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By J. B. Gwiazda

Maintenance of a constant load setting throughout longwall support units and selection of the proper initial bearing capacity depending on the type of roof strata are the basic factors that ensure good performance of roof support in a longwall. These requirements can only be met by hydraulic support. The greatest advantage of hydraulic support is achieved when uniform pressure is imposed on the roof throughout the length of the longwall. Such support, however, is provided only if each of the support units acts with the same force against the roof, i.e., has the same setting load. In such cases, the roof behaves like a uniform plate without bending and shearing stresses, thereby ensuring an undisturbed structure. Without a positive set value system, achieving an equal setting load for all units of the longwall support is impossible. Due to alterations of the feed line pressure of support units as well as some reasons related to man's psychology, operators extend the height at different setting loads. This produces nonuniform roof stress, and disturbs the structure. Consequently, the roof usually cracks. The author has developed a positive set value system, which is described in this Technical Note. Selection of the setting load Two pressure values in the feed lines are usually applied in longwall hydraulic systems. Lightweight support is fed by 25 MPa (250 bar) liquid, while heavy duty units receive a nominal pressure of 31.5 MPa (315 bar). Such pressure is required not only for the props but also to power the adjust¬ment jacks and the advancing ram. If the feed pressure is too low, there will be difficulty in shifting the unit despite the inversion system of the advancing rams. On the other hand, for many roof types, the feed pressure often appears to be too high when applied as the setting load pressure. An excessive setting load acts too strongly against the roof, crushing weak strata close to the roof. The author has recognized a case where an excessive setting load destroyed not only the nearby roof strata but also the strata above a 2-m (6.6-ft) sandstone layer. In addition, an excessive setting load relieves the side¬walls, increasing the resistance when using cutting machines. As a result, the yield of coarse coal is diminished, and increased fines dominate in the final product, lowering its economic value. As indicated, selection of the proper setting load, depend¬ing on the mining and geological conditions of the extracted seam, is extremely important. In some mines, measures applied to prevent disturbances include reduction of the feed line pressure by adjusting the feed pump valve. The disadvantage accompanying reduced feed line pressure is more difficult operation in advancing the ram. Due to the reduced feed line pressure, the force of the advancing ram is much lower than the designed value. Other designs suggest using a third feed line. However, installation of supplementary valves on the support units is required, a time-consuming and expensive procedure. The disadvantages of the powered supports are eliminated by a system designed by the author. So far, such a method of setting load control has not been used in any type of support. Setting load control unit The designed positive valve set for prop loading and the setting load control correspond to existing control systems for hydraulic powered support. The layout of the unit connected to the hydraulic prop control is presented in Fig. 1. The unit is marked LIDS. It incorporates three valves that may operate separately or connected. Valve A automatically opens and closes with liquid flow in the prop feed circuit. The valve is opened when the canopy touches the roof and closed when the support unit is withdrawn. Valve B serves as the setting load control. Valve C automatically opens and closes the flow in the line connecting the under-piston space of E to the prop F with the separator G. The valve block of each support prop is marked BZ. The UDS unit is connected by the hydraulic lines to the F prop control circuit. A valve is connected by H to pressure line J and by K to the G separator. In the UDS-3 version, line L is connected to M, linking the over-piston space of F prop with the G separator. Valve B is linked with valve A by a connector; it is also connected to the under-piston space E of prop F by line P. Valve C is fixed between lines K and P, connecting space E of prop F with the separator G. When setting the support, the liquid flows from line J through separator G, the BZ valve, and valve C to the space E of prop F. When reaching the roof with the canopy, valve A is opened and C closes. In this way it is impossible for the operator to cause the liquid pressure in space E of prop F to reach the level of line J. The prop pressure is set by valve B of the UDS unit. When withdrawing the support, valve C is automatically opened and A closed. Three UDS units have been fabricated and are designated

