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By R. J. Roman

Introduction This paper attempts to describe the characteristics of the industrial minerals segment of the minerals industry that make it different from the other segments. Also described are the trends, strengths, and shortcomings in the industrial minerals segment. This may help professionals in the minerals industry determine if there is a place for them in industrial minerals. Industry Overview It is difficult to generalize when describing the industrial minerals industry. Some products, such as cement and potash, are produced exclusively by large, well capitalized and technically sophisticated companies. Others, such as sand and gravel, are generally produced by small, family-owned companies operating on a shoestring budget. Still other products, such as feldspar, are produced by both types of companies. Not only are the companies vastly different, the products and marketing of their products are also different. Industrial mineral products are sold based on their physical characteristics (abrasives), or their chemical characteristics (trona), or both (glass sand). Marketing can be direct by the producing plant, a corporate office handling products form several plants, or through distributors. The geologic, mining, and processing skills needed can be minimal or sophisticated. There can even be large differences in the technical skills needed in two plants producing the same product. One feldspar producer may only need to mine and grind his ore, while another may use a delicate flotation process. One easy generalization is that there are industrial mineral operations in all geographic locations. A second is that industrial minerals offers more opportunities to entrepreneurs than the metals field because of the variety in the sizes and types of operations. Industry Growth Industrial minerals must be considered a mature industry. Consequently, little exciting growth can be expected. Over the past 30 years, industry sales have grown at a rate of 6.2%, compared to a 6.7% growth rate of the GNP (Kliff, 1978). A 3.8% growth rate has been forecast for the remainder of this century. The fortunes of those plants producing a product for local consumption will depend on the prosperity of the local economy. The fortunes of those plants that sell into the national market will depend on their competitiveness in their industry. Major industrial mineral products are shown in Table 1. The first three commodities account for 66% of the industrial minerals market. Typically, deposits of stone, sand and gravel, and the limestone, silica, and clay needed for cement are easy to find and mine. Role of Mining Engineers and Geologists Because a sophisticated techcical knowledge of geology or mining is not required, very few geologists and mining engineers are employed in industrial minerals. Geologists working for these companies function less as geologists than managers and executives dealing with permitting and land acquisition. Likewise, mining is done mostly by open-pit methods. It is generally supervised by an individual experienced in the operation and maintenance of heavy equipment, not a mining engineer trained in rock mechanics or computer methods of mine planning. What of the geologists and mining engineers employed by the industrial minerals industry? Many gained their positions as a result of their experience dealing with the regulatory agencies that control the operations of all mines and mills. Their technical expertise in geology or mining becomes only a side benefit. Finding and mining industrial minerals does not generally require the concentrated efforts of geologists and mining engineers. But the processing of the ores does

