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By Robert J. Mac Kinnon

A moving grid finite element model for the analysis of transients in a rock mass surrounding an underground coal conversion cavity is presented. Major model components are outlined, including a special element formulated to account for internal discontinuities. A numerical example is described.

Jan 1, 1986

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION This method is a variation of cut-and-fill mining where either the wall rock or the vein is too unconsolidated to permit safe mining by conventional cut-and-fill methods. Basic development consists of driving a 2.4 x 2.7-m (8 x 9-ft) haulage heading laterally in the footwall side of the vein. Crosscuts are driven over to the vein at 91-m (300-ft) intervals. Then the vein, at each inter¬section, is mined out to a distance of about 9 m (30 ft ) above the track level. After the installation of an ore chute, scrams or intermediate drifts are driven in both directions along the vein. Mining then starts imme¬diately below the level bounding the upper limit of the stope. A cut varying between 3 and 4.2 m (10 and 14 ft) in depth and mined for a distance of 45.7 m (150 ft )along the strike is taken from the orepass in each direction (Fig. 1). In this mined-out void, the floor of the cut is equipped with 203 x 203-mm (8 x 8-in.) caps hitched into the wall with steel saddle hitches installed on 1.5-m (5-ft) centers. Double wood stringers, 152.4 x 152.4 mm (6 x 6 in.), are then installed on 1.2-m (4-ft) centers longitudinally down the length of the stope. Reinforc¬ing wire mesh is also placed, overlying each mat of timber. Directly above the stringers, stulls are installed from hanging wall to footwall to hold the timber in place during the sand-filling operation. Then a cement¬sand mixture in a ratio of about 1:10 weight is poured to a depth of about 2.4 m (8 ft). The entire cycle is then repeated down to the scram drift. Another variation of timbering is illustrated in Fig. 2. There is a vein width of approximately 5.5 m (18 ft) with the same unconsolidated ground conditions. The primary difference between handling this situation and the previously described undercut-and-fill method is that standard timbered sets are installed in each excavated stoping void rather than stulls. Once again the basic method is to work down the ore block from an upper timbered drift to a lower timbered drift. Cuts are taken from a timbered raise with a chute on the stope side. Each piece of ground excavated is essentially two sets high, 5.5 m (18 ft), by two sets wide, 5.5 m (18 ft), by the length of the cut along the strike of the vein. The timbered sets are 2.4 x 2.4 x 2.4 m (8 x 8 x 8 ft). The net result is, therefore, two sets of timber on a top floor separated by doubled-up 76-mm (3-in.) lagging from

Jan 1, 1982

By Robert C. Dunne, Jeffery P. Vaughan

The Archean ores of Western Australia are in total the largest producers of gold in Australia. In addition to numerous small and medium sized deposits, the province also contains the giant Kalgoorlie Golden Mile deposits. A variety of different gold ore mineralogical associations has lead to the development of a number of different processing routes for both refractory and non-refractory ore types. Intense weathering and lateritization of Archean greenstones in the semi-arid Yilgarn region have resulted in the development of a deeply oxidized profile (60-100m) above most sulphide orebodies. Residual and re-deposited supergene gold within iron oxide/silica gangue of this zone responds well to CIP, CIL, heap and vat leaching techniques. Although gold is easily liberated in principal from oxidized gangue, in practice recoveries are sometimes affected by iron oxide coatings on gold grains, and by clay minerals which can interfere with plant performance. Refractory ore types in the Eastern Goldfields fall into three main categories - pyritic, arsenopyritic and telluride-bearing. Pyritic ores may be either free milling or require roasting. The difference appears to depend largely on whether gold occurs along grain boundaries or fractures, which can be liberated by grinding, or whether gold occurs as inclusions within pyrite grains and therefore cannot be satisfactorily liberated by grinding. Arsenopyrite bearing ores are not well understood and are complicated by what appears to be a variable degree of gold solid solution, in addition to included gold. Gold occurring in complex gold-silver telluride assemblages has traditionally been recovered by roasting in the Eastern Goldfields.

