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By V. R. Degner

Comments Summary It is very interesting to review the technical changes which have occurred in the industrial practice of minerals flotation over the past two decades. In 1972, worldwide industrial flotation was dominated by 'mechanical' flotation machines having an individual cell volume less than 500 cu.ft. Manufactures, such as DENVER Eqpt. Co., OUTOKUMPU, WEMCO, GALIGHER (AGITAIR), AKKER, and later DORR-OLIVER, produced flotation machines featuring a mechanical rotating impeller which served to mix pulp and air, prior to flotation, and to keep the solid ore particles in suspension. While sharing this common mechanical impeller feature, the machine design of the various manufactures differed in detail; particularly in regard to the vertical location of the impeller in the cell and the means by which flotation air was brought into contact with the pulp within the rotor mechanism. A review of the existing technical literature in 1972 would identify academic research studies, particularly at Columbia University (N. Arbiter, C, Harris ...) aimed at comparing flotation cell design concepts, of the various machine manufactures, on a hydrodynamic basis. These research studies were confined to very small, laboratory bench size, flotation machines, and the benefits of incorporating hydrodynamic considerations in the scale-up of the existing 300 cu.ft. mechanical flotation technology to larger cells was not commonly practiced industrially. Technological advances in minerals flotation, over the past two decades, have been influenced by a variety of economic factors including the increasing development of lower grade ore bodies, the attractions of producing a higher grade final concentrate, and the economy of tails retreatment to improve overall plant recovery performance, and flotation system “optimization" became increasingly important through the 1970-80 time period. The trend toward processing lower grade ores led to higher tonnage mill operations, and stimulated the development of the very large capacity mechanical flotation machine for roughing and scavenging duty. In 1972, the 500 cu.ft. size flotation machine was being pilot tested and introduced to the industrial market. Today, 1993, individual cell volumes of 1500-3000 cu.ft. are operating successfully throughout the world, and well planned plant test programs have concluded that the very large flotation machine can achieve the same metallurgical performance level of its smaller predecessor. The large flotation machine development record clearly credits the incorporation of internal cell hydrodynamic considerations in guiding the scale-up of the mechanical flotation machine to large sizes. This guidance eliminated, or reduced significantly, costly in-plant prototype machine changes which increase Dramatically with increasing flotation machine size. Early in the 1980's, at approximately the time that flotation machines in the 1500 cu.ft. class were being introduced into the market, it became evident that the availability of the very large mechanical flotation machine brought with it the need for a systematic means to "optimize" the flotation system by relating flotation cell size, row length (i.e. number of cells per row), and flotation rougher, scavenger, and cleaner system configuration to the economic objectives of the concentrators' operating management. Flotation system optimization consists of quantifying the flotation process performance using a suitable model which relates the kinetic response of the trace specie and the gangue material to the cell overflow, and produces a numerical relationship between product grade and value recovery. The computer analysis technique has proven to be an invaluable tool in determining the kinetic model coefficients which characterize a given application, and is also used to apply the resulting model to a proposed minerals beneficiation situation seeking to "optimize" the flotation system by relating metallurgical grade-recovery performance, flotation cell size, and circuit arrangement, to plant space, installation, and operating costs. It is interesting to note that these same computer based kinetic flotation analysis techniques can be easily applied to quantitatively evaluate the effect of different chemical reagent strategies, and flotation air transfer, on process performance for a variety of candidate feed ore types.

