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By Richard R. Klimpel

Introduction Froth flotation is one of the most common unit engineering operations in use today. Its aim is to upgrade the quality of coal by removing ash/pyrite and separating selected minerals from undesired gangue materials. Flotation has been successfully practiced at the industrial scale for more than 50 years. Its use is increasing due to the ever decreasing feed grades of the coals and minerals being processed. Flotation has also shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a given equipment configuration. The operation must exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the mechanical flotation cell is scaleable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required - pH regulators, frothers, collectors, activators, and depressants - are all relatively inexpensive and common. They are not usually used in large quantities. Froth flotation is certainly one of the most versatile and forgiving unit operations in the chemical, mining, and agricultural industries. The problem with the industrial flotation process is that it is fairly easy to get reasonable results and relatively difficult to get excellent results. Why is it so difficult to fine tune the froth flotation process? There are some obvious reasons and a few not-so-obvious factors. The obvious category would certainly contain the tendency of natural feed materials (ores and raw coals) to have variable physical, metallurgical, and surface properties. This occurs even when special care is taken to minimize such differences with blending. There is another obvious reason for difficulty. An inherent broadness of the feed particle size range to flotation results from using any of the common industrial-scalefine grinding devices - ball-, rod-, or autogenous mills. Research has been done into the chemical reagent and equipment aspects of flotation. It has shown strong interactions (limitations) of chemical reagents and equipment on larger and smaller particles. The point is that current flotation practice works best on average- or medium-sized feed particles. The greater the amount of large or small particles, or of both large and small, the more difficult it is to achieve excellent flotation results. Other reasonably obvious factors include: • variations in mineral type and liberation sizes; • water chemistry variations such as those involving Ca", C03, or Fe", or various soluble metal cations; • the presence of stray organic materials such as returning soluble reagents or partially degraded reagents in reuse water; • the lack of reliable on-line instrumentation for monitoring and control; • frequent equipment malfunctions of all types; • excessive feed rates for installed plant capacity; • poor operator interaction with the process; and • poor reagent dosage control. There are less obvious, but real limitations to flotation. These are inherent in the design and operating characteristics of the most common cell design in use today - the mechanical flotation cell. Individual performance attributes of the chemical reagents used, along with their manner of use, also carry limitations. The bulk of this paper will deal with chemical reagents and their use. Chemical reagents In 1979, the author and his co-workers were asked to invent and implement, at the industrial scale, new flotation reagent chemistry. After reviewing some of the literature (Sutherland and Wark, 1955; Klassen and Mokrousov, 1959; M. Fuerstenau, 1962; D. Fuerstenau, 1976; and Leja, 1982), a series of carefully controlled plant scale tests were set up. These tests were designed to understand existing reagent scale-up from the laboratory to the plant. They were to quantify at the plant level the effect of some of the more important, controllable chemical reagent factors (frother and collector type, frother and collector dosage, pH effects). This is a learning experience that all researchers involved in flotation theory should experience. It quickly became apparent that: • because a factor is important

Jan 12, 1988

Power cable costs are only a small part of total mining costs. So many mine operators consider power cable failure and resultant downtime as part of the cost of doing business. But, viewed in terms of lost production, these costs can be quite significant. Now one company, Cablec, seeks to cut cable costs by upgrading the polymer compounding process used to make cable insulating and semiconducting materials. Cablec is the leading manufacturer of electrical power cables in North America. And with about a third of the market, Cablec is the largest supplier of power cable to the mining industry in the United States. To improve its products, Cable has entered the polymer compounding business. In July, it began producing insulator and semiconductor polymer compounds at its plant in Indianapolis, IN. "This new facility provides a quantum leap over conventional compounding methods," said Harry C. Schell, Cablec's president and chief executive officer. "The Cablec polymers plant is producing a dramatically higher standard of polymer compounds that provide significantly higher levels of performance and improved life cycle costs for power cable." Cablec faces tough foreign competition in the wire and cable business. Competing on price alone is difficult, particularly when foreign producers are state subsidized. So Cablec feels the best way to compete is to establish new quality production standards. The company's new polymers plant is one way to do this. By increasing purity control and uniformity in polymer compounding, Cablec says its power cables will last longer and fail less often. A typical medium voltage cable consists of a conductor, conductor shield, insulation, insulation shield, metal shield, and jacket. The conductor shield and the insulation shield are conducting polymers. Contaminants and imperfections can occur within the insulation, at the conductor shield/insulation interface, or at the insulation shield/ insulation interface. Over time, these contaminants and imperfections can decrease the electrical strength of the cable or cause premature cable failure. The effort to minimize the number and size of any possible contaminants begins with pure polymer compounds mixed in a clean facility. However, most power cable manufacturers manually handle raw materials, use ethylene/propylene (EP) in bulk bales, and mix polymercompounds in open Banbury mixers. The quality and uniformity of polymer compounds is also impacted by temperature variations in the mixing process. This results in wide gradations of product consistency from batch to batch and ultimately contributes to power cable failure. Cablec says the improved polymer compounds from its state-of-the-art plant will be the purest and most consistent insulating and semiconducting materials available. The plant itself RCA spent $18 million to build Cablec's Indianapolis plant. RCA used the facility to mix specialty polymer compounds used to make video disks. RCA had two considerations in mind for the plant, cleanliness and uniformity of the compounds. However, when the video disk market failed to materialize, RCA sold the 46.5 dam 2 (50,000 sq ft) plant to Cablec for $3.1 million. Cablec invested an additional $3 million for modifications and increased production capabilities. Today's replacement cost for such a facility is estimated at $30 million. Cablec says the plant will set a new standard for performance and be economically difficult to duplicate anywhere. One of the essential elements of the plant's clean process environment is the air intake system. It filters contaminants greater than 2 um, less than one-fiftieth the current industry standard. All material handling and conveying areas in the facility are air-locked. This keeps out contaminants such as smoke, dust, and pollen. Banks of pneumatic pumps move polymer components through the system and continually filter the air. The plant also has a backup air intake system. No process downtime due to pump failure here. From the time raw material enters the plant, it is stored, transported, and processed in filtered air by an airtight stainless steel system. The stainless steel resists rust and corrosion. This further eliminates the danger of contamination from paint or rust particles in the conveyance network. A computer system allows a single operator in a central control room to monitor every aspect of the compounding process from air quality to line speed. The computer

