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By Carl Erickson, James Hislop

INTRODUCTION As practiced today, longwall mining is a mechanical version of an established mining system. The equipment for longwall mining consists of four basic units: roof supports, face conveyor, winning or mineral-getting ma¬chine, and stage-loader conveyor. The system mines relatively thin slices of mineral across a predetermined face length in a virtually continuous sequence. As the winning machine cuts, the mineral is loaded automatically onto the face conveyor. In turn, the face conveyor is pushed over into the newly cut track by hydraulic rams attached to the conveyor and the roof supports. The roof supports then are lowered individu¬ally from the roof, advanced to the repositioned con¬veyor, and reset to the roof. At the end of the pass, the longwall cutter is pushed over, and a slice is taken in the opposite direction, so the system repeats itself. Fig. I illustrates a typical longwall mining system. The principle behind the longwall mining system is that the roof is supported only immediately adjacent to the worked seam, keeping the blocks of immediate roof clamped securely from the face to the rear of the sup¬ports. At the rear of the supports, the roof is allowed to fall and break up, with the swell of the broken material eventually establishing an equilibrium in the strata be¬hind the mined area. Fig. 2 illustrates the principle of longwall support. There are two distinct types of longwall mining. The longwall retreating system is the most common in the United States, South Africa, and Australia. The longwall advancing system is the most common in the re¬maining industrialized countries such as Britain, Ger¬many, Poland, and the Soviet Union. The longwall retreating system comprises a mining panel, where the roadways that form the head entry (end at which mineral is delivered from the face con¬veyor to a stage-loader conveyor) and the tail entry are predriven before the longwall units starts production. In the longwall advancing system, the longwall unit makes the head and tail entries as the face advances, utilizing various techniques to prevent the entries from being destroyed as they are left in a consolidating gob area. Fig. 3 illustrates the comparative approaches of the longwall retreating and advancing operations. As with most other underground mining systems, there are many variations on these two basic systems, using combinations of the systems that are dictated by