Jan 2, 1985

By F. Feichtinger, M. Riess, A. Lammer

Introduction This article discusses the unique range of properties of platinum and its applications as a high technology metal. World reserves of this strategic and precious metal are confined mostly to South Africa and, to a far lesser extent, the Soviet Union. An analysis is given of past trends in world platinum output. The timepath is computed for the world platinum output rate, covering a period of more than 350 years. This is based on the postulate that the industrial conversion of a virgin resource progresses according to a Sigmoid curve. This computation is performed by means of a nonlinear, three-stage optimization model employing a multiform Sigmoid function. Finally, the price dynamics are examined. Platinum is the most important of the group of the platinum metals to which it gives its name. Other platinum group metals are iridium, osmium, palladium, rhodium, and rhutenium. The metals occur in nature in close association and are chemically very similar to each other. Three members of the group - platinum, iridium, and osmium - are the heaviest metals known to man. Platinum's density is 11% heavier than gold and about twice the weight of the same volume of silver or lead. Properties Platinum is a typical grayish-white metal. Its special properties were known 1000 years ago to the American Indian craftsmen of what is now Ecuador and Colombia. From those distant beginnings, platinum is now firmly established as the most precious and distinctive metallic element known to mankind. It is shown in Table 1, platinum has a unique wide range of physical, chemical, and mechanical properties including high melting point, resistance to chemical attack, high catalytic activity, mechanical strength, and malleability. These properties make it indispensable for use in many industries. Application Platinum is first and foremost a high technology metal. During the last few decades, fertile scientific minds searched for and developed widespread cost-effective uses for this precious metal. Its unique properties render it irreplaceable in many key industrial processes. Important major fields of application include: • protection of man's natural environment (pollution control), • defense, • fuel production, • fertilizer manufacture, • glass and glass fiber manufacture, • pollution-free energy generation - replacement of nuclear and fuel operated power plants, • electronics, • dentistry, • medicine, and • jewelry. Besides these industrial markets, platinum is increasingly finding a place in investment portfolios around the world as investors switch from gold and silver holdings to this relatively scarce and strategic commodity. Platinum has a growing popularity among private investors. This is being accelerated by the predicted "big crash" of the US dollar and its associated weak currencies, which American financial analysts are expecting towards the end of the decade. Figure 1 shows the demand for platinum in various applications throughout the western world in 1986. The total demand amounted to 88.6 t (97.6 st) of which autocatalyst usage accounted for 32.8 t (36 st) or 37%, the biggest share. Catalytic converter technology is increasingly centered on rhodium-platinum three-way catalysts. These catalysts are able to deal with all three pollutants - carbon monoxide, hydrocarbons, and nitrogen oxides - in a single catalytic unit. The introduction of these three-way converters will result in a further steep rise in the platinum demand for autocatalysts. Among all the metals used in the antipollution devices for automobiles, platinum remains

Jan 2, 1988

By Tim Neil, O&apos

Large is the applicable adjective here. The largest hydraulic shovel in the world, the largest copper mine in the world, the largest copper producing company in the world. The O&K RH 300 hydraulic excavator has been put into service to move overburden and mine ore at the Chuquicamata open pit mine of the Chilean state copper mining company, Corporation Nacional de Corbre de Chile (Codelco). A large shovel: Equipped with a 26-m3 (34-yd') rock shovel, the RH 300 is able to fill a 170-t (187-st) truck in three passes with cycle times of 30 seconds. A large mine: At present, the Chuquicamata pit is 4 km (2.5 miles) long, 2 km (1.2 miles) wide, and 520 m (1700 ft) deep from rim to bottom bench. The current 450 kt/d (496,000 stpd) of material mined will increase to 600 kt/d (661,000 stpd) when the expanded production rate is reached in 1989. (Chuqui Expansion details, p. 20). A large mining company: Codelco's total output from its two open pit and three underground copper mines is more than 1.1 Mt/a (1.2 million stpy) of copper. This makes Codelco by far the largest copper producer in the world. It accounts for about 15% of the Western World's copper requirements. Chuquicamata is located in the hot and arid Atacama desert of northern Chile. The mine is located 240 km (150 miles) northeast of the port city of Antofagasta. Mine elevation is 2700 m (8850 ft). Chuquicamata's known reserves of oxide and sulfide ores amount to more than 450 Mt (496 million st) of contained copper. There are also reserves of more then 1.4 Mt (1.5 million st) of contained molybdenum. Originally, Chuquicamata was a rail pit. The shovels loaded rail cars hauled by locis on tracks along the benches. During the 1970s, the rail transport was replaced by truck haulage. Crushing and conveying rock waste in the pit is the latest step in the continuing evolution of more productive and economical methods of materials movement. In 1990, the in-pit crushing plant will be in operation. The waste will be mainly transported out of the pit on belt conveyors rather than trucks. The system will incorporate a 9.6-kt/h (10,600-stph) semi-portable crushing station, a 14-km (8.7-mile) conveyor using 1.8 m (6 ft) wide heavy duty belts, and a tripper-spreader to stack the waste dumps. Since Chuqui went into production in 1915, more than 2.5 Gt (2.8 billion st) of material has been excavated. This includes nearly 1.7 Gt (1.9 billion st) of ore from which some 16.4 Mt (18 million st) of copper has been recovered. The Chuquicamata porphyry type ore body is an altered and fractured intrusive granodiorite or quartz-monzonite porphyry that carries the mineralization. The action of ground water on primary sulfide minerals in geo¬logic time has created oxide ores and a zone of secondary enrich¬ment. The oxides occur above the sulfide zone and are also mixed with the upper parts of the latter. At Chuquicamata, the current grade of 1.5% copper is expected to decline to about 1% by 1995. To maintain metal production at cur¬rent levels, it will be necessary to increase ore production rates by about 50%, as mentioned, to 153

