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By B. C. Darst

Intensive efforts to develop potash use in agricultural production were begun in the US 50 years ago. Early growth was rapid and continued well into the 1960s. Many soils were inherently low in K, so crop responses to fertilization were often dramatic. Deficiency symptoms were common. Per-acre use has continued to increase in recent years, but annual growth rates have slowed. Soil buildup has occurred in some areas. Poor economics in the 1980s have been difficult to overcome. Efforts are now well under way to develop potential markets around the world.

Jan 1, 1987

By J. D. Miller

Reaction kinetics for gold dissolution in acid thiourea solutions were investigated using formamidine disulfide as an oxidant. For the conditions studied it appears that a surface reaction limits the rate of gold dissolution. The rate was found to be first order with respect to thiourea concentration and half order with respect to the concentration of formamidine disulfide; the apparent activation energy for the reaction was 10.40 kcal/mol (43.40 kJ/mol); and the reaction rate was independent of pH for the pH range 1.0 to 3.0.

Jan 1, 1999

By K. N. Han

The adsorption of gold onto activated carbon in iodide solutions was studied. The adsorption rates were determined as a function of temperature, salt concentration and pH, as well as activated carbon concentration. A kinetic model was developed for the adsorption. The adsorption rate was believed to be controlled by a combined mechanism of film diffusion and pore diffusion with the film diffusion having the most significant effect. Experimental findings were in good agreement with the model developed. The apparent activation energy was found to be 14.9 kJ/mol.

Jan 1, 1995

By K. N. Han

The kinetics of gold dissolution in iodide solutions with oxygen in an autoclave was investigated. The leaching process was described by a gas-liquid-solid heterogeneous reaction. The effects of stirring speed, temperature, oxygen partial pressure and lixiviant, concentrations on the overall leaching rate was investigated. It was found that the dissolution of gold could be represented by a heterogeneous chemical reaction. Under the conditions studied, the apparent activation energy ranged from 34.7 to 54.3 kJ/mol. Experimental findings were in good agreement with the kinetic model proposed.

Jan 1, 1998

By L. Zhou, J. P. Rider, T. W. Beck, J. A. Organiscak, Y. Zheng

The National Institute for Occupational Safety and Health’s Office of Mine Safety and Health Research recently developed a series of models using computational fluid dynamics (CFD) to study the gas distribution around a continuous mining machine with various fan-powered flooded bed scrubber discharge configurations. CFD models using Species Transport Model without reactions in FLUENT were constructed to evaluate the redirection of scrubber discharge toward the mining face rather than behind the return curtain. The following scenarios are considered in this study: 100 percent of the discharge redirected back toward the face on the off-curtain side of the continuous miner; 100 percent of the discharge redirected back toward the face, but divided equally to both sides of the machine; and 15 percent of the discharge redirected toward the face on the off-curtain side of the machine, with 85 percent directed into the return. These models were compared against a model with a conventional scrubber discharge, where air is directed away from the face into the return. The CFD models were calibrated and validated based on experimental data and accurately predicted sulfur hexafluoride (SF6) gas levels at four gas monitoring locations. One additional prediction model was simulated to consider a different scrubber discharge angle for the 100 percent redirected, equally divided case. These models identified relatively high gassy areas around the continuous miner, which may not warrant their use in coal mines with medium to high methane liberation rates. This paper describes the methodology used to develop the CFD models, and the validation of the models based on experimental data.