Jan 1, 1990

By D. Koch

Bentonites and bentonite containing sealing materials are used since many years for sealing purposes in foundation, dike construction, hydraulic engineering and landfill construction, especially for encapsulation of old waste deposits In the near future compacted bentonites will be important components of multi-barrier systems for dry charge of old mines, for underground disposal of hazardous or radioactive wastes. Because of their exceptional physical, structural and chemical properties, bentonites are offering manifold possibilities to protect the environment against the negative effects of dumping grounds. Often the compliance with legal regulations for waste dumping or the isolation of contaminated areas is possible only by using bentonite containing sealing systems. The essential properties of bentonite, bentonite containing building materials and the different techniques of making horizontal technical base liners and vertical cut-off walls are explained. For horizontal technical base liners the adsorption ability, the ion exchange capacity and the swelling behavior of bentonites are important. For cut-off walls bentonite controls the rheological behavior of the slurry and is responsible for the low permeability of the hardened bentonite-cement-wall. In the multi-barrier systems of repositories for chemical or radioactive wastes compacted bentonites are considered to be favourable sealing and backfilling materials because of their high swelling capacity, their adsorption and ion-exchange ability and their plasticity in water saturated state.

Jan 1, 2005

By H. M. Giegerich

About Cominco Only the original Broken Hill zinc deposit in Australia is larger than Cominco's Red Dog deposit. Indeed, the third largest zinc deposit in the world is Cominco's own Sullivan mine in British Columbia. The production plan at Red Dog is based on mining 1.9 Mt/a (2.1 million stpy) or ore, to produce 635 kt (700,000 st) of zinc and lead concentrates. At full production, Red Dog will be the largest zinc mine in the Western world. It will also be one of the lowest cost producers in the Western world. And Red Dog will be one of the first major hardrock mines to operate in Alaska in modern times. The Red Dog project will be the third and largest of Cominco's Arctic mining developments in the past 13 years. The company's Black Angel mine in Greenland started up in 1973. And the Polaris mine, high in the Canadian Arctic, began production in 1982. Red Dog is one of the world's great mineral deposits. It is a unique combination of large size, high grade, and a waste-to-ore ratio less than 1:1. The present mineral reserve (Table 1) is estimated to be 77 Mt (85 million st), at a grade of 17.1% zinc, 5% lead, and 82 g/t (2.4 oz per st) of silver. Mine life is estimated at nearly 50 years. Cominco is the largest producer of lead and zinc in the Western world. Cominco's smelter in southern British Columbia is the largest lead and zinc plant in the west. Cominco also produces copper, steel, tin, gold and silver, along with potash, ammonia, urea, and phosphate fertilizers. But Cominco's major products are lead and zinc. At present, Cominco produces more than 10% of the Western world's needs for these metals. Background-Red Dog The Red Dog mineral deposit is located in northwestern Alaska, about 960 km (600 miles) by air from Anchorage, Alaska's largest city. The Red Dog region is home to more than 5000 people living in 11 villages. The largest village is Kotzebue, with a population of about 2500. The village is located about 144 km (90 miles) southwest of the mine. The country around Red Dog features rolling hills, leading into mountains to the north and east. The deposit, at an elevation of 305 m (1000 ft), is in permafrost. But the area is not nearly as cold as that encountered in the Canadian Arctic, or on Alaska's North Slope. At Red Dog, the permafrost is relatively fragile. It can be thawed fairly easily, a factor in engineering designs. Red Dog Creek flows through the mineral deposit. One of the natural phenomena at Red Dog is the presence of significant heavy metal contamination in the creek. This has eliminated all fish in the vicinity of the deposit. The contamination, and resulting mineral stain, led to the discovery of Red Dog. The staining is caused by rain and snowmelt. This leaches the heavy metals from the mineral zone, which comes right to the surface in many places. As mining progresses, this surface mantle will be removed. This will