Jan 12, 1988

By S. Hodgson

Introduction The Red Dog zinc/lead deposit in northwest Alaska is the second largest zinc deposit ever discovered. Being developed by Cominco Alaska, with startup scheduled for early 1990, it will ultimately be the largest zinc mine in the Western World. Table 1 shows the reserves of the deposit. Bob Baker, a local bush pilot, discovered Red Dog in 1968. The US Geological Survey assessed the find. But the Red Dog site had been withdrawn from staking by provisions in the Alaska Native Claims Settlement Act of 1971 (ANCSA). Under the terms of ANCSA, the deposit was selected by the NANA Regional Corporation, one of 13 native corporations established by the Act. Then, in 1980, the US Congress confirmed this selection, in the Alaska National Interest Lands Conservation Act. In 1982, Cominco and NANA reached an agreement that has led to the development of the mine. Under the agreement, Cominco leases the property from NANA, develops and operates the mine, and markets the concentrate produced. In return, NANA receives royalties and, after capital recovery, participates in the net proceeds from the operation. The Alaska State Government is also playing a role in the development of the mine. Under the auspices of the Alaska Industrial Development Authority (AIDA), the state will construct, own, and operate the DeLong Mountains Regional Transportation System. This consists of a 84-km (52¬mile) road from the minesite to the Chukchi Sea. Located here are the port facilities necessary to store and ship the annual concentrate production and receive the mine fuel and supplies. For using this transportation system, Cominco will pay an annual toll fee. Over the anticipated 50 year life of the operation, the fee will provide a 6.5% return to the State of Alaska. In late 1986, Cominco's Board of Directors approved the Red Dog project. And, during the summer of 1987, road construction began. In February 1988, initial stripping began at the ore body and construction started at the millsite. Project setting The project site is approximately 1000 km (600 miles) northwest of Anchorage, 84 km (52 miles) east of tidewater on the Chukchi Sea, and 192 km (120 miles) north of the Arctic Circle (Fig. 1). The countryside consists of low, rolling hills that lead into mountains north and east of the deposit. The vegetation is typical tundra mat. And the site is in an area of permafrost. However, the permafrost is relatively warm and some areas have been found to be thawed. Examples would be creek bottoms and the site of the main waste dump.

Jan 12, 1988

BP Minerals America (BPMA) has completed its $400 million modernization of Kennecott Utah Copper's Bingham Canyon mine. Modernized facilities include the in-pit crusher, conveyor, ore storage area, concentrator, ore pipe-line, and tailings pipeline. With modernization, Kennecott will achieve substantial cost reductions. Bingham Canyon is now one of the lowest cost copper producers in the world. And the mine will continue to provide jobs and income in Utah for another 30 years or more. The Modernization Project began in February 1986. The work took 33 months. In late 1987, the construction force peaked at 1100 workers. Kennecott Utah Copper now employs a permanent work force of 2100 employees. Less equipment and more efficiency The modernization work has increased the scale of the facilities that transport and process copper ore. A large, continuous transport system has replaced multiple haulage units. Fewer but larger pieces of equipment have replaced many crushing, grinding, and flotation units. The result is less maintenance, fewer employees, and improved metal recoveries. About one-third of the $400 million modernization project is directed at improving ore movement by eliminating rail transport. The new concentrator accounts for the remaining two-thirds of the modernization investment. The modernized plant was started up in stages. In February, ore was fed to the crusher. By the end of March, the three grinding lines were operating and the flotation circuit was producing copper concentrate. In July, the design through-put rate of 70 kt/d (77,000 stpd) was reached. October saw the achievement of targeted metal recoveries that corresponded to the fully modernized operation. It is estimated that Bingham Canyon will annually produce 181 kt (200,000 st) of refined copper, 9.3 t (300,000 oz) of gold, 71 t (2.3 million oz) of silver, and 5.4 kt (12 million lb) of molybdenite. Current Bingham Canyon production is about 97 kt/d (107,000 stpd) and a similar amount of overburden. Some 70 kt/d (77,000 stpd) of ore are processed through the in-pit crusher and 27 kt/d (30,000 stpd) of ore are transported by rail. The Bingham Canyon mine has yielded more than 11 Mt (12 million st) of refined copper. In 1863, precious metals - silver-bearing lead - were first discovered in Bingham Canyon. What was once a mountain is now a huge open pit. In fact, the mine is the world's largest man made excavation. The pit is some 4 km (2.5 miles) wide and 0.8 km (0.5 miles) deep. Since the pit began operating in 1906, more than 4.5 Gt (5 billion st) of material have been removed. Modernized facilities Electric 21 m3 (27 cu yd) or 23 m3 (30 cu yd) shovels are used to load the haulage trucks. Both 154 t (170 st) and 181 t (200 st) trucks are used to haul the ore to a mobile, in-pit crusher. The crusher replaces three crushers located at the old concentrators. There are two dumping locations at the new crusher, which can handle two trucks a minute. The 1.5 x 2.7 m (60 x 109 in.), 745 kW (1000 hp) Allis-Chalmers gyratory crusher reduces the ore to 254 mm (10 in.) size or less and delivers it to the conveyor. The crusher and conveyor systems are capable of handling 9 kt/h (10,000 stph) of ore.

Jan 11, 1988

By T. B. Jarboe, H. Wen

Introduction The 150 MW circulating fluidized bed (CFB) power plant described in this article is the product of recent Bechtel experience in the industrial power and steam market, where the economic feasibility of fluidized bed projects has been proven. The ability of CFB boilers to burn low grade fuels makes it possible to burn low ranking coals or coal wastes. These include anthracite culm, coal mine rejects, and coal wash-plant-waste slurries. This article also addresses the impact of fuel characteristics on plant design. This includes solids handling, steam generation, sulfur capture, and auxiliary load requirements. The uncertain economic environment in which utilities must now operate has shifted the industry's interest away from large, base-loaded generating units to small, more versatile installations. Utilities now consider purchasing new generating capacity in smaller increments and staging its deployment to correspond with demand increases. In this way, they lower their risk in the increasingly hazardous art of long-range load forecasting. This strategy has many advantages. Multiple, smaller units have inherently greater turndown capability and better availability than a single larger unit of the same size. In addition, the smaller units will each have better cycling capability. And, with CFB combustion, the smaller units will have greater fuel flexibility. General plant description The 150 MW powerplant described here is a grassroots CFB unit. Fluidized bed technology differs from conventional combustion techniques. With CFB, solid fuel, together with inert material such as sand, gypsum, limestone, or ash from the fuel, is kept suspended in the combustion chamber. This is done by using the action of the fluidizing air distributed below the bed. Fluidization promotes the turbulent mixing conditions required for good combustion. The turbulence causes the entire mass of solids to behave like a liquid. The result is improved mixing. This allows heat to be generated at a lower and more uniformly distributed temperature that is below the ash fusion temperature of most coals. The fluidizing mechanism offers several other advantages over conventional combustion. There is less volatilization of alkali compounds, which cause deposits on boiler tube surfaces. And there is less sensitivity to the quantity and nature of the ash in the fuel. The low combustion temperatures prevent vitrification of the ash particles and minimize problems associated with furnace slagging. The ability to select from many fuels is one of the major advantages of CFB combustor technology. Fuels that can be burned in CFB plants to produce steam and electricity include: low grade coals with high ash or sulfur, lignite, peat, wood and pulp products, biomass, refuse, anthracite culm, petroleum coke, and oil shales. The plant's turbine cycle uses superheated steam at 12.4 MPa (1800 psi) and 538°C (1000°F), with single reheat to 538°C (1000°F). Two high pressure and three low pressure, closed, feedwater heaters are furnished. So is a deaerator, which is an open, direct-contact heater. The CFB plant exceeds EPA's New Source Performance Standards (NSPS) for air emissions. The design limits sulfur dioxide (SO,) emissions to 0.25 kg/GJ (0.6 lb per million Btu). SO2 reduction occurs in the combustor by reac-