Jan 1, 1982

By S. Horton, F. L. Stanley, P. J. Bethell

Introduction The McClure River Processing Plant is located at McClure in Dickenson County, VA. The plant was originally constructed in 1979 as a 600 st per hour facility. The plant is fed by coals from numerous mines in the area as well as a major longwall mine onsite. Coal is processed predominantly for the metallurgical coal market. The original plant was comprised of three separate processing circuits incorporating a heavy medium vessel for the coarse material (+10 mm). Intermediate sized material (10 x 0.6 mm) was beneficiated in heavy medium cyclones, with the remaining fines (0.6 mm x 0) passing to froth flotation for upgrading. The products from all three circuits were combined prior to thermal drying and then stored in silos or ground storage prior to shipment to the metallurgical coal markets. A middlings heavy medium cyclone circuit was also present to rewash the vessel and primary heavy medium cyclone rejects. The original circuit is as shown as Fig. 1 (Editor's note: All figures are reproduced at the end of this paper.) In reasonably close proximity to the McClure Plant is the much older, less efficient Moss 1 Plant, (Chance cone, table, water only cyclone, froth flotation, Fig. 2), which prior to the McClure Plant expansion was processing considerable tonnages of coal. At that time, the McClure River Plant was processing coal at the maximum possible feed rate, and the Moss 1 Plant was finding increasing difficulty in maintaining the stringent quality specifications on its major metallurgical coal accounts. Given the fact that many of the feeder coals to Moss 1 could be moved with minimal difficulty and cost to McClure, an exercise was conducted to evaluate the potential of expanding the McClure Plant. This would result in transferring coals from Moss 1, thereby increasing yield, reducing operating costs, and enabling customer specifications to be met more easily. After evaluating many possible ways of expanding the McClure Plant, we eventually settled on an approach based on the idea of Gallagher at German Creek in Australia (Gallagher, 1986). Our objective was to increase plant capacity by removing the coarser size fractions (0.60 x 0.15 mm) from froth flotation and at the same time eliminating the finer fractions (1.25 x 0.6 mm) from the heavy medium cyclone circuit, thereby allowing these circuits and therefore the plant as a whole to increase its capacity. These size fractions (0.60 x 0.15 mm from froth flotation and 1.25 x 0.6 mm from the heavy medium cyclone circuit) would be treated in a new spiral circuit. After desired evaluation of the various circuit capacities, including desliming screen, clean coal filter and thermal dryer capacities, etc., it was concluded that we should be able to maintain high levels of processing efficiency in the existing circuitry while at the same time expanding plant feed rate from a nominal 600 st per hour to 1000 st per hour, thereby achieving the objectives The circuit that was finally settled on is shown in Fig. 2. Several unique features were incorporated in this circuit in an attempt to maximize processing efficiency and reduce maintenance costs. First, the idea of a single-stage spiral circuit was abandoned in favor of a two-stage circuit incorporating rewashing the primary spiral middlings in a secondary spiral (Bethell, 88). The spirals selected were Mineral Deposits Model LD4 units. It was felt, after much previous testing, that spiral middlings rather than being a true middling gravity fraction were in fact a combination of true middlings with a preponderance of misplaced light coal and high-density rock. Consequently the insertion of this material directly into either clean coal or refuse streams (per normal practice) would only lead to loss of efficiency. The operating philosophy behind the design of the circuit was to produce the purest coal and rejects possible by adjusting the spiral splitters. Misplacement of refuse into the clean coal and vice versa would be minimized by taking a fairly large middlings cut. By reprocessing this material (10%-15% of primary spiral feed) in a secondary bank of spirals the misplaced material in the primary middlings would be reprocessed and its probability of reporting to the correct stream greatly enhanced. To minimize fines loading on the spirals, fine aperture sieve bends were to be included. A further novel feature of the circuitry is the desliming/ dewatering system. A philosophy of accepting inefficiency but forcing it to work in our favor was developed. The dewatering screens (Linatex Combis) were chosen not for their ultimate screening efficiency but for their ability to remove slimes (tolerating a fair quantity of oversize in the effluent) and their enhanced maintenance characteristics. Any oversize passing into the effluent, together with the fine coal centrifuge effluent, would pass to classifying cyclones, the objective of which was to reject to overflow (and hence froth flotation) as many fines as possible (-150 µm) without running the risk of losing coarse (1.25 x 0.15 mm) material to oversize. Hence, by a process of producing a "slime free" screen product coupled with a "coarse free" cyclone overflow, sizing efficiency could be maximized. Obviously, recirculating load had to be factored into the Combi Screen design capacity. The spiral circuit as shown in Fig. 2 was installed at McClure River in January 1988. The spiral circuit efficiency was very high immediately, and, after the normal debugging, plant capacity has been increased to the point at which we are currently processing of the order of 1100 st per hour. Test program The performance of the spiral circuit has been evaluated by two major samplings of the entire new circuit, coupled with routine bi-monthly tests on all other major components of the plant circuitry.

Jan 1, 1992

By T. H. Eyde

Estimated natural zeolite production worldwide during 2004 remained at about 3.6 Mt (4 million st), about the same as in 2003. China continues to be the largest producer of natural zeolites followed by Cuba, Germany, Japan and South Korea. This may be changing, however. Since 1988, zeolite production from Australia, Indonesia, Malaysia, New Zealand, Philippines and Thailand has increased rapidly. Most of the production is consumed in aqua culture. Global demand for seafood has exceeded the ability of the oceans to supply it without over fishing. Zeolite minerals remove ammonia compounds from the ponds in which fish and shrimp are grown. It is estimated that, in Thailand and Indonesia more than 55,000 farms produce260 kt (286,500 st) of shrimp. Therefore, as shrimp production increases, so does zeolite consumption.

Jan 1, 2005

By Eleanor Sonley

In October 2000, temporary seismographic stations were installed above the Trail Mountain Mine, Emery County, Utah, USA to supplement a regional network. Over seven months, 1826 mining-induced seismic events were recorded with a magnitude range of M 0 to M 2.2 (Arabasz et al., 2002, 2005). Routine event locations for this dataset cluster around the longwall mining activity during the recording interval. However, applying a double difference relocation technique to the routine locations significantly improves the correlation between mining activity and event location. The resulting relocations depend greatly on station distribution and the number of arrival time picks. We use this dense temporary array to assess station distribution requirements for applying the double difference technique at other mines by systematically removing data from available stations, thus simulating sparser networks. The insight gained using the Trail Mountain dataset is used to develop a metric for assessing the usefulness of above-mine networks and to develop a tool for expanding seismic network coverage. To verify the appropriateness of these tools for wider application, we examine data from two additional mines located in the western United States. Improving the relative locations of mining-induced seismic events highlights the association of these events with active portions of the mine. Clusters, regions of quiescence and relative timing of events can be used by mine operators to assess hazard, optimize production, and evaluate performance of mine plans. The lessons learned from this study may be used in future network design to improve accuracy of event locations at other mines.

