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By Manuel Villanueva Bullón

El presente texto describe el proceso de optimización en las operaciones mineras del yacimiento minero San Vicente a través del empleo de las microcomputadoras, las cuales al ser fácilmente transportables, de bajo costo y no requerir de ambientes especiales, permiten brindar in situ con amplitud al soporte que se requiere. Por ello, esta red de microcomputadoras desarrollan sistemas de carácter técnico y administrativo que controlan y retroalimentan información de todas las operaciones, optimizando el control de la producción en la mina y planta concentradora.

Nov 24, 1986

By Oscar Marreros Daza

Dada la complejidad de las rocas y cuerpos mineralizados en el Tajo Mc Cune de Cerro de Pasco, es más difícil calcular los parámetros óptimos de perforación y voladura por métodos manuales. En este sentido, el presente texto describe un programa en Basic para la computadora IBM Pc, donde se puede hacer un análisis de estándares de voladura y costos para cualquier diámetro de explosivo, tipo de roca, explosivo y altura de banco. Asimismo, se considera el método de cálculo de los estándares de perforación y voladura perteneciente a Richard L. Ash y se realiza el análisis de voladura con ANFO para los diferentes tipos de roca dado su bajo costo y rendimiento

Nov 24, 1986

By Luis de Montreul

El presente texto describe la mineralogía del oro que comprende sus asociaciones con otros minerales metálicos, relaciones con las gangas no metálica, granolometría, abundancia y las características textuales de los minerales. Para ello, es vital conocer la selección del método de tratamiento a emplear para obtener una óptima recuperación del oro.

May 1, 1986

By Joseph Werner

El presente texto define la función y metas del supervisor de plantas concentradoras a través de tres términos: análisis(A), balance (B) y Control (C ). Asimismo, lo novedoso de la supervisión consiste en la sistemática integral del método aplicado, basándose en la formulación de modelos matemáticos de "simulación manual" orientado al tipo de plantas de mediano tamaño que se instalaron en el pasado por etapas sucesivas.

May 1, 1986

By María Jesús Ojeda

El presente texto describe la aplicación de los principios de exploración geoquímica en un sector de la faja aurífera Nazca -Ocoña correspondiente a la mina "Las Agripiinas". El objetivo es determinar anomalías de oro en base a un muestreo sistemático de afloramientos y niveles ya en desarrollo y así ubicar nuevas áreas favorables para la realización de labores mineros que conduzcan a aumentar las reservas de la mina. La exploración por oro se realizó en rocas metamórficas esquistosas con vetas de cuarzo-pirita, intrusivos granodioríticos y volcánicos andesíticos caracterizados por procesos de dispersión primaría. La dispersión secundaria se determina en óxidos, producto de la alteración de los ferromagnesianos y sulfuros.

May 1, 1986

By Luis Mercado

El presente texto analiza el panorama global nacional del carbón en el Perú, así también señala las posibilidades de este como alternativa energética frente al petróleo, debido al alza de precios de producción y venta. Para ello, se revisa las reservas de carbón, el contexto geográfico y geológico de las mismas y los métodos de explotación de los países industrialmente desarrollados.