Jan 1, 2015

By J. Lovering, H. Wilhelm, P. Siewert

Introduction The Afton property is located 290 km (180 miles) by air east-northeast from Vancouver and 14 km (8.7 miles) west of Kamloops, a city of 60,000 people, in south central British Columbia, Canada. The mine is adjacent to the Trans-Canada Highway at an elevation of 670 m (2198 ft) above sea level. The ore body is a porphyry copper deposit that has undergone supergene alteration. The major economic minerals in the supergene zone are native copper and chalcocite with chalcopyrite and bornite in the hypergene areas. The grade is 1% with an overall copper distribution - 70% native, 25% chalcocite, and 5% chalcopyrite with bornite and covellite. The ore also contains important but variable amounts of gold and silver. The mill was designed to treat 6350 t/d (7000 stpd). Semiautogenous grinding was selected to minimize capital cost and because of the expected high clay content of the ore, which would have caused problems in a conventional crushing and screening plant. Test work indicated that a recovery of 87% was possible in a circuit incorporating both flotation and gravity separation. Flowsheet Run-of-mine ore is crushed in a 1.06 x 1.65-m (3.5 x 5.4-ft) Allis Chalmers gyratory crusher set at 228.6 mm (9 in.), closed side setting. The surge pocket, below the crusher, is emptied by a Hydrastroke feeder onto number one conveyor, which discharges onto a 180,000-t (198,416-st) coarse ore stockpile. Six Hydrastroke feeders on two conveyors withdraw the crushed material from the bottom of the pile. These two conveyors, in turn, discharge onto the belt feeding the semiautogenous mill. The live storage in the stockpile is approximately 22,000 t (24,250 st), sufficient for three days' mill feed. Primary grinding is accomplished in an 8.5-m (28-ft) diam by 3.7-m (12-ft) long Koppers (Hardinge Cascade) mill (Fig. 1) containing a 10% ball charge and driven by a 4000-kW dc variable speed motor. The mill dis¬charge is pumped by a 10 x 12 G.I.W. pump to a 1.22 x 4.88-m (4 x 16-ft) stationary screen sloped at 20°. Screen oversize returns to the semiautogenous mill (SAM), and the undersize flows by gravity to the ball mill discharge pump box. Secondary grinding is performed in a 5-m (16.4-ft) diam by 8.84-m (29-ft) Koppers overflow ball mill driven by a 3430-kW synchronous motor through an air clutch. The mill is in closed circuit with a Krebs Cyclopac containing 10 635-mm (25-in.) cyclones and the cyclone overflow, at 35% solids and 65% to 70% -200 mesh, is flotation feed. In order to limit the buildup of native copper, circulating in the secondary grinding circuit, a portion of the underflow from the cyclones is processed in a circuit containing screens, cyclones, and shaking tables to produce a finished metallic copper concentrate. Primary mill variable speed drive The overall waste to ore ratio at Afton was 4.5:1. The mining was to be done with only three shovels, which meant that it was highly unlikely that more than one of them would be in ore at any one time. The resulting inability to blend the mill feed made it impossible to prevent wide swings in the grade and grindability. The variable speed do drive motor installed on the semiautogenous mill was selected because of the extreme variability of the Afton ore body. This variability has persisted throughout the lifetime of the mine. There are times, however, when due to ore conditions, the mill is operated at full speed (78% of critical) for extended periods of several shifts duration. There are other times when the mill speed may be changed several times in a 12-hour shift due to changing ore conditions. When ore is processed that contains a fairly large proportion of fine native copper, the primary mill speed and, consequently, the tonnage may be reduced to improve the secondary grind and to maintain an acceptable grind and recovery. High clay ores require less mill speed and more dilute grinding densities. In the latter case, the slower primary mill speed also helps to minimize damage to the mill liners. Approximately 57% of the time the mill operates between 90% and 100% of full speed or between 71% and 78% of critical. The variable speed is also used for inching during mill relines.

