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By John W. Lane

Climax Molybdenum Company, a Division of American Metals Climax, Inc., operated a hydrometallurgical plant* at Climax, Colorado from August 1966 through August 1968 to recover molybdenum from a mixed oxide and sulfide ore. The ore treated for this addition to the existing plant contained sufficient Molybdenite values to be mined for Molybdenite recovery alone, so that the oxidized molybdenum recovery could be regarded as a byproduct operation. Climax is located at an elevation of 11,300 feet. All process equipment, except thickeners, was enclosed in buildings. Annual snowfall is approximately 250 inches per year. Paved highways and rail service to the plant are maintained and kept open all year. Most of the employees live in the town of Leadville, which is approximately twelve miles from the plant site. The sulfide flotation plant handles 40,000 to 45,000 tons per day of ore. The mine is one of the world's largest underground mines.

Jan 1, 1971

By David B. Spencer

This paper describes a process for separating non-ferrous metals such as aluminum, copper and zinc, from other residential and industrial waste by means of permanent magnets. Although the separator may be used in the recovery of non-ferrous metals from auto scrap, most of the testing to date has been aimed at recovery of non-ferrous metals from municipal solid waste, especially recovery of aluminum from the heavy fraction of a resource recovery plant. In most resource recovery plants the incoming solid waste material is first shredded and then air classified into a light and a heavy fraction. The light fraction which contains mostly paper, cardboard, etc., can be used either as a fuel for power generation or in the production of new paper products. The heavy fraction which remains consists mostly of glass, ceramics, wood, ferrous metals and non-ferrous metals. The ferrous metals may be removed from the heavy fraction by conventional magnetic separators. After primary shredding, most of the glass and ceramic is fine in size and may be screened away leaving a residue of wood, cardboard, and non-ferrous metals. This residual heavy fraction is the input material for the separator described in the present paper.

Jan 1, 1975

By E. Martinez

A heavy mineral sand sample was tested containing 13% heavy mineral consisting of rutile, zircon, and ilmenite. Eleven tests were run at a mill site with a Mineral Deposit Limited MG 4 spiral to determine the effect of the addition 0; magnets to the spiral compared to conventional gravity separation. Different magnet configurations and field strengths were tested. Optimum results were obtained with the spiral retrofitted with magnets. The heavy mineral content of the concentrate was increased from 55.5% in the control to 64.2% with the magnets. The 8.7% improvement in heavy mineral recovery was due not only to the weakly-magnetic ilmenite, but also to increased recoveries if the non-magnetic zircon and rutile. The increased recoveries were achieved with no loss of concentrate grade.

Jan 1, 1992

By Felicia F. Peng

One of the serious environmental problems related with the extraction of coal for carbon and energy resources is that upon preparation of coal a large quantity of fine coal has been discarded and/or lost to tailing impoundments. These problems present hazardous conditions to the local community. Also, the precious heating value of energy resources has been wasted and Left unrecovered in the tailing impoundments. Both RCR and SMRC ACTs recognize and mandate an investigation into slurry impoundment problems, including consideration of potential use of those coal fines in the tailing ponds. Bituminous coal producing district 46 encompasses the northern panhandle of West Virginia. The major producing seam in this district is the Pittsburgh coal, 1-7 meters thick. 75 percent of the coal producted in this district is deep mine. This district produces approximately 3 million tons of coal waste materials annually. One of the coal preparation plants in the district which we have visited has been processing Pittsburgh too. 8 seam coal for the past ten years. The characteristic data of the Pittsburgh raw coal is shown in Table 1. The raw coal produced in this district is characterized by its high heating value, high ash and high sulfur contents. The coal preparation plant flowsheet designed and used to process this Pittsburgh raw coal prior to 1979 is shown in Figure 1. It was a partial coal cleaning plant' with 1 092 mt/hr designed capacity, which consisted mainly of a coarse coal cleaning circuit. The rejected fine particles in the size range of -600µm which consisted of vorsiv underflow and cyclone underflow from clean coal streams and spiral classifier overflow from refuse stream were settled down in a thickener with the aid of anionic flocculant prior to being pumped to the tailing ponds for disposal. Figure 2 shows that the same coal preparation plant in Figure 1 was modified to a semi-complete coal cleaning plant and expanded its designed capacity to 1 820 mt/hr, when market of coal fines becomes feasible in 1979. Tabling in the clean coal circuit and thermal dryer in the dewatering circuit were added to the flow sheet. The purpose is to recover more coal fines, reduce more pyritic sulfur and ash contents in coal and also to reduce solid waste material to be stored into the slurry impoundment, giving them a Longer Life. The rejected fine particles fed to the thickener are clean coal fines from two cyclone overflows and the fine refuse from the rufuse cyclone overflow. The clean coal fines presently lost to the tailing impoundments are in the size range of -150 µm and constitute about 5 percent of the overall plant feed. Total amount of tailing waste material accumulated in a tailing impoundment is estimated at 1.3 million metricton for the past ten years.

