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By B. E. Davis

A technique was devised by the Bureau of Nines to recover coarse (plus 1.7 m) North Carolina phosphate now discarded as waste. Concentration utilizes the difference in particle sphericity between the relatively flat calcite and the rounded phosphate pebbles. Sizing, treatment in a continuous hydraulic classification system, and grinding upgraded a 17.0-pct-P205 reject to 24.5 pct P205 with a phosphate recovery of 63 pct.

Jan 1, 1985

By P. Geoffrey Feiss

The Piedmont province of the southern Appalachians is becoming the focus of interest for many exploration geologists. In the past, only those deposits with significant surface exposure have been exploited. Thus, few deposits have been found and relatively little exploration has been conducted in recent years. Modern geochemical and geophysical techniques can alleviate this problem of concentrating on exposed ore bodies by allowing us to "look through" the soil and saprolite to detect the presence of significant mineralization at depth. This paper will attempt to evaluate the feasibility and effectiveness of several geo¬chemical exploration techniques in the Piedmont, of Cabarrus and Stanly Counties, North Carolina. The area chosen for the study is shown in Figure 1, lying almost wholly within the Mt. Pleasant quadrangle, North Carolina. The area is on the eastern boundary between the Charlotte belt, a zone of plutonic and intermediate grade metamorphic rocks, and the Carolina Slate belt, a series of complexly interbedded volcanic and sedimentary rocks of predominantly low grade metamorphism. The area is now considered to be a 500 million year old active, continental margin.1 As is typical of most of the piedmont, outcrop is sparse. Figure 2 is a sketch map of the general geology of the area under study, showing the Charlotte belt rocks in the northwestern quadrant and slate belt rocks, to the east. Within the area are a number of old mines and prospects, the most famous of which are the Furniss and the Phoenix mines in Cabarrus County. All are reported to have contained Au and Cu. Ag and Pb were reported from some. A summary of the deposits is given in Table 1. Data used in this study are from three Master's theses conducted at the University of North Carolina at Chapel Hill. 3,3,6 In separate studies, well waters5, B-zone soils3, and vein and float quartz6 from the area described were sampled and analyzed. Each study concluded separately that geochemical anomalies in the vicinity of the ores justify use of that specific sample type in a geo-chemical exploration program. However, no comparative study of the methods has

Jan 1, 1977

By V. L. Ketellapper

This paper discusses the technical issues and stakeholder collaborative processes which resulted in achieving a balance between the reclamation of abandoned mine sites, providing affordable housing, and preservation of open space at the French Gulch Site. Through this process, it was discovered that mine site reclamation would need to be conducted in conjunction with two priorities for community sustainable development: providing affordable housing for locals who work in the resort town of Breckenridge, and open space preservation. It was found that conducting these activities concurrently would result in providing resources necessary to complete mine site reclamation. The first phase of the affordable housing development is almost completed. This development resulted in redevelopment of mining impacted lands and was recognized by the U.S. Environmental Protection Agency (EPA) with a National Smart Growth Award in 2002.Negotiations are also underway to preserve the remaining 1,800 acres of land held by the mining company as open space. This transaction will in turn provide resources to address water quality issues associated with mine discharges.

Jan 1, 2004

By Joseph E. Worthington

The Haile Gold Mine is one of the earliest gold discoveries in the eastern United States as well as one of the largest gold producers from that region. Gold was first discovered in the gravels of what is now known as Haile Gold Mine Branch in 1828 on lands then owned by Colonel Benjamin Hale (spelling not the same). Small placer workings were immediately started and it was not long before the surrounding residual soils and saprolites were also found to contain gold. Following the presence of gold in the residuum eventually lead to the discovery of the underlying gold-bearing lodes that actually outcropped in saprolite form where not masked by heavy vegetation or a thin veneer of Coastal Plain sand. The earliest attempts at production, immediately after discovery, utilized slave labor, washing gravel and residuum from leased tracts that were 15m (50 ft) square. Several at- tempts at small milling operations were made during the en- suing decades, but there was really no substantial or successful production until the 1880s.