Jan 11, 1988

By Kelly F. Bailey

Introduction On June 20, 1986, the Occupational Safety and Health Administration (OSHA) published revised asbestos exposure standards for general industry and construction. The standards reflect OSHA's attempt to adequately control workplace exposures to minerals it considers carcinogenic - minerals capable of causing or contributing to cancer. These standards specifically identify asbestos as: chrysotile, an asbestiform serpentine mineral; and the amphibole minerals amosite, crocidolite, tremolite asbestos, actinolite asbestos, and anthophyllite asbestos. Each of these has a more common nonasbestos mineral analog that exists in nature in a crystalline, blocky shape rather than the hair-like or fibrous shape of asbestos. The mineralogical names for three of these nonasbestos minerals are unique: antigorite for chrysotile, cummingtonite-grunerite for amosite, and riebeckite for crocidolite. The other three nonasbestos analogs do not have unique mineralogical names. They are simply designated as actinolite, tremolite, and anthophyllite without the word asbestos following their names. The 1986 OSHA standards not only cover exposure to the six asbestos minerals, they also cover specifically the nonasbestos forms of actinolite, tremolite, and anthophyllite (AT&A). The new standards regulate these minerals exactly like asbestos (OSHA, 1986). The construction aggregate industry views this as a major problem because these nonasbestos minerals are common amphibole rock-forming minerals in the earth's crust. They exist in small quantities over large areas of the United States (Kuryvial et al., 1974). These minerals, unlike asbestos, are not mined for a specific commercial purpose. They are unavoidable components in much of the aggregate used for construction throughout the US. They are also common in the gangue material of metallic ores. There are areas of the US where amphibole-bearing bed¬rock is common. Not every rock mass in these areas contain amphiboles, however. It does mean, though, that amphiboles are physically compatible with many of the rocks in those areas. And given the correct geochemical conditions, they will be present primarily in the nonasbestiform variety. In addition, these amphiboles will probably exist in the natural drainage system, sand and gravel deposits, stream sediments, lake shores, valley basins, or ordinary beach sand within these areas. There has been little quantification of nonasbestiform AT&A in dusts and soils in the US. This is not surprising since these nonasbestiform minerals are not commercially valuable. However, an example of the pervasive nature of these minerals can be found in a 1981 Geological Society of America publication where about 0.7% tremolite-actinolite was found in the desert dust in and around Tempe, AZ (Pewe, 1981). Since OSHA standards treat these common nonasbestos minerals as carcinogens in the same way as asbestos, large natural areas in the US are implicitly being labeled as hazardous by OSHA. When a substance is identified as a carcinogen, another OSHA standard comes into play, the Hazard Communication standard. There are also right-to-know laws in 9 states that essentially duplicate this federal standard. These standards require that a product containing 0.1% or more of an OSHA-designated carcinogen be labeled as such (OSHA, 1983). This means that much of the stone and sand gravel products occurring naturally and mined in the US could be labeled a carcinogen when, in fact, they are not. The National Stone Association (NSA) and the domestic construction and mining industries believe that OSHA has seriously erred. The NSA has studied the health, mineralogical, technical, economic, and legal basis for OSHA's action. These studies concluded that there is no justification for the agency regulating nonasbestos minerals as if they were asbestos. Health issues The preamble to OSHA's 1986 asbestos standard states that evidence for asbestos-like health effects from exposure to nonasbestiform varieties of AT&A is inconclusive (OSHA, 1986). The fact is, not only are the data inconclusive, they are nonexistent. During 1986-1987, NSA's occupational health and epidemiology consultant, Environmental Health Associates (EHA), reviewed all available health studies related to AT&A. EHA found evidence that malignancies in both experimental animals and humans are associated with the asbestos forms of these minerals. No experimental or epidemiological evidence was found that indicated such pathogenic effects occur from exposure to nonasbestiform varieties of these minerals. There are relatively few scientific studies of the health effects of exposure to nonasbestiform varieties of AT&A. In three different animal studies, exposure to either nonasbestiform tremolite or actinolite did not result in pulmonary fibrosis on in excess tu-

Jan 11, 1988

By Tim Neil, O&apos

Finnish mining traditions date back 400 years. The country was then one of Europe's major iron producers. Most of the mines in current operation, however, were started during the last 20 to 30 years. The Finnish mining industry must make do with limited resources. Deposits are usually low grade and small in size. Open-pit methods are used where possible. But thin ore bodies and harsh winters can impose severe restrictions. Underground mining is often the only option. Despite these handicaps, Finnish metal miners produce copper, nickel, zinc, vanadium, iron ore, cobalt, and ferrochrome. Ironically, the success of Finnish mining is rooted in a crisis the industry faced several decades ago. Mining costs were rising. High grade mineral deposits were nearly exhausted. It was vital that Finland develop new mining machinery and methods that afforded higher productivity. Cooperation between the Finnish mining industry and its equipment manufacturers was good. Partly as a result, the Finnish mining industry has turned itself into one of the most modern and efficient in the world. The Finns did this while continuing to meet the stringent safety and environmental requirements in force in the Nordic countries. Today, the share of domestic raw materials used by the Finnish metallurgical industry is decreasing. So the country increasingly depends on imports of ores and concentrates. To better provide these imports, Finnish mining companies have expanded their foreign operations. And, while importing raw materials, the mining industry has been exporting its technological and equipment know-how and expertise. Mining methods Conventional mining methods and machinery are used in surface mines. Drilling is the exception. It is done with Tamrock hydraulic track drills and 100 to 150 mm-diam (4 to 6-in.-diam) drill holes. Pumpable slurry explosives are used in larger open pits. Underground, sublevel stoping with open stopes is the most common mining method. It accounts for about 80% of all underground stoping. The main benefits of this method are low production costs, high mechanization, and easy control. The best productivity figures exceed 10 t (11 st) per man-hour, including all mining personnel. The development of cable bolting has made it possible to use sublevel open stoping even in relatively weak rock. Cut-and-fill methods, such as concrete-pillar stoping, inclined-wall stoping, and traditional cut-and-fill have also been used, combined with sublevel caving and room-and-pillar methods, according to local demands. Various fill materials and stabilization components have been investigated and tested for cut-and-fill methods. Sublevel open stopes are often later filled, to stabilize the surface or to allow pillar recovery.