Jan 12, 1986

By Reul A. Dejú

The mineral olivine has been the subject of continued research by various authors. Paik (1967) conducted infrared and flotation studies of the adsorption of sodium alkyl benzene sulfonates on olivine. These studies are of interest from two points of view. First, in the metallurgical field it is of interest to understand the mechanism of silicate flotation using sulfonate collectors and second to a hydrologist studying the pollution of surface or subsurface waters it is of interest to study the reactions taking place when water plus sulfonates encounter a porous (silicate) medium. On the basis of Paik's study he was able to postulate a possible mechanism for the attachment of RSO3 (sulfonate radical) to the olivine surface. Parks (1969) has been studying the important influence of the structure of a silicate such as olivine on its dissolution mechanism and its related surface charge. Also Luce, a student of Parks (1964, 1965) has been studying the kinetics of dissolution of olivine using the techniques of direct solution analysis and infrared analysis. Certainly a classic in the study of silicate dissolution is the work of Murata (1940) on the reaction of orthosilicates (olivine in particular) with strong acid solutions. Many authors (Dejú and Bhappu, 1966. 1967. 1968, Dejú, 1969; Rhoades and Krueger. 1968, and Faust and blunter, 1967) have reported studies of the behavior of silicates immersed in water (especially orthosilicates) and have found that there is an exchange of cations from the silicate surface for H+ from the water. Studies by Dejú and Bhappu have also included the system olivine-water-amine.

Jan 1, 1971

By Leons Kovisars

Two-thirds of the world copper produced in 1979 incurred cash costs between $0.35 and $0.89 per lb, averaging $0.62 per lb. North American costs are higher at $0.48, $0.88, and $0.68 per lb, respectively. Variable cash costs are $0.20, $0.58, and $0.39 per lb for the world; $0.29, $0.63, and $0.46 per lb for North America. Cash costs are relevant for comparing existing operations and evaluating market pricing during periods of excess supply. Analysis of full costs of production (cash costs plus income taxes, profit, and non-cash items) becomes relevant in evaluating the incentive price required to develop new production plants. In mid-1979 this cost (incentive price) was about $1.10 per lb; it is expected to approach $2.00 per lb by 1985.

Jan 1, 1979

By Guy G. Jr. Granger

INTRODUCTION Due to the current depressed state of the mining industry, many mining companies have been forced to operate at less than maximum capacity. This paper will describe how the AMAX Henderson Molybdenum Mine and Mill near Empire and Parshall, Colorado respectively have utilized operations scheduling and budgeting to achieve production targets at minimum operating costs The AMAX Henderson Molybdenum Mine and Mill have been operating most recently at approximately sixty percent of maximum capacity. OPERATIONS SCHEDULING Production Targets. The production targets are tons, operating days, tons per opera- ting day, head grade, recovery, and pounds of molybdenum. One specific objective of operations scheduling is to meet the production target for pounds of molybdenum. Present market conditions dictate that production be adjusted to match sales of molybdenum. Over-production and creation of molybdenum inventories is unacceptable. In summary, the product ion target for pounds of molybdenum must be met, but not exceeded. Capacity. The AMAX Henderson Mill single stage open crushing circuit capacity is 48,000 dry short tons per day. The AMAX Henderson Mill single stage semi-autogenous primary grinding mill circuit capacity is 38,000 dry short tons per day. All other AMAX Henderson Mi 11 circuits such as rougher flotation, regrind, cleaner flotation, lead leach, filtering, drying and concentrate handling are sized to accommodate the 38,000 dry short tons per day and pro- ducts from same. The tailing delivery system is designed to accommodate tailing from a 38,000 dry short tons per day operation. Constraints. One mill production scheduling constraint is that the mill operating schedule must be the same as mine production, ore haulage and crushing. There is a total crushed ore stockpi1e storage capacity of 240,000 dry short tons. This breaks down to 60,000 dry short tons of crushed ore over the three reclaim pan feeders for each of the four primary grinding mill circuits. However, there is only approximately 10,000 dry short tons of live crushed ore available over the three reclaim pan feeders for each of the four primary grinding mill circuits. Dozers must be utilized to re- claim and push the other crushed ore. Primary grinding mill throughput is