Jan 1, 1987

By Lawrence D. Ramspott

A test of retrievable geologic storage of spent fuel assemblies from an operating commercial nuclear reactor is underway at the Nevada Test Site (NTS) of the U.S. Department of Energy his generic test is located 420 m below the surface in the Climax granitic stock. Eleven canisters of spent fuel approximately 2.5 years out of reactor core (about 1.6 kW/canister thermal output) were emplaced in a storage drift along with 6 electrical simulator canisters. Two adjacent drifts contain electrical heaters, which are operated to simulate within the test array the thermal field of a large repository. Fuel was loaded during April-May 1980 and initial results of the test will be presented.

Jan 1, 1981

By R. S. Conti

The US Bureau of Mines (USBM) has developed an inflatable feed-tube seal (IFTS) for high-expansion foam generators. The IFTS is a lightweight, portable, rectangular inflatable bag that can be used by firefighters to rapidly seal large openings, such as those in underground mines, and to simultaneously provide a feed-tube for a high-expansion foam generator. Thus allows fire fighting foam to freely flow to the fire site and control or extinguish the fire. Studies indicated that a high-expansion foam plug will travel 183 m (600 ft) down an entry with a 4.5% rise in elevation, before leakage of foam from around the inflatable partition.

Jan 1, 1996

By K. Saberyan

A technique is described for the solvent extraction of zirconium. In this method, zirconium has been extracted quantitatively from 0.005 M sodium salicylate solution at pH 2.5 - 3 using 0.01 M triphenylphosphine oxide dissolved in toluene as an extractant. Zirconium has also been recovered quantitatively from 0.2 M potassium thiocyanate solution, but extractions in nitric and hydrochloric acids (as media) were negligible. The resulting extracted metal ion has been stripped with sulfuric acid (0.5 M) and determined with inductively coupled plasma atomic emission spectrometry. This method has been applied successfully to achieve 95% recovery of zirconium from zircon concentrate.

Jan 1, 2010

By CHENGLONG XU, HAORAN LI, YALI FENG, JINRONG JU, Shunliang Liu

With the depletion of mineral resources and the growing demand for metal resources from industrial development, the recovery of metal resources from secondary resources has attracted increased attention. Among all secondary resources, spent catalyst is one of the most important as it contains valuable metals such as vanadium, nickel, molybdenum, cobalt, and aluminum. Compared with the extraction of metals from minerals, the recovery of valuable metals from spent catalysts has the advantages of a simple process, a short production cycle, lower investment cost, and less environmental impact.

Apr 1, 2023

By A. H. Shapiro, W. –F. Leung

A decanter centrifuge of novel design was developed for the dewatering of fine-solids slurries having micrometer-range particles that form pastelike, flowable, hard-to-dewater cakes. The design combines a compound beach with a cake flow control structure that allows expressed liquid, from the cake in the dry beach region to be returned to the liquid pool while maintaining a high-solids throughput. To provide optimal performance with various operating parameters (i.e., feed rate, centrifugal gravity, pool, differential speed, concentration of feed solids and slurry type), the cake flow control is adjustable. The second part of the beach is preferably at a zero angle. Tests showed that the new design allows increases in solids dryness and solids throughput. Key words: Dewatering, Centrifuge, Slurry, Fine particles, Beach decanter, Cake-flow control

Jan 1, 2003

By Felicia F. Peng

An economic analysis for evaluating coal preparation plants in the Appalachian Region is presented. A complete coal preparation plant which includes multi-streams and/or multi-stages coal cleaning circuits, complete closed water circuits, environmentally acceptable refuse disposal and reclamation systems is used to determine the capital and operating costs. The plant is capable of producing metallurgical grade clean coal or deep cleaned steam grade coal for future power generating plants use. Utilizing a profit maximization as a decision criterion, an optimal yield, and optimal coal clean¬ing circuits, closed water circuits, and disposal and reclamation systems for a given coal characteristic may be selected among the process alternatives by using a process simulation-economic model of the coal preparation plant. A deterministic analysis is used to calculate the discounted-cash-flow interest rate of return (DCFIROR) and project the selling prices, and determine the Return on Investment (ROI) for the cases of 32.65% and 52.74% plant utilization with producing two cleaned products. Sensitivity analyses are performed to quantify the effect of total capital investment, operating cost, run-of-mine cost and yield on an economic criterion-ROI and selling prices. A probabilistic evaluation using the Monte Carlo simulation technique is further made to quantify the risk involved in obtaining the desired profit levels. The results of the economic analysis for a given coal preparation plant configuration and characteristics of raw coal indicates that the selling prices of producing two cleaned products are greatly affected by the change in percent of plant utilization (production schedule), yield, and run-of-mine cost, and total capital cost. The sensitivity analysis also shows that the total capital investments has a three times greater effect on required selling prices than operating costs. The risk analysis shown that the overall economics look very attractive with current coal sale prices and 52.74% plant utilization.