Jan 1, 2009

By Charles Morris, Patricia Petty

Charles Morris Minerals engineers have a new opportunity to shape their industry's future. They can help secondary school teachers in their own communities improve their scientific knowledge and thus be better prepared to teach their students. Local sections of AIME and other groups are being asked to sponsor expense-aid scholarships for teachers who complete the summer field course A Total Concept of the Mining Industry. Now in its 18th year, the six-week summer course is presented by the Colorado Mining Association Education Foundation Inc., in cooperation with the Colorado School of Mines Office of Special Programs Continuing Education. The course will be held June 9 through July 18. "In the past, we have been successful in soliciting support for teacher-scholars from many organizations, corporations, and individuals within the industry," said John Robertson, Jr., Foundation chairman. "Now we hope to enlist the help of engineers at the grass roots level in identifying qualified local teachers and giving them what it takes to participate in this continuing education program." Robertson said the course provides secondary school teachers, counselors, and administrators with a working knowledge of the mining and metallurgical industries. It also helps teachers recognize career opportunities for their students. "With the industry facing a challenging new era, we must have teachers who can relate the excitement of this challenge to their students. They will inspire future generations of resource finders, developers, and managers," he said. More than 670 teachers in 27 states, including 45 graduates of the "class of '85" have completed the course and gained six semester hours of graduate credit. The Foundation's Planning and Curriculum Committee meets regularly to develop the course outline and recruit the faculty. Lecturers are selected on their ability to communicate with people not familiar with the industry. Last year, 67 faculty members came at their own time and expense to discuss their specialties, ranging from exploration to marketing. Normally, teachers receive scholarship aid in the form of tuition and fees, and transportation and meals while on field trips. Expense-aid scholarships cover room and board on campus for those beyond the Denver-Metro area, and travel to and from Golden. "We hope to broaden the appeal and the geographical range of the program," Robertson said. Inspiration for the field summer program for teachers came in 1967 from a mining engineer, the late C.L. (Lindy) Barker, a 1931 graduate of the Colorado School of Mines. Barker and a group of concerned individuals convinced the Colorado Mining Association to accept sponsorship of the course in 1968. The Colorado School of Mines and its Office of Special Programs and Continuing Education agreed to be a co-sponsor to give it accreditation and provide physical facilities and coordination. The first course was held in 1969. Operating funds were solicited from industry to pay tuition, course material, the cost of a course coordinator, field trips, and overhead. Through the years, personal contact by an industry representative has been the most effective teacher recruiting tool. Through expense-aid scholarships provided by local sections of professional organizations, such as AIME, as well as CSM alumni and other friends, there will now be a greater incentive for more teachers to get to know the minerals industry and pass that knowledge on in the classroom.

Jan 3, 1986

By N. Finkelstein, C. G. Palagi, S. S. Stillar

The C. E. Weed concentrator treats 51,000 tons per day of 0.5% to 0.7% copper ore received from the Berkeley Pit, the Continental East Pit and the Butte underground mines. Improved operating techniques and conversion from semi-autogenous secondary grinding to ball milling have increased tonnage from an original design of 42,000 tpd. Separate treatment of sands and slimes, with recovery of soluble cop- per from the slimes, has increased total copper recovery. Concentrator tailings are pumped 600 vertical feet and 20,000 horizontal feet for disposal.