Jan 1, 1988

By Eivind B. Jensen

Introduction Though temporarily shutdown due to the depressed economy, the Henderson mine is a large underground molybdenite producer, using a continuous panel-caving mining system. The location of the ore body requires access from the surface through three shafts, the collars of which are located about 3.15 km (10,350 ft) above sea level. At this altitude, normal temperatures are such that heat must be added to intake-ventilation air for about half of each year. This prevents ice from building up in the intake shafts, waterlines from freezing underground, employee discomfort, and equipment-related problems such as pneumatic rock drill icing and cold starting of diesel engines. Because of topographic and environmental constraints, the Henderson mill and concentrator are located 24 km (15 miles) from the mine, on the west side of the Continental Divide. All ore from the mine is transported to the mill by an electrified rail haulage system, with 15.4 km (9.6 miles) of the trip being underground through a double-track tunnel. A shaft at the midpoint of the tunnel exhausts air drawn from the mine and from the portal that is at an elevation of 2.7 km (8,950 ft). Intake air at the portal must be heated for part of the year to prevent ice buildup. Mine Ventilation System The mine ventilation system uses three sources of fresh air and two exhausts. The primary freshair intake is No. 3 shaft. It is concrete lined to a diameter of 7 m (23 ft) for a total depth of 698 m (2,291 ft). An airflow of 755 m3/s (1.6 mil¬lion cfm) is supplied to this shaft by two vane-axial 932-kW (1,250-hp) fans on the surface. The air is distributed within the mine by means of fresh-air distribution levels connected to production and development levels by raises. A secondary source of fresh air to the mine is provided by No. 2 shaft, the primary man and material access. This shaft is concrete lined to a diameter of 8.5 m (28 ft) for a total depth of 946 m (3,105 ft). Downcast airflow is induced by a pressure differential produced by the fans in the system such that net intake and exhaust volumes are equal. The No. 2 shaft flow of 165 m3/s (350,000 cfm) enters the mine network from the production and haulage levels. Exhaust air is moved from the various working levels through connecting raises to exhaust col¬lection levels that return it to No. 1 shaft. No. 1 shaft is the primary mine exhaust. It is concrete lined to a diameter of 7 m (23 ft) for a total depth of 788 m (2,584 ft). Three vane-axial 932-kW (1,250-hp) exhaust fans are located at the collar of No. 1 shaft, producing a flow of 850 m3/s (1.8 million cfm). In addition, an airflow of 71 m3/s (150,000 cfm) from the intake distribution level is exhausted via the haulage tunnel, by means of No. 4 shaft located midway in the tunnel. The main features of the mine ventilation network are summarized in Fig. 1. Ventilation Air-Heat Plants No. 3 Shaft In the early design phase of No. 3 shaft, it was recognized that air heating would be required to maintain underground temperatures above 0°C (32°F) during cold winters in the Colorado mountains. When the air reaches the mining