Jan 1, 1979

By Robert A. Jankowski, Ellsworth R. Spencer, Jay F. Colinet

Longwall mining is the most productive method of under- ground coal mining in the United States (US.), and record levels of production have been reached in recent years. These ever-improving production levels have the potential to generate significantly higher quantities of respirable dust. Consequently, the longwall industry continues to struggle with controlling dust liberation and maintaining compliance with the federally-mandated respirable dust standard. In response, longwall operators have in- creased the application of primary dust controls, airflow and water flow, in an effort to improve respirable dust control and reduce worker exposure. Pittsburgh Research Center (PRC) personnel conducted dust surveys at 13 longwalls operating throughout the U.S. and collected information to: identify current operating practices on these longwalls, identify the types of dust controls in use, document the levels at which these controls are being applied, and measure the respirable dust levels present. This data was analyzed to quantify the dust levels generated by the major sources on the longwall section and to evaluate the relative effectiveness of the control technologies in use.

Jan 1, 1997

By P. D. Pattillo, J. B. Cheatham, Y. -H. Lin

The stability of a hemispherical cavity, intended to model a perforation opening in the vicinity of a borehole, is analyzed for the case of poroelastic-plastic material behavior. Two plasticity models of differing complexity are considered, with the emphasis in either model being the use of strain softening to calculate inelastic stiffness. Sample results illustrate the effect of steady fluid flow on the response, and possible collapse, of the near perforation region.

Jan 1, 1986

By John Carr

The tremendous growth in grinding mill size and horsepower during the '70s demanded advancements in gear cutting technology and equipment for the '80s. Computers are used as an aid in the design and selection of the mill drive system; sophisticated computer graphics are used to create manufacturing drawings; and a new generation of gear cutting machines are used to manufacture the gears. These new gear cutters can generate-cut gears over 40' (12.2 m) in diameter at an AGMA quality level exceeding the currenty acceptable standards. The precision of the resulting gear assures maximum tooth contact, which will result in longer gear set life.