Jan 1, 1976

By John Frostiak

The Campbell gold operation of Placer Dome Inc. produces an arsenopyrite-sulfide concentrate that must be treated by roasting, biooxidation or pressure oxidation to obtain high gold recoveries. Pressure oxidation is a proven process that also meets Ontario's stringent environmental regulations. A comparison of biooxidation and pressure oxidation was completed early in 1990. Placer Dome selected pressure oxidation as the preferred process to replace the roaster. The commercial plant design was based on technology supplied by Sherritt Gordon. In July of 1991, the existing roaster operation was shut down and a new pressure oxidation plant was commissioned. The Campbell Red Lake Mine, in Balmertown in northwestern Ontario, is the first plant in Canada to use Sherritt Gordon's pressure oxidation technology to recover gold from a refractory ore body. A similar process has been used at Sao Bento in Brazil since 1986. The Porgera complex in Papua New Guinea also uses Sherritt's technology. At the Campbell Mine, gold occurs as either free gold or is encapsulated with sulfide minerals, mainly arsenopyrite, pyrite and pyrrhotite. These sulfide species are recovered from the ore by crushing and grinding to 80% -75. µm (200 mesh), followed by flotation. About 40% of the gold in the ore is recovered by gravity separation. The remainder reports to the ground slurry feed to flotation. The flotation concentrate averages about 55 to 60 t/d (60 to 66 stpd). A typical analysis is shown in Table 1. Before the installation of the pressure oxidation plant, the flotation concentrate was treated in a two-stage, fluid bed roasting circuit. Roaster off-gases passed through two parallel sets of cyclones, followed by electrostatic precipitators. Arsenic trioxide was captured in a four-chamber baghouse before the gas was discharged to a stack. The arsenic trioxide was sold or conveyed to an underground storage vault. The roasted calcine was fed to a gravity separation circuit and then subjected to cyanide leaching. Gold was recovered from the filtered leach solution in a Merrill-Crowe zinc precipitation circuit.

Jan 1, 1992

By Elif Çevik

This study deals with the mineralogical-geochemical characteristics and the industrial usage of the olivine from the dunites that crop out on the west of the Topuk village (Orhaneli). Mineralogical compositions of non-altered dunite samples are averagely as: Olivine (>90%), serpentine (1-4 %), opaque minerals especially chromite, magnetite, pyrite (1-3 %), and secondary minerals as magnesite, chlorite (1-2 %). X-Ray Diffractograms of the dunite samples indicate that olivines are in forsterite (93 on an average) composition, and chlinochrysotile and lizardite accompany to the forsterite. Chemical analyses of the dunites show that MgO, SiO2, Fe-,03 and LOI contents vary from 40,99 to 50,2 %, 34,67 to 40,6 %, 6,8 tol3,7 %, 0,11 to 11,27 %, respectively. SE-TAT Mining Company states that the study area has 400.000.000 tonnes proved, 800.000.000 tonnes probable and totally 1.200.000.000 tonnes olivine reserve. Consequently, all studies indicate that Orhaneli Olivin has convenient mineralogical and chemical properties and the grain size (0.04-40 mm) for industrial usages such as abrasive, olivine sand products, steel slag conditioner, balast and covering material in industry: Orhaneli olivines are demanded especially from the firms related to refractory industry, iron-steel factories, naval forces and shipyards. SE-TAT Mining Company also exports these olivines to different countries such as Macedonian and Spain.

Jan 1, 2005

By John C. Draper

The principal object of an underground coal mine transportation sys- tem is to move coal from the face where it is produced to the outside of the mine where it is prepared for market. In addition to the movement of coal, a transportation system must also move supply materials to the inside; it must move men in and out of the mine; and it must move rock and other debris out of the mine for disposal. Usually, mine haulage is divided into three categories named for the various areas of the mine served : 1) Pace Haulage. This is used at face from the mining machine to the intermediate transport system. 2) Intermediate or Secondary Haulage. Generally this is used between face haulage and main haulage. Sometimes this is merely an extension of the main haulage. 3) Main Haulage. This extends from the secondary haulage to the outside and is usually either a track or conveyor system, sometimes a combination of both. Transportation modes of coal in the various areas pose different problems, but the whole system must be synchronized. For instance, at the face the coal should be removed quickly from the various places to facilitate the movement of machines, whereas the intermediate system just passes the same coal on to the main transportation system. Selection of the transportation system must be based on the seam height, mining methods used to extract the coal, capacity required, ease of coal handling, and haulage efficiency. When driving rooms, mine cars can be used. They often simplify supply haulage and rock handling because they can also handle these items. Conveyors could be used and would have a greater haulage capacity but they cannot handle supplies easily or segregate the rock. When using a different mining method, such as longwall, mine cars cannot be used because there is not sufficient space to move the cars at the face nor can the cars haul the coal without delays which would cause excessive losses in production. As for ease of handling at a longwall face, conveyors are convenient because of their low height, ruggedness, and ability to operate while being advanced. Similarly, in low seam heights, conveyor systems are easier and more economic to operate than the regular track haulage because roof has to be taken down for handling mine cars of adequate size.