Jan 10, 1988

By R. N. Chakravorty, R. J. Kolada

Introduction Spontaneous combustion in coal mining is recognized as a major hazard in terms of safety and economics. A survey by Olaf (1979) revealed that 80% of coal mine fires were caused by spontaneous combustion. Modern mining methods demand a high capital investment. Closure of a working face may then result in a huge loss. Even a short production interruption can prove expensive. Spontaneous combustion may also produce lethal amounts of carbon monoxide, thus posing a serious threat to the health and safety of underground miners. The worldwide recession in the early 1980s resulted in a number of mine closures and production cutbacks. So large stockpiles at mine sites were common. Experience has shown that stockpile fires are difficult to extinguish. Critical factors influencing spontaneous combustion There is no simple method of predicting whether a particular mining method applied to a specific coal seam will result in spontaneous combustion. Under favorable conditions of heat accumulation, spontaneous combustion can occur with any type of coal. It is known that subbituminous coals are more susceptible compared to low volatility bituminous coals, and that a high ratio of exinite increases the risk. However, the inherent moisture content of the coal and the exposed surface area may be more significant. Actual mining and ventilation conditions also strongly influence susceptibility. In practice, the experience gained with the mining method applied to a particular type of coal seam must be carefully evaluated with respect to the factors responsible for spontaneous combustion. Feng, Chakravorty, and Cochrane (1973) summarized the critical factors influencing spontaneous combustion in coal mines (Table 1). Prevention methods and strategies Mine layout is the most important factor preventing spontaneous heating. Working districts should be designed so that a particular section can be isolated on short notice without affecting production in others. Sites for preparatory seals should be identified and marked during planning. With coal seams that are highly susceptible to spontaneous combustion, the retreat method of working is preferred. This is because the gob is not subjected to a large difference in ventilation pressure. Coal pillars should be designed to resist excessive crush. The panel system of working is preferred, particularly for mining coals susceptible to spontaneous combustion. The size of each panel is calculated based on the incubation period and the rate of extraction. Controlling air leakage is also important for controlling spontaneous combustion. This is usually achieved by minimizing ventilation pressure difference, proper siting of doors, and avoiding unnecessary stopping and starting of fans. Air leakage is also reduced by sealant coating, injection of cements in cracks and crevices, and by maintaining good gateside packs. Application of inhibitors to control spontaneous combustion met with limited success. Chamberlain and Hall (1973) concluded that borate solution or Montan powder sprayed in the gob area or infused into a low rank coal would reduce the possibility of spontaneous combustion. Early detection of heating Significant research is being conducted worldwide to develop methods for early detection of heating. Monitoring combustion products during the early stage of heating is the most suitable method for detecting spontaneous combustion. The major gaseous products resulting from low temperature oxidation of coal are carbon dioxide, carbon monoxide, and water vapor. In mining operations, carbon dioxide monitoring is considered unsuitable because there are sources other than spontaneous heating from which this gas may originate. Carbon monoxide, however, is a more specific product of combustion during low temperature oxidation of coal. Over the years,

Jan 10, 1988

By Steven J. Luzik

Introduction From 1973 through 1986, a total of 12 fires involving oil-flooded rotary screw compressors (OFRSC) were reported to the Mine Safety and Health Administration (MSHA). Most of these incidents were attributed to faulty or by-passed safeguards, or improper maintenance. The cause of a fire in an OFRSC is unique when compared to other types of compressors. A compressor inspection checklist is provided here that can be referenced when conducting regularly scheduled maintenance operations. This checklist and a well-planned program of periodic training and maintenance should help minimize the probability of fires and explosions in OFRSC's. Operation A rotary screw compressor operates on the principle of positive displacement of air. Compressed air in this type of compressor is produced by the meshing of two helical rotors, or screws, as they rotate within a steel housing called a stator. The male rotor normally has four lobes and is driven either directly or indirectly - through a gear assembly - by a motor. It intermeshes with a female rotor having six cavities. Figure 1 explains the compression mechanism: (1) Atmospheric air is drawn into the stator by the suction action of the two rapidly rotating rotors. (2) As the rotors rotate past the intake port, air is trapped in the stator and in the cavities between the rotors. (3) Further movement of the rotors causes an increased reduction in volume, compressing the entrapped air. (4) Compressed air is expelled at the discharge port in the bottom of the stator at a pressure and quantity determined by the particular compressor rotor design parameters, such as length, diameter, and speed. Radial bearings on the male and female rotors compensate for radial loading on the shaft and position the rotors within the stator. Thrust bearings on the discharge end of the compressor prevent rotor misalignment due to thrust developed during the compression cycle. This design prevents the male and female rotors from making contact with each other or the stator. A clearance of 0.07 to 0.1 mm (0.003 to 0.005 in.) exists between the two rotors and the stator. For this reason, there is virtually no wear on the rotors or stator, and the unit is relatively maintenance-free. Oil is injected into the compressor and serves to cool, seal, and lubricate. When air is compressed from atmospheric pressure to some higher discharge pressure, a specific amount of heat is produced. This heat is sometimes referred to as the adiabatic heat of compression. It is partially removed by injecting oil into the compression chamber. Premachined strips - a raised surface on the male rotor and a

Jan 9, 1988

By H. Colijn

Introduction Bulk commodity prices are ever increasing. This forces buyers and sellers in world markets to look more carefully at methods for obtaining an accurate accounting of commodity transactions in commerce and materials utilization in processes. Weighing and sampling inaccuracies can result in loss of income, product quality, and cost controls. So weighing and sampling accuracies are extremely important. They are the key components of managing the quantity and quality control aspects of bulk materials handling. This article will confine itself to the weighing systems and will deal with various aspects of system design, calibration, and operations. Emphasis will be placed on new developments and requirements in this field. Weighing methods To measure the gravimetric pull on a body being weighed (the load), the gravimetric force must be countered by an opposing force. Measuring this counter-balancing force thus becomes the measurement of weight. The conventional even-arm balance shows the concept of balancing an unknown mass with an equal known counterbalancing mass. But the loads encountered in industry generally range up to many thousands of pounds. So it is obvious that the even-arm approach is not practical. Thus, with the exception of direct load-bearing load cells, most of the weighing equipment uses some form of lever system. This is usually a combination of first- and second-class levers, to gain a mechanical advantage. These levers permit the load force to be counter-balanced and hence measured within a practical range. Lever systems not only perform this force-reducing action, they also serve to focus the load force at a single point, for convenient measurement. They align the force vertically in one direction so that measurement accuracy will not be affected by unrelated side forces. There are five basic methods of gravimetric weight measurements. That is, there are five basic types of opposing forces. • The first is the standard weight counterbalance. It uses a known weight to counterbalance the load directly or through a lever system.