Jan 1, 1987

By J. D. Miller, R. J. Smith

Earl C. Herkenhoff Publication of the article "The Flotation Behavior of Digested Asphalt Ridge Tar Sands" by R.J. Smith and J.D. Miller has confirmed and again spotlighted a most significant and important development in the field of tar sand and heavy oil mining. This is the ability to recover 95 % of the bitumen in dry tar sands by a hot water process. The initial investigation by Miller and Sepulveda in early 1978, "Separation of Bitumen from Dry Tar Sands," was the first blockbuter. It is disappointing that there has been so little scientific and industry ackowledgement that the authors laid to rest the long entrenched belief that the sand grains had to be "water wetted" for a hot water process to be applicable. Dr. Miller and his researchers at the University of Utah used slightly different mineral dressing terminology in describing their process but it is essentially the old, established technique of high solids attrition scrubbing at elevated temperatures with a conditioner or modifier, followed by dilution and froth flotation. Similar techniques were described in detail for other nonmetallic minerals in the mid-1940s by Stuart Falconer of American Cyanamid Co. The authors used a 15-minute "scrub" period (or in their terminology "digester step") at 73% solids and 95°C (203°F) in a 0.3 m solution of sodium carbonate -about 4 kg/t (8 lb per st) of tar sand. Flotation was at 10% solids. Obviously, the scrub is an expensive step with respect to energy input but the high recovery of bitumen appears to justify it. Two basic questions then arise: • What are the effects of shorter scrubbing time, lower temperature, and less sodium carbonate? • Would a two-stage approach be advantageous? That is, could you get the major proportion of the bitumen with a mild scrub and apply the high power and temperature only to the refractory portion? These variables should be evaluated. Apparently, all results were only for rougher flotation. What would the grade and recovery of bitumen be if multiple cleaner flotation steps were employed? We must concede that a 60-62% grade of bitumen leaves considerable room for upgrading. Finally, it should be noted by all executives who approve funds for heavy oil recovery schemes, most of which are variations of in-situ methods, that mining can yield a 95% recovery of bitumen. That is correct, 95%. Imagine what that does to the total barrels recovered from a heavy oil resource compared with the scrawny 30-35% maximum recovery we are getting now. The conclusion-let's mine heavy oil and tar sands. Reply by R.J. Smith and J.D. Miller We, too, share Mr. Herkenhoff s enthusiasm for the mining of tar sands-heavy oil and continue the development of related processing technology. Specifically, it will be noted that by appropriate control of bitumen viscosity in the feed (probably the most important variable for successful separation) the recovery can be accomplished at low digestion temperature and low addition of alkali. A 91 t/d (100 stpd) pilot plant based on our technology is now being operated by Enercor in Salt Lake City (Hatfield, Oblad, and Miller, 1982). In these same pilot facilities, rougher concentrates containing more than 80% bitumen on a dry basis are being produced from Asphalt Ridge Feed. However, the processing of certain other tar sands results in incomplete digestion of bitumen/sand aggregates as discussed elsewhere (Miller and Misra, 1982). Such undigested aggregates are shown in Fig. 1. In this regard, it is intended that a recycle stream will be installed in the pilot plant to provide for redigestion of middling products. Further details regarding surface chemistry aspects of this tar sand technology have recently been published (Misra, Aguilar, and Miller, 1981). References Hatfield, K.E., Oblad, A.G., and Miller, J.D., 1982, "Pilot Plant Recovery of Bitumen from Oil Wet Tar Sands," Processing, Second International Conference on Heavy Crude and Tar Sands, Caracas, Venezuela, Feb. 7-17. Hatfield, K.E., Oblad, A.G., and Miller, J.D., 1982, "Pilot Plant Processing of Oil-Wet Tar Sands," Symposium, Rocky Mountain Fuel Society, Salt Lake City, Utah, Feb. 18-19. Miller, J.D. and Misra, M. 1982, "Hot Water Process Development for Utah Tar Sands," Fuel Processing Technology, Vol. 6, No. 1, April, pp. 27-49. Misra, M., Aguilar, R., and Miller, J.D., 1981, "Surface Chemistry Features in the Hot Water Processing of Utah Tar Sand, Proceedings, "Symposium of Separation Science and Technology for Energy Applications," Gatlinburg, Tennessee, May 5-8.

Jan 11, 1982

By F. J. Sotillo

Mining in Peru is mainly concentrated on metallic and polymetallic ores of gold, copper, lead, zinc, silver, tin, iron, etc. However, little has been done to develop extensive and rich non-metallic or Industrial Minerals deposits with the exception of some silica and phosphate deposits. This paper presents the characterization and metallurgical studies for the development of a feldspar deposit. These studies showed that the special characteristics of the ore, such as high porosity, high Al2O3 content, low clays content, and sized material allowed us to obtain commercial products applying relatively simple and low cost concentrating unit operations. The characterization studies showed that the feldspar products contained no quartz and about 90% of albite (NaAlSi3O8). The 14x40-mesh size fraction product analyzed 5.91% Na2O, 22.56% Al2O3, 65.73% SiO2, 0.65% K2O, 6.50% CaO, and 0.40% Fe2O3; whereas, the 40x150-mesh size fraction product analyzed 6.22% Na2O, 24.30% Al2O3, 60.80% SiO2, 0.91% K2O, 7.06% CaO, and 0.54% Fe2O3. These products could be used for manufacturing glass, ceramics, refractory, and filters in different applications. In addition, Middlings, could be used as insulation material due to its porosity (bulk density of 0.828 Kg/dm3), analyzing 4.46% Na2O, 15.7% Al2O3, 67.2% SiO2, 2.57% K2O, 3.94% CaO, and 4.24% Fe2O3. The material showed a sphericity of 0.77, a roundness of 0.68, and acid solubility of 13.3025% on only washed and 2.72% effective for clean material suitable for use in these applications. No selective accumulation of the valuable mineral was observed at any specific size fraction. However, iron and titanium bearing impurities trended to concentrate in fine fractions (minus 70 mesh -212 µm-). Metallurgical studies demonstrated that this feldspar ore required sizing at 14x40 mesh and 40x150 mesh to render commercial products. In the case of the 14x40-mesh size fraction, the process involved magnetic separation, gravity separation, and flotation for feldspar (direct flotation) using a blend of primary and condensate amines. However, the reverse flotation for mica-biotite and iron bearing minerals was also possible using petroleum sulfonate. The metallurgical process for the 40x150-mesh size fraction required magnetic separation and reverse flotation of iron and titanium bearing minerals, the feldspar being reported in the sink material. Commercial products from this feldspar deposit for glass, ceramics, refractory, and filtration; and for the manufacturing of insulation material were obtained.