Jan 1, 1993

By Patzen Wu, Chao-Ming Tsai, Shaw-Chii Wu, Yuan Lin

INTRODUCTION The development of nuclear industry in Taiwan requires an increasing supply of uranium. Uranium recovery from phosphate ores, becomes an important item of the R & D programs. Many countries have already started such a program. USA has set up an industrial-scale production plant; South Africa, Israel, Morocco and other countries have also engaged in the development of extraction technology of uranium from phosphate ores and started setting up the production plant. In Taiwan, a feasibility study of establishing the production plant was started in 1977. A pilot plant was completed, tested, and operated in 1979. The main source material. - phosphate ores is obtained from China Phosphate Company. The annual yield of the pilot plant has been planned to be one ton of ammonium diuranate (ADU). The Health Physics Division of Institute of Nuclear Energy Research (INER), Atomic Energy Council of Republic of China (ROC AEC), is responsible for the radiation safety control of the plant. The purpose of the radiation safety control is to collect radiation safety information of the plant of this type, to evaluate possible radiation hazards to the personnel, and to analyze the radiation safety of the immediate environment adjacent to the plant. Furthermore, the information and experience obtained are used as the basis for applying radiation safety control to the production plant. The main emphases of the radiation safety control are therefore the arrangement of radiation safety control areas and the collection and evaluation of the health physics data. This report includes the past one and half years' experiences of the radiation safety control of the pilot plant, the planning of the production plant with an annual yield of 10 ton ADU, and considerations given in improving the radiation safety control of the production plant. PILOT PLANT AND ITS OPERATION PROCESS The pilot plant occupies an area of 350 m2 and is a two-story rectangular building. Exits are provided by a front and a rear door. An additional small door at the central right side of the building is connected to the maintenance area located outside the main building. The interior of the building is divided into interconnected offices, laboratories and process area (Figs. 1 and 2). A stairway to the process area of the second floor is located in the central part of the first floor. The second floor offices are accessible by the additional front and rear stairways. The main process area of the first floor is furnished with plastic floor, and the second floor with poriferous iron-grid floor. The building is not furnished with centralized air-conditioning system. Individual air conditioners are, however, available to some laboratories. Because the back door of the building is frequently open, the product storage room is provided with a blower to maintain good ventilation. Natural uranium contents of phosphoric acid solution vary in the range from 0.07 to 0.11 kg/m3. The process consists mainly of the uranium concentration cycle and the purification cycle. A flow diagram of the process is shown in Fig. 3. CONSIDERATIONS GIVEN FOR RADIATION SAFETY CONTROL AND THE ESTABLISHMENT OF RADIATION CONTROLLED AREAS The radiation safety control of the pilot plant is carried out on the basis of the considerations listed below:1,2 1) Since the wet process is adopted, and the flow of the radioactive material is confined to closed pipes and containers, air contamination is therefore not too much of concern. 2) The concentrations of radioactive material of the whole process are mostly less than 10 ppm. An exception is the final ADU product after precipitation and centrifugal separation, which contains up to 50% natural uranium. It is therefore necessary to apply radiation safety control to the process area where

Jan 1, 1981

Microorganisms have been widely used for hydrometallurgical operation and effluent treatments(Smith, et. al., 1991a; Gadd, et. al., 1986; Darnall, 1986; Paponettis, et. al., 1991; Mihaylov, et. al., 1991; Rusin, et. al., 1991; Noble, et. al., 1991; Yang, et. al., 1991). In recent years there is a growing interest in using microorganism for flocculation of minerals fines and flotation collector (Smith, et. al., 1991b, 1991c; Misra, et. al., 1991a, 1991b, 1992a, 1992b). It is envisioned that the utilization of nonpathogenic microorganism and/or extracellular surfactant derived from microorganism as a flotation collector will replace some of the environmentally toxic flotation collectors and at the sometime improve flotation efficiency. The selection screening, and judicious selection of appropriate microorganism as flotation collector has been a challenge to process engineers. Over the past five years, continued research at the University of Nevada has shown that certain microorganism Mycobacterium phlei (M. phlei) can be a good flocculating agent and has potential to be used as a flotation collector (Dubel, et. al., 1992; Misra, et. al., 1991a, 1991b, 1992a, 1992b; Smith, et. al., 1991d; Chen, et. al., 1992a, 1992b)).