Jan 1, 1986

By Marvin Barnes

The Sacaton mine is a relatively new open¬pit copper mine owned by Asarco, Inc., which has been in production only a few years. The mineral deposits consist of two ore zones; a western, near-surface zone minable by open-pit methods, and a deeper eastern zone which will have to be mined underground, probably by block caving. The copper deposits are related to small quartz monzonite intrusives implaced into Precambrian granites. Both rock types act as hosts for the copper mineralization, which has been secondarily enriched sufficiently to constitute ore. In the western zone the enriched supergene chalcocite has been subjected to postenrichment oxidation processes with partial destruction of the chalcocite blanket. The deeper eastern chalcocite zone has not been as severely affected by later supergene processes. Several mineralogical populations are represented in the total copper deposit. Leached capping overlies the chalcocite enriched zone, and in the western ore body there is a very irregular interface between the leached and enriched zone, which is complicated by oxide copper minerals irregularly distributed. Primary lower-grade chalcopyrite mineralization occurs below the enriched zone. When mining commenced near the top of the enriched zone, considerable difficulty was encountered in correlating the actual grade of ore in place with the estimated grade of the mining blocks. The mineralogical complexity and fairly wide spacing of diamond drilling made reliable grade estimation of individual ore blocks a difficult procedure.

Jan 1, 1979

By S. K. Bhatt

This paper reviews major developments in continuous haulage technology for underground coal mines. Haulage systems in use and development are investigated through a comprehensive literature search and visits to mines and manufacturers. The systems include flexible conveyor train, mobile conveyor, multiple- unit continuous haulage, extensible and belt turning systems. Strengths and weaknesses of the systems are assessed in light of their operating under remote control and their application to the computer-assisted continuous mining machine being developed by the US Bureau of Mines. Through mining scenarios, candidate haulage systems are conceptualized to be consistent with the capabilities of the mining machine.

Jan 1, 1991

By Chee Theng

"In 2013, Coeur Mining had more than $55 million in general and administrative (G&A) costs and less than $100 million in cash flow. This year, its G&A costs stand at $30 million, down by more than 40 percent, its all-in sustaining costs have shrunk by more than 30 percent, and its cash flow is above $200 million. To realize this, the company aggressively reduced both its corporate overhead cost and operating cost and repaid more than $300 million in debt.The more robust balance sheet meant the precious-metals producer was well positioned to reap the rewards of an upsurge in commodity prices in the early part of 2016. Its stock price soared from $1.75 in January to $15.98 in August. Although the stock price has since dropped to around $9 due to a stronger dollar and a more hawkish Federal Reserve that recently raised U.S. interest rates, shifting the opportunity-cost argument away from buying gold and silver, Coeur remains attractive to analysts in light of its prudent spending and improved fundamentals.Focus on balanced portfolioCoeur is the largest primary silver producer and one of the largest gold-producing companies in the United States and one of the largest silver companies in the world. Incorporated in 1928, the U.S.-listed company has a total of five operations: three in the United States, one in Mexico and one in Bolivia. About two-thirds of its revenue comes from the United States.The company is conscientious about having a balanced portfolio, so no single mine accounts for more than about 30 percent of its business. Being in three different jurisdictions provides further diversification, and its revenue split of about 60 percent from gold and 40 percent from silver gives it some metals mix. This year, it expects to produce about 466 t (15 million oz) of silver and 10.3 t (330,000 oz) of gold."

