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By David T. Hobbs

Large quantities of highly alkaline, radioactive, liquid wastes are stored in underground tanks throughout the Department of Energy complex. These wastes will be pretreated to separate radionuclides and the remaining decontaminated liquid wastes incorporated into solid waste forms for permanent disposal. Significant savings in disposal costs could result by recovering .the sodium as sodium hydroxide for reuse or disposal as nonradioactive waste. Recent testing demonstrated an electrochemical separation process for the recovery of sodium hydroxide from alkaline waste solutions; The separation process uses ion-selective membranes to separate sodium from the bulk of the waste constituents under an applied electrical field. The sodium is recovered as a 10-15 wt % solution of sodium hydroxide. Results of bench-sc3Ie and pilot-scale tests with simulated and radioactive waste solutions are presented.

Jan 1, 1999

By J. W. Spyker, D. H. Filson, W. H. W. Husband

"Large deposits of sodium sulphate exist in southern Saskatchewan in the brines or muddy crystal beds of alkali Jakes. Currently, more than 300,000 tons per year of sodium sulphate are produced from these lakes and sold as salt cake to the paper industry. Most of the production is obtained from the brines of a limited number of Jakes.The Saskatchewan Research Council was requested to develop a process for recovering sodium sulphate from the muddy crystal beds so that this type of deposit could be exploited in the future. A flowsheet was developed with the aid of phase diagram data and the process was studied on a pilot-plant scale using raw material from three different lakes. Satisfactory yields and product purities were obtained in all cases, and an economic analysis indicated that a 100,000-ton-per-year sodium sulphate plant could produce salt cake at a cost of $12.26 per ton for a rate of return upon investment of 16 per cent at current market prices."

Jan 1, 1966

Some design and feasibility studies have been made into possible methods of economic recovery of sulphur from smelter gases emitted at Western Mining Corporation's Nickel smelter at KalgoorIie in Western Australia. The smelter discharges some 330 000 tonne/yr of sulphur dioxide in a stream of approximately 6 per cent by volume of S02.Three possible recovery methods have been considered:1. The manufacture of sulphuric acid using the contact process.2. The concentration of sulphur dioxide to a pure form using dimethyl aniline as an absorbing medium.3. The reduction to elemental sulphur using natural gas as the reducing agent.Design studies were carried out by final year chemical engineering undergraduate students at Royal Melbourne Institute of Technology (RMIT) in 1981 under the supervision of the author and other teaching staff of the chemical engineering department. These studies indicate that sulphuric acid manufacture is the economically preferred recovery route.Markets for sulphuric acid include several potential mineral processing applications and the established fertiliser industry.Sulphuric acid is currently manufactured in a number of plants located at coastal regions of Western Australia using imported sulphur. The sulphuric acid is then used for superphosphate or titanium dioxide pigment manufacture at adjacent plants.The possibility is considered of manufacturing superphosphate at Kalgoorlie from locally produced acid obtained from the smelter gases and from phosphate rock railed to the site from the coast. The advantages of such a scheme include improved scale and technology of operation compared with some existing coastal operations, use of a waste material to replace imported sulphur, and avoidance of costs and hazards associated with railing large quantities of sulphuric acid.

Jan 1, 1983

Recovery of Tantalum Capacitors from Waste Printed Circuit Boards Using a Pneumatic Separator

Sep 13, 2010

By Wenming Wang

Recycling of retired cadmium telluride (CdTe) photovoltaic (PV) modules has attracted attention due to new European environmental regulations. Recycling tellurium from CdTe PV modules is becoming more important because of the low availability of this metal. We investigated several treatment paths for reclaiming tellurium from solutions obtained from leaching fragments of CdTe PV modules. Our experiments showed that conventional soda ash neutralization/hydrolysis methods recovered only 50~60% of tellurium. However, by appropriately combining neutralization and sulfide precipitation, almost all the tellurium present in the leaching solution was precipitated and recovered; the concentration of tellurium in the processed solution was reduced from 900-1000 mg/L to less than 1 mg/L, yielding a tellurium recovery greater than 99%. This method can be incorporated in the scheme of recycling spent CdTe PV modules and manufacturing wastes. It may be also an alternative to the current operations for removing tellurium from copper anode slime leaching solutions.