Jan 1, 1993

By William L. Jude

INTRODUCTION Although the mining method described in this chapter is locally designated "modified Mitchell," there are so many features which are characteristic of square sets that the designation "modified square set" would describe the method equally well. The method was developed as a result of local operating experience with various mining methods, observing operations in mines in the southwest where Mitchell stoping had been thoroughly developed, and an exhaustive study of min¬ing methods literature. Since 1941, modified Mitchell stopes have con¬tributed over 50% of the total stope production and were largely responsible for supporting production dur¬ing the war years and period of declining manpower and increasing costs. GEOLOGY The community of Gilman is situated on Battle Mountain at an elevation of 2743 m (9000 ft) in the Battle Mountain mining district, about 45 km (28 miles) by highway north of Leadville, CO. The sur¬rounding area, including the Mount of the Holy Cross and Eagle River Canyon, is noted for its beauty and recreational facilities, both summer and winter. Massive sulfide ore bodies of the limestone replace¬ment type occur as chimneys and mantas or as blankets in strata dipping 0.19 rad (11°) northeasterly. A chimney is shaped like an inverted cone with shale, quartzite, and porphyry caprocks forming its base. This relatively impervious capping apparently blocked rising ore constituents which formed the chimneys and mantas. At the apex or lower extremity of the chimneys, the ore follows small fractures and bedding planes in under¬lying quartzite. Dimensions of chimneys vary, ranging up to 61 m (200 ft) vertically and 76 m (250 ft) horizontally near capping. The manta or blanket ore bodies vary in vertical cross section from one hundred to several thousand feet. One such manta was developed from its outcrop to its chimney source approximately 610 m (2000 ft) downdip. Another manta of substantial cross section and length did not follow shale but extended updip with limestone capping from its source at a lower horizon. All chimneys, however, make up to the shale¬quartzite-porphyry capping. Chimney ore minerals are mainly pyrite and chalco¬pyrite with accessory copper and silver minerals, surrounded by, and in places overlain with, marmatite, pyrite, and galena. Manganosiderite extends into the ore. It may or may not be impregnated with sufficient marmatite and galena to be of commercial grade. Ore may be massive, coarsely crystalline, or sandy, some¬times rubbly. Generally speaking, chimney ore masses are incompetent and of variable grade. Caprock is heavy and caves readily. Manta ore minerals are mostly marmatite, pyrite, and galena associated with manganosiderite. Structural characteristics of the ore are similar to those found in chimneys and where the manta contact shale capping the ground is unusually heavy. However, the large manta with limestone capping mentioned previously consisted of a hard strong ore mass in fairly strong host rock. Commercial ore contacts are fairly regular in this manta but in others the ore waste contacts along the walls are irregular with shaly limestone pendants extend¬ing downward into the ore at capping. Dolomite sand is also quite frequently encountered along wall contacts and less frequently overhead. MINING METHODS Practically all stoping in the Gilman ore bodies has been carried on with timber and fill methods and the features which have been briefly described have dictated the choice. Although at the present time modified Mitchell and underhand square-set stopes account for the total production, considerable production has been obtained from alternate horizontal cut-and-fill stopes with Mitchell truss stopes in the pillars. This method was highly satisfactory in the hard manta ore body mentioned earlier and where ore excavations 11 m (35 ft) wide and 23 m (75 ft) long required no auxiliary timber support. It was a pillar recovery application of the method devised by M. W. Mitchell at the Calumet

Jan 1, 1982

By D. A. Elgillani

Oxidation potentials have been measured in the presence of various concentrations of cyanide, ferrocyanide, and ferricyanide and ethyl xanthate at various values of pH and related to flotation response. Eh-pH diagrams are constructed and show that the formation of surface ferric ferrocyanide is probably responsible for depression when cyanide is added. The influence of cyanide in the depression of pyrite with xanthates as collector has been the subject of many investigations (1,2,3,4,5,6), and from these studies several theories on the mechanism of depression have evolved. Wark and Cox (7) and Gaudin(8) have suggested that the depressing effect is due to a competition of cyanide ion with xanthate ion for the surface. Cook and his colleagues (9,10,11) have explained this phenomenon in terms of competition between hydrocyanic acid and xanthic acid. Sutherland (12) has shown that both of these theories describe accurately the relation between pH value and cyanide addition at constant collector addition but fail to describe the relationship, between pH value and the amount of collector required to cause flotation. Taggart(12) suggested that depression in these systems is due to the formation of a reaction product between ferric ion at the pyrite surface and ferrocyanide ion derived from solution. Majumdar(4,6) has attempted to prove this hypothesis by measuring the contact angles of pyrite in the presence of 25 mg/liter ethyl xanthate and different concentrations of potassium ferrocyanide and ferricyanide. In all cases the contact angles were quite high up to pH 10; these results indicate that pyrite should not be depressed by either potassium ferrocyanide or ferricyanide. In view of these facts, Majumdar has assumed that the compound Fe(CN)2 forms at the surface.