Jan 1, 1983

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By Bradley Johnson

Introduction Since its creation in 1910, the Bureau of Mines has conducted a wide-ranging program of research on mining technology. Over the years, the Bureau has perhaps become best known for its accomplishments in mine health and safety technology, a field that has been the object of much public attention. Nevertheless, the Bureau's program in other areas of mining technology has always been strong and remains so today. A distinguishing feature of the current effort is its tendency to concentrate on high-risk, high-gain projects. Throughout the federal research establishment, the Administration's policy has been to phase out short-term research that could be performed by private industry, in favor of long-terms projects that industry cannot usually afford to fund. This tendency is evident in the following account of the highlights of the Bureau's current mining technology research program. Underground Mining Technology Fragmentation. Research studies on fragmentation have focused on nonexplosive concepts. If successfully developed, these concepts would lend themselves to continuous automated operation and reduce time and costs of fragmentation in underground mines. A special laboratory-scale device to investigate ripper cutting was designed and fabricated in 1982. Experiments were conducted with large drag cutters to develop an efficient cutting technique for hard rock. Tests were successfully conducted with 76- and 152-mm (3¬and 6-in.) bits in rocks up to 138-MPa (20,000-psi) compressive strength. Because of the success of these initial tests, larger-scale proof-of-concept tests are planned. To prepare for these tests, a tester capable of cutting a 1.8-m (6-ft) opening was fabricated. This device will allow ripper cutting to be carried out at full scale, and will also allow mucking and holddown requirements to be investigated. Preliminary estimates indicate that this concept has the potential to double excavation rates while reducing costs up to 30%. Mobile Hydraulic Monitor. A mobile hydraulic monitor system was designed, fabricated, and used to mine the St. Peter sandstone at an average rate of 22 t/h (25 stph) from a face in Martin-Marietta's foundry sand mine at Clayton, IA. A slurry transport system worked in tandem with jet cutting, resulting in a continuous mining and transport system. Problems with spray from the jet hitting the sandstone showed that a remote-controlled mobile monitor was needed at a station 15 m (50 ft) from the face, so the operator would have a clear view of the face and be safely out of the backwash. Crushers. Research on crushing systems to improve efficiency in underground mines is continuing. Final testing of a low-profile, portable crusher for underground applications has begun. A system for handling oversize feed to the crusher was conceived and tested. Underground crushing systems, portable enough to be kept near the face, offer considerable improvement in materials handling because main haulage by conveyor is often less costly. Hoists. Research to increase hoisting efficiency and speed and to reduce energy requirements has led to experimental ore skips. The skips feature guide rollers to improve riding characteristics, bearing improvement at critical locations, and weight savings of about 1.8 t (2 st). An experimental aluminum skip was installed last year at Kerr McGee's Church Rock mine. Performance is being monitored under production conditions. A portable service hoist for low-cost, efficient movement of men and supplies in raises between mine levels was designed, fabricated, and tested on the surface. The hoist has maximum equipment payload of 570 kg (1,250 lb) at

Jan 11, 1983

By M. J. dos Santos

"INTRODUCTION In his landmark Society for Mining, Metallurgy and Exploration (SME) 2000 writing, “Theory and Design of Sandwich Belt High Angle Conveyors According to The Expanded Conveyor Technology” J.A. Dos Santos generalized the belt conveyor technology to include all possible variations and augmentations including, particularly, multiple belts along a common carrying and/or return path as is the case most notably in the sandwich belt high angle conveyor technology and belt-on-belt type intermediate (booster) drives. Figure 1, excerpted from his writing, represented the proposition of the generalized “Expanded Conveyor Technology,” abbreviated ExConTec.This formed the basis for the generalization which was completed in theory and in math before 2000 and presented by J. A. Dos Santos at the Y2K SME Annual Meeting. The writing can be found in the SME library (Dos Santos J. A., January 2000). While Figure 1 with the supporting mathematical development solved the generalization, it failed to identify the variation that would prove to have the greatest significance. As with most significant developments, necessity is the mother of invention. This is the case with the Adder Snake belt technology, the subject of this writing. BACKGROUND The concept of elevating bulk materials at a high angle using the sandwich belt concept was first introduced in 1951. That introduction did not produce any lasting success. Commercial success was achieved in the 1970s with the Loop Belt, a sandwich belt elevator of C-shaped profile developed by Stephens Adamson of Canada. The Loop Belt was and is the vital element of a conveyor based self-unloading ship system. Such self-unloading ships have had great success achieving unloading rates above 10,000 t/h with Loop Belts of width to 3,048mm (120”). Being strictly of C-profile, Loop Belts could not be adapted to the general high angle conveying path which is predominantly along a straight incline. It was the success of the Loop Belts that inspired the Dos Santos Sandwich Belts of the 1980s. These expanded the Loop Belt capabilities by producing endless elevating profiles that could take the most direct and/or conforming path between the loading and discharge points. Sandwich Belt high angle conveyor systems have seen great success through the years, with more than 200 installations throughout the world."