Jan 1, 1990

By Boghani. A. B.

A computer program has been developed to simulate evacuation of a mine in various configurations. The simulation input variables include characteristics of the mine, expressed, in the form of a network of passages, properties of hoists and hoisting strategies, initial miner positions, ventilation and compressed air velocities, stench warning system and self rescuer capabilities. The simulation can be carried out for cases in which there is no fire, i. e. , fire drill simulation, or for cases in which there is a fire at any arbitrary location. The fire is assumed to generate gases which impede movement of miners in the gaseous region. However, the thermal effects of the fire are not considered. The simulation output include the evacuation time for each miner, movement of the hoists, propagation of smoke through the mine passages and smoke exposure of the miners. Several illustrative simulations have been carried out to verify validity of the program and to demonstrate its capabilities. The computer program has been able to predict evacuation time, within 5 percent, for an actual fire drill. It is believed that this program can be used to evaluate evacuation performance of a mine, identify problem areas and demonstrate improvements due to the recommended corrective actions.

Jan 1, 1977

By Anthony G. Fonseca, W. W. Wen

About 45-50% of the coal mined in the U.S. passes through coal preparation plants; east of the Mississippi River this number increases to about 75-80%. Although the cost for coal preparation is worthwhile to some, the coal industry faces the challenges of continuing downward pressure on the price of coal and the impact of new environmental regulations. Coal preparation, as commercially practiced today, is an effective process achieving 75-803 ash reduction, 15-8m trace element reduction, and 85-90% Btu recovery; it is less effective for pyrite reduction (35-70%), and on-line operating time (40-60%), and suffers from obsolete control systems. Methods will be discussed for reducing costs of coal preparation and improving the performance of coal preparation plants. Comments are included on equipment selection, especially for -28 mesh coal, and prep plant operation and control practices. Btu recovery, ash and pyrite reduction, fines processing including dewatering and slurry fuel use options are emphasized. Trace element removal and expert control systems for maximization of prep plant operation also will be highlighted.

Jan 1, 1995

By W. H. Schiffbauer

The U.S. National Institute for Occupational Safety and Health (NIOSH) developed an active proximity warning system called hazardous area signaling and ranging device (HASARD) for warning workers as they approach known dangerous areas around heavy mining equipment and other dangerous work zones. HASARD is composed of a transmitter and a receiver. The transmitter generates a 60-kHz magnetic field using one or more wire loop antennas. Each antenna is adjusted to establish a magnetic field pattern for each hazardous area. The receiver, which is worn by the worker, is a magnetic field strength meter. The received signal is compared against preset levels that are calibrated to identify levels of danger. The receiver outputs can include visual, audible and vibratory indicators, and it can also be made to disable machine functions. HASARD was field tested on a Joy 12 continuous miner, a Komatsu 210 M Haulpak and on a high wall launch vehicle. Slight modifications were made for each application. HASARD provided warnings as designed, and it proved to be rugged enough to withstand the harshest of production environments.

