., canvas), porous plastic, woven wire, or punched metal supported on a wire mesh or grid. As with a shaking table, the device produces a concentrate, middlings, and tailings. The middlings are recycled or are subsequently treated on another concentrating device. As with many wet separations, presizing by screening is necessary. The oft-repeated admonition of Professor Richards, "Separation without classification is damnation," is cer¬tainly true here. In this way a small heavy particle which might weigh the same as a large light particle and thus report to the same place on the air table has already been removed in the sizing step. Air tables are almost universally encased with a canopy to remove dust by aspiration. Some concentration may also be achieved in this step when favorable specific gravity, size, and shape differences are present. Applications Ore Separations. Air tables were originally developed for ore separations, and they find application in and regions or where water is deleterious or inconvenient to use or remove (e.g., small tonnage materials already dried). They have been used for asbestos, bauxite, calcite, cassiterite, columbite-tentalite, diatomaceous earth, fluorite, gilsonite, graphite, kyanite, manganese minerals, mica, managite, perlite, pyrite, pyrrhotite, vermiculite, and uraninite ores. More recently, air tables have found favor in separating a wide variety of secondary materials such as abrasive grains, bone char, catalysts, fiberglass, scrap glass, scrap wire from its insulation, prills from slag, dross and coke from metal or from each other, metal from crushed crucibles, lead from plastic in old batteries, and cubic particles from flat ones. The devices also finds considerable application in the preparation of food¬stuffs. As stated before close sizing before separation is desirable, and if the specific gravity difference is slight, very tight

Jan 1, 1985

By S. P. Mehrotra, P. C. Kapur

A phase-species classification is suggested for the mathematical models of flotation kinetics. Important models are discussed and the structure of their rate constants and parameters is analyzed. Split or enhancement factors are formulated for different kinetic models. The steady state dynamics of the flotation circuits can always be described in terms of these factors. For simulation of circuits, feed is characterized in a formal manner starting from a multivariate distribution of particles in flotation rate, size and mineral content. Synthesis of optimum circuits which simultaneously searches for the best process variables and structural parameters from a generalized master network is described. Trends in the areas of search algorithms, role of middlings and feed identification etc. are indicated.

Jan 1, 1989

By John L. Bolles

The mineral processing engineer must have some knowledge of mineral identification, and the greater his knowledge in this field, the greater will be his potential to do his own specialty well. Most metal-bearing ores and industrial minerals require processing of some kind to produce a salable product or a material suitable for further processing. The types of treatment are many, but each one is aimed at separating a mineral or family of minerals from others. Obviously, then, it is basic to all mineral processing to identify the minerals present in the raw material and in intermediate and final products resulting from processing. Very often the mineral processing engineer will enter planning while exploration of a mineral prospect is going on. He will examine hand specimens, drill cores and cuttings, bulk samples, and materials in place. He must identify the valuable minerals, their occurrence and associations with other minerals, and the identity of diluents or host rock. An exact and detailed knowledge of the raw material will usually suggest the treatment processes that should be investigated in the development of the most suitable scheme for that particular raw material in that particular location. Physical properties of a mineral often lead directly to a process for separating that mineral from others. Some well-known examples of these are: magnetic separations for such minerals as magnetite, rutile, ilmenite; gravity separations for gold, silver, galena. cassiterite; flotation separation for most sulfides; distillation of mercury from cinnabar; and selective dissolution in specific solvents such as halite or sylvite in water, gold and silver in cyanide solutions, copper oxides, carbonates, sulfates and silicates in dilute sulfuric acid. These processes and many others are described in detail elsewhere in the Hand¬book and so it is not appropriate to give details here, except to emphasize that processing is aided greatly by complete mineral identi¬fication. Developing a treatment scheme for a new mineral discovery is an exciting assignment fag the mineral processing engineer, but much more of the usual work relates to quality control in existing plants. Here too, mineral identification is important. Some processing is aimed toward producing an end product which will meet exacting specifica¬tions regarding purity or allowable diluents. In this situation it is vital that contaminants be identified so that the process can be adjusted

Jan 1, 1985