Jan 1, 2006

By C. J. Chindgren

Two air classification systems were built and tested as separatory devices for recovering the nonferrous metal values from shredded automotive nonmagnetic reject scrap. A horizontal air system with a capacity up to 16 tons per hour was used to obtain preliminary separations, and a smaller capacity vertical system was used as a cleaning device. Shredded automobile reject material containing aluminum, copper, and zinc values of about $55 per ton were air-classified in the horizontal elutriator. Ninety-six percent of the metal was recovered in a 74-percent metal concentrate. Tandem operation of both classifiers recovered 92 percent of the metal in an 80-percent concentrate while rejecting 87 percent of the nonmetallics. Heavy media separation of the air-classified concentrate produced an overall recovery of 91 percent of the metal in the form of a 99-percent metal concentrate.

Jan 1, 1971

By N Rivera, A Music, C Cisternas

In an underground mine operated with panel caving and subjected to high stresses, geotechnical events such as collapses or rock bursts are fairly recurrent. Both situations can involve risks for people, considerable production losses and increasing operational costs.This paper discusses an operational methodology for the recovery of the undercutting front carried out at El Teniente mine, which uses panel cave mining with post-undercutting, after the occurrence of severe damage due to a seismic cluster. The main objective of the recovery of the undercutting front is to prevent the spread of the collapse to the production level, risking the operational continuity. The methodology is the result of repairs that were carried out in the year 2014 at the North Reserves sector from El Teniente mine. Its characteristics are an undercutting front of almost 1000 m long, with the presence of relevant structures and faults, high stresses and high seismic activity.The practical implementation of the proposed methodology was successful since it prevented the spread of the damage, completely recovered the undercutting front and avoided production losses during the execution of the repairs. Also, considering the origin of the damage and the operations associated to the subsequent recovery of the undercutting front, in this paper we discuss the main control variables to be measured. These variables should allow any operator and short-term planner to take action in mines operated with conventional panel caving in a complex geotechnical environment.CITATION:Rivera, N, Music, A and Cisternas, C, 2016. Recovery of the undercutting front in a mine with conventional panel caving – case study at North Reserves sector, El Teniente Mine, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 713–722 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By L. C. Bird, N. T. L. Landgren, M. W. Mikhail

"Through funds made available in 1972 by the Nova Scotia Department of Mines and Energy (DME) and the federal Department of Regional Economic Expansion, a program was initiated to recover saleable coal from mine waste dumps adjacent to the existing Acadia wash plant in Stellarton, Nova Scotia.Following evaluation of these dumps in 1972, the Acadia wash plant was modified to treat high-ash material. The first production was recorded in late 1974. By the end of 1979, a total of 738,202 t of dump material had been processed, yielding a total of 205,212 t of thermal-quality coal.The successful Stellarton operation led to considering the possibility of recovering thermal coal from other dumps. Evaluation of the Westville dumps indicated total recoverable reserves of 1,346,617 t averaging 53.8OJo ash.Parameters for plant design were defined from the washability characteristics of the dump material and the results of a bulk test. Two options were proposed; one was to use the existing plant with some modifications, the other was to build a modular mobile plant that could be used at future dump sites in the province after Westville. At the 25% ash level, the potential recovery of coal from the dump was estimated at 41% yield (by weight) or 549,679 t."

Jan 1, 1981

A modified liquation technique was developed for recovering tin from tinhouse scruff metal or hardhead containing about 10 per cent iron and 15 per cent occluded flux, palm oil, and moisture. This was effectedby heating the scruff metal to 950°C followed by rapid cooling to suppress the formation of FeSn2 and FeSn, and then subjecting it to the ordinary liquation and drossing processes of refining. By this method it was possible to recover about 83 per cent of the original tin content and, with subsequent smelting of the irony drosses, an overall recovery of 92 per cent was obtained.