Jan 9, 1988

By G. T. Newell, J. G. Stone, V. M. Mejia

Introduction Advances in drilling have reached a point where large diameter cores can be recovered from "tight," or weakly indurated placer gravels. In such ground, core drilling can provide more reliable data regarding tenor than can be obtained using churn drilling or similar classical techniques. It can also provide metallurgical and geological information that is not available from samples obtained through alternate methods. In 1985, Coastal Mining Co, a subsidiary of M. A. Hanna, and Western Gold Reserves began to review a Tertiary placer deposit owned by San Juan Gold at North Columbia, CA, about 14 km (9 miles) northeast of Grass Valley. The deposit is one of the largest remaining unmined portions of the formerly extensive early Tertiary ancestral Yuba river system. It has been known since the 1850s, has been the subject of much technical literature, and has been the object of at least four previous drilling programs. The eastern one-third of the 6 km (3.7 mile) stretch of the channel between North Columbia and Badger Hill was partially stripped by large scale hydraulic mining in the late 1870s and early 1880s. Mining ceased in 1884 when the Sawyer Decision prohibited further discharge of hydraulic tailings into the Sacramento and San Joaquin Rivers. By that time, about 30 to 45 m (100 to 150 ft) of relatively low grade upper gravels had been removed over some 81 hm2 (200 acres). About 90 to 105 m (300 to 350 ft) of higher grade middle and lower gravels were left at least partially stripped. In 1914, a few churn holes were drilled along a widely-spaced line. In 1938-1939, Selection Trust conducted an extensive drilling campaign to evaluate the deposit. Particular attention was directed toward the partially stripped eastern portion. In 1968, the US Geological Survey drilled three churn holes in the eastern part of the deposit. The US Bureau of Mines conducted experimental mining and drilling in the Badger Hill area. In the late 1970s, Placer Service Corp. acquired a lease on the deposit. Between 1979 and 1984, Placer Service drilled 28 large diameter BADE (a German-manufactured machine) drill holes on the eastern portion of the deposit. The surviving records from the widely-spaced 1914 drilling program are fragmentary and the reported grade not well substantiated. The 1968 holes were drilled for scientific purposes. Again, drilling details are not available. However, detailed records for both the churn drilling program and the BADE program were available and formed the basis for the initial evaluation of the property. Geology The geology of the auriferous Tertiary gravels of California have been described by Whitney (1880), Lingren (1911), and, more recently, Yeend (1974). In general, the Tertiary gravels in the North Columbia area occupy a broad channel cut into pre-Tertiary igneous and metamorphic rocks. The upper, or white gravel is overlain conformably by volcanic tuffs and volcaniclastic rocks. A middle gravel is characterized by the presence of silicified and carbonized wood. A lower blue gravel unit has relatively coarser cobbles and contains a higher proportion of igneous and metamorphic cobbles than the other units. The upper gravel consists of interbedded pebbly sand and silty, or clayey sands with prominent cross bedding. Most of the pebbles are well rounded and consist mostly of white vein quartz and quartzite. The upper unit is moderately well compacted. Exposures in the walls of the old hydraulic mine pits stand at 45° and 50° angles. The gold content of the unit is well below an economic cutoff. The middle gravel - included with the upper unit by Yeend (1974) - is coarser grained, with carbonized wood, and 75 to 100 mm (3 to 4 in.) cobbles of metased-imentary and metavolcanic rocks in a sandy matrix containing abundant lithic fragments. The upper contact appears to be conformable, but the lower portion of the unit appears in places to consist of reworked lower gravels. The unit contains less clay than the upper unit and is somewhat more friable than the underlying lower gravels. The gold content, while somewhat higher than the upper level, is too low to be of ore grade. The lower gravel averages between 30 to 45 m (100 to 150 ft)

Jan 9, 1988

By Louis J. Sousa, Elizabeth H. Yaremchuk, Annette P. Graham

Introduction Responding to Congressional inquiries regarding the purchase of a number of major domestic mining and metals companies by foreign firms, the US Bureau of Mines examined trends in foreign direct investment in the US. mining and mineral processing industries1. "Foreign Direct Investment in the US Minerals Industry" analyzes the extent of foreign investment in several major mineral commodity industries. It identified the countries of origin of foreign investment in US mineral assets. The study also described several reasons for the growth of foreign investment in US minerals and discussed some potential implications of this activity. Foreign investment trends in the economy and minerals For the economy as a whole, between 1977 and 1985 the number of US affiliates of foreign investors more than doubled - from 4046 to 9824 - as did employment at these affiliates - from 1.2 million workers to 2.9 million (Table 1)2. During this same time period, the total value of assets of all US affiliates throughout the economy rose from $239 billion to $736 billion (1985 constant dollars). In the domestic minerals industry, foreign investment also grew substantially during that same time as the number of US minerals-related affiliates increased by 74% - from 125 to 217 - between 1977 and 19853. During the same period, employment at all foreign-owned affiliate mineral operations in this country increased by 78% and totaled 198,000 workers at the end of 1985. Parallelling the growth in the number of and employment at foreign-owned US mineral and metal operations, the total value of US mineral assets owned by foreign firms (or individuals) has nearly doubled since 1977 and amounted to $30.9 billion at the end of 1985. The rate of increase in foreign investment in US mineral assets was slightly higher in the metals industries than in the nonmetallic groups. Foreign investment more prevalent in minerals Perhaps a more useful way of evaluating the growth in foreign investment is to analyze growth in the "intensity" of foreign investment - the proportion of total US industry assets held by foreign investors. Although insufficient data on total industry assets prevented the development of such a measure, employment data provide an indication of relative differences in the intensity of foreign investment among the mineral industries and other industries. For example, as noted in Fig. 1, the 198,000 minerals industry workers employed at US-affiliate companies in 1985 represented 12.7% of total employment in this sector of the economy. In comparison, for the mining and manufacturing sectors of the economy as a whole, 7.8% of all workers were employed at US-affiliate firms in 1985. For the total economy, 2.7% of the entire civilian employed workforce worked at US-affiliate firms that year. Using employment as an indicator, foreign investment has been more prevalent in the minerals industry than in most other industries. About one worker in eight in the minerals sector was employed at a US-affiliate firm in 1985. In comparison, for the goods producing sectors of the economy - all mining, petroleum, and manufacturing industries - about one worker in 13 was employed at a