Jan 1, 2018

By M. S. Bickers

Prior to 1970, the environmental liabilities associated with mine construction and management were few. Mostly, they were aimed at health and safety matters. At that time less than two dozen codes were incorporated into the engineering design and they were somewhat static and well known by the engineering designers and managers. These codes issued by governmental agencies and professional associations are listed in Table 1

Jan 1, 1988

By D. J. Shalanski

In 1962 a computerized data processing system was first installed and programmed for the grading of direct-shipping iron ore from the Sept-Iles terminal of the Iron Ore Company of Canada. Over the years several changes and improvements have been incorporated into this system. It is the purpose of this paper to describe its present operation and features. Sept-Iles is located on the north shore of the St. Lawrence River, 500 miles north east of Montreal. The direct-shipping ore is received by railroad from Schefferville, 320 miles to the north in the Knob Lake Iron Range of New Quebec and Labrador. The shipping season extends from mid-March to December. The maximum tonnage handled in this period being 13,045,000 long tons in 1959. Crude ore receipts for 1974 amounted to 6,568,000 tons of direct-shipping and 4,700,000 tons of plant feed material from Schefferville. This was augmented by 4,447,000 tons of concentrate and 7,572,000 tons of pellets from Labrador City 200 miles north of Sept-Iles. The direct-shipping ore has come from as many as ten mines, and grade variations, even-within the same mines, have been sufficient to require blending for uniform, acceptable cargos. Shipments of low phosphorous, non bessemer and manganiforous iron ore are made, to Canada, the United States, Europe and Japan with grade variations to meet customers requirements.

Jan 1, 1975

By R. R. Johnson

The Hama Mining Company owns and operates The Groveland Mine near Iron Mountain, in the Upper Peninsula of Michigan. The orebody is part of the Menominee Range of the Lake Superior District. The ore (generally known as Jasper taconite) contains an average of 35% iron ranging from 28% to 39% with an overall split of 60% magnetite and 40% hematite. The variations encountered throughout the orebody require a certain amount of blending to reduce feed changes to the mill. One of the variables which affects the crushers, as well as the mill, is the range of hardness which has a low of 7.9 and a high of 24.6 as measured by the Bond Twin pendulum impact and crushing technique. Reference is made to hardness only to point out its effect on crushing and wear rate variations. Over a period of time, these variables do, of course, average out. Mining of the low grade ore started in 1959 and has operated continuously to date. Preparation of the feed for conventional rod and ball mill grinding is accomplished by three stages of crushing. Reduction of crude ore is first passed through a 54 inch gyratory crusher; then through a secondary crusher; followed by two or three tertiary crushers in closed circuit with screens as production quotas require (Figure 1). Current demand to the crushers is about 5,200,000 tons of crude ore per year.