Jan 1, 1968

By D. A. Robbins

The application of remote sensing for environmental measurements is already being utilized. Aircraft mounted imaging spectroscopy has been utilized to evaluate areas of historical mining operations as well as active mine sites. To the extent that remote sensing can provide more sensitive, cost effective, and timely monitoring potential, the technology is of interest to the industry. Questions remain about the application and validation of the data and the availability of the technology in the private sector. This presentation will discuss some of the potential uses, benefits, and concerns related to remote sensing and efforts to monitor the technology's development.

Jan 1, 1999

By S. Bhamidipati

Abstract. Controlled flow is an aspect often desired in mine ventilation. This is normally obtained by using fans and regulators to achieve the required pressure balance in the system. Preference for a certain network layout of these control devices is influenced by technical as well as legal and economic factors. In this study, a model is developed based on an economic criterion by which an optimum layout of fans and regulators for a controlled flow system can be determined. The concept is analo¬gous to the critical path crashing. However, a linear programming package is used to demonstrate the solution procedure for a hypothetical venti¬lation network.

Jan 1, 1983

By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores, such as cobalt, platinum, gold, silver, etc. Flotation of metal sulfide ores has been successfully practiced at the industrial scale for more than 50 years. It is the relative technical/economic ease of floating metal sulfide minerals that has been a major driving force behind the rapid expansion in the supply/demand of metals over this 50 year period. It is estimated that 1.6 billion tons of feed sulfide ores are processed each year in the free world based on the froth flotation process, Klimpel (1987) and the manufacture/use of associated chemical reagents is a major portion of the 1.5 billion dollars (U.S.) free world mining chemicals business, Klimpel (1988b). The use of froth flotation appears to be increasing partially due to the ever decreasing feed grades and liberation sizes of the feed minerals. Froth flotation has shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a variety of equipment configurations. Such operations exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the widely used mechanical flotation cell is scalable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required -- pH regulators, frothers, collectors, activators, and depressants -- are all relatively inexpensive and common. They are not usually used in large quantities. While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very ample. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic sources having suitable-grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 987 were known and manufactured in some form 25 years ago. This is clearly not typical of reagent development in other process-related industries. In addition, the industrial-scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully on an industrial scale for many years. Thus the industrial process of froth flotation, especially as applied to the recovery of sulfide minerals, poses a dilemma: such a process is widely used, economic, versatile, forgiving, etc. and yet is still not very well understood in a mechanistic sense leading to prediction of results from fundamentals even after 50 years of usage. Lack of such predictive ability has not limited the general over-all industrial use of froth flotation in sulfide mineral recovery. Rather the lack of