Jan 1, 2003

By J. W. Jang, I. Iwasaki

Introduction Grinding ball wear in wet grinding is due to a complex combination of abrasion, corrosion, impact and their interactions. The ball charge is continually falling from a height greater than half of the ball mill's diameter, and contact is made with ore particles or other balls. Abrasive wear occurs through penetration and plowing of media materials from the ball surface by hard ore particles (Gangopadhyay and Moore, 1985). Electrochemical interactions between the abraded and unabraded areas of the ball surfaces, as well as between the grinding media and the minerals, occur in grinding mill pulps (Adam et al., 1986). The wear depends on the media composition, the hardness and the corrosive/abrasive characteristics of the slurry. The marked-ball-wear test method was used to evaluate the performance of grinding media under nitrogen and oxygen atmospheres. Also, the wear mode was studied by observing the grinding media surface under a scanning electron microscope. In general, grinding media of high hardness and high chromium contents were found to have better corrosive wear resistance than those of low hardness and low chromium contents (Jang, Iwasaki and Moore, 1988). The purpose of this investigation was to determine the size distribution and the shape of wear debris in a study of the wear rates and wear behaviors of grinding balls. Experimental Materials The grinding ball alloys used in this investigation were: •martensitic stainless steel (MSS; 0.96% C and 16.2% Cr), •high-carbon, low-alloy steel (HCLA; 0.89% C and 0.53% Cr), and •mild steel (MS; 0.22% C and 0.09% Cr). The mineral used in this study was synthetic corundum, because it was on hand and contained no magnetic impurities. Debris A 203- mm (8-in.) diam porcelain mill was used for grinding the corundum. The mill charge consisted of: •1200 g of 12-grit synthetic corundum, •770 mL of 0.05 mole/L Na2SO4 (61% solids), and •125 steel balls of nominally 25-mm (1-in.) diam. To minimize the debris corrosion, nitrogen was introduced into the mill. Grinding with MSS, HCLA-steel or MS balls was conducted for 2 hr at 51 rpm. The grinding time using the MS ball was varied from 0.5 hr to 2 hr in an attempt to study the debris amount and shape as a function of time. The debris was collected with a hand magnet from the ground slurry and separated from the corundum particles using a Davis magnetic tube. After the separation, the debris particles were immediately washed with acetone and then filtered and dried. Debris size analysis Buckbee-Mears 3-in.-diam, micro-mesh sieves and an electromagnetic shaker were used for determining the debris size distribution. In sieving, isopropyl alcohol was used. The debris fraction collected from each sieve was washed with acetone, dried and then weighed on an analytical balance. [ ]

Jan 1, 1993

By Don Macfarlane, Brian Heer, Tom Martin

The Manapouri Power Station, commisioned in 1969, is located in Fiordland National Park in the South Island of New Zealand. It is the country’s largest hydro power scheme and can generate 700MW of electricity. The power station utilizes the hydraulic head of Lake Manapouri, using an underground power station and a 10 km-long tailrace tunnel was constructed by drill-and-blast to allow discharge to Doubtful Sound. Power generation fell short of expectations due to higher than expected friction in the tailrace system. To remedy this a second tailrace tunnel was constructed between June 1998and May 2002. It consisted of a 10 m diameter, 10 km long tunnel in predominately gneissic rock and was excavated using an Atlas Copco Robbins TBM. High water inflows, Owner/Contractor production incentives and a major claim settlement were handled during the 5 1/4 year project. This paper discusses TBM advance rates, utilization factors that increased in the second half of the project plus an innovative method for sealing Tunnel 1 to construct the bifurcation.

Jan 1, 2003

By Irvine G. Reilly

Results of a preliminary study of the ammoniacal leaching of covellite (CuS) in an aqueous oxidizing system in the presence of a non-miscible organic sulfur solvent are described. Up to 60% of the sulfur can be recovered from the solvent as elemental sulfur 1 while achieving high degrees of solubilization of the mineral. The effects of temperature, particle size, oxygen pressure, ammonia concentration and ammonium ion concentration on the degree of sulfur recovery were investigated. Maximum recoveries were achieved at moderate temperatures and in well-buffered systems, with a large excess ammonia concentration. Oxygen pressure and particle size were found to have little effect on fractional recovery of sulfur. A simple empirical model based on electrochemical theory correlating the effects of some of the parameters cont-rolling the degree of sulfur recovery is presented. The results of this work also confirm that in any proposed reaction mechanism for the oxidative ammoniacal leaching of copper sulfides, the primary oxidation product of the sulfide ion in elemental sulfur.