May 1, 1986

By L. J. Buter, D. J. Collins

Introduction FMC's Paradise Peak gold mine was discovered by staff geologists in July 1983. Located south of Gabbs, in western Nevada, the deposit contains about 11 Mt (12 million st) of ore grading at 3.6 g/t (0.105 oz per st) of gold and 129 g/t (3.75 oz per st) silver. It is estimated that the ore contains 34 t (1.1 million oz) of recoverable gold and 933 t (30 million oz) of recoverable silver. The 3.6 kt/d (4000 stpd) mill began operation April 14, 1986, just 18 months after Davy McKee completed a feasibility study and construction, and two-and-a-months ahead of schedule. It also came in about 10% under budget at a cost of about $100 million. Crushing Run-of-mine ore is delivered from the mine in 45-t (50-st) haul trucks to a 272-kt (300,000-st) capacity open air stockpile. Ore is reclaimed using 5 m3 (7 cu yd) front-end loaders that feed a 91-t (100-st) hopper equipped with a 610- x 610-mm (24- x 24-in.) grizzly. A 1.4- x 11-m (54-in. x 35-ft) apron feeder delivers ore from the hopper by way of a vibrating grizzly screen with 127 mm (5 in.) openings to a 1- x 1.3-m (43- x 51-in.) Rexnord jaw crusher. Crushed material is conveyed to a 1.8- x 4-m (6- x 14-ft) Link Belt vibrating screen equipped with a 50-mm (2-in.) manganese grizzly bar deck. Screen oversize is fed to a 1.7/0.6-m (5.5 ft) Rexnord standard cone crusher. Product from this crusher joins the -50 mm (-2 in.) screen undersize and is conveyed to a 136-t (150-st) surge bin. Three hydraulically driven belt feeders deliver ore from the ore bin to three 2 x 6 m (7 x 20 ft) Link Belt single-deck tertiary screens fitted with 6 x 25 mm (0.25 x 1 in.) slotted polyethylene screens. Screen undersize - -6 mm (-0.25 in.) rock - is conveyed to a 3.6-kt (4000-st) ore storage bin. Screen oversize is delivered by way of a conveyor to a 91-t (100-st) surge bin. From this bin, two belt feeders feed ore to two 1.7 m (66 in.) Rexnord Gyradisc crushers, which are in closed circuit with the tertiary screens. Design features include magnetic protection for tramp iron removal and sloping floors and pump sumps for ease of clean up. Dust scrubbers keep the working environment clean and in compliance with state dust emission regulations. Grinding Two 1 x 4.5 m (3.5 x 15 ft) belt feeders deliver ore from the fine ore bin to a conveyor belt that feeds it to the 5 x 7 m (16 x 22 ft) Allis-Chalmers ball mill. The conveyor belt is equipped with a Merrick weightometer. This controls feed rate to the ball mill by speed control on the hydraulic motors driving the feeders. The 2.6 MW (3500 hp) powered mill grinds 3.6 kt/a (4000 stpd) to an average grind of 85% -75 µm (-200 mesh). The mill is in closed circuit with four 660 mm (26 in.) Krebs cyclones. Three of these are normally in operation while one is a standby. Warman pumps feed the cyclones through a cyclone distribution manifold. The cyclone underflow returns to the ball mill, which has a 300% average circulating load. Mill discharge pH is 11, and that is obtained through pebble lime addition to the feed belt. Cyclone overflow is sampled using a Heath & Sherwood two-stage sampler. It then flows to a 27-m-diam (90-ft-diam) Eimco thickener. Thickener overflow is returned to the grinding circuit, while underflow at 50% solids is pumped to the leaching circuit. Grinding-circuit design features include a steeply sloped floor for ease of clean up, an open area over the pumps for maintenance accessibility, and a ball hopper with a self discharge bottom to speed ball addition. Other features include a mill liner handler and an air clutch that connects the mill motor to the mill.

Jan 12, 1986

By G. H. Lewis, C. L. Krcil, N. S. Harding, D. E. Havekotte

Introduction Limestone injection work was completed in a 440-kW (1.5-million Btu per hour) pilot-scale combustor in preparation for a 50,000-kg/h (110,000-lb per hour) industrial steam boiler limestone injection test program. The main objectives were to: (1) identify the important variables that affect sulfur capture, (2) determine the maximum SO2 removal, and (3) determine the effect of limestone addition on furnace operability and heat transfer. A more detailed report is given by Lewis et al. (1985). The test program was divided into two phases. The objectives of Phase I were to determine the important variables that affect SO2 reduction and to investigate the effect of limestone addition on combustor operation and heat transfer surfaces. The objective of Phase II was to determine, using only the important variables identified during Phase I, the conditions that resulted in the greatest sulfur dioxide reduction. Phase I Test program Using one limestone, five variables were identified as having an effect on SO2 reduction. These are noted with their respective ranges: heat release rate [103 to 166 kW/m3 (10,000 to 16,000 Btu per hour-cu ft)], Ca/S molar ratio (0 to 3), flue gas oxygen concentration (3% to 5%), additive particle size [90% -150 µm (-100 mesh) and 90% -45 µm (-325 mesh)], and burner stoichiometry (SR = 0.75 and >1.15). A 2.3% sulfur northern West Virginia coal, pulverized to a nominal grind of 70% -75 µm (-200 mesh), was burned for the tests. The limestone used was obtained from a Maysville, KY, quarry and contained 51.75% (wt) CaO. An 11-run statistically designed program (plus three duplicates) was used to screen the five variables. The effect of limestone injection on the combustor heat transfer surfaces was determined by comparing the heat transfer rates through a slag panel and fouling probes during periods with and without limestone injection. Combustor operation was determined by noting any operating problems that occurred during the tests. Test procedure After furnace wall temperatures had stabilized at operating conditions, heat transfer surfaces (slag panel and fouling probes) and the convection section were blown clean, and a four-hour test was begun. The