Jan 1, 1989

By Richard E. Hucko

The Coal Preparation Division of the Pittsburgh Mining Technology Center conducted an investigation to assess and compare the pyritic sulfur reduction potential of coal-pyrite flotation and high-gradient magnetic separation (HGMS). Tests were run on three coals--an Upper Freeport bed coal, a Lower Kittanning bed coal, and a Pittsburgh bed coal--to establish grade- recovery relationships characteristic of both processes. Representative portions of each coal sample were crushed to two top sizes--nominal minus 35 mesh (0.5 mm) and nominal minus 100 mesh (0.149 mm)--and subjected to conventional froth flotation. The froth product was then beneficiated by coal-pyrite flotation or HCMS to provide a direct technical comparison of the two processes. At minus 35 mesh (0.5 mm), the size normally employed for coal flotation, both coal-pyrite flotation and HGMS effectively separated pyrite from coal, and the performance of the two processes was essentially the same. At a size of minus 100 mesh (0.149 mm) , however, the performance of HGMS exceeded that of coal- pyrite flotation. The separations effected by HGMS were more efficient at minus 100 mesh (0.149 mm) than at minus 35 mesh (0.5 mm). On the other hand, it was concluded that either the effectiveness of coal-pyrite flotation decreases with diminishing top size, or the process has not been optimized to treat material finer than 35 mesh (0.5 mm).

Jan 1, 1981

By D. W. Alberter

I consider it a great pleasure and distinction to be invited to appear before your organization to discuss our experiences with a comparatively new system of what is known as a "Full Dimension" extensible conveyor system. One of these systems has been operating in our Diamond Smokeless Mine, Boltz, Pennsylvania, since 1959 servicing a continuous mining machines For those of you who are not familiar with this method of transportation, I would like to offer an explanation.

Jan 1, 1962

By Andrew Cho, Changsoo (Kevin) Moon, Richard Flanagan, Brian Larsen

"INTRODUCTION Project Background The U.S. Army Corps of Engineers (USACE)/Port Authority of New York & New Jersey (PANYNJ), as part of their New York and New Jersey Harbor Deepening Project, will dredge this shipping channel to a depth of 16.5 m below mean low water (MLW). By deepening the channel, the new generation of larger, “post-panamax” container ships coming to U.S. waters as a result of the expansion of Panama Canal can be accommodated in local marine terminals, helping to retain the nearly 300,000 jobs and $12 billion in annual wages the PANYNJ provides to the region. The existing siphons are situated at a depth that could be compromised by the planned dredging of the channel (nycedc.com). The purpose of this project is to build a new 1.83 m diameter water main inside a 3.66m diameter tunnel replacing the two existing siphons in order to provide the necessary backup water supply to Staten Island across the Anchorage Channel within the Upper New York Bay from Brooklyn. The proposed tunnel is approximately 2.9 km long and at a depth of at least 26 m below MLW. The project is funded jointly by New York City’s Department of Environmental Protection (NYCDEP) and PANYNJ and is managed by New York City Economic Development Corporation (NYCEDC). The contractor of the project is Tully/OHL JV, the designer is CDM/HMM JV, and the CM is LiRo/PB JV. History of New York City’s Soft Ground Tunneling The New York City area has a relatively long history of soft ground tunneling, commencing in the late 19th century. Construction of the first Hudson River tunnel crossings was begun (but not completed) in 1874 for the Hudson & Manhattan Railroad using compressed air supplying a shield. The work continued intermittently until the project finally opened for service in 1907. These 1.72km long tunnels (maximum depths of 29 m) are still in operation and now operated as the PATH Tunnels. They are approximately 5.5m diameter utilizing cast iron segments. The river bed consisted of primarily silt with shallow rock near the New York side. Air “leaks” occurred during the compressed air excavations resulting in some serious “blows” with water and soil entering the tunnel, including one incident with 20 deaths."

Jan 1, 2016

By Robert J. Pruett

The Georgia kaolin District is the world's largest producer of kaolin. Georgia production tripled from 2,344 million metric tons in 1964 to 7,570 million metric tons in 1994. Development of the Georgia kaolin industry can be traced to the mineral resource, the adoption of innovative process technologies, and the market growth for kaolin. Georgia kaolin is an abundant and high-quality resource that shows a range of advantageous properties. Technical developments in wet processing, delamination, and calcination allowed production of high-value products. The application of kaolin pigments to the paper industry provided a market that could absorb rapid growth.