Jan 1, 1981

I was very interested in Professor Park's data on subsidence over longwall panels. His paragraph labeled "Conclusion," however, does not seem appropriate for the article. Nothing in the article gives any discussion of how knowledge of subsidence characteristics can be used in mine design to minimize damage. That thought may have been a reason for undertaking this study, but not a conclusion of it. From his article, I would conclude that subsidence movements last considerably longer at great depth. And the National Coal Board's subsidence prediction model represents this area reasonably accurately, though most literature claims their model predicts far greater disturbances than actually occur in most areas of the United States. Also, I would like to know, is the method used to determine angle of draw? The criterion used can make a significant difference in the value obtained. With 30 m (100 ft) spacing and a 122-m (400-ft) depth the angle between stations can be up to 14°.

Jan 1, 1989

By A. G. Moon

Kennecott Copper Corporation's Chino Mines Division concentrator is located in Hurley, New Mexico. Construction of the concentrator was started in 1910 and its capacity was progressively increased over the years to the present nominal 20, 425 t/d of ore (22, 500 stpd). The principal copper minerals present in the ore are chalcopyrite and chalcocite with much of the chalcotite intimately associated with pyrite. Lime is used to depress pyrite, and flotation at a high pH or alkalinity is required for optimum recovery of chalcocite with good rejection of the pyrite. Regrinding is also required to liberate chalcocite-pyrite middling grains. Accurate control of the lime addition is necessary to obtain both optimum metallurgy and minimum lime usage. At low pH levels, flotation cell control is difficult due to excess froth, but excessively high pH levels can go unnoticed because froth characteristics do not change from those experienced at the deisred pH. With manual control of the lime addition, the desired pH can be exceeded for periods of up to two hours. In 1972 a major renovation of the flotation and regrinding circuits was com¬pleted and at that time a new lime distribution and control system was placed in operation. Limestone is obtained from a nearby quarry west of Hurley. The limestone is crushed, sized and fed to a rotary kiln and the quicklime product is stored prior to slaking. The lime is slaked, as required, in a mixing tank-classifier type slaker and the lime slurry is pumped to a concentrator storage tank from which it is distributed through a conventional closed loop system. The current average lime usage rate is approximately 6 kg/t of ore (12 lb per st).

Jan 1, 1976

By Ellis G. Goree, George A. Simmons, Andrew B. Pervis

The pneumatic classifiers are widely used for fine dry powders separation These classifiers can be used for either the removal of fines from coarse feeds or the removal of coarse particles from ultrafine feed To optimize the operation of pneumatic classifiers, a thorough understanding of the impact of operational parameters on the efficiency of classification is essential This paper describes the design features and effects of operational parameters followed by illustrations of pneumatic classifiers usage in two case studies

Jan 1, 1983

By Douglas R. Cook

The Ertsberg East discovery was announced by Freeport in 1979 with proven geological reserves of 45 Mt (49.6 million tons) with a grade of 2.65% copper and recoverable gold values. This recently discovered deposit is one of a cluster of skarn-type copper occurrences found within the limits of Freeport Indonesia's mining con- cession at an elevation of about 3600 m (12,000 ft) in the extremely rugged Jaya Wijaya Mountains of Irian Jaya, the Indonesian half of the island of New Guinea. The Ertsberg East lies about 1.5 km (0.9 miles) east of the Ertsberg ore body, which is also known by the Indonesian name "Gunung Bijih" or "ore mountain." For general background, the original Ertsberg ore body was first discovered by the Dutch geologist Jacques Dozy in 1938 while on a mountain climbing expedition. It was mapped and sampled by Forbes Wilson and Del Flint after a two-week harrowing trek into the mountains from the coastal swamps. A drilling program in 1967- 1968 established that the Ertsberg was a viable mining venture, containing approximately 33 Mt (36.4 million tons) of skarn-type copper ore. The deposit averaged 2.5% copper, with significant values in gold and silver. The first shipment of copper concentrates was made in December 1972, following a capital investment of $200 million. Following the routine production of ore from this deposit, it became necessary to consider additional sources of ore to extend the life of the mining operation. An exploration program was thus started with the initial objective of finding shallow or surface ores which could be blended with the material being mined at Ertsberg. Project geologist Frank Nelson was transferred from Freeport Exploration Co. to the Freeport Indonesia organization at the Ertsberg mine. However, the exploration work was the functional responsibility of Freeport Exploration. Del Flint had overall responsibility for the program from the Reno headquarters office and provided liaison with management in the USA and Indonesia.