Jan 1, 1981

By Joseph W. Murray

GENERAL DESCRIPTION The undercut-and-fill method of mining was devel¬oped by Inco Ltd. in the Sudbury district of Ontario, Canada, to deal with abnormal ground conditions en¬countered in pillar recovery. Although it is now some¬times utilized as a primary stoping method, the main application continues to be in pillar recovery. In any mining method in which fill is employed, pillar systems tend to fail gradually as stoping pro¬gresses. As a result, most pillars are completely destressed and broken by the time pillar recovery com¬mences. Control of the back in such ground becomes a major problem and is contingent on local conditions as they exist in the workplace. In most mining methods, operations such as drilling, blasting, slushing, and tim¬bering can be readily assessed and the mining cycle pre¬dicted. However, in the mining of pillars with overhand methods, the time and effort required to maintain a safe back varies to a quite unpredictable extent. For years the square-set method was the most effective means available for pillar recovery. However, in areas where bad ground was encountered, square setting was a costly and sometimes hazardous method. As the tonnage mined from pillars increased, it was apparent that a safer and less expensive approach was required. Past experience indicated that the best way to deal with broken ground was to mine from the top down. Top slicing systems were examined but were rejected due to the major sub¬sidence which resulted. However, ideas generated from studying top slicing finally developed into the present method of undercut-and-fill. This method, which has almost completely replaced square-set mining, exhibits reduced costs, has improved safety and efficiency, elimi¬nates the factor of bad ground overhead, and allows a closer prediction of the actual mining cycle. The result is more accurate scheduling of pillar production. This factor is of increasing importance as the tonnage from pillar recovery increases in relation to total mine pro¬duction. Undercut-and-fill is a method of extracting a block of ore by mining successive cuts, working from the top down. After a cut of ore is completely mined out, lami¬nated timber stringers are constructed along the sides for the full length of the cut. Round logs covered by a wire screen are laid across the stringers to form a mat. The cut is then tightly filled with hydraulically placed cemented sand fill. Mining is then resumed on the cut immediately below the mat. As the cut advances, the timber stringers are supported by round timber posts seated on the solid bottom of the cut. Drilling, blasting, muck removal, and timbering operations are repeated until all the ore in the cut has been removed. At this time, another log mat is laid and the cut is filled. This sequence is continued until the block has been mined from level to level. The basic principles of undercut-and-fill can be adapted to a variety of mining situations. It can be applied transversely or longitudinally in ore bodies of varying width, that dip from vertical to flat lying. The method can be applied using conventional equipment or it can be mechanized through the introduction of small diesel or air-powered units for muck removal. Undercut¬and-fill is a selective method, ore recovery is high, and dilution can be controlled to an acceptable level. How¬ever, it is a high-cost, labor-intensive method and as such its use is restricted to areas where the less costly meth¬ods of ground support are not successful or where bulk mining methods are not adaptable. One basic require¬ment for undercut-and-fill is an adequate source of hydraulically placed cemented sand fill and a dependable sand plant and distribution system. The success of the undercut method is to a large degree dependent on the competence of the fill placed in each of the successive cuts. In theory, the weight of the fill on each cut is supported in part by the cemented sand itself as it binds to the rough sides of the mined¬out areas next to the pillar walls, and in part by the tim¬ber support. As mining progresses beneath the mat, a downward pressure is exerted within the cemented fill and a voussoir arch develops above the mat, across the full width of the cut. This supports the major portion of the load. The log mat and posts have a double func¬tion in that they carry the remaining weight not sup¬ported by the voussoir arch and in addition prevent any pronounced settling of the fill mass. Excessive settling breaks the bond between the cemented fill and the pillar walls, eliminating the arch effect and transferring the entire load to the mat and posts. The theory of a voussoir arch forming within the cemented sand fill mass is substantiated in practice by the fact that pillars have been mined from level to level with no more significant weight on the bottom cuts than was experienced on the top cuts. In effect, as mining progresses downward, there is no additional transfer of weight from previous cuts to the cut being mined. APPLICATION OF UNDERCUT-AND-FILL Because of their size, the Sudbury ore bodies are, for the most part, mined transversely and, to a considerable extent, by filling methods. Transverse stoping blocks are designed with 9-m (30-ft) crown pillars established be¬low each level and with stopes separated by vertical rib pillars. In past years, floor pillars were left above the level as well, but present practice favors the silling of primary stopes at the base of the rail. Rib pillars are designed to meet the ground control requirements of the primary stoping methods. Local conditions and past experience generally dictate their size. Historically, since the square-set method was to be used for recovery, rib pillars were designed to accommo¬date standard 1.7-m (5'h-ft) sets. A typical rib pillar is four sets wide [6.7 m (22 ft)] when located between two square-set or cut-and-fill stopes having a width of 8.4 m (27.5 ft). When the undercut-and-fill method was de¬veloped, a mat width of two sets was adopted. This di¬vided each cut of the common four-set pillar into two