Jan 8, 1988

By Robert D. Coyle, David L. Emmons

Introduction The Cove gold-silver deposit is located in Lander County, NV about 50 km (30 miles) southwest of Battle Mountain, NV (Fig. 1). The deposit is in the north-central part of the Fish Creek Mountains within the McCoy mining district. Gold was first discovered in the McCoy district in 1914 by Joseph H. McCoy (Vanderburg, 1939). From 1928 to the early 1930s, minor production occurred from shallow underground workings. Total gold production was less than 311 kg (10,000 oz; Stager, 1977). The district was inactive until the mid-1960s when it was explored by Bear Creek Mining, Summa Corp., Houston Oil and Minerals Corp., Gold Fields, and Tenneco Minerals. Tenneco began a detailed reevaluation of the McCoy deposit in the fall of 1984. Tenneco's decision to place the deposit into production was based on a geological reserve of 8.7 t (280,000 oz) of gold (Kuyper, 1987). Subsequent exploration and development drilling of the McCoy skarn deposit has more than doubled the mineable skarn reserves to 19 t (615,000 oz) of gold. In September 1985, Tenneco began development work on the McCoy gold-skarn deposit. The project is an open-pit, heap leach mine. The first gold was poured in April 1986. Echo Bay Mines acquired Tenneco's precious metals properties in October 1986. The mining operation and district exploration started by Tenneco is continuing under the Echo Bay's ownership. The Cove deposit was discovered in January 1987 and production began in February 1988. The deposit will begin as an open-pit, heap leach operation. It will be converted into an open-pit and underground mine with a 9-kt (10,000-stpd) mill. Geologic setting The Cove gold-silver deposit is situated on the east flank of the Fish Creek Range. Surface elevations range from 1460 to 1680 m (4800 to 5500 ft). The deposit is 1.6 km (1 mile) northeast of the McCoy gold-skarn deposit. The Cove deposit is hosted by the Augusta Sequence of Triassic age. Three formations of the sequence are mineralized at the Cove. In ascending order they are the Favret, Panther Canyon, and Augusta Mountain formations. The Favret formation is a gray, locally silty limestone that is commonly fossiliferous. The base of the formation is not exposed. Exposed and drilled intervals are more than 60 m (200 ft) thick. Conformably overlying the Favret formation is the Panther Canyon formation. This formation is 215 m (700 ft) thick and consists of a siliclastic and carbonate sequence of conglomerate, quartzite (sandstone), mudstone, limestone, and dolomite. It is a distinctly coarsening-upward sequence of rocks. The top of the unit is conglomerate with chert and quartzite clasts up to 150 mm-diam (6 in.-diam). The Augusta Mountain forma-

Jan 8, 1988

By G. Johnston, P. K. Chatterjee, S. Holguin

M. Gershon The article is an excellent one, providing a basis for the conduct of the major tasks in underground mine planning and scheduling. I object to only one word. Since that word is featured in the title, it is a strong objection. The word used erroneously is optimization. There is no optimization performed in the article. By definition, an optimization is an approach that yields the best of all possible solutions. We can limit this definition to its narrowest sense, the optimization techniques of operations research such as linear programming, or broaden it to include less sophisticated search procedures that find the best mine plan or schedule. No such claim is made in the article. No optimization of any kind is performed. The major tool used is a simulation, which, in truth, is probably more appropriate for this problem than any optimization. But simulation does not optimize. It only allows for quick evaluation of many options, leading to the selection of the best plan "of those tried," not "of all possible plans." This article, of course, is not the first mining paper to make this error. It seems that the word optimization is used more to impress people than to refer to any act of optimizing. It is used in a way that implies "try to optimize"or "come close to optimize," but not to really do it. It is usually appropriate to quietly ignore this error. But when Dr. Chatterjee, who has done so much good work in true optimization projects and whose work I have come to admire, makes such an error, it is time to publicly dispute it. Having discussed this very point with him in the past, I am sure that he will agree. If he can support this position in his response, perhaps people will start to listen and we can reduce the number of false claims of optimization. This should be easy for him in this case since the approach he used, simulation, is the more appropriate one anyway.

Jan 8, 1988

By Barton J. Suchomel

Introduction Despite perennial economic turbulence, Argentina continues to push ahead with efforts to develop its fledgling hardrock mining industry. The country lacks the infrastructure and mining history of neighboring Chile, but contains vast areas of great resource potential. The eastern slope of the Andean Cordillera in west-northwest Argentina is a relatively little-explored region offering geology highly conducive to the occurrence of base and precious metals deposits. Past exploration has concentrated on locating and evaluating porphyry copper-molybdenum (Cu-Mo) type occurrences. Comparatively little attention has been given to precious metals potential. An interesting and important group of base and precious metals prospects in northwestern Argentina occurs within and adjacent to the Upper Miocene Farallon Negro volcanic complex in Catamarca province. The volcanic complex contains numerous base and precious metals vein occurrences, high gold porphyry copper prospects, and possible disseminated gold targets that have attracted relatively minor attention in the past. Various geologic aspects of the region have been reported by several workers (Llambias, 1970, 1972; Caelles et al, 1971; and Sillitoe, 1973). Regional geology The Farallon Negro volcanic complex is located within the Sierras Pampeanas morphostructural province about 200 km (125 miles) east of the main Andean Cordillera (Fig. 1). The Sierras Pampeanas is characterized by north-south-trending basin and range topography that developed during the Miocene and Pliocene. Uplifted blocks are composed mainly of lower Paleozoic metasediments and granitoid plutonic rocks unconformably overlain by Tertiary redbeds and minor carbonates. A consistent landward migration of volcanic and intrusive ac¬tivity occurred in this region of the Andes from the early Mesozoic through the mid-Tertiary. An eastward extension of magmatism during the Miocene and Pliocene resulted in the formation of the Farallon Negro volcanic center. Volcanism at Farallon Negro ini-