Jan 1, 1998

By A. G. Oppegaard

Most iron ores treated by the R-N direct reduction process yield metallic iron products considerably lower in titanium, manganese, and chromium oxides and alkaline metal sulfates/sulfides than were contained in the original ores. These metallic iron products are suitable as feed stock for the open hearth and electric steel-making furnaces. The elements which are rejected in part are acceptable in the product since the product is designed to by-pass the blast furnace. R-N reduction conditions (PCO+H2-1 atmosphere, 1830° to 2150° F.) are adjusted for each ore. The iron ore which is reduced to 90 to 95 percent metallic iron is then subjected to the conventional beneficiation techniques. Under these conditions, 55 to 80 percent of titanium oxides, 75 to 95 percent of manganese oxides, 60 to 90 percent of chromium oxides and up to 97 percent of alkaline metal sulfates contained in the ore are rejected with the gangue to the non- magnetic tailings. The metallic iron products illustrated in this work contained 89 to 96 percent total iron, one to five percent TiO2, 0.2 to one percent MnO, 0.3 to 1.6 percent Cr2O3 and/or up to one percent alkaline metal sulfide. Usually only one of these components was present. Recoveries ranged from 80 to 96 percent of the iron in ore.

Jan 1, 1958

By A. C. Bigley

Before going into this subject, I would very much like to call to your attention some past history on patents granted by the United States Patent office that involved ammonium nitrate. On February 10, 1885, a patent was granted to Russell S. Penniman. "The object of this invention is to securely protect ammonium nitrate against deliquescent and to maintain it in substantially its normal disintegrated condition, so as to afford a new product suited for ready use in the manufacture of high explosives, and enable the immediate admixture thereof with any of the well known explosive substances or compounds heretofore employed. 'Whether the nitrate be thus granulated or otherwise reduced to a granular or divided condition, it is then thoroughly mixed with about twelve (12) percent of the fluid or semi-fluid carbonaceous protecting medium, which forms a readily applied jacket or coating for the granules of nitrate and also serves as a carbonaceous element in such explosive compounds as it may thereafter be employed in."

Jan 1, 1958

By S. R. Rao, J. A. Finch, E. El-Ammouri, K. Ngoviky, P. A. Disten

Sludges produced by lime treatment of acid mine drainage (AMD) can contain significant quantities of base metals which represent both a potential hazard and a loss of values. A new technique for metal recovery from sludge is proposed by combining selective acid leaching and adsorption of metal ions by activated silica sol. The process is demonstrated on two AMD sludges, one containing primarily nickel and the second zinc. They were first treated with sulphuric acid (pH 3.5) to selectively leach the target metals and then activated silica sol was added to the leach solution to adsorb the nickel and zinc. Thickening and centrifuging gave about 90 percent rejection of the input volume to effluent. The adsorbed metals were redissolved by strong acid into a concentrated solution relative to the feed and the silica sol was recycled.