Jan 1, 1978

By R. L. Atwood

The structure, properties, and composition of commercial chelating-type copper extractants are examined. Characterization of these extractants was accomplished using organic separation techniques and analytical tools such as NMR and IR. Results indicate that only the trans isomer of the diphenylhydroxyoxime (LIX (R) 64N) is capable of copper chelation. The effect of the structural modification of the extractants on the extraction and stripping reactions are discussed.

Jan 1, 1972

By E. Baliktsis

PPC (Public Power Corporation) operates a new power plant at the South-West part of Crete. Recently the extension of the Power Plant started. ATHENA SA the main contractor of the project assigned to EXORIXI SA the implementation of a blast vibration monitoring program and prevention-evaluation of risks caused by construction blastings inside or very close to the operating part of the power station. The blasting parameters and characteristic elements that had been evaluated, the measures that had been taken and materialized, the overall risk management and the results of the whole program are presented in this paper.

Jan 1, 2006

By B. J. Arnold

Utility coal specifications usually contain a requirement for a maximum moisture content in order to prevent coal handling problems and to minimize evaporative losses in the boiler. As most coal cleaning processes are water-based, dewatering and sometimes thermal drying steps are required to meet moisture specifications. Screens are often used for dewatering, with centrifuges and filters used to further reduce moisture in many cases. Algorithms to predict the performance of commercial dewatering devices were developed for the Coal Cleaning Simulator running under Aspen Plus? during a project sponsored by the U.S. Department of Energy and the Electric Power Research Institute. These algorithms include calculations for size degradation in the devices and fines loss to effluents, where appropriate, and a calculation for moisture content of the product based on particle size and specific gravity distribution. This paper discusses the algorithm for coarse and fine coal centrifuges and gives comparisons between predicted moisture and solids loss and published moisture and solids loss values for various particle size ranges.

Jan 1, 1996

By Donald P. Richards, Patrick Ramond, Martin Herrenknecht

Phase 1 of the second line of the Cairo Metro includes approximately 5.83km, of double track tunnels excavated in water bearing sand of the Nile flood plain. Closed face bentonite slurry shield tunnel boring machines 9.43m in diameter were used. State-of-the-art technology was used in TBM operations and guidance, with surface settlements averaging less than 14mm, and alignment deviations averaging less than 25mm. Production rates up to 580m/month were achieved, with average production of 4.8m/shift, 13.4m/day and78m/week. The completed tunnels, lined with precast concrete segments, are completely watertight under 20m of water pressure.

Jan 1, 1997

The US Bureau of Mines' dictionary of mining terms defines productivity as "the efficiency with which economic resources (men, material, and machines) are employed to produce goods and services." Or, the relationship between the physical outputs and the physical inputs used to produce those outputs. As demand for mine materials dropped, mining companies and equipment manufacturers discovered that ratios between the inputs and the final products were disproportionate to the profits realized. Some companies are fighting for survival. The US, while still the most productive country in the world in absolute terms, is looking over its shoulder as other industrialized nations continue to exceed its yearly productive growth rate. According to John Belcher, Jr., of the American Productivity Center in Houston, TX, the US has seen a steady decrease in productivity growth since the mid-1960s. During the post World War II period, the US labor productivity growth rate was 3.2% annually. The slowdown began in 1965 and by the mid-1970s, the country's annual growth rate was 1%. By 1978, labor productivity growth ceased to gain at all and from 1978 to 1980 a 2% decline occurred. Solutions to the productivity decrease are increasing product prices or improving productivity. By increasing prices, though, the producer runs the risk of pricing himself out of the competitive market. To maintain an acceptable profit level, prices must increase proportionately higher than required to offset production cost increases, Belcher says. Many companies are turning to in-house productivity improvement programs as an answer to increased costs and declining profits. These programs are geared toward management commitment and employee awareness of what productivity means. Companies are finding ways of measuring productivity levels in financial reporting, budgeting, and planning. Organizational teams have been formed to coordinate activities and oversee product management, and improve productivity of materials and capital investments. This special issue of Mining Engineering examines productivity in the mining industry. First, there are articles about various, and different, factors that affect productivity in each of ME's five main readership areas - exploration, mining, processing, coal, and industrial minerals. There is also a look at mining industry productivity from the manufacturer's view. The next section of this special issue is devoted to some topics that affect all areas and sectors of mining: research, management practices, computers, maintenance, and energy management. Then follows case studies in exploration, coal, and iron and steel. These studies provide examples of practices and strategies used by companies and operations to boost productivity.