Jan 8, 1988

By H. Lyn Bourne

This is the sixth annual tabulation that gives mineral royalty information. This tabulation, like the previous editions, includes all of the earlier data plus more than 100 new entries. In addition to the royalty information, some entries show the cost of purchased reserves. An article in the December 1984 Kentucky Coal Journal could have far reaching applications. It explains that a banking firm examined the "reserve risk" as one of six factors for analysis when financing a coal project. The firm concluded that coal reserves in Kentucky are worth an average of $0.29/t ($0.26 per st) in-place. The banking firm based this in-place value on its examination of more than 50 leases. In-place value is especially important when trying to establish a mineral value for a given deposit. Ernie Lehman wrote to this author after last year's column appeared and he quoted the lease rates for lead and zinc on federal land. He states, from first hand information, that the leases are based on a percentage of net smelter return as follows: 4% for the first five years, 4.5% for the next five years, and 5% thereafter. The figures used in last year's column ($ per st) came from dividing reported royalty income by reported production. The information reported last year merely translates the results of the net smelter return (as written in the lease) to royalty income per ton to the US Government for the various years. Statistics from the Minerals Management Service of the Federal Government show that oil and gas account for 40% and 53%, respectively, of the Federal royalty income. Coal accounts for about 4% and all other minerals, the remaining 3%. Oil and gas leases are mostly at 12.5%. Most of the coal leases are at 12.5% for surface mining operations and 8% for underground mining. Many of the other commodities are leased at 5%, which has been true for decades. Relative to Federal leases, this year's table shows some percentages and some costs per ton. This was done to give information about the leases and also to show income produced by various commodities. Note that some commodities have a variation in $/ton even though the royalty percent is the same. These variations result from a combination of factors that influence the value of a mineral commodity. The same caveats that have appeared in previous columns relative to factors that influence the value of reserves are restated. The data in the table serve as a general reference and are not intended to present an absolute scale of reserves costs. Several factors influence the value of any mineral commodity. The factors include: • location/transportation (especially significant for most industrial minerals); • market conditions both for the commodity and the company acquiring the right to mine it; *quality or grade of the deposit as it reflects the amount of processing/beneficiation necessary to produce the finished product; and • legislative restrictions and conditions for mining a given material at a given site. These and other factors will significantly influence the cost of reserves. The table shows the range of royalties and the variations in the basis of computing the fees. In a few instances, it gives the cost of acquired reserves based on purchase price and quantitative estimates of the material in place. The table lists the commodity in column 1, and the subsequent columns give the location - either by state, US geographic area or Canadian province; the cost - R for royalty and P for purchased cost; and the year in which the lease or purchase agreement occurred. The last column offers either comments or gives a footnote number for more information. The cost column requires more explanation. The units may refer to cubic yards (/yd), cubic feet (/cu ft), tons (/T, 2000 lbs) or acres (/A). 1 cu yd = 0.7646 m3 1 cu ft = 0.2832 m3 1st=0.9078t 1 acre = 0.4047 hm or m2 In several cases, the royalty is a percentage of the sales price of the finished product and is levied against either the amount produced or the amount sold. We thank those who contributed information for the table: G. Anderson, B. Brown, R. Ganis, K. Hauser, E. Lehman, and S. Sims. Their help is appreciated as are the contributions by those who asked to remain anonymous. This column requires additional information each year. That data must come from people in the minerals industry who have royalty information. What appears in the table is only a very small fraction of the cost information relative to the various commodities and the geographic areas for which there is royalty information. If you find this column of interest and possess information write the author at the address given in the author's box or call 313-462-0005.

Jan 7, 1988

By Barbara S. Parrish, I. S. Parrish

For two consecutive years the author and his wife have been fortunate enough to have been invited by the Peoples Republic of China to work with Chinese geologists of the Geological Bureau of the Ministry of Metallurgical Industries (MMI). Our visits, each of which lasted 7-8 weeks, took us to northwestern Sichuan, central Anhui, and much of Hebei Province, as shown in the accompanying drawing. We were asked to evaluate, criticize, and offer suggestions on the exploration and mining techniques used by the Chinese. Our commission both years was centered on gold, although we were also to visit other operations when the opportunity arose. In 1986, the visit primarily concerned with disseminated "Carlin" type gold deposits; in 1987, with "Greenstone" type gold deposits. Gold has been targeted by the Chinese as a metal of importance to the nation. Several governmental agencies have received direction and budgets to seek and develop gold properties. Besides the Ministry of Metallurgical Industry, gold is being sought by the Ministry of Geology, the China Gold Corporation, and reportedly the Peoples Army. There is considerable competition among these agencies. It is not uncommon to find two or more agencies working in the same general area. However, communication between the groups is minimal. Last year, we toured a small mine at Bambishan. We could see a diamond drill across the valley. The rig belonged to another Ministry and my guides voiced their curiosity to know what their competitors were finding. Mining, too, is not the monolithic bureaucracy that most Westerners anticipate in China. There are various levels of mining. These differ not only in sophistication, but in ownership. First, for small deposits the state allows "Peoples Mines". These can be owned and worked by individuals. The individual owner-miner may hire relatives or neighbors and give them a wage or he may turn the mine into a joint ownership situation. They

Jan 7, 1988

By Charles L. Kimbell

Introduction Despite the focus on the so-called "high-tech" and service components of the economy, the minerals industry in 1987 continued to provide the world the overwhelmingly dominant share of the total raw material supply for all manufacturing operations. And, in the form of fertilizers and other soil treatment materials, minerals also supplied indispensable raw materials to ensure continued high production of the agricultural and forestry sectors of the world's economy. Moreover, the role of minerals as a source of construction raw materials was essential to that sector. On the basis of preliminary and partial statistics on world output of crude mineral commodities, 1987 appears to have been a year of further recovery from the recessional years of 1981-1983. The value of world crude mineral industry production (all mineral commodities including fuels) apparently advanced about 5% to nearly $1.1 trillion in terms of constant 1983 dollars.' That is significantly above the recently revised 1986 level of nearly $1.05 trillion, but still considerably below the record-to-date high of $1.27 trillion set in 1980, as shown in Table 1. That growth in value took place despite relatively flat world oil prices - oil being the most important component of the aggregate, as it has been for decades. There is insufficient information available to assess the total value of processed mineral commodities (refined metals, refined oil, manufactured fertilizers and the like). But a figure of $2.6 trillion (constant 1983 dollars) would be a quite conservative estimate for those materials derived only from primary or newly mined materials. And an additional, as-yet-unestimated increment should be added for processed materials derived from secondary raw materials - scrap and other reclaimed substances. From a world trade viewpoint, minerals and mineral products dominate in terms of tonnage. For example, in 1985, on a weight basis, mineral commodities including fuels constituted more than 58% of all goods moving through the Suez Canal and 53% of all goods passing through the Panama Canal. These waterways cannot accommodate giant tankers and ore carriers. So an even larger share of total oceanic trade is composed of minerals and mineral products. Yet, in terms of value, the share of the total accounted for by these industrial essentials is much lower than that accounted for by other commodities, most notably sophisticated high-tech equipment and other goods with a relatively high per-pound value. Worldwide data on the value of mineral export trade for 1987 is so incomplete at this writing that it is impossible to assess that actual level. Indeed, the compilation of similar available data for 1986 covers only the exports of nations with developed market economies. In 1985, these countries accounted for only 67% of total world exports of all goods. They accounted for only 41% of world exports of major mineral commodities: crude nonmetals; metal ores, concentrates, and scrap; mineral fuels; nonferrous metals excluding gold; and iron and steel. For these developed-market-economy countries, the value of exports of these major mineral commodities declined by more than 10% between 1985 and 1986. Most of the shortfall came as a result of a 26% drop in the value of fuels exports. If these patterns of