Jan 1, 1976

By R. W. M. Lai

The study of flotation kinetics was first reported by Garcia-Zuniga1 in 1935. Since then, literature on the subject has grown with the time. 2-19 Although there is an abundance of flotation kinet¬ics models available in the literature to describe flo¬tation systems, they are mostly of academic interest and are seldom applicable for use in the industrial environment. Several reasons contribute to the wide gap in the industrial -application of the academic models. These include: 1) the lack of generality in the model, e.g., the model is limited to certain size ranges; 14 2) the academic models involve complicated mathematics; 5,6,10,11 3) confusion about the order of the flo¬tation rate, e.g., several authors proposed a first order rate model, 1,15,16 whereas the others propagated a second order rate model: 2,17,18,194) the use of an unrealistic rigid value, such as the use of a maximum recovery of 100%, 5,8,19 which is unrealistic in industrial operations; and 5) the use of graphs that confuse the reader rather than interpret information. 2,3,10 This paper intends to introduce a simple and easy to understand empirical flotation kinetics model along with a way of presenting the model in an informative graphical form, which should stimulate the interest of other practical engineers to become concerned with practical application within the mineral industry.

Jan 1, 1982

By C. J. Bise

Virtual Reality (VR) is a concept in which the human experience of perceiving and interacting with a computer-modeled environment is achieved through the use of sensors and effectors. It is an advanced fonn of computer graphics where the images are generated based upon where the user looks and how the user moves. As personal-computer-based (PC-based) instruction gains acceptance as an innovative approach to providing safety training, VR offers the capability of providing a multisensory, rather than merely a multimedia, mode of presentation. VR has been effectively used by the military for training tank commanders and by hospitals for training surgeons. However, these applications have been limited to small numbers of highly skilled participants. The purpose of this paper is to examine VR's potential for enhancing the hazard awareness of larger numbers of individuals in a broad-based approach to mine safety training, such as new-miner, annual-refresher, and independent-contractor training.

Jan 1, 1996

By W. A. Zingg

Impervious zones in copper waste heaps create a problem in contacting the rock with leaching solutions. These zones are normally associated with compacted layers .formed on the surface of each lift by dump truck traffic and by disintegration of the rock exposed to atmospheric conditions. These layers are usually only a few feet in depth--but, nevertheless, act as barriers to vertical flow causing the leach solution to perch in the heap, Consequently, much of the rock remains untouched and the leach solution breaks out the side of the heap instead of continuing downward to bedrock. One solution to this problem is breaking the hard pans with explosive charges. However, to use explosives means adopting new explosive materials and methods. To use this process successfully requires finding answers to operating conditions inherent to waste heaps. Foremost among these are: (1) type and size of casing that can be used because of limitations brought about by drilling problems, caving and, corrosive conditions; (2) perforating the casing at required levels; (3) an explosive capable of being pumped; (4) obtaining detonation of the, main charges through, the wail of the casing; (5) an explosive with the ability to propagate through the voids of the rubble and one which is not affected by water, acidic solutions, or back pressure conditions from hydrostatic heads; (6) to heave and rupture the compacted zones without chambering or recompressing the waste material.

Jan 1, 1970

By R. Woods

INTRODUCTION Gaudin (1), considered that "the mechanism of mineral collection is the central problem of flotation theory". From their work on adsorption phenomena in flotation for more than fifty years (2), Gaudin and his co-workers have made major contributions to the elucidation of this mechanism, a fact acknowledged by their contemporaries in the field (e.g. (3)). Their investigations in a number of areas, e.g. demonstrating that submonolayer quantities of adsorbed collector species can induce flotation, or identifying the influence of substances which can activate or depress (1), have helped to provide the basis for and the development of a theory to account for the interaction between mineral and collector. It was established in the late fifties (4) that the presence of oxygen is a prerequisite for the flotation of metal sulfides with thiol collectors. This fact rendered inadequate the concept of simple ion adsorption being responsible for rendering these minerals hydrophobic. The requirement for oxygen has been explained (1) in terms of this species being a component in chemical reactions between the sulfide surface and the collector. Gaudin has been instrumental in identifying the products of chemical reactions at sulfide minerals (1.5), a line of work which has been continued particularly by Finkelstein and co-workers in South Africa (6). In presenting a detailed mechanism of the mineral-collector- oxygen interaction, it has often been considered that the mineral surface - must first be oxidized to a metal oxy-sulfur species which