Jan 12, 1986

By D. M. Avant, A. J. Solomon, S. M. Pickering, W. R. Fox

Introduction Most high purity silica is produced from crystalline quartz, generally from sand. However, there is a small area of southern Illinois where a porous, friable white chert rock is found that is very different in physical form from quartz sand. This microcrystalline material, commonly referred to as tripoli, is more easily ground to micron sizes than sand. Although of similar Moh's hardness, microcrystalline tripoli is less abrasive than either ground quartz sand or many synthetic abrasives. Location The southwestern corner of Illinois is underlain by extensive deposits of very porous chert; in some places the material is quite white. Two companies, Illinois Minerals Co. and Tammsco Inc., mine and process this white material. The Illinois Minerals plant is located at Elco, and the Tammsco plant is in Tamms, IL. Both companies mine from various locations within a 10-mile truck haul of their plants. Production from this district has been continuous since at least 1915. Malvern Minerals Co. mines and processes a generally similar microcrystalline, friable novaculite near Hot Springs, AR. Geology The southern Illinois white tripoli area is on the southern flank of the Illinois Basin. The commercial silica zone is a facies of the Clear Creek chert of Lower Devonian Age. To the north, where it occurs deep in the basin, the Clear Creek is a siliceous limestone. In the mining area, thin overburden has allowed access by surface ground water, which has leached away the carbonate content and left a quite pure silica residue. Fossils characteristic of sea bottom deposition, such as brachiopods and trilobites, are occasionally encountered in mining. The structure of the mining area is poorly known, due to infrequent outcrops. Strata in most mines are inclined at dips of 5° to 20°. There appear to be occasional high angle normal and reverse faults, but these are seldom of any significance in mining. The friable white silica zone mined in the Clear Creek chert is overlain by dark brownish-red chert and siliceous shale. Exploration drilling to date has not encountered the base of the microcrystalline tripolitic zone. Chemical content Leaching by ground water solution has removed virtually all the carbonate from the original siliceous limestone. Impurities encountered are red to yellow iron oxides, aluminous clays, and occasional black manganese oxides. The silica has no detectable formula water. A typical chemical analysis and physical properties are listed in Tables 1 and 2. Mineralogy and crystallinity Southern Illinois friable tripoli appears amorphous under the polarizing microscope. No crystal structure and only occasional, vague, patchy anisotropism is evident. However, the material gives an excellent quartz X-ray diffraction pattern. Thus, so far as X-ray mineralogy is concerned, the material is clearly

Jan 12, 1986

By J. B. Gwiazda

S.S. Peng The highlights of this paper are : • The shield supports with lemniscate links will exert a tangential force on the roof strata, causing unnecessary and harmful roof movements, thus damaging the roof structure. The principal source of the tangential force is the horizontal movement of the point O, the pivotal joint between the canopy and the caving shield. When the support height is changed during setting, the point O will move along a curve upward and rearward toward the gob. As a result, it produces a vertical movement, causing the roof to move up a distance of ?d, and a horizontal movement toward the gob, creating vertical fractures in the roof. • The major purpose for employing the lemniscate-link system is to protect the hydraulic legs against the action of the bending moments induced by the tangential force that causes deformation or destruction of the legs. • Based on these two points, the author suggested that, in order to eliminate the harmful effects of the lemniscate-link system and strengthen its advantageous features, a shield support with a new design of that mechanism has been developed. With the new design, the conventional rigid lemniscate links are replaced by hydraulic rams. When the pressure inside the rams reaches a certain level, the ram will yield.