Feb 1, 2017

By Shahzad Moosa, John M. Pitt

Statistical variability in rock strength can be evaluated by rapid, inexpensive ultrasonic pulse velocity measurements, an important consideration since rational probabilistic design in rock is often hampered by the cost of obtaining sufficient strength data. Experimental ultrasonic pulse velocity measurements on a set of 96 limestone specimens showed two distinct statistical populations. Both populations followed Weibull distributions, with flexural strength distributions paralleling pulse velocity distributions. Thus, experimental results suggest that strength variability may be predicted from a few rationally selected strength measurements and several ultrasonic tests. The correspondence between strength and velocity may then be used to develop an experimental transformation function, allowing the development of a statistical distribution function for strength based on the known statistics of velocity. Comparison of experimental groups with full sets of corresponding strength1 velocity data showed the technique offers a fairly accurate description of the statistics of strength.

Jan 1, 1986

By J. D. Raulerson, R. P. Farias, G. A. Gruber

A flotation process has been developed to successfully treat ore from the Santo Domingo phosphate deposit in Baja California Sur. The ore contained less than 4% P2O5 with quartz, silicates, and carbonates as gangue minerals. Flotation of the low- grade feed in seawater required that the conventional Crago process, comprising direct anionic flotation, deoiling, and inverse cationic flotation, be augmented by an anionic cleaner flotation step, phosphate depression, and a calcite flotation step.

Jan 1, 1988

By H. S. Kibler, A. E. J. Gallagher, J. J. Sebesta

A review will be made of the nature, value, costs, and factors associated with industrial minerals production in general. Comparisons will also be made between some of these data and similar criteria for metallic mineral production.

Jan 1, 1987

By F. Calizaya

This study summarizes the ventilation requirements for the highly mechanized PT. Freeport Indonesia's underground mine operation in Irian Jaya, Indonesia. The deep ore zone (DOZ) is a coppergold deposit that is being developed underneath the current intermediate ore zone (IOZ) mine. One of the company's current plans is to prepare this mine for a steady ore flow of 25 kt/d (27,500 stpd) by 2003. This study includes the estimation of airflow requirements for various mining activities, the design of major intakes and returns, the determination of airway resistances and specifications of fan duties. It also includes the results of air pressure/quantity measurements in existing airways and the standards stated by the current Indonesian mining law.

Jan 1, 2000

By B. Seegmiller, W. A. Hustrulid, O. Stephansson

Introduction The concept of in-the-wall haulage was formally introduced to the mining community in the early 1970s. The basic concept is simple. Rather than the traditional in-the-pit haulage/access ramp, single or twin ramps driven in the pit wall are used with crosscut connections on the working levels. The in-the-wall pit haulage system could be considered as part of the opening of a new mine, a way of completing to final depth an existing pit, or part of the transition from open-pit to underground mining with the ramp remaining as part of the access. This paper is intended to acquaint the reader, particularly mine planning engineers, with some of the possibilities presented by the concept. Nordstrom (1983) said, "In many countries, mining engineers are either open-pit or underground specialists and, hence, often don't consider the possibilities for combining the best aspects of each into a hybrid mining system." This is believed to be true. There are few "hybrid mines" employing, for example, open-pit methods with underground haulage systems. One reason is the variance with the experience of the mining company and the engineers involved. However, with today's low metal prices, innovative and unconventional thinking may be just what a property requires to remain in production. All the possible situations where an in-the-wall ramp might be considered cannot be presented here. Rather, a few examples will be discussed for illustration purposes. Potential savings in stripping costs A major cost in open-pit mining is associated with the removal of overburden and waste rock. Although much of this stripping cannot be avoided, the volume (V) associated with the presence of an in-pit haul road can be quite large, V 100 (pit depth) a (road width) 2 x road grade (%) As an example, consider a vertical ore body 400 m (1312 ft) long, 60 m (197 ft) thick, extending to a depth of 160 m (525 ft). Cross sections are shown in Fig. 1. Other data are: ore tonnage = 11 Mt (12 million st), ore density = 2.9 t/m3 (0.09 st per cu ft), waste rock density = 2.7 t/m3 (0.08 st per cu ft), road width = 20 m (66 ft), road gradient = 10%, and level interval = 10 m (33 ft). The evaluation has been made assuming that the pit slope (0) is constant around the pit except on the side with the haul road. On this side, the slope is decreased to allow for the road. Slope angles of 30° to 60° have been selected to indicate the dependence of stripping volume (V) on slope angle. The stripping at the ends have been calculated assuming representation by cones. The volume associated with the haul road is 2.6 Mm3 (92.8 million cu ft). A comparison of the strip volumes involved is shown in Table 1. As can be seen, the percent savings in strip volume varies from about 4% to 17% and depends greatly on the slope angle. It is perhaps more interesting to examine the savings in stripping costs that could then be used to cover the cost of the access ramp and the crosscuts for stripping costs of $3/m3 ($0.08 per cu ft). The cost with the road volume is 2.6 million x 3 = $7.8 million.