Jan 1, 1976

By R. S. Rickard, B. Ball

INTRODUCTION Pyrite responds to the flotation reagents normally used in sulfide flotation systems. For example, it is well known that xanthates are good collectors for pyrite, and cyanide, sulfide and hydroxides are depressants. The particular response of pyrite to these and other reagents is used to advantage in several areas. By careful control of depressant and collector additions, pyrite can be separated from other sulfides by flotation. This is of obvious importance in base metal sulfide flotation where pyrite depression is often required to produce a marketable grade concentrate. Flotation of pyrite has long been employed in gold ore processing, and it appears that pyrite flotation from coal may become important in the future (1). It is probably a direct consequence of this relative ease of separation, plus the low economic value of pyrite, that detailed investigations on the chemistry of its flotation and depression are relatively limited. Gaudin (2,3,4) and others, in the mid-1950's published data on the flotation of pyrite, but detailed or consistent mechanisms were lacking. In 1955, Sutherland and Wark's book (5) was published, in which a large amount of empirical information, especially earlier work of Wark and Cox, was presented. More recently, in 1968, significant contributions have been made by Fuerstenau et al. (6,7) and Majima and Takeda (El. These papers introduced electrochemical explanations of the effects observed with xanthate and cyanide ions.

Jan 1, 1976

By T. J. Close

The Silver Peak Range North Wilderness Study Area was deleted from the list of US Bureau of Land Management lands in Nevada to be considered for wilderness designation. This decision was based in part on the US Bureau of Mines mineral investigation of the wilderness study area which found indications of precious- and base-metal-resources. The study area lies between the Silver Peak and Dyer mining districts of west-central Nevada. The Bureau examined 13 mineralized sites, most of which were previously unknown and unclaimed. These sites are along two zones that cross the study area from the Silver Peak mining district to the Dyer mining district. The two zones transect a volcanic center of fractured and altered rocks that contain epithermal veins of quartz and calcite with gold, silver, and base-metals. At the surface, these veins are poorly defined. Exposures of the more than forty veins examined were as thick as 17 ft and as long as 1000 ft. Samples from these veins contained significant amounts of precious- and base-metals-as much as 0.02 oz per ton gold, 2.1 oz per ton silver, 1.0% copper, 2.6% lead and 0.4 % zinc. No vein large enough or of high enough grade to be classified a mineral resource was disclosed by the sample data. However, most do have indications of precious- and base-metal-resources at depth. Subsurface exploration of some of the veins is recommended.

Jan 1, 1995

By S. G. Hardcastle, Akbar Sheikh

In Canadian underground uranium mines, bacterial leaching is being increasingly applied to low-grade ore reserves. At Denison Mines Limited uranium operations, old stopes are used as leaching vessels; these contain a leaching muck pile brought down by drilling and blasting from the back which contained a low-grade reef. The bacteria used require atmospheric oxygen to oxidize the pyritic ore. A study was initiated to evaluate the ventilation, including oxygen supply, specifically, of flood leaching stopes during their rest period. The method of ventilation investigated was a single ended compressed air system which feeds airlines laid under the muckpile and carried through the inlet pressure bulkhead. The air delivery system was tested for post- blast integrity with a sulphur hexafluoride, SF6, tracer gas technique. This was injected into the intake compressed air line and collected at the stope's borehole exhaust in grab-sampling bags. The samples were analyzed with a laboratory gas chromatograph and results provided information on the stope's airflow, average residence time, overall clearance time, volume of the airpath and the airflow distribution. The borehole exhaust was monitored by anemometer and psychrometer and indicated that the psychrometric conditions in the stope were independent of the area's ambient climate. Evaluation of the exhaust's standard gaseous composition showed the oxygen consumption of the bacteria to be independent of the investigated flow range. Field derived results are compared with computer models, theoretical predictions and laboratory investigations.