Jan 12, 1986

By P. D. Chamberlain

Agglomeration for the heap leaching of gold and silver ores is the process of attaching fine ore particles-less than 150 µm (100 mesh)-to coarser particles. The fines are thus immobilized and uniformly distributed throughout the heap. This is needed to make heaps more porous and uniformly permeable to the flow of a leaching solution. This article describes liquids and binders used to form agglomerates. Agglomeration techniques and equipment are also described. These include drum and disk agglomerators, stockpile and belt agglomeration, and the use of vibrating deck and reverse-running conveyors. The need for agglomeration The main reason for agglomerating heap leachable ores is to prevent percolation problems in a heap. This is different than pelletizing iron ore. In the latter case, the purpose is to bond very finely ground material into coarser, strong particles. These particles can then withstand the rigors of transportation and the loading forces in a blast furnace. Most percolation problems in heaps are caused by the segregation of coarse and fine particles during heap construction. This segregation creates areas with significantly lower permeability because fines plug the channels between coarse particles. Consequently, the leach solutions are subject to channeling, following the path of least resistance. They percolate downward through the coarse ore areas and bypass or barely wet areas that contain large amounts of fines. This results in lower extraction, longer leach time, and higher reagent consumption. Permeability problems in heap leaching are compounded because of the vertical migration of fine particles after the heap has been built. These fines eventually accumulate in pockets or layers and impede uniform solution flow. The fines move because they are not attached to a coarser particle. Moreover, the sluicing action of leach solution flowing over the surface of particles may break the bond between fine and coarse particles. Some fines are also created by the decomposition of ores after being wetted. Percolation problems are minimized if the fines can be attached to the coarser particles. Fines are thus uniformly distributed and rendered immobile within the heap. Heinen (1979) conducted comparative tests of nonagglomerated and agglomerated gold and silver ores. He showed that percolation rates through agglomerated material usually are improved by 10 to 100 times. This reduces leach time for economic recovery to less than one-third of that for nonagglomerated material, and total recovery may be increased. Cyanide consumption is also reduced in most cases because of the shortened leach time. Even run of mine (ROM) ore should be agglomerated. This can be achieved with moderate success by passing each truckload of ore under a water spray to bring total moisture of the ore up to about 8% to 10%. The ROM ore will mix when the ore is dumped, when it cascades down the slope of a heap, or when it is pushed up with a dozer. Overall, agglomeration allows low-cost heap leaching technology be applied to gold and silver ores and tailings that could not be economically processed by conventional methods, according to Phariss (1982). How agglomeration works Agglomeration means that the clays and fine particles contained in the ore adhere to the coarser particles. This creates a coating of fines around the coarse particles. When the ore has few fines and the percolation problems are not severe, the addition of a liquid may be all that is necessary to make the few fines adhere to coarser particles. However, when the quantity of fines is substantial, e.g., +10% -75 µm (-200 mesh), a binding agent may be needed. Use of solution only for agglomeration There are many forces that cause agglomeration and one of them is the surface tension of water. Water-saturated particles collide with each other or with