Jan 1, 1982

By P. Zhang

Phosphate mining in Florida generates approximately 100,000 tons/day of phosphatic clays (slime). This waste slime not only creates one of the most difficult disposal problems in the mining industry, but also causes tremendous loss of P2O5. About 20-30% of the phosphate mined end up in phosphatic clays. Recovering the phosphate from this waste is difficult for two main reasons: 1) the slime is extremely fine in size, and 2) clay minerals account for almost half of the material. A half dozen dispersants were evaluated for keeping the clay minerals suspended for subsequent separation from non-clay minerals by either settling or cycloning. The separation efficiency and economics are discussed.

Jan 1, 1998

By P. Zhang

Previous studies have shown that separation of clay minerals from non-clay minerals is the key to economic recovery of the phosphate from the waste phosphatic clays (slime) which represent about 20- 300/0 of the phosphate mined. Small hydrocyclones of different sizes (4" and 6") were tested for sizing the slime at +10 microns in order to remove most of the clay minerals. The effect of design variables and operating parameters were evaluated. Multi-stage cycloning and the effect of dispersant were also investigated. Chemical and mineralogical analyses of different size fractions of both the underflow and overflow samples indicated satisfactorily separation efficiency.

Jan 1, 1998

By C. E. Jordan

The Tuscaloosa Metallurgy Research Center of the Bureau of Mines, U.S. Department of the Interior, is investigating the recovery of phosphate from beneficiation slime, as part of the Bureau's efforts to conserve mineral resources through increased use of wastes. Phosphate equivalent to about two-fifths of U.S. production is discarded annually in the wastes. Exploratory flotation tests on the plus 5-um fraction of a slime sample showed promise for a classification-flotation technique. Onsite classification tests at six phosphate washing plants were conducted with a two-stage hydro cyclone system designed to recover the plus 5-um material. The cyclone underflows contained 8.3 to 19.1 percent P2O5, and the recovery of the plus 5-um PZOS values ranged from 81 to 95 percent. The percent P2O5 recovery from the total slimes ranged from 38 to 47 percent. Batch-flotation studies are currently being conducted.

Jan 1, 1978

By S. A. Beeson, C. Poole

This paper describes the development of a novel selective flocculation technique for recovering phosphate values from ores containing silicate or carbonate gangue minerals. Data from initial flocculation response and reagent adsorption studies on component mineral suspensions and synthetic mixtures were used to assess the influence of various process parameters (e.g., reagent dosages, particle size, solids concentration, and conditioning regimes) on single-stage selective flocculation tests involving natural silicate and carbonate ores. In these tests a high degree of selectivity was achieved using low dosages of a low-molecular-weight cationic dispersant and a mid-range molecular weight cationic flocculant. Testwork was then extended to the evaluation of several floc recovery systems (sedimentation, hydrocyclones, and floc flotation) in a multi-stage process involving redispersion and reflocculation. Results from these tests clearly demonstrated the feasibility of producing marketable phosphate concentrates from both ore types.