Jan 1, 1987

By R. E. Brewer

Pilot scale flotation tests were conducted with a 1-ft-diam column at the Barrick Resources (USA) Mercur mine, Tooele County, UT, to scavenge unleached (0.01 to 0.018 oz/st) gold values from the tailings of a carbon-in-leach (CIL) circuit. The optimal recovery was 35 pct of the unamenable-to-leaching gold, at a grade of 0.353 ounces of gold per short ton, using a mixture of dithiophosphate, xanthate, and diesel fuel as a collector. This was achieved by floating approximately 1.6 wt pct of the CIL tailings. The gold values are believed to be contained in arsenic-bearing minerals, attrited carbon, and carbonaceous material in the ore. Redox potentials were measured with a gold electrode during the conditioning stage prior to flotation. The correlation between flotation recovery and redox potential values could be used to control reagent additions. Treatment of a previously leached and weathered tailings material, averaging about 0.055 ounces of gold per short ton, was also investigated using the column. The tailings were first water leached at approximately 600 C, and the water-leached tailings, which contained approximately 0.03 ounces of gold per short ton, were then fed to the column flotation unit. The effects of pH and sodium sulfide addition on gold flotation recovery and redox potentials were examined. maximum gold recovery, 36 pct, was achieved at pH 8.5 with a sodium sulfide addition of 0.18 lb/st.

Jan 1, 1989

By George Aul, Richard Ray

Short-hole horizontal degasification systems have been utilized in Island Creek's Virginia Division mines since 1986. These systems remove methane from the coal seam in advance of longwall mining. Without an in-seam methane drainage program, methane liberation rates from the longwall faces may exceed 0.47 cu m/s (1,000 cfrn), requiring 47 cu m/s (100,000 cfm) of air for adequate dilution capability. Optimization of the degas systems has enabled methane liberation levels and the associated ventilation requirements on the longwall faces to be reduced by 50 to 80 percent. Since the inception of the first degas system at the Virginia Pocahontas No. 6 Mine, Island Creek has developed an overall methane drainage program designed to maximize gas production and minimize ventilation requirements for methane dilution. Gas production from the degas systems is affected by four main design criteria: (1) hole location, (2) hole length and spacing, (3) collection system capacity, and (4) time available for drainage. Drilling equipment and materials for the gas collection system are selected to meet the necessary criteria while minimizing manpower requirements. Benefits gained from reducing methane liberation rates from the longwall faces are fourfold: (1) improved safety due to lower methane concentrations in the face area, (2) fewer production delays caused by excessive methane concentrations, (3) lower fan operating costs resulting from reduced air requirements, and (4) enhanced worker comfort resulting from lower air velocities and less dust generation.

Jan 1, 1991

By M. V. Chari

A prime objective of physical coal cleaning is the separation of raw coal into a clean coal product with the required low ash and pyritic sulfur content and a high ash refuse. The refuse should also contain a minimum amount of combustible matter. In order to be effective, cleaning processes require that the mineral matter in the feed, including pyrite, be largely discrete and free particles, liberated so that the particles are not attached or embedded in coal. Liberation of mineral matter from coal, particularly pyrite, was investigated by Bechtel as part of two fine coal cleaning research programs sponsored by the Department Of Energy (DOE), Pittsburgh Energy Technology Center (PETC). Two proof-of-concept (POC) plants, each with a capacity of 900 kg per hour, were installed and tested to evaluate production of ultra-clean coal using microbubble flotation and spherical agglomeration technologies. The plants were designed to produce clean coal with 2 percent ash and pyrite content lower than that achievable by conventional methods. Also, the quality targets were to be attained at energy recoveries above 80 percent. Bench-scale flotation and agglomeration tests indicated that it would be practical to grind the 6 mm top size feed coal to a mean particle size (050) of 10 micrometers in the POC test plant. In this part of the paper, the steps taken in the grinding circuit to maximize liberation of mineral matter are described.

Jan 1, 1991