Jan 12, 1986

By H. M. Giegerich

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

Jan 12, 1986

By V. A. Evans, D. L. Hellier

Introduction Preparing a report on fluorspar economics in the US could have been somewhat complicated at one time. This was because the number of producers, both large and small, derived production from different locations through-out the western US and the southern Illinois and western Kentucky fluorspar districts. Companies, such as Allied Chemical and Williamson Mining, operated in the Jamestown district of Colorado. In Walden and Salida, CO, the Colorado Fluorspar Co. operated mines and mills. In New Mexico, both Du Pont and Allied had mines and mills. In the Illinois-Kentucky district, US Steel, Rosiclare Lead and Fluorspar, Minerva Oil Co., and Marathon Oil all produced spar for the US market. Now all of the companies are out of the fluorspar business and, with the closing of the Seaforth Mineral Co.'s Inverness Mining Co., only Ozark-Mahoning Co. survives as a domestic producer. History Ozark-Mahoning Co., as it now exists, was formed in 1946 by the merger of Ozark Chemical Co. and Mahoning Mining Co. Ozark Chemical Co. was started in the 1920s to produce sulfuric acid for two refineries located in Tulsa, OK. Mahoning Mining was originally formed by the founders of Ozark Chemical Co. and the Youngstown Sheet and Tube Co. to produce fluorspar and associated minerals from properties in the southern Illinois and western Kentucky fluorspar district. In 1937, Mahoning Mining started an exploration program, employing contractor-operated churn drills to look for reserves in Hardin and Pope Counties in southern Illinois. Based on known reserves in the area and the results of the exploration program, work was started in the fall of 1938 to build a selective flotation plant at Rosiclare, IL. The mill was completed and put into operation in August of 1939. The company's first mining efforts were in the Cave-In-Rock district just north of Cave-In-Rock, IL. The deposits in this area were of the Manto or bedded type. They have been the mainstay of the company's operations in the Illinois-Kentucky fluorspar district. In 1950-51, the company expanded its operations by purchasing fluorspar properties from Colorado Fluorspar Co. at Walden, CO. This location was operated from 1952 through 1958 and from 1968 through 1973 when it was closed permanently. In 1974, Ozark-Mahoning Co. was purchased by Pennwalt Corp. Presently, it is a wholly-owned subsidiary of that corporation. As the only major domestic producer of fluorspar in the US today, the company's ability to survive under the flood of imported material is directly related to the four major segments of its operations. These are: exploration, mining productivity, maintenance, and marketing. All are done in-house. Exploration The Southern Illinois fluorspar district is located in the most complexly faulted area of the central US. The formations are sedimentary in nature and range in age from Devonian through Pennsylvanian. The deposits are small, they range in size from a few thousand tons up to 900 kt (1 million st), with the average minable deposit being around 544 kt (600,000 st). The ore occurs in both nearly vertical vein deposits and in Manto or flat lying bedded deposits. Due to the relatively small size of the deposits and the depth at which they occur, a good low-cost exploration program is essential to ensure that new finds of ore are equal to the mining extraction rate. Ozark maintains a leased land position in Pope and Hardin Counties that will fluctuate somewhere between a low of 40 km2 (10,000 acres) up to a high of 129 km2 (32,000 acres), depending on total size of tracts leased. Most all lands are obtained through a standard lease purchase option, with a small minimum royalty per acre during the normal 10-year option period. All leases are standard and without deviation to control land cost. All land leasing is done through Ozark's geological staff, who also

Jan 11, 1986

By Lawrence Pelham

Introduction Fluorspar demand or consumption has been taken for granted in recent years, except for those that produce, market, or consume the commodity. Thinking used to be that as the steel and aluminum industry goes, so goes fluorspar. This is much less true today. The industry is becoming more dependent on chemical industries that feed primarily off hydrofluoric acid (HF) production. The US Bureau of Mines (USBM) published an assessment of world fluorspar reserves and resources. It created a foundation of valuable data that can be used to monitor fluorspar availability. Celebrations were held this year commemorating the 100th anniversary of the isolation of fluorine. The US and other countries continue to struggle with the controversy of ozone depletion and the effect of fluorocarbons on the atmosphere. Goals for the government's strategic stockpile are being reevaluated. The market for fluorine chemicals is changing and diversifying, making it a challenge to assess and forecast fluorspar/fluorine demand. US production and distribution Fluorspar Fluorspar shipments from US mines reached an all-time high near the end of World War II in 1944 at 375 kt (413,800 st). Eighty-two mines produced more than 453 t (500 st) each, and 10 mines produced more than 9 kt (10,000 st.) each. Illinois supplied 43% of total shipments. Consumption in 1944 was 372 kt (410,200 st), of which 56% was used by steel producers and 32% by HF producers. This was the last year that US production exceeded US consumption. The post-World War II record for fluorspar shipments occurred in 1951 at 315 kt (347,000 st) 58 mines produced more than 453 t (500 st) and eight mines produced more than 9 kt (10,000 st).