Feb 23, 2014

By P. Shang

Fly ash and zeolite have strong absorption to cations since their porous microstructure. The porosities of fly ash and zeolite can also be increased by modification, which improves the absorption of the materials. This research showed that the effects of modified fly ash and zeolite to remove ammonia nitrogen from rainwater. The optimal reaction time, additive amount of materials, and pH value of the stormwater to remove ammonia nitrogen were experimentally determined. For modified fly ash, the optimal removal rate of ammonia nitrogen from 100 ml stormwater was 74.41% with the addition of 10.0g fly ash, 1.5 hours of the reaction time, and 8 of pH value. To the modified zeolite, the optimal removal rate of ammonia nitrogen was 51.22% at the reaction conditions of 2.0g of zeolite, 1 hour of reaction time, and 7 of pH value.

Jan 1, 2010

By J. Kallioinen

Flotation of low grade Ni ores with serpentinite host rock tend to yield low recoveries of Ni With high MgO contents in the concentrate. The recovery can be enhanced by optimizing the sub-processes in the pulp and in the froth. The new OF-100 flotation cell has been designed to be adaptable to different feed fineness (Coarse, Fine). The losses can according the research conducted at Outokumpu's Hitura plant be substantially decreased by OF-100's new rotor and stator design and froth zone geometry. The froth residence time is used to adapt for feed particle size changes measured by the Outokumpu PSI- 200. Models for the behaviour of different particle Sizes in the sub-processes have been developed. The modified froth zone geometry and the capability of maintaining a froth of desired thickness allows efficient washing of the obtained concentrate producing low levels of MgO.

Jan 1, 1994

By Bobby Michael (Mike) Roberts, G. M. (Mike) Bernard, Ronald R. Kelly

This paper describes Cananea's successful implementation of the world's first mobile high angle conveyor (MHAC™) to achieve multiple-lift stacking of their primary crushed (2,500 mtph) low-grade ore. The mobile high angle conveyor answered Cananea's needs to stack a second lift, 30-meter high, on their Quebalix I pad. The mobile high angle conveyor is a custom unit that was designed by Terra Nova Technologies (TNT) in conjunction with Continental Conveyor and Equipment Company (CCE's 100th HAC). The stacking system expansion project also includes a rail mounted tripper on the existing overland conveyor and a custom designed "corridor conveyor" to build the corridor for the upper mobile HAC tracks and tail of the existing mobile stacking conveyor. Mexicana de Cananea, part of Grupo Mexico, is a large copper producer in Sonora Mexico and has been in operation continuously for over 100 years.