Jan 11, 1983

By Y. H. Alghamdi, A. D. S. Gillies

"INTRODUCTION Background Diesel Particulate Matter (DPM) has been a critical issue during the last decade in underground mines. This problem is threatening the health of mine workers, as well as the mine’s productivity. It has been confirmed that the diesel engine exhaust is the main source for the diesel particulate matter (DPM), and that it has the highest impact in raising carbon monoxide, carbon dioxide, oxide of nitrogen, and hydrocarbons in the underground environment. Miners who are exposed to this kind of emission are at high risk for developing lung cancer and other lung diseases. The Mine Safety and Health Administration (MSHA) has issued standard rules to regulate diesel emissions in underground mines. The final ruling confirmed in 2008 that the total carbon should not exceed 160 µg/m3. (MISHA, 2008) Monitoring DPM in underground mines has become important to ensure compliance with the MSHA regulations. There are many different types of DPM monitoring systems that are currently being used in underground mines. Two methods are typically used for DPM sampling: the personal sampling method or the direct exhaust sampling method, which takes a direct sample from the engine’s tailpipe. Many mines do not recommend using the direct exhaust sampling method since it involves complex measuring equipment and procedures, as well as it is not regulated in the United States or Canada. Thus, the personal sampling method is the preferred method in many mines when taking DPM samples. Even this method includes many variations, such as the Respirable Combustible Dust (RCD) method and the method devolved by the National Institute for Occupational Health and Safety NIOSH 5040. The NIOSH 5040 method is the preferred method overall because it gives more accurate measurements compared to the other methods. This method does not allow organic carbon sources other than those coming from diesel engine exhausts to interfere with the measurements, which otherwise would lead to overestimating the actual amount of the DPM in the mine environment. An impactor feature in the cassette, which holds the samples in the instrument, characterizes this method. Since the diesel particles are less than 0.9 micron, this impactor does not allow particles more than 0.9 microns to be collected in the cassette filter. The cassette is then analyzed in a lab and the concentrations of organic carbon (OC), elemental carbon (EC), and total carbon (TC) are determined. Lately, the National Institute for Occupational Safety and Health (NIOSH) has developed a new instrument that can use the same concept of NIOSH 5040, but with instant measurements of elemental carbon. It utilizes laser radiation with the cassette so it can give a real time measurement. The real time measurements give instant information about the underground environment that will allow the mining engineer to take immediate action and enact plans instead of waiting for results from the laboratories where NIOSH 5040 method may take several weeks to get the measurements."