Jan 11, 1986

By S. G. Malghan

W.H. Moyer In reference to the section “Excess pyrite, ” the author discusses difficulties in separating chalcopyrite from pyrite. A technique does exist, and has been used commercially, for the flotation, separation, and recovery, if desired, of both chalcopyrite and pyrite. The method was described in Mining Engineering, March 1966, “An Improved Flotation Method for Cornwall Sulfides. ” While it was not used on a porphyry copper ore, it was used on a chalcopyrite - pyrite-containing ore and might well have application. Both sulfides, chalcopyrite and pyrite, were floated simultaneously using xanthate as a collector and MIBC frother to produce a cleaned sulfide product. This concentrate was then thickened and conditioned at high solids with aeration and lime to a pH above 12.0 for a period of 1.5 to 2 hr. A supplemental suppressant, such as cyanide, could be used if necessary to sharpen the subsequent separation; however, it was not required at Cornwall. The conditioned sulfides were floated with two stages of cleaning to produce a highgrade chalcopyrite froth product and a pyrite sink product barren in copper. The separation at Cornwall gave a 28.6q. Cu grade chalcopyrite product and a pyrite analyzing 0.0670 Cu. A desliming step on the raw ore initially used at Cornwall for a period of time was later discontinued. Advantages to be found in using a strong sulfide collector, such as a xanthate, is the rapid flotation it induces and, generally, an ability to float at a coarser grind. A following regrind circuit could be used to give maximum liberation and produce a clean sulfide concentration. The “differential” separation is concerned primarily with grade and with assuring a barren pyrite for maximum recovery.

Jan 11, 1986

By J. B. Hiskey

Introduction The domestic copper industry has suffered for some time from declining ore grades; rising costs associated with conventional mining, milling, smelting and refining; tightening environmental restrictions; and shrinking markets due to materials substitution. Ore grade depletion is a major problem. It is illustrated by the decline in average copper yield shown in Fig. 1. From 1979, copper grade has dropped 60%, lowering yields from 1.2% to 0.49% (Sousa, 1981). The US copper industry is taking drastic measures to enhance its competitive posture and to guarantee its future worldwide position. In a sense, this involves a general restructuring of the processing philosophy of US copper producers. Judging from recent trends, this restructuring relates to the application of solution mining techniques to the treatment of copper ores and low grade waste material. Recovering copper by leaching techniques has been practiced for centuries. As early as the mid-16th century, some Hungarian copper mines were recycling copper bearing leach solutions through waste heaps (Nash, 1912). In the US, recovering copper from dilute mine waters has been carried out for more than 100 years. There are, however, certain technical innovations that have elevated the importance of solution mining as a process for copper recovery. These include: • Advances in solvent extract-electrowinning (SX-EW); • Developments in acid and acid ferric cure processes for oxide and mixed oxide sulfide ores; • Improvements in heap and dump construction; and • Advances in in situ leaching technology. Generally, solution mining enjoys certain intrinsic advantages over conventional mining and milling. Compared to conventional milling, the combined capital and operating costs of leaching facilities are normally lower, start-up times are faster, and the leaching operation usually has less impact on the environment. Furthermore, solution mining represents an expedient way of extracting metals from small shallow deposits, and is particularly suited to the treatment of low grade sources. And, in the case of waste dump leaching, there is a tremendous resource in-place that has the mining cost already off the books. This paper updates current US copper solution mining trends. Solution mining systems Schlitt (1980) summarized the characteristics of the principle leach systems in the US. He pointed out that the wide variations in copper deposits make solution mining a more site-specific activity than conventional mining and milling. Heap leaching is usually selected for moderate to high grade ores having a copper mineralization that is amenable to acid leaching (oxides, silicates and certain sulfides). Higher grade ores mined specifically for heap leaching are usually treated to optimize copper extraction, therefore, crushing and some pretreatment (acid curing) may be justified. On the other hand, dump leaching practice usually involves leaching low grade waste rock placed on large dumps as dictated by topography and haulage costs. Mine waste is therefore treated in an uncrushed run-of-mine condition. Finally, in situ leaching is appropriate for either very deep ore bodies or for recovering copper from low grade rock left behind in abandoned pit walls, stopes, and subsidence zones from earlier mining activities. True in-situ solution mining should be defined as the leaching of ore in its original geologic setting. Heap leaching Heap leaching has a long history as a hydrometallurgical process for copper recovery. The methodology of heap leaching established at Rio Tinto in Spain more than 300 years ago is basically the same as that used today. Processing concepts, such as solution management (sprinklers and leach/rest cycles) and copper