Jan 1, 1969

By B. D. Pandey, M. S. Kim, B. S. Kim

The recovery of Sn from used Pb-free solder (Sn, 3~4 % Ag, < 0.9 % Cu, < 1.2 % Pb) has been investigated by electrorefining in sulfuric acid solutions containing small amount of hydrochloric acid. This involved preparation of the anode material by melting and casting the used solder followed by refining. During electrorefining, dendritic tin was deposited on stainless steel cathode. Increasing the amount of hydrochloric acid up to 0.3 mol /L in the electrolyte enhanced the current efficiency to 92.2 % and the enrichment ratio of silver to 94.3 % in the anode slime which was otherwise 84.7% and 81.6 % in absence of HCl, respectively. The electrolyte containing hydrochloric acid has also prevented the precipitation of tin oxide in the bath. In the scale-up experiment using 1.6 kg anodes, 84.3 % current efficiency and close to 100 % Ag enrichment ratio were obtained, while the purity of electro-refined tin was over 99.9 %. The energy consumption of the winning process was found to be 0.151 kWh /kg-Sn which is competitive as compared to commercial tin electrorefining processes.

Jan 1, 2014

By Zhenglong Li

A process has been developed to recover titanium from a blast-furnace slag which contains about 20-25% Ti02 and 12% Fe. In this process. "hot" slag from the blast furnace is directed to a separate furnace where it is reheated with additives to form two simple slag phases: one is titanium enriched anosovite (Ti3O5) and the other is a titanium lean non-crystal glass phase. The modified slag is then processed by conventional beneficiation techniques to produce a marketable titanium concentrate and a usable slag material. This paper focus on separating anosovite from glass by electrostatic separation and froth flotation.

Jan 1, 1996

By Gyeonghye Moon, Inhyeok Choi, Jyothi Rajesh Kumar, Jin-Young Lee

"This paper deals with the recovery of the tungsten and vanadium from spent SCR catalyst leach liquors by different hydrometallurgical methods. The first part consists of pressure leaching of spent SCR catalyst using caustic soda followed by metal recovery by liquid-liquid extraction. In a leaching process, the following experimental parameters were tested and optimized: NaOH, particle size of feedstock, temperature, and solid to liquid ratio. The optimum condition for leaching of tungsten and vanadium was under the following conditions: NaOH 3 mol/L, temperature 250°C, particle size <300 µm, and solid to liquid ratio 0.4. From these optimum conditions in this experiment, the dissolution efficiency of vanadium and tungsten was 91.5and 87%, respectively. Liquid-liquid extraction following commercial extractants were tested: TBP, Cyanex 272, D2EHPA, Aliquat 336, Primeme JM-T, Alamine 308, Alamine 304-1, and methyl-tri-octyl-ammonium chloride. The conditions for both valuable metal recoveries were optimized.INTRODUCTIONSelective Catalytic Reduction (SCR) catalyst has become the most effective to reduce nitrogen oxide (NOx) emissions from stationary segments such as power plants and waste incinerations around the world due to the reinforcing regulatory trend and emerging environmental issues (Forzatti, 2001; Spivey, 2002; Li et al., 2011).Currently, the most widely used SCR catalyst is V2O5-WO3/TiO2 system due to its great catalytic properties (Li et al., 2011; Marshneva et al., 1995; Marbergeret al., 2015). In general, the composition of the SCR catalyst used in the stationary application is 0.5 to 1 wt% of V2O5and 7 to 10 wt% of WO3(Marberger et al., 2015; Chen &Yang, 1992, Vuurman et al., 1991; Alemany et al., 1991). Currently spent SCR catalysts are treated as waste and disposed of in landfills. This causes a lot of waste of secondary resources of vanadium and tungsten. In terms of industrial value of vanadium and tungsten (Gupta &Mukherjee, 1990; Moskalyk &Alfantazi, 2003) and reduction of pollutant emissions, it should not be disregarded for their recovery for their material recycling from spent SCR catalyst. However, investigations for recovery of metal values from spent SCR catalyst has been rarely carried out."