Jan 1, 2016

By E. Milosavljevic, C. H. Bucknam, K. L. Gabby, T. C. Eisele

"Since mercury is chemically similar to silver, it tends to dissolve along with the silver and gold during cyanide leaching. Ultimately, mercury is recovered along with the silver as mercury/silver/gold amalgam, which requires retorting to remove the mercury prior to refining the silver and gold. A possible approach to separating mercury from silver and gold in cyanide leach solutions is to take advantage of the replacement reaction between [Hg(CN)4]2- and Ag2S, which results in precipitation of mercury as HgS while dissolving silver as [Ag(CN)2]-. Results are presented demonstrating that mercury can be removed from cyanide solutions with high effectiveness while, at the same time, recovering additional silver in solution. Initial experiments with static flasks in the laboratory showed a high degree of mercury removal could be achieved. Results using actual metallurgical plant products in bottle roll tests indicate the process works on a larger scale, but also results in significant gold losses if the solution being treated has high gold levels, such as carbon stripping pregnant eluent solution. IntroductionThe dissolution of mercury along with precious metals during cyanide leaching is a long-standing problem, which can result in significant mercury toxicity issues (Washburn, 2003). Mercury is first dissolved by the cyanide solution, and then, since it has a great deal of chemical similarity to silver, it follows silver throughout the concentration and recovery processes, ultimately being recovered as a silver/ mercury amalgam when the metal is precipitated by either the Merrill-Crowe process or by electrowinning.While it would be desirable to either prevent mercury dissolution entirely while still allowing silver dissolution, or to selectively capture mercury before the precious-metals recovery step, there is not yet an effective and economically practical approach to doing this (Grosse, 2003; Ravichandran, 2004; Misra, 1998; Pai, 2000). There are a few processes that have been proposed, such as differential adsorption. While activated carbon, as used in the carbon-in-pulp process, is not highly selective between mercury and silver (Washburn, 2003), it may be possible to achieve some separation by selective elution of the carbon to produce high-mercury and low-mercury eluates (Bunney et al., 2010). Certain synthetic resins were shown to be selective for gold but are more expensive than carbon, requiring special equipment, and have not yet been shown to be selective for mercury (Grosse, 2003). There may be potential for differential dissolution approaches, as thiosulfate leaching does not dissolve mercury sulfides as readily as it dissolves gold (Oraby et al., 2010), and it is possible that it may be similarly more selective for dissolving silver relative to mercury. Some work has also been done with selective precipitants, such as sodium polythiocarbonates (Bucknam, 2007), and 1,3-benzenediamidoethanethiol (Matlock, 2002). Common problems with these approaches include introduction of a solid:liquid separation step, reagent costs, the need to measure the mercury concentration for process control for minimizing silver losses, and regulatory requirements to retort otherwise stable mercury concentrates."

Jan 1, 2015

By L. Xu, R. Bhaskar

The trends in dust exposure levels and longwall production over the last 20 years were examined. The study was conducted for nine major underground coal mining states: Alabama, Colorado, Illinois, Kentucky, Ohio. Pennsylvania, Utah, Virginia and West Virginia. The dust exposure and shift production data were obtained from the Mine Safety and Health Administration (MSHA). Trends were identified and analyzed for the headgate, tailgate, jack setter and shearer operator positions. The relationships between average scam height and dust exposure levels as well as the percentage of dust samples exceeding the dust standard for the four occupations were examined. The concentration to production ratio, a better indicator of the success of dust control measures, was calculated for the 20-Year period. The data show that, in general, dust exposures over the last two decades hare decreased. However, it appears that much of the decrease was in states that had significantly high dust exposure levels (such as Alabama and Utah), while for states such as Virginia, West Virginia and Pennsylvania, the decrease was not as dramatic. Major reductions were made in the period starting in 1982, when a host of longwall dust control techniques was introduced. However, nearly 35% of jack setter and shearer operator dust exposures were still above 2 mg/m3 in 1990, with nearly 20% of tailgate operator exposures and 8% of headgate operator exposures being above the federal standard. The concentration/production ratio has shown a remarkable decrease since 1981. This indicates that long wall operators have been able to meet increased production as well as, to a large extent, control chest production at the face. An analysis of the relationship between average seam height and dust concentration for the years 1984 and 1990 shows that while in 1984, the dust concentration could be correlated to seam height, in 1990 the relationship was less definitive. This indicates that technological factors may be more important than seam height.