Jan 1, 1988

By Gordon A. Gross

Banded siliceous manganese-iron bearing sediments recognized as distinctive facies of Algoma type iron-formation constitute low grade resources of manganese in many countries. Manganiferous facies associated with oxide and carbonate facies of iron-formation have iron to manganese ratios ranging from 0.2 to 2 and commonly about one. Manganiferous facies up to 30 meters thick occurring separately or with other iron-formation facies are associated with cherts, carbonates, shales, turbidites, tuffs and lavas. Primary cherty manganiferous facies and their gondite metamorphic equivalents represent significant manganese resources and are protores for higher grade deposits derived by secondary enrichment processes.

Jan 1, 1982

By J. G. Wargo

In situ leaching of disseminate gold deposits has been studied from a theoretical perspective to outline the operating environment for the process. Rock permeability and the relationship of permeability to the site of gold particle deposition are key factors in predicting the efficiency of in situ leaching. Permeability measurements, however, have rarely been made during gold exploration, so that secondary geologic features of the deposit must be used. This paper examines some of the geologic characteristics of disseminated gold deposits and comments on the problems and possibilities of in situ leaching of gold.

Jan 1, 1987

By Hugo J. Garcia, M. C. Fuerstenau, D. M. Lupan, J. L. Hendrix

Biodegradation of cyanide by microorganisms isolated from solids previously exposed to cyanide has been investigated. The anaerobic cyanide biodegradation showed good, short-term removal of cyanide in the presence of nutrients (medium A). This system retained metabolic activity at cyanide concentrations up to 300 ppm at pH 10.5. Microscopic examination of stained specimens revealed that the bacteria were Gram-negative, rod-shaped bacilli. Strains capable of cyanide decomposition were identified as Pseudomonas and Alcaligenes. The kinetics of biodegradation were found to be first order with respect to cyanide concentration, and the overall reaction rate constant is k = 0.102 day-1. Studies with Na14CN showed that 17 % of 14C transformation went into the cells, 18 % into carbonate and 47 % into bicarbonate.