Jan 1, 2017

By P. B. Altringer

Searles Lake, CA, contains the largest known domestic tungsten resource. The brines in this near-dry lake bed, located in the Mojave Desert 130 mi northeast of Los Angeles, contain an estimated 135 MMlb of tungsten. However, the low concentration of tungsten (0.080 g/L W03) in the supersaturated brine makes selective tungsten extraction very difficult. Although many brine chemicals are extracted from the lake, tungsten recovery was an elusive goal prior to Bureau of Mines development of an ion-exchange technique. This publication describes the Bureau process. The following subjects are ad-dressed: ? The Searles Lake tungsten resource. ? General problems associated with tungsten recovery. ? Bureau research objectives. ? Synthesis of ion-exchange resins. ? Development of a two-stage ion-exchange tungsten extraction and concentration procedure. ? Product recovery and purification. ? Economic evaluation. Flowsheets for recovering three products (calcium tungstate, sodium tungstate, and tungstic oxide) are presented and evaluated, as well as a summary of the research effort and potential applications.

Jan 1, 1985

By Robert A. MacGregor

"Following several months of test work, water leaching in mined-out areas at Stanrock commenced on a production basis in February, 1963. A combination of fast high-pres-sure washing and rewashing of stopes on a predetermined cycle, and the intermittent sprinkling of low-grade heaps, using both fresh water and recirculated acid water, has proved most successful. There was also a correlation of daily sampling data points to a relationship between the leach water grade and the grade of the material being washed, the leaching time and the location along the dip. An increasing tolerance for uranium by the acid-producing bacteria is also indicated. A steady production rate of 10,000 to 12,000 lbs. of u.08 per month is being maintained, with an over-all mine water grade of 0.25 lb. per ton and a pH of 2.3."

Jan 1, 1966

By R. E. Brewer

Pilot scale flotation tests were conducted with a l?ft-diam column at the Barrick Resources (USA) Mercur mine, Tooele County, UT, to scavenge unleached (0.01 to 0.018 oz/st) gold values from the tailings of a carbon-in-leach (CIL) circuit. The optimal recovery was 35 pct of the unamenable-to-leaching gold, at a grade of 0.353 ounces of gold per short ton, using a mixture of dithiophosphate, xanthate, and diesel fuel as a collector. This was achieved by floating approximately 1.6 wt pet of the CIL tailings. The gold values are believed to be contained in arsenic-bearing minerals, attrited carbon, and carbonaceous material in the ore. Redox potentials were measured with a gold electrode during the conditioning stage prior to flotation. The correlation between flotation recovery and redox potential values could be used to control reagent additions. Treatment of a previously leached and weathered tailings material, averaging about 0.055 ounces of gold per short ton, was also investigated. The tailings were first water leached at approximately 60? C, and the water-leached tailings, containing approximately 0.03 ounces of gold per short ton, were then fed to the column flotation unit. The effects of pH&apos; and sodium sulfide addition on gold flotation recovery and redox potentials were examined. Maximum gold flotation recovery, 36 pct, was achieved at pH 8,5 with a sodium sulfide addition of 0.18 Ib/st. Overall gold recovery by leaching and flotation was in the range of 70 to 80 pet.

Jan 1, 1989

By R. E. Brewer

Pilot scale flotation tests were conducted with a 1-ft-diam column at the Barrick Resources (USA) Mercur mine, Tooele County, UT, to scavenge unleached (0.01 to 0.018 oz/st) gold values from the tailings of a carbon-in-leach (CIL) circuit. The optimal recovery was 35 pct of the unamenable-to-leaching gold, at a grade of 0.353 ounces of gold per short ton, using a mixture of dithiophosphate, xanthate, and diesel fuel as a collector. This was achieved by floating approximately 1.6 wt pct of the CIL tailings. The gold values are believed to be contained in arsenic-bearing minerals, attrited carbon, and carbonaceous material in the ore. Redox potentials were measured with a gold electrode during the conditioning stage prior to flotation. The correlation between flotation recovery and redox potential values could be used to control reagent additions. Treatment of a previously leached and weathered tailings material, averaging about 0.055 ounces of gold per short ton, was also investigated using the column. The tailings were first water leached at approximately 600 C, and the water-leached tailings, which contained approximately 0.03 ounces of gold per short ton, were then fed to the column flotation unit. The effects of pH and sodium sulfide addition on gold flotation recovery and redox potentials were examined. maximum gold recovery, 36 pct, was achieved at pH 8.5 with a sodium sulfide addition of 0.18 lb/st.