Jan 1, 1982

By M. R. Yalamanchili, C. A. Young, J. J. Kellar, W. M. Cross, J. D. Miller

Introduction Researchers have long speculated regarding the adsorption state of collectors on mineral surfaces. Only recently have analytical techniques become available that allow direct in situ analysis of adsorbed layers. With reactive internal reflection elements (IREs), Fourier transform infrared spectroscopy coupled with internal reflection spectroscopy (FT-IR/IRS) allows for unobtrusive investigation of the adsorbed layer (Harrick, 1979). This technique has been used both for the in situ identification of adsorbed species and for the direct calculation of collector adsorption density in real time from infrared spectral data (Miller and Kellar. 1988). The reactive mineral IRE serves both to internally reflect the infrared light and to act as a substrate at which collector adsorption can be monitored. Shown in Fig. I is a schematic of light being internally reflected in an IRE. As the light is totally reflected at the interface, an evanescent wave propagates a short distance (typically a few microns) away from the IRE, and sampling of the interfacial region occurs. Because of multiple internal reflections, great sensitivity can be obtained. Flotation systems for which this method has been demonstrated include CaF2/oleate, Al2O3/dodecylsulfate, CaCO3/oleate, ZnS/xanthate and KCI/octylamine (Miller and Kellar, 1988; Kellar et al., 1989, 1990; Miller et al., 1990). Of particular interest to flotation chemists is the conformational behavior of the alkyl chain of adsorbed collector molecules. In the simplest sense, conformation can bethought of as the relative flexibility of the hydrocarbon chain caused by rotations about carbon-carbon single bonds. The flexibility is described by the amount of gauche/trans conformations in the alkyl chain. Figure 2 shows examples of gauche and trans conformations (rotational isomers or rotamers) in a hydrocarbon chain. These two conformations are observed when the terminal methylene groups in any four-carbon subchain are on the same side (gauche) or opposites sides (trans) of the plane of the middle carbon-carbon bond. A gauche conformer (rotamer) is thus a state of maximum potential energy, whereas a trans conformer is a state with minimum potential energy. The difference between these states has been found to be approximately 0.5 to 0.8 kcal/mole (Mead et al., 1986). The relative number of butane molecules present as gauche conformers with respect to trans conformers can be calculated from the Boltzmann equation (Mead et al., 1986). At 25° C approximately 66% of the butane molecules are trans rotamers. [ ] However at 100° C, only 33% of butane molecules are trans rotamers. Exposing the hydrocarbon chain to a polar, rather than an a polar, medium should also cause an increase in the number of gauche butylene groups in a hydrocarbon chain, although this effect has been difficult to observe experimentally (Mead et al., 1986). Forming a crystalline hydrocarbon is equivalent to decreasing the temperature, which greatly increases the amount of trans butylene groups present in the chain. From the preceding discussion, a picture of the solution species of concern can be developed. The collector monomer should exhibit the greatest number of gauche rotamers and most likely will look crumpled or balled-up to reduce the surface area in contact with water. The micellar form is expected to be aggregated, with the hydrocarbon chain existing in a less polar environment. These characteristic conformational features have been examined by FT-IR spectroscopy for various liquid phases of alkanoic acids (Umemura et al., 1981; Cameron et al., 1982a; Yang et al., 1986; Wong, 1987; Wong and Mantsch, 1983; Cross et al., 1992), phospholipid membranes (Fringe Ii, 1980; Casal and Mantsch, 1985: Wong et al.. 1988;

Jan 1, 1994

By W. H. Andrews

DEFINITION OF BAUXITE The term "bauxite" (1) dates back to 1821, when it was assigned to a deposit of red earthy material outcropping near the village of Les Baux in Southern France. Chemical analysis indicated the presence of aluminium oxide, iron oxide and water. Deposits of bauxite are widespread throughout the world although the major developments occur within 30 degrees of the Equator. Essentially bauxite is the product of weathering and leaching of a variety of sediments or rocks, generally in a high-rainfall tropical or sub-tropical climate, resulting in the enrichment of the aluminium oxide content. It is characterised by the presence of one or more of the minerals, gibbsite, boehmite, and diaspore. The other constituents are usually clay, (typically kaolin), hematite and/or goethite (iron-oxides), anatase (titanium oxide), and possibly silica as quartz. A variety of minor elements may be present, such as calcium, magnesium, phosphorus, vanadium, gallium, manganese, and many others in very small quantities. Organic compounds are normally present from development in association with a soil horizon where vegetation may be encouraged by the high rainfall necessary for bauxitisition. The physical form depends upon the geological history of the deposit and correlation also exists with the mineralogy. Older deposits which have been subject to burial are invariably hard and compact with the alumina predominantly occurring as boehmite or diaspore, while the more recent, surface deposits are frequently less compacted and gibbsite is the dominant alumina mineral. In these later deposits, the physical form can vary from the fine red soil (terra rosa) of Jamaica to the Weipa pisolites to the cemented Arkansas type. The presence of concretionary structures is an important aid in identification in the field.