Jan 1, 1995

By John C. Gould

GENERAL DESCRIPTION The Climax mine is located 161 km (100 miles) southwest of Denver on Fremont Pass astride the Con¬tinental Divide. Climax employs 2800 people to pro¬duce 42 670 t/d (47,000 stpd) of molybdenum ore. The Storke level, the main underground level, produces 19 970 t/d (22,000 stpd); the 600 level, the lowest ac¬tive underground area, produces 7260 t/d (8000 stpd); and the open pit presently produces 15 430 t/d (17,000 stpd), Fig. 1. A panel caving method is possible because of the favorable geology and geometry of the Climax ore body. Geology The Climax deposit is composed of three separate ore bodies formed in a massive Tertiary stockwork which penetrates Precambrian granites and schists. The ore bodies are elliptical in plan with overall dimensions of 914 x 1219 m (3000 x 4000 ft). In section the ore bodies are arcuate with an overall height of 457 m (1500 ft), although the depth is not firmly ascertained. Ore grade decreases rapidly into a highly siliceous cen¬tral core and more gradually toward the outer bounda¬ries of the deposit. The mineralization is relatively continuous in the ore zone and contains few dilution-causing waste inclusions. Molybdenite is contained in pre-ore fractures which were largely recemented during the three mineralization pulses. Postore fractures, with a spacing of a few inches to several meters (feet), render the deposit amenable to caving when it is undercut. The peripheral rock caves similarly to the ore, but the central core is competent and is difficult to cave. Caving Feasibility Successful caving action requires an undercut area of approximately 122 by 122 m (400 by 400 ft). When the undercut area exceeds this minimum distance, the arching effect is lost, and the ore column collapses. Ore extensions from the main cave area less than 122 m (400 ft) wide usually cannot be caved. Waste or low¬grade areas within the deposit can be left if they are large enough to carry weight around openings passing through them. Experience indicates that these areas should be over 30 m (100 ft) in length and width. Smaller unprofitable units of waste, such as postore dikes, are mined to avoid weight problems and result in ore dilution. Successful caving and ore recovery are dependent upon timely ore removal from the undercut zone because columns of immobile ore inhibit caving and concentrate weight onto the underlying workings. Climax experience indicates that a finger (drawpoint) spacing* averaging 10.4 m (34 ft) yields an excellent caving action and a 92.5% ore recovery at the estimated grade. An additional 15% of the ore is recovered-at a diluted grade (Fig. 2). Mine Layout The slusher drift is the primary production facility (Fig. 3). The ore is undercut and caved into the fingers,

Jan 1, 1982

By F. Calizaya

A booster fan is an underground fan installed in series with a main surface fan and used to boost the air pressure of the ventilation air passing through it. To accomplish this objective, the fan is installed in a permanent stopping and equipped with airlock doors, a monitoring system, and interlocking devices between main and booster fan controls. The stopping and doors are used to prevent flow recirculation, and the monitors to ensure safe operation of the fan. A booster fan, properly sized and sited, can be used to create safe work conditions and allow the extraction of minerals from areas that would otherwise be uneconomic to mine. In deep and large mines with heavy emissions of air contaminants, the required quantities of air can only be supplied by using high pressure fans. These fans will inevitably induce significant losses of fresh airflow through the stoppings and doors. For well defined ventilation circuits, booster fans can be used to decrease the main fan pressures and reduce the leakage flows. However, this will require a good understanding of the correct ventilation practice, from planning to operation and maintenance of booster fans. The objective of this study is to investigate the conditions under which booster fans can be used safely and efficiently in underground coal mines. Specifically, the study is directed at (1) collecting reliable information on airway resistances and flow requirements from two large U.S coal mines, (2) collecting fan performance data from one or more existing coal mine where booster fans are used regularly, (3) monitoring the performance of booster fans in a laboratory model and an experimental mine, (4) developing a booster fan selection method to assist the ventilation engineer at the planning stage, and (5) training six M.S. or Ph.D. graduate students in advanced mine ventilation.

Jan 1, 2010

By Robert F. Conley

In the mining, manufacturing and application of inorganic pigments, dispersion contributes significantly. This is especially true with mineral pigments because of the variety, in slurry operations required for their processing. Such operations include any of the following: (1) mine-transport (2) wet grinding (3) flotation/beneficiation (4) wet magnetic separation (5) selective floculation (6) size classification (7) chemical processing (8) spray drying Dosage levels and general economics of dispersant use are broadly variant, being dependant upon the agents employed, from the simplist, i.e., sodium silicate or sodium carbonate, to the more complex, as polyacrylates and organo-phosphates. In certain unit processes the dispersant must exhibit unusual chemical stability and persistence, e.g., in protracted pigment slurry shipments. In others, the agent should be fugitive to prevent secondary reactions from occurring, e.g., with high temperature calcination where alkali salts can act as aggregation fluxing agents.

Jan 1, 1985