Jan 11, 1986

By Syd S. Peng

Introduction The most commonly used ground control design in coal mining is the determination of pillar size by various pillar strength formulae for mine layout (Bieniawski, 1983; Holland, 1973; Hustrulid, 1976; Obert and Duvall 1967; Wilson and Ashwin, 1972). However, there have been few documented case histories concerning the validity of those strength formulae. Recently there are indications that those strength formulae tend to overdesign considerably. From a safety point of view, there is nothing wrong with overdesign. But an overdesign with large chain pillars reduces the flexibility of machine movement and decreases production. It also reduces coal recovery and, thus, shortens the reserve's life. This is especially true for small operators who own a limited reserve. Those small mines, through past practices or out of necessity, experiment with smaller pillars than those predicted by the strength formulae. Conclusive ev¬idence has been collected that points to the need of a new concept of coal pillar design. Two case studies will be cited to illustrate this point. 3-D US pillar strength formulae Figure 1 shows a sectional layout of a one-unit mine. The seam was about 1.3 m (50 in.) thick with a uniform cover of 64 m (210 ft). A house was located near the center of the section. The homeowner demanded that no surface subsidence and domestic well dewatering should ever be allowed as a result of underground room-and-pillar mining. Physical properties - strength and Young's modulus - of the rock strata including coal were obtained by testing small samples in the laboratory. The uniaxial compressive strength of the coal was 3660 kPa (823 psi). Using various pillar strength formulae, the homeowner recommended that square chain pillars be at least 11 m (35 ft) wide. The coal company, however, desired to pursue optimum coal recovery in the safest and most economical manner. They engaged in three-dimensional finite element analysis of the section taking into account the complete stratigraphic sequence of the overburden. Results indicated that the section can be developed by 5.5 m (18 ft) wide entries at 21 m (70 ft) center-to-center. In retreat, the pillars can be splitted along both centerlines leaving a stump pillar of 5.5 m (18 ft) square at each of the four corners. The 5.5 m (18 ft) square stump pillars still maintained a safety factor of 1.5, although the safety factors calculated by various pillar strength

Jan 11, 1986

By A. G. Sciulli, G. P. Ballock, K. K. Wu

Waste disposal underground is the most effective method when viewed from an environmental aspect. As costs of other methods of disposal continue to escalate, the cost effectiveness of underground disposal during the mining cycle becomes increasingly attractive. The use of concrete for sealing abandoned mines has proved to be suspect when associated with a relatively incompetent adjacent strata due to its shrinkage characteristics. Complete failure of the plug has been known to occur in some instances. A more effective, cost competitive seal can be obtained using pneumatically stowed material and incorporating a polyurethane grout, which has greater adhesion characteristics than concrete. Injection of the surrounding strata and of the stowed material helps to ensure an impervious barrier and prevent the plug from moving when subject to high pressure heads. When properly placed, impervious seals cause the old workings to flood, raise the mine pool level, and thus eventually eliminate the potential for oxidation, which is the primary cause of acidity in the water.

Jan 11, 1986

By W. M. Luff

Introduction The Brunswick No. 12 mine is located about 27 km (17 miles) southwest of Bathurst, New Brunswick. The deposit is a stratiform massive sulfide deposit occurring in a series of sedimentary and volcanic rocks of the Tetagouche Group. These rocks are generally considered to be Ordovician in age. Ore reserves at Brunswick No. 12 are 104 Mt (115 million st) grading 3.72% lead (Pb), 9.01% zinc (Zn), 0.32% copper (Cu), and 97.9 g/t (2.8 oz per st) silver (Ag). Production in 1984 was 3.5 Mt (4 million st) at 3.57% Pb, 8.91% Zn, 0.32% Cu, and 100 g/t (2.9 oz per st) Ag. Total payable silver production that year was 211 t (6.8 million oz). This made Brunswick one of the largest silver producers in Canada. Bathurst-Newcastle area and Brunswick Mining Interest in the Bathurst area was aroused in 1897 when W. Hussey discovered an iron formation at Austin Brook, just south of Brunswick No. 6 deposit. This area, and the current No. 6 area, was explored from 1897 to 1907. Iron ore was mined at Austin Brook from 1907 to 1913 and from 1942 to 1943. Any occurrence of sulfides was overlooked until 1952. In 1952, M. J. Boylen Enterprises investigated the small sulfide body associated with the Austin Brook iron formation. Other electromagnetic anomalies associated with the iron formation were investigated. The No. 6 deposit was found late in 1952, when hole B12 intersected 97.5 m (320 ft) of massive sulfides. The No. 12 deposit, 10 km (6 miles) to the north, was found in 1953. More than 30 massive sulfide deposits have since been identified in the Bathrust area. In the 1950s, St. Joe Lead Co. did extensive test work on the No. 6 and No. 12 deposits, including shafts at No. 12, lateral development at No. 12, and testing at both deposits.