Jan 1, 1991

By W. H. A. Timm

Lake Dufault Mines, Limited, is located ten miles north of Noranda, Quebec. Following discovery of a copper-zinc orebody late in 1961, metallurgical investigations of diamond drill core were conducted by Lakefield Research Limited during 1962 and 1963. These were followed by pilot-plant tests in May and June of 1964, utilizing underground development ore. Milling operations commenced on August 16, 1964, and reached rated capacity of 1,300 tons per day by October 1, 1964 (Figure 1). Since that time, tonnage was gradually increased to a maximum of 1,400 tons per day and since January 1, 1968, gradually decreased to the present rate of 1,100 tons per day due to an increase in the hardness of the ore. The ore milled from start-up until January, 1968, was mainly massive sulphides, consisting of a complex mixture of iron, copper, zinc and minor amounts of lead sulphides with important associated values in gold, silver and cadmium (aver- age specific gravity 3.8). From January, 1968, until the present, the ore being processed is described geologically as a boxwork of sulphide stringers of chalcopyrite, pyrite and pyrrhotite and some local dissemination in siliceous rhyolite (average specific gravity 2.9) supplemented by the massive copper-zinc ore. The limited size of the orebody and the mining method employed (open stoping by blasthole and slash and bench methods) would have made grade control from the mine difficult and expensive. It was decided to try to cope with grade variations in the mill. The resulting rapid and wide swings in mill heads, hardness and specific gravity have presented many and varied challenges for the mill personnel.

Jan 1, 1970

By Roy E. Williams, Muriel S. Robinette

Ground-water contamination from uranium operations could be avoided with proper site analysis techniques. Contamination from tailings disposal sites is carried downgradient within permeable hydrostratigraphic units which can be located and delineated with a combination of drill holes and borehole geophysical logs. Engineering safeguards can be designed into a disposal site by determining if permeable hydrostratigraphic units exist which will dominate the flow system and provide seepage paths for contaminates. It is more cost-effective to accurately assess seepage potential with geophysical logs before contamination becomes a problem than to be faced with a major clean-up effort after contamination has occurred. If a uranium waste impoundment already has a ground-water contamination problem, then borehole geophysical logs may be used to delineate the contaminated zones which need to be cleaned up. It is less expensive to withdraw the contaminated plume from only within the affected zones rather than attempt to dewater the entire area. Consequently, the geophysical logs are invaluable in assessing which zones should be pumped and where packers should be set for construction of piezometers to allow monitoring of water-quality changes. A surface geophysical technique is being tested and developed for monitoring of seepage frum uranium tailings impoundments. Geoelectric soundings and profiles from direct current resistivity surveys can locate anomalous areas for the placement of drill holes to better assess seepage. This reconnaissance technique should eliminate random hit-or-miss drilling at sites where monitoring is required. Tests and experiments are being done on developing the capabilities of this technique to incorporate complex resistivity data to assess cultural impedence noise which complicates simple resistivity data.

Jan 1, 1980

By E. Finch

Beneficiation tests were conducted on samples from three ore zones in the Los Monos, Navay orebody in Táchira State, Venezuela as part of an overall feasibility study. The scope of work included ore sample characterization, flowsheet development, and optimization of beneficiation variables on a bench scale followed by confirmation on a pilot plant scale. Phosphate in the ore samples occurred primarily as francolite and was associated with quartz, kaolinite, and feldspar. Liberation of the phosphate began at 589 microns (28 mesh) for the main ore zone sample and at 295 microns (48 mesh) for the lower-grade ore zone samples. Development tests on simulated mining composites demonstrated that the samples responded well to 589 micron (28 mesh) grinding, two stages of high intensity scrubbing, conditioning at high % solids with caustic and equal amounts of tall oil and fuel oil, and separate coarse and fine flotation circuits with one stage of roughing and two stages of cleaning (reflotation). Concentrates in the range of 31.5-34% P205 with recoveries from 64-70% were obtained from the mining composites. Product quality for all concentrates in the range of 31.5 % P205 was good; 1.40 CaO/P205 and 0.04-0.05 minor element ratios. Confirmation testing in a continuous pilot plant produced results similar to those achieved in the laboratory.