Jan 1, 2002

By Ron Robinson

Booster fans have been in regular use in British coal mines since the first half of this century, but a large increase in their use did not take place until after the Nationalization of coal mines in 1947. From that time the need was identified to improve standards of ventilation in general and mining engineers specializing in ventilation were appointed at all levels in the mining industry. A comprehensive approach to ventilation surveys and planning, together with the realisation of the benefits of booster fans, led to their more rapid introduction. Many coal mines in the U.K. are at moderate depth and have virgin strata temperatures of 30°C or less. The ability to control climatic conditions by increased ventilation rather than resort to cooling systems has led to the introduction of booster fans and recirculation systems in these mines. Large makes of methane and high levels of coal dust production due to increased mechanization have also required enhanced ventilation standards. This paper deals with the reasons for selecting booster fans and the specific safety considerations involved in their use. Numbers, types and sizes of booster fans in the U.K. are outlined, together with notes on installation and performance. Remote monitoring and control from the surface is dealt with, and current legislation. The evolution of controlled recirculation is also covered starting with auxiliary fan systems in 1970 and leading to very large district schemes in the late 1980s. The research at Newcastle upon Tyne University and other establishments in the climatic conditions associated with recirculation is summarised and its use in assisting existing and proposed conventional and recirculation systems is indicated.

Jan 1, 1989

By C. E. Semler

A multiyear program involving online, in-vessel exposures at Barrick Goldstrike Mines Inc. (BGMI) has provided information and experience about the materials and designs for brickwork linings in pressure oxidation autoclaves. A variety of brick and mortar products was evaluated. Properties before and after exposure were used to document the performance. Some benefits include the development of a new silicate mortar, a better understanding of brick requirements and increased autoclave operating time between maintenance shutdowns. BGMI operates six pressure oxidation autoclaves at its Eureka County, NV gold operation. These autoclave process 15.8 k/ld (17,500 stpd) of gold ore. Figure 1 is a flowsheet of the six autoclave circuits. The gold ore is refractory, in which the gold particles are mineralogically encapsulated on a micro- and submicroscopic level. The autoclaves are 4.4- m- (14-ft-) diam and 26 m (85 ft) long (tip-to-tip). Inside of each is a full brickwork (brick plus mortar) lining that divides the vessel into five compartments. Each compartment has a motor-driven agitator that keeps the slurry suspended and disperses the oxygen to enhance reaction. During autoclave operation, the conditions are steady state at 210° to 232° C (410° to 450° F) and 61 to 62 kPa (420-430 psi). As shown in Fig. 1, there are other accessory vessels (splash and flash towers) in the autoclave circuits that are also brickwork-lined. The linings used to date, other than small test sections, have all been acid-resistant fireclay bricks, bonded with alkali silicate mortar, PbO- based mortar and/or vinyl ester mortar, but mainly alkali silicate mortar.

Jan 1, 1997

By Clément F. Bernier

One of the chief sources of the world's supply of asbestos lies in a belt of ultrabasic rocks which runs northeastward through an area called the Eastern Townships, south of the St. Lawrence River in the Province of Québec. The productive portion of this belt extends from the town of Asbestos in the south, for a distance of some 9,5 km to beyond the village of East-Broughton. The first discovery of asbestos in this region was made in 1874. Only crude fibre was produced until 1888 when the first mill was erected near the town of Black Lake. Since the inception of the industry, almost 100 years ago, no less than 100 different compa¬nies have been formed and 46 different mines have been operated. There are presently 5 active companies in the area providing direct employment for 4100 workers, down from 6000 during the seventies.

Jan 1, 1984

By D. D. Harbuck, G. K. Green

The Bokan Mountain area on Prince of Wales Island, AK, contains vast quantities of heavy rare earth minerals. To lessen the US dependence on foreign sources, the US Bureau of Mines (USBM) investigated yttrium recovery from this ore. By sulfating the ore at 250 ° C for 1 hr at an H2S04:ore ratio of 1:1 and then water leaching the sulfated material for 1 hr at 20% solids and 25ºC, over 74% of the yttrium, cerium, and lanthanum were solubilized. Solvent extraction (SX), using a synergistic combination of OPAP and DEHPA, selectively loaded 100% of the yttrium. Yttrium was stripped directly as a solid using a 2M, 35 ° C NaOH liquor. When the OPAP/DEHPA mixture was tested in a continuous SX circuit, yttrium loading and stripping were 100% and 96%, respectively. The organic showed excellent recyclability properties with minimal solvent loss.