Jan 1, 1994

By James M. Barker

No commodity agreements now exist for sulfur and a lack of reference to sulfur commodity agreements exists in the literature. Law (1975) made a passing reference to sulfur as one of a group of commodities where "there have been additional control efforts of various types." Clearly, sulfur was a candidate, at least conceptually in the past, for some form of commodity agreement and the following discussion centers on the world sulfur market of the early 1980's in this context. GENERAL THEORY OF COMMODITY AGREEMENTS The upsurge in demands for commodity agreements seen in the last decade or so stems largely from agitation by the less developed countries (LDC) for better economic treatment by more developed countries (MDC). Several economic factors, real or imagined, adversely affect the LDC's as follows (MacAvoy, 1977): 1) variable demand in MDC's causes wide fluctuations in commodity prices which hurt the economies of commodity producers, who are often LDC's although consumer economies may also benefit from more predictable prices; 2) a general decline of commodity prices relative to the price of manufactured goods which hurts commodity producers in their balance of trade while their stocks and overcapacity rise; 3) instability of LDC economies enhanced by the above factors lowers internal stability, inhibits foreign investment, and retards economic growth; and 4) a lack of self respect when dealing with developed countries o, following the OPEC lead, LDC producers want the stature related to cartel market power.

Jan 1, 1985

By Tom Plouf, Deepak Malhotra

INTRODUCTION The U. S. mining industry experienced a severe recession in the early 1980's. This resulted in the realization that the domestic mining industry had to compete on a global basis and only the strong or the low-cost producers would survive. The economic viability of a mining company not only depended on maximizing mineral recovery but also on its ability to do so at minimum cost. Companies had to become more productive not only by cutting costs, but also by developing and/or evaluating new technologies in existing or new plants. Several innovative concepts, including in-pit crushing and conveying, fine crushing, improved classification equipment, flotation, column flotation, pressure filtration, etc., have been developed in recent years. This paper addresses the impact of new flotation concepts and equipment on mill design. MILL DESIGN PHILOSOPHY Modern mill design, especially for large capacity plants, has its foundation in minimizing capital and operating costs. Capital costs can be minimized by minimizing the various components of the capital cost: infrastructure, equipment, instrumentation, structural, initial spare parts costs, etc. All these costs are minimized by simplifying the mill design and/or process flowsheet. The simplified process flowsheet, where applicable, consists of a single stage primary crusher followed by a semiautogeneous - ball mill grinding circuit to produce feed for a flotation circuit. Multiple flotation circuits have been eliminated in most cases. These concepts minimize the cost and encourage the use of large capacity equipment while maintaining optimum metallurgical recovery. The philosophy of mill modernization/modification is generally driven by the desire to increase the mill capacity within the limits of the existing mill building. This need is facilitated by using large size flotation equipment and eliminating multiple parallel flotation circuits in order to minimize capital and operating costs. Conventional flotation machine development has paralleled this mill design philosophy. FLOTATION CELL DESIGN Since froth flotation has been the most widely used separation process for over half a century, one would expect that it would be possible to design and select flotation machines for specific applications based on scientific principles. But this has not been the case. Best judgment, experience and personal prejudice have played a key role in equipment selection. These factors, along with the difficulty of evaluating new concepts in commercial plants, has slowed progress in implementation and commercialization of innovative ideas and/or equipment. Conventional Flotation There have been no major technological advances in conventional flotation cell design in recent years. A recent flotation cell survey indicates that 'larger flotation cells are still the order of the day.”(1) A comprehensive review of design of flotation machines indicate that they have been mechanically redesigned and are generally simpler and easier to maintain and less expensive to manufacture. (2,3) The approximate chronological development of conventional cells from the 25 cubic foot cells to the 3000 cubic foot cells, which were recently installed at the Copperton concentrator of Utah Copper, is given in Table 1. Larger flotation cells are the obvious choice for new plants or plants looking for expansion because of the following reasons: • They have lower capital cost per unit volume • Installation costs are generally lower because fewer units require less space • Maintenance costs are lower because of fewer mechanisms • Power requirements (HP/ton dry feed) are generally lower • with increased flotation cell size as illustrated in Table 2 [Table 1. Conventional Flotation Cell Size Development Year Individual Cubic Cell Size, Feet 1925 25 1935 50 1960 100 1965 300 1970 500 1975 1,000 1980 1,500 1985 3,000]

Jan 1, 1989