Jan 1, 1993

By S. Di Muzio

Sulphate Reducing Bacteria (SRB) are hetero-trophic anaerobic micro-organisms that can be used to remove sulphates and heavy metals from Acid Mine Drainage (AMD), heavy metal-bearing streams with high iron and sulphate contents and low pH, generated by sulphur bio-oxidation in abandoned mine districts. After the preliminary chemical precipitation of iron with carbonate stones (I step, chemical precipitation) SRB can be used to reduce sulphates to sulphurs then favouring heavy metal removal as sulphur precipitates (II step, biologically mediated precipitation). In this work the experimental results of heavy metal and sulphate removal from column tests were reported. Sulphate and heavy metal abatement in fixed bed column was performed using chemically pre-treated samples of synthetic AMD. Attenuation of toxic effects of treated samples was also ascertained by toxicological tests of germination and radical growth with Lepidium sativum.

Jan 1, 2006

By L. A. Arrington-Webb

Forty-four states produced clay products during 1994 for a total value of $1.6 billion. This was a 3% increase in total tonnage and a 7% increase in value from 1993. Kaolin Domestic production of kaolin during 1994 was estimated to be 9.3 Mt (10.2 million st), according to the US Bureau of Mines (USBM). Georgia remained the leading producer with 7.5 Mt (8.2 million st), followed by South Carolina with 535 kt (589,000 st). Kaolin used as a paper coating or filler pigment accounted for 51% of domestic consumption. Other uses included grogs (14%), fiberglass (5%) and plastics (4%). A reclassification by an Alabama producer of fire clay to kaolin accounts for the introduction of kaolin used in grogs. This also accounts for the apparent decrease in the use of kaolin in paper from 60% in 1993. However, the actual amount of kaolin used in paper was the same.

Jan 1, 1995

By A. Probst, N. Moores, J. A. Finch

In 1994, McGill University and INCO Ltd. (Sudbury operation) initiated the transfer of technology originally developed to detect levels in equipment ranging from flotation columns to thickeners. A conductivity probe was installed in a concentrate thickener to monitor the mud line and the solids load. The probe consisted of a vertical tube with 32 equally spaced electrode rings connected in pairs to give 31 conductivity cells. Conductivity was measured as a function of depth to produce a conductivity profile. From the conductivity profile, the mud line and the percent solids can be inferred. The probe was installed in October 1994, and it has been successfully detecting profiles since then. This paper includes a description of the incidents that occurred and the experiences gained during this technology-transfer process. Discussed is probe installation and startup; software modifications to accommodate both operating conditions and the needs of the operators; and calibration procedures to adjust to changes in performance as a consequence of long-term use. Also included are preliminary observations of the still-in-progress last stage, i.e., signal interpretation and use to facilitate operation, particularly by providing early reaction to upstream disturbances.

Jan 1, 1999

By Stephen H. Stow

Economically important heavy minerals have been studied in the sediments of offshore Alabama and Mississippi. Sandy samples have been collected from a 200 square mile area from Dauphin to Horn Islands. The minerals concentrate in sediments under 10-35 feet of water and may reach 2.4% by weight. The distribution of total heavies is related to the hydrodynamics of offshore Alabama-Mississippi; the greatest concentrations are in the Gulf where the longshore current is diverted from its westerly flow. Preliminary observations indicate that the distribution of individual minerals can be correlated with wave energy, particle size, specific gravity, and morphology of the mineral grains. The coarser, lighter minerals--kyanite, sillimanite and staurolite--are concentrated in relatively low energy areas in the Mississippi Sound and at the end of Dauphin Island in the Gulf; maximum concentration for each is 34.5%, 6.0% and 43.2%, respectively, of the heavy minerals. The smaller, more denseminerals--ilmenite, rutile, leucoxene, zircon and monazite--are concentrated in high energy areas in the Gulf; maximum concentration for each, respectively, is 42.0%, 5.9%, 9.652, 10.7% and 1.3% of the heavy minerals. On the basis of a minimum concentration of 1% heavy minerals, two areas have been identified and the total amounts of different minerals have been calculated to a depth of 25 feet of sediment. The data are preliminary and in no way indicate recoverable reserves, but they do show that on the order of 20 million tons of heavies exist. These deposits should be termed "indicated resources".

Jan 1, 1975