Jan 1, 1999

By N. Abdel-Khalek, A. Youssef

The tailings from beneficiation of Florida phosphate ore contain considerable phosphate values. These tailings are currently discharged into large disposal areas. More than three million tons of P2O5 are discarded with the phosphatic slimes (clays) in Florida each year. These phosphatic clays not only create one of the most difficult disposal problems in the mining industry, but also represent a tremendous loss of P2O5. As the quality and quantity of Florida’s phosphate reserves diminish, recovery of phosphate values from these wastes becomes more vital. This will also significantly lengthen the life of these important phosphate resources. Thus, it is the goal of this study is to recover more than 80% of the phosphate values in phosphatic clays. A systematic study to characterize samples taken from inactive disposal area has been conducted. The data suggest that natural segregation takes place in the settling pond. In other words, coarser particles settle closer to the feeding point of the pond. Flotation tests were conducted using the +325 mesh fraction of samples obtained from sites close to the feeding point (sample# A) and further from that point (sample# B) and (sample# C) which is even further away from the feeding point. The results show the following: Sample# A gives the highest grade and recovery, followed by that of # B and then of # C. There is a great potential to upgrade sample# A where concentrates of high grade (68.9-70.1 % BPL or 31.5- 32.1% P2O5) and recovery (~86.6-95.4 %) can be obtained under the investigated operating conditions (1.0 – 2.0 kg/ton sodium silicate and 1.5—3.0 kg/ton collector mixture of fatty acid and fuel oil). The mineralogical and chemical analysis of all samples is presented in this paper for comparison. Only in this paper, the flotation data of (site # A) are discussed below.

Jan 1, 2002

By C. M. Rastle

Resource Technology Associates (RTA) has developed a flocculation-flotation process that increases phosphate recoveries from current operating plants. A major advantage of the process is the ability to float the phosphate from the matrix without a preliminary separation to remove the slimes. A majority of the material that is now discarded in the slime fraction will be recovered in the new process. The process can also be used to recover phosphate material from current slime ponds either as a separate feed or mixed with matrix material. Flotation tails from the process demonstrate improved dewatering characteristics then those seen from conventional float tails.

Jan 1, 1988

By Sylvain J. Pirson

A fluid mining process has been developed by the writer whereby phosphate deposits are leached in place through the injection of a weak and recoverable acid solution much in the same manner as in the recovery of oil by water-flooding. The phosphate mineralization is dissolved In place, selectively treated in a surface plant and a high grade diacalcium phosphate substatially free of fluorine (hence of feed grade quality) is recovered. Because of its finely divided and powdery state, the plant product is highly desirable as stock material for the manufacture of triple-superphosphate and of high grade phosphatic chemicals. Details of well completion, fluid injection, areal sweep and solution efficiency, plant design and economics are presented The rate of return on investment is particularly attractive under conditions of shallow occurrence of the phosphate mineralization such as in the Bone Valley formation of Florida. The process may also be applied economically to the Phosphoria formation of wide- spread occurrence in the Northwest United States through the application of selective formation fracturing (or Hydrofrac) previous to fluid mining operations.

Jan 1, 1959

By James C. Irvine

Pea Ridge Iron Ore Co., previously Meramec Mining Co., a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp., mines and pelletizes iron ore at the Pea Ridge mine. The Pea Ridge property, now wholly owned by St. Joe, is located near Sullivan, MO, about 112 km (70 miles) southwest of St. Louis. The ore body was delineated in the mid-1950s by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drill¬ing in 1956 and 1957 confirmed that the ore body was minable. Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Pro¬duction commenced in April 1964. The ore body is overlain by about 396 m (1300 ft) of flat-bedded sediments. It is tabular, about 792 m (2600 ft) long, and up to 182 m (600 ft) thick. It dips about 1.39 rad (80°) and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 45-m (150-ft) intervals. Crosscuts were driven across the ore body on 58-m (190-ft) centers. Banks of stopes were mined between the crosscuts by a modi¬fied shrinkage stoping method (Fig. 1). This was done by undercutting a 12 m (40 ft) wide by 45-m (150-ft) long block and blasting horizontal "lifts" drilled on 1.5-m (5-ft) intervals. A 20-m (65-ft) sill was left between levels which contained the slushing drift, fin¬gers, crown pillar, and adequate thickness for support. By 1971 a "lattice" of pillars had been left and mining had progressed to the point that an orderly pillar re¬covery program was necessary (Fig. 2). INITIAL PILLAR RECOVERY PROGRAM The program was started in the upper western por¬tion of the ore body, furthest from the shafts. Due to the fact that the ore body narrows in the western extremity, the stope orientation was changed to mini¬mize required development. This left pillars 18 to 24 m (60 to 80 ft) wide and up to 79 m (260 ft) in length. The area selected for the start of the program had been developed between the uppermost level at 419 m (1375 ft) below surface and the 510-m (1675-ft) level. Due to the 91-m (300-ft) difference in depth, sublevels were driven on 15-m (50-ft) intervals and the ore was taken by rather classical sublevel stoping methods (Fig. 3). This left a structure with accesses compatible with drilling with 114-mm (41/2-in.) bore drifters mounted on columns. Most holes were drilled 63.5 mm (21h in.) diam and reamed to 100 or 127 mm (4 or 5 in.). Practical depth capability was about 21 m (70 ft) for holes above horizontal and 13 to 15 m (45 to 50 ft) for holes below horizontal. The interdependence of these long pillars necessitated that several of them be blasted simultaneously, making large blasts the only practical approach. Due to the required length of such loading cam¬paigns, it was necessary to select an explosive which could stand in the hole in an underground mine for a 4 to 6 week period. After some consideration a pump¬ able water gel was selected for both uphole and down¬hole loading. This afforded a high velocity, high density explosive which could be handled in large quantities. Loading the downholes was a fairly straightforward process. The holes were lined with an 8 mil polybutylene plastic sleeve and water gel was pumped into them with a double diaphragm pump mounted on the bottom of a stainless steel tub with a 81-kg (180-1b) capacity. Potential leakage into cracked areas of the pillar was