Jan 10, 1986

By P. D. Chamberlain, R. M. Nadkarni, D. J. Kinneberg

Introduction The relationship between precious metal miners and gold refiners has been uneasy over the years. Reasons for this uneasiness and lack of trust include technical factors, economics of scale, and cultural differences. Technical factors take two forms. First, concentrates, precipitates, and sponge produced at the mine site are notoriously difficult to sample. A nonrepresentative sample can be the cause of misunderstanding and distrust. Dote, however, is relatively homogeneous and can be correctly sampled and analyzed with greater reliability. Second, the technology and chemistry of fine gold and silver production is complex and quite different from dote production. Most mining companies are content not to be involved with additional processing steps. From the standpoint of process economies, it is easy to see why miners and refiners are separate groups. Mine location is dictated by the natural occurrence of ore bodies. Heavy capital investment requirements for infrastructure in remote locations usually limits companies to produce the first marketable product from a mine. Beyond that point, the economies of scale and large capital investments for refineries generally dictate that the outputs of several mines be combined to feed a single refinery. The refin¬ing process is so complex that refineries are generally built in larger cities with a skilled labor pool. Therefore, only the largest mines can afford to refine their own output. Lastly, there are cultural differences between the miner and the refiner. The miner, particularly the small gold miner, is by definition a risk taker in a risky business. Refineries, on the other hand, have historically been operated under legal or implied monopolies approved by the government. The government is interested in owning gold, either for currency reasons or because many national mints are major users of gold coins. It is not surprising, then, to find that miners may regard refiners as part of an unresponsive government bureaucracy or, worse, as "city slickers." It is easy for the two groups to distrust each other with these different backgrounds. The interface between miner and refiner does exist and affects nearly all precious metal mining companies and most refineries. This paper discusses the technical and economic aspects of the interface. It points out how both parties can improve the relationship. Technical interface Precious metals selectively dissolve to produce a pregnant solution. They are then recovered from solution by zinc precipitation (Merrill-Crowe process) or by absorption onto activated carbon. This is followed by stripping and electrowinning. At this point, the mining company must decide whether to ship precipitates (or sponge) to refiners or to produce a dote. Some factors the company must consider are the ability to sample precipitates, concentrates or sponge, security of storage, and production rates. An important factor is the incremental cost (capital and operating) of converting precipitates or sponge into dote at the mine. Generally, mines prefer to melt precipitates or sponge to dote to avoid prob¬lems associated with sampling them. There are three types of fur¬naces used by mining operators to melt precipitates to dote. The first and most widely used is a sil¬icon carbide crucible heated ex¬ternally with a gaseous liquid fuel. The crucible is easy to use and slag can be skimmed before pouring dote. This simple system is capable of handling 1 t/a (32,000 oz per year) gold. It could cost as little as $20,000. More complex systems could cost up to $200,000. The second type of furnace is the gas- or propane-fired rever¬beratory furnace. This is usually used only by large-scale opera¬tors in the silver business. Precip¬itates in reverberatory furnaces come in direct contact with the flame. The large volumes of flue gas can carry significant amounts of dust and fume that must be re¬moved from the gas before dis¬charging it to the atmosphere. In the days when the Rand re¬finery in South Africa used reverberatory furnaces, the silver leaving the furnace in the form of vapor and fume was estimated at 2.2% of the total silver charged to the furnace. The gold in the gases was about 0.03% of the gold charged. Most of this silver and gold was recovered in the exhaust gas scrubber system. These scrub¬ber systems can be more costly than the furnace itself.

Jan 10, 1986