Jan 1, 1984

By J. D. Miller, C. L. Lin

The appropriate software and hardware necessary to adapt an X-ray CT analyzer for the determination of coal washability was developed and successfully tested for well-characterized samples under laboratory conditions. Evaluation of the CT-based analyzer under simulated on-line conditions is also presented to verify the on-line capabilities of the newly developed X-ray CT analyzer for the control of coarse coal cleaning circuits. The investigation shows that the data obtained from CT washability analyses compare favorably with experimental sink float data. Further, image frequency was considered to evaluate the CT-Based coal washability analyzer for simulated on-line conditions. Specifically, the number of observations (image spacing) needed for accurate sampling and the total time required to perform the calculations for a representative data set were determined. Key words: Coal washability, Computer simulation, X-ray CT analysis

Jan 1, 2003

By Robert A. Cummings

Disintegration of the immediate shale roof in coal mines interferes with ventilation, production, and safety. The effect is seasonal and correlates with high levels of moisture in the intake ventilation air. Disintegration, though not completely understood, is caused by shale dilation and weakening upon exposure to atmospheric moisture. Fluctuating moisture conditions are particularly deleterious, and may be at least partially controlled by "tempering" the air before it is brought to the working places. Roof behavior was studied in an operating mine during a 12-month monitoring period. This approach can be used in developing criteria to reduce or prevent shale roof deterioration.

Jan 1, 1981

By Martial P. Corriveau

The preparation of coal is essentially the same for export as it is for the domestic market. The big difference, I would say, is that, while coal for certain domestic uses (utilities, for instance) need not always be of the highest quality, as long as it is uniform, the buyer of coal for export invariably demands a well-prepared, top-quality product. And he is willing to pay for it, as long as he knows what he is buying. This is where quality control enters the picture. The day is long past since the export market was the dumping ground for coals which could not be sold on the domestic market. Today, the export buyer wants American coals not so much for a price but for the necessity of blending the American coals with foreign coals to achieve the desired chemical and/or physical characteristics, To blend intelligently, he must know in detail what are the qualities of the coals to be blended. Sampling and chemical analysis facilities at the piers provide the information on the coals exported, These facilities are maintained for the export buyer by independent laboratories; however, in one instance, the buyer operates his own laboratory. With very few exceptions, and Clinchfield?s Moss No. 3 mid-volatile coal is one, cargoes shipped overseas are loaded as blends of two or more coals. Up to now, foreign requirements have not been too clearly defined; therefore, shippers must be prepared to blend mixtures of the desired combination. Successful blending requires an accurate knowledge of the coals shipped. And this makes it necessary for the coal producer to maintain his own control laboratory.

Jan 1, 1959

By David A. Rice, Scott A. Shuey, William K. Tolley

The U.S. Bureau of Mines (USBM) is developing new dewatering technology to permit the recycling of iron values contained in wet steel plant sludges. In laboratory tests, Q-BOP fume sludge was dewatered to 12% moisture by displacing water with acetone in a plate-and-frame pressure filter. Acetone is reclaimed through distillation at lower cost than conventional thermal drying of the sludge. The projected additional energy requirement for acetone treatment of the sludge, including acetone recovery, is $5 to $7 per dry ton.

Jan 1, 1995

By Cesar A. Busnardo

The carbonatite orebody in Jacupiranga is situated some 230km (140-mi) southwest of Sao Paulo and is mined by Serrana S/A de Mineragao, producing an apatite concentrate for fertilizer production and a calcite tailing for cement manufacture (see Figure 1). The orebody (Table 1) is mined by open pit and the total production is 4.5 million mt/year of carbonatite ore with a further 2.2 million mt/year of overburden. The carbonatite low-grade ore (Table 2) is fed to the concentration plant at 460 mt/h , comprising of four parallel lines of grinding, by rod mills 2.7m x 4.5m (9 x 15-ft), operating in closed circuit with hydrocyclones. The grinding circuit product is deslimed in four stages using hydrocyclones and the deslimed product is conditioned with corn starch and subsequently floated, using fatty acids as collector to, recover the apatite.

Jan 1, 1985