Jan 1, 2000

Ion exchange (IX) has been in use for a long time, references to IX are made in The Bible, Exodus Chapter 15 vs. 22-25 ?and the waters were made sweet?.1, 2 Aristotle noted a reduction in the salt content of water when passed through certain types of sand, a phenomena probably associated with ion exchange characteristics of zeolites. Some major milestones in IX development include; ? ~ 1400 BC Moses ? ~ 330 BC Aristotle ? 1850 the discovery of the IX process3 ? 1909 Gans proposed zeolites for gold recovery ? 1935 sulfonated coals used for water softening ? 1939 copper recovery from Rayon Production ? 1940 The Rare Earth Project at Ames ? 1952 first South African U IX mine at West Rand (first SX of U using amines 1957) ? 1954 first U IX in USA at Shiprock Mine N.M. USA. ? Late 1950?s U IX over 50 fixed bed systems installed in North America ? 1950?s-1960?s FSU introduce U IX both fixed bed and RIP ? 1968 Gold IX (RIP) Muruntau Uzbekistan (CIL wasn?t commercialized until 1973) ? Mid 1970?s FSU & USA independently develop ISL combined with IX for U recovery ? 1978 Chinese Tungsten IX commercial Plant ? 1984 Uranium extraction from seawater ? 2004 Recovery of Gallium using hydroxyamic acid ligands ? Current ? Commercialization of IX for recovery of gold thiosulfate. Ion exchangers can be naturally occurring or synthetic; the majority of ion exchange materials in use today are synthetic resins based upon organic polymers such as polystyrene or methacrylate. Into more modern times the use of ion exchange for metal recovery was first proposed for the recovery of gold using zeolites in 1909.4 In 1939 copper was being recovered from the manufacturing of Rayon using ion exchange5 a development that continued applicability through to current times in the production/recovery of copper and nickel in ammoniacal leach processes. In early 1949 researchers at the Rohm & Haas Company discovered that hexavalent uranium existed as an ionic complex in sulfuric acid leach liquors, and that quaternary ammonium anion exchange resins exhibited a high selectivity for this ionic species6 The first commercial application of this technology went on-stream in October 1952 at the West Rand Consolidated Mines, Ltd operation in South Africa. 7 Ion Exchange has been applied for the recovery and purification of a range of metals;8 ? Uranium, Thorium and transuranic elements ? Rare Earths ? Gold, Silver and platinum group metals (PGM?s) ? Chromium, Copper, Zinc, Nickel, Cobalt ? Refractory Metals; Molybdate, Vanadate, Tungstate, Rhenium Mercury.

Feb 27, 2013

Fluorspar was not mined in the United States during 2004, although a small amount of usable synthetic fluorspar (CaF2) was produced from industrial waste streams. The majority of fluorspar consumed in the United States came from imports or stockpiled material purchased from the National Defense Stockpile (NDS). In addition, consumption is supplemented by use of by-product fluorosilicic acid produced at phosphoric acid plants. In 2004, annual production of fluorosilicic acid was 50kt (55,000 st) or about 88 kt (97,000 st) of 92 percent fluorspar equivalent. In 2004, annual production of fluorosilicic acid was 50kt (55,000 st) or about 88 kt (97,000 st) of 92 percent fluorspar equivalent.

Jan 1, 2005

By P. A. Plunkert

Metallurgical-grade bauxite was imported into the United States from 11 countries in 2004. The principal suppliers of this 10 Mt (11 million st) of imported ore were Jamaica (33 percent), Brazil (25 percent), Guinea(23 percent) and Guyana (11 percent). The average value of imports, cost, insurance and freight (c.i.f.) U.S. ports, was $31/t ($28/st). Calcined refractory-grade bauxite imports were 165kt (182,000 st) with an average value of $146/t ($132/st), c.i.f. U.S. ports. China supplied 70 percent; Guyana, 22

Jan 1, 2005

By Francisco A. Fonseca

The Carajás Hills are in Brazil's Eastern Amazon, between the Araguaia and Xingu rivers, and parallels 5° and 7° Latitude South. A location map is given in Fig. 1. It is an area of low mountains (maximum altitude 900 meters), covered by the Amazon rain forest. Annual average temperature is 25°C, ranging from 31°C (maximum) to 19°C (minimum). Annual rainfall averages 2,000mm. Although it rains throughout the whole year, heavy rains occur from December through March. At the beginning of the 20's, Brazilian explorationists made brief geological descriptions of the area. Due to the thick soil, scarce fresh rock outcrops and the tropical forest covering a completely unpopulated area, detailed geological studies were difficult and reconnaissance expeditions were limited to the river margins.

Jan 1, 1982

More than 1,200 professionals from the global tunneling and underground space industries are expected to be on hand for the 2009 Rapid Excavation and Tunneling Conference to be held at Caesars Palace in Las Vegas, NV. The conference, set for June 14-17, is sponsored by the Society for Mining, Metallurgy, & Exploration, Inc. (SME) and the Construction Institute of the American Society of Civil Engineers. The four-day conference will feature more than 100 technical papers in 19 sessions. Two one-day short courses are also planned, as are a pair of field trips. And the show?s exhibit will feature more than 120 of the industry?s leading equipment and service suppliers.

Jan 1, 2009