Jan 1, 1982

By S. K. Kawatra, T. C. Eisele

Under normal conditions, a significant amount of pyrite is recovered in the froth during flotation of high-sulfur coal. To reduce this pyrite recovery, it is first necessary to determine the primary recovery mechanism, because different measures are required for reducing entrainment, locked particle flotation or true hydrophobic flotation. In this paper, evidence is presented that suggests that hydrophobic flotation is not the major mechanism for recovery of liberated pyrite; instead, the bulk of the recovered pyrite results either from simple entrainment or from being physically locked with floatable coal particles.

Jan 1, 1993

By S. K. Kawatra

Under normal conditions, a significant amount of pyrite is recovered in the froth during flotation of high-sulfur coal. In order to reduce this pyrite recovery, it is first necessary to determine the primary recovery mechanism, as different measures are required for reducing entrainment, locked-particle flotation, or true hydrophobic flotation. In this paper, evidence is presented which suggests that hydrophobic flotation is not the major mechanism for recovery of liberated pyrite, and that the bulk of the recovered pyrite is there either as a result of simple entrainment or by being physically locked with floatable coal particles.

Jan 1, 1991

By Harold L. Lovell

The current interest in sulfur recovery from coal measures is indicated by a recent report prepared for the Public Health Service (1) It was noted that the concentration of pyrites to a product containing 42-48 percent sulfur should be relatively easy with present equipment. It was also pointed out that a utility potential exists for the high-grade iron concentrate which results as a by-product of the consumption of pyrites in the acid-making process. However, economics, marketing characteristics, location, transportation, pyrite levels and coal characteristics all strongly influence the feasibility of fostering industrial development of such processes. The extent of pyrite recovery and product quality is intimately related to pyrite particle size and its degree of liberation, a consideration usually given greater emphasis in metal ore preparation than in processing coals. Consequently, fine crushing to nearly 20 mesh probably will be essential for maximum practical processing. It appears that any process separating pyrites from a coal must involve a reprocessing of the normal preparation plant refuse, although some yields can be attained from certain existing operations as products from, a shaking table or a jig. Further, the existence of millions of tons of coal refuse cannot be ignored as a mined source of raw materials. The recovery of saleable products from such wastes, although untried on a large scale under current conditions, not only offers a realistic source of sulfur but also would achieve a secondary refuse with the minimal contents of substances which incite and propagate the burning of organic and inorganic combustibles. These silicate materials, thus beneficiated, also offer some potential for economic utilization. Some significant tonnages of pyrites have been produced from coal measures in the past, but the experience was, in general, not consistent with current coal preparation practice. The processes involved did not include cyclones and heavy media; nor were they carries out under the sociological environment concerned with pollution as exists today.