Jan 7, 1988

By L Baumgardner, A. V. Castelli

Barite In 1987, United States mine production of barite increased 15.870, consumption (sold or used by grinding plants) increased by 6.97, and imports are estimated to have fallen by 19.57. World mine production decreased by 137, according to the US Bureau of Mines. The value of US produced barite, fob mine, decreased 1.4%, according to the Bureau. The declared value for cif US port of all imported unground barite during the first 10 months of 1987 decreased from $39.02/t ($35.40 per st) to $37.16/t ($33.71 per st), the Bureau stated. Nevada continued to be the leading US producer of' barite with 637r of the total. It is followed by Georgia, Missouri, California, Tennessee, and Montana. The Bureau of Mines estimates 657 of the US mine production was used as a weighting agent in drilling fluids. The remaining 35°/r was used in barium chemicals, glass or as a filler. Most of the production from Missouri, Georgia, and Tennessee was used in the non-oilfield sector. This sector continued to increase. Of the total consumption (used by grinding plants and chemical manufacturers), it is estimated that 8070 was used in drilling fluids and 2017( in the non-oilfield sector - barium chemicals and fillers in paper, plastics, and glass. The tonnage consumed in the drilling fluids market increased slightly. But the average number of rotary rigs operating decreased from 964 in 1986 to 936 in 1987. This is the lowest number of operating rigs reported since 1949. The increase in consumption in this sector is due to an increase of rigs operating in the US Gulf and other higher consumption areas over 1986. Crude barite imports were estimated to be down by 19.5% in 1987. Still, they accounted for 637 of the barite consumed in the US. During the first 10 months, China continued to be the leading exporter to the US with 86.970, followed by Mexico with 10.997, and Morocco 2.107(. About 5Y of the imported ore was used in the nonoilfield sector. Ground barite imports dropped from 19.8 kt to 8.8 kt (21,800 st to 9800 st) in 1987. Of this amount, 1.4 kt (1598 st) was used in drilling muds and the remainder of the non-oilfield use. During 1987, a few more grinding plants that supplied the drilling fluid market were closed. Along the US Gulf Coast, there are 13 operating plants versus a high of 31 in 1981. Cameron Offshore reopened its grinding plant at Cameron, LA. New Riverside Ochre Co. of Cartersville, GA, purchased a jigging-washing plant in Missouri and plans to place it in operation this year on a new deposit near Cartersville, GA. Milpark purchased Hughes Drilling Fluid in 1987 and operates under Milpark. US mine production will probably stay about the same as 1987. Price increases in the 10 % range for imported ore are expected due to water freight and the weak dollar. Baker Hughes estimates the average number of operating rotary rigs at 1070. A slight increase is expected in the non-drilling fluid market (chemicals, glass, and filler). Bauxite and alumina L. Baumgardner, US Bureau of Mines Domestic bauxite production in 1987, expressed as dried equivalent, was 581 kt (640,400 st), a 14% increase over 1986. US mine production amounted to less than 170 of total world production. Virtually none of the domestic bauxite was consumed in the production of primary alumina metal. About 30% of the bauxite produced by US mines was calcined to produce 117 kt (128,970 st) for use in abrasives, catalysts, chemicals, proppants, and refractories. Most of the uncalcined bauxite was refined to specialty aluminum oxide (alumina) forms. Metallurgical grade bauxite was imported from 10 countries in 1987. Principal suppliers of the 9.8 Mt (10.8 million st) were Guinea (4270), Jamaica (3501,), Australia (127), and Brazil (57). The average value, delivered U.S. ports, cif, as reported to US Customs Service, was $29.50/t ($26.76 per st). Calcined refractory grade baux-

Jan 6, 1988

Crushing and grinding The Crushing and Grinding Technical Committee There was substantial activity in the area of crushing and grinding with the improved outlook in the industry. The Crushing and Grinding Committee's theme of 1987's Annual Meeting in Denver was Fundamentals of Mechanical Design for Crushing and Grinding Mills. At the meeting, the concept of gearless sag mill variable speed drives was taken out of the closet, dusted off, and a wraparound motor was put on hard copy with the ordering of a 9.7- x 4.5-m, 8.2-MW (32- x 15-ft, 11,000-) sag mill. Additionally, comparisons were provided to the mill design of the gearless drive sag mill and the twin 4.5 MW (6000 hp) do variable speed drives placed on the drawing board for the Kennecott UCD Expansions. This involved a 10- x 4.5m, 8.9 MW (34- x 15 ft, 12,000 hp) sag mill. Differing opinions were also expressed in technical presentations pertaining to the need for quality mill specifications in design and manufacturing. J. Berney-Ficklin of Bechtel presented a paper outlining the need for quality design and manufacturing specifications in large scale grinding mills, and the overall cost reduction in capital and operating costs. Others presented differing specific views, but all emphasized a central theme. Quality and reliable designs in large mills should not be compromised. In-pit crushing applications continue to increase, overseas and in the US, in both mining and quarrying. Different in-pit crushing units have been classified as mobile, with integral transport systems; semi-mobile, typically modular systems that use separate transporters for repositioning; semi-stationary plants that must be dismantled before transporting and typically require earthwork and concrete foundation work; and stationary in-pit crushing plants, which remain in place for the life of the pit. The majority of equipment sales of conventional crushing and grinding circuits in 1987 were to gold mining operations. There have been two areas of end use: • conventional jaw/gyratory or impact crushing circuits preparing low grade ores for heap leaching; and • small scale sag/ball mill circuits preparing higher grade ores for classical cyanide leach/ recovery circuits. There was also continued interest in ABC and SABC circuits for both base and precious metals plants. High pressure roll crushers continue to attract considerable interest in the cement industry. There, they are used to prepare raw materials or clinker for subsequent ball/tube milling. However, they have apparently been less successful in the mineral industry where the energy savings/ capacity gains have been much less dramatic. Several papers covering the theory and application of these machines were presented in an industrial minerals session on cement at the SME Annual Meeting. The Third International Conference on Hydrocylones, held in England in October produced several papers of interest on classification and screening to the mining industry. In particular, a new cyclone apex, developed by Mt. Newman Mining, shows promise of reducing water flow to the cyclone underflow stream and, therefore, increased underflow pulp densities and reduced fines recycle in grinding circuits. A twin vortex hydrocyclone was also described and shown to give a sharper split than conventional units. Similar double cyclones were introduced in the 1950's, but failed to achieve widespread acceptance. This was due mainly to a proneness to blockage in the transitional zone. It will be interesting to see whether this latest variant can overcome this problem in the typical operation. A number of North American iron ore plants installed the capability to produce fluxed iron ore pellets. Most chose to grind the stone on site with circuits designed to grind taconite ore. Methods employed to grind this material included single-stage ball mills closed with fine screens, rod and ball mill circuits - both open and closed with cyclones - and semiautogenous mills, followed by ball mills closed with screens or cyclones. Dewatering and tailings disposal B.M. Moudgil and D.L. Sober, University of Florida Due to tightened environmental regulations and the need for more efficient land use, a greater interest in dewatering and waste disposal has developed. As a result, research efforts in flocculation, surface chemistry, and polymers science have focused on the problems related to the dewatering of solid mineral wastes. A few studies have been conducted to examine the flocculation process. Hogg et al., (1987) discussed the formation and growth of flocs. They determined the size and density of flocs was controlled by the physical conditions of the system (e.g. agitation,

Jan 5, 1988