Jan 1, 1996

By Erik Westman, Brandon Stables, Benjamin Mirabile

"Standing support is often used in conjunction with underground retreat mining. Knowledge of the load-displacement behavior of standing support and loading induced by the mine opening is critical to proper support selection. The NIOSH STOP database contains load-displacement laboratory test data for most commonly used standing supports. Hydraulic load cells currently used to measure in-situ loading of standing supports have exhibited leakage under load, producing irregularities within the dataset. An improved hydraulic load cell eliminates leakage and produces more consistent data.INTRODUCTIONA hydraulic load cell, more commonly known as a flat jack or JackPack, is a device installed to determine the loading characteristics of the roof-to-floor supports. These standing or roof-to-floor supports include, but are not limited to pumpable cribs, cans, wooden cribs (4-point, 9-point, etc.), metal props, wood timbers, concrete donut cribs, and other variations of these standing supports.Flat jack devices have been used previously to determine the loading of such roof to floor supports. The most commonly used load cell for standing supports is based on a design intended for pre-stressing standing supports, not load measurement. Preloading helps to keep the supports, such as timbers props and cribs, in place during the dynamic loading associated with blasting (Barczak and Tadolini, 2004).Flat Jacks are manufactured in various shapes and sizes depending on the application; common shapes are square and round that are 25.4 mm (1.0 inch) thick when collapsed, as shown in Figure 1. These load cells are produced in South Africa and can be formed into several different shapes depending on the application (Barczak, 2003). The sheet metal that makes up the bladder is comprised of 19-gauge or 1.11 mm (0.04 inch) thickness; the top and bottom pieces of sheet metal are resistance-seam welded around the perimeter. There are two different ports installed on one of the bladder skin corners of the flat jacks; these ports allow fluid to fill the unit, and the other port has a relief valve. The fill port is comprised of a check valve. The pressure relief valve is located in the second port; preset to 1,724 kPa (250 psi), the pressure relief valve keeps the load cell from bursting. Bursting of these load cells typically causes a small tear in the metal skin allowing the fluid to leak out without a large catastrophic failure.The effective inflation range of these load cells is between 25.4 mm (1.0 inch) to 254.0 mm (10.0 inches) thick. Different types of fluids can inflate these load cells but the most common and easily accessible fluid for testing applications is water. Water pressure from the mine supply line can inflate these load cells to the desired load. Depending on the pre-stressing requirements, a pneumatic powered pump, hydraulic pump, or hand pump may be required to inflate the load cells."

Jan 1, 2019

The segregation process has been known for many years and applied industrially for the treatment of various refractory copper ores. The principles of the process are briefly reviewed; the process was tested in the laboratory on two copper-cobalt oxide ores from Gecamines, Shaba Province, Zaire. Various parameters of the process, such as segregation temperature, dosage of reducing and chloridizing agents and flotation conditions were inyestigated. Experimental results have shown that the segregation of copper could proceed satisfactorily at low temperature (1000° K) but that of cobalt required higher temperature and higher additions of reducing and chloridizing agents.

Jan 1, 1992

The adsorption of gold onto activated carbon in iodide solutions has been studied. The adsorption rates were determined as a function of temperature, salt concentration and pH as well as activated carbon concentration. A kinetic model was also developed for the adsorption. The adsorption rate was believed to be controlled by a mixed mechanism of film diffusion and pore diffusion, but the film diffusion had a major contribution to the overall rate. Experimental findings were in good agreement with the model developed. The apparent activation energy was found to be 14.9 kJ/mol.

Jan 1, 1993

By Chelsea Woodward

The Spokane Research Laboratory of the National Institute for Occupational Safety and Health has demonstrated a prototype hydrogen-fueled mining vehicle as an alternative to diesel equipment in underground mines. Safety and control systems were added to allow a 70-kW rubber-tired hydrogen vehicle, originally developed by the U. S. Bureau of Mines and Hydrogen Consultants, Inc. (now Hydrogen Components, Inc., of Littleton, CO), to be tested. This work began in late 2000 and culminated with a Mine Safety and Health Administration fire safety review and field tests at the Stillwater Mine in Nye, MT, in 2004.

Jan 1, 2006

By J. S. Jaspal

Mining of thin coal seams is both difficult and uneconomical. As a result, most of thin-seam coal reserves in the U. S. are not being mined. It is estimated that nearly 910 Gt (K. Unrug, 1985) of coal reserves exist in 355 to 710 mm thick seams. The Bureau is attempting to develop new techniques for mining 150 to 1065 mm thick coal seams. This paper summarizes the research work on existing thin-seam mining techniques, testing on small and full scale physical models of the Bureau slot mining concept, and computer modeling of the full-scale process.

Jan 1, 1992