Jan 1, 1967

By John Litz

The term "refractory gold ores" refers to a wide number of ore types which are not readily amenable to direct cyanidation. A number of processes have been used commercially, and a number of others have been proposed and in many cases pilot tested. The selection of the appropriate process for an ore should not be arbitrary. Ore grade may preclude some processes, but a good understanding of the ore mineralogy can guide the selection. Many present gold operations and many potential gold operations have faced or are facing problems with refractory gold mineralization. We at Hazen Research have had the opportunity to perform bench testwork and in some cases, pilot plant testing on a number of refractory ores. The data included in this paper are excerpted from the testing of fifteen ores which cover a wide range of mineralogical compositions. This paper is not intended to be a treatise on the processing of refractory gold ores. In many of these gold projects, not all of the standard process alternatives were evaluated, nor was adequate mineralogical work performed. Such is the dilemma of contract research. "If I believe that I don't need it, why should I pay to have it done." The intent of this presentation, therefore, is to show how different ores respond to the various processes, and to demonstrate that one must have an open mind when developing a process for a specific refractory gold ore.

Jan 1, 1998

By John B. Goddard

The discovery of rhenium on the ion exchange resin used at the uranium in-situ leach operations at Palangana (Texas) led to a laboratory study on the possible methods of recovery. Rhenium, present in small amounts in the ammonium carbonate leach liquors as the perrhenate anion, Re04-, concentrated on the anion exchange resin, which was a factor of ten more selective for rhenium than for uranium. Various elution systems were investigated for stripping the rhenium from the columns after the uranium had been eluted with chloride or nitrate solutions. Although the perchlorate ion was most efficient on a molar basis, nitrate ion combined the benefits of efficiency and low cost. A 3[N] or stronger concentration of nitrate containing up to 0.5[N] H+ was found to be the system of choice. A rhenium sulfide concentrate containing up to 14% Re was obtained from these effluents by hydrogen sulfide precipitation. Preliminary experiments were run to produce rhenium metal powder by oxidation of the concentrate to NH4Re04 followed by hydrogen reduction at elevated temperature.

Jan 1, 1982

By B. Pesic

The recovery of silver and gallium from flue dust with acidified thiourea solutions was examined. The studied parameters were temperature, particle size, thiourea concentration, acid concentration, and the presence of oxidants. The study of the extraction of these two metals was undertaken in sulfuric acid medium. Silver dissolution was dependent on all examined parameters, whereas gallium extraction was only dependent on the initial acid concentration. The experimental results suggest a new very efficient method for recovery of both silver and gallium. Almost complete recovery of silver and gallium can be accomplished within 90 minutes. A technological scheme for recovery of both metals is also proposed.

Jan 1, 1986

By T. G. Carnahan

The U.S. Department of the Interior, Bureau of Mines, performed a laboratory-scale investigation of technique for recovering silver from pregnant solutions generated in the hydrometallurgical treatment of complex base-metal sulfide concentrates with oxidizing chloride media. This provides a means for recovery and simplified reduction of accessory silver to metal from ions. The method consists of adding NaI or KI to the solution to precipitate AgI, which is then contacted with Na2S to produce Ag2S and to regenerate the iodide salt solution. Silver recoveries range from 99 pct with 20 pct excess iodide precipitant, to-92 pct when the theoretical stoichiometric amount of precipitant is employed. The Ag2S product can be reduced to silver metal by contact with aluminum chips in NaOH solution followed by purification with conventional fire refining techniques.

Jan 1, 1980

By Charles A. Rhoades

The Bureau of Mines is conducting research on a dual leaching method that offers the potential for the economic recovery of the silver as well as the manganese contained in some of the domestic manganiferous silver deposits. The method is an initial leach of the manganese with aqueous SO2 followed by a neutralization rinse, and then by a second leach for the extraction of silver with cyanide solution. Refractory silver in these ores is either bound by the high-valence manganese, chemically resistent to cyanide leaching, or is rendered inaccessible to cyanide leach solution by the surrounding manganese mineralization. Less than 10 pct of the silver was extracted when -1 in +1/2 in (-2.5 cm + 1.3 cm) ore pieces were leached directly with cyanide. However, extractions of 25 to 82 pct for silver and 65 to 96 pct for manganese were attained in laboratory column leach tests using the dual method on the same size ore pieces from six domestic deposits.

Jan 1, 1984