Jan 1, 1989

By Fathi Habashi

"A process is proposed for the treatment of phosphate rock for the recovery of uranium and the lanthanides. The process assures the production of phosphatic fertilizers without polluting the environment with radioactive material.IntroductionWhile many processes for uranium recovery from phosphoric acid produced by the H2S04 route are known and utilised commercially, little is known about uranium recovery from phosphoric acid produced by the HN03 route1. Advantages of the latter route are both the elimination of the gypsum disposal problem, and the possibility of radium disposal in a controlled way2. Further, the lanthanides originally present in the rock can be recovered if the nitric acid route is used; they will be mostly lost in the gypsum produced by the sulphuric acid route3. In the present work a scheme for the recovery of uranium and the lanthanides from phosphate rock without interfering with the normal route for fertiliser production is proposed using HN03 as a leaching agent. Attention was also given to uranium decay products and their separation so that they might not contaminate the fertiliser4. Further, an attempt was made to defluorinate the leach solution in order to upgrade the fertiliser and at the same time recover the fluorine5. Rapid analytical methods for the determination of uranium, radium, and lanthanides in phosphate rock were developed6,7."

Jan 1, 1997

By F. J. Hurst

Phosphate rock, which is a prime source of phosphate used in fertilizer manufacture in the United States, contains uranium. In general, the uranium content of the rock increases with the phosphate content and most high-grade phosphate deposits (>30% P2O5) contain 0.01- 0.02% U3 8. The large tonnages of rock being mined represent a significant potential source of uranium that lends itself to recovery as a by-product of the wet-process phosphate industry. For example, in 1953, the U. S. Geological Survey estimated that recoverable phosphate reserves in the U. S. totaled about 5 billion tons and contained 600,000 tons of uranium. I This estimate did not include any rock from North Carolina which has since been estimated at about 2 billion tons. For the most part, the known, economically mineable U. S. reserves - estimated at about 7 billion tons - are concentrated in a few areas: 43% in the western states of Idaho, Montana, Utah and Wyoming; 28% in each of Florida and North Carolina and less than 1% in Tennessee. In the past, Florida has supplied about 80% of the total U. S. production, practically all of which is land pebble phosphate. All forecasts indicate increased production rates from North Carolina and the western states.

Jan 1, 1976

By M. Fan, Z. Long

"This paper presents some studies on recovering valuable elements from Chinese coal by-products, including coal combustion fly ash, stone coal ash, and coal flotation tailings. Both commercialized technologies and economic potential recovery methods are described in this study. Magnetically stabilized air-dense medium fluidized bed separation, tribo-electrostatic separation, froth flotation, and hydrometallurgical methods are discussed in recovering various valuable elements in Chinese coal by-products from different coalfields. These valuable elements include iron, carbon, aluminum, rare earth elements, lithium, gallium, germanium, uranium, titanium, scandium, silicon, vanadium and phosphorus, etc. The ultimate goal is to fully utilize all valuable elements in coal by-product. Each component within coal by-product should be efficiently segregated to its highest value. INTRODUCTION Coal is mainly composed of energy-producing elements carbon and hydrogen. In addition to these elements, coal usually contains many other valuable elements, including iron, aluminum, silicon, rare earth elements, lithium, gallium, germanium, vanadium, uranium, titanium, scandium, and phosphorus, etc. These coal-associated elements are significantly enriched in coal combustion by-product fly ash because of majority of main components, carbon and hydrogen elements, being removed. Table 1 provides the chemical composition range of the main elements oxides in 12 Chinese fly ash samples collected from coal power plants located in Northeast China, Northwest China, Southwest China, and Southeast China. It can be seen from Table 1 that the major mineral components of different ashes are similar, but the bulk chemistry of different fly ashes may vary widely due to the changes of ratios of the mineral components in different coals. The major mineral components in various Chinese coal combustion fly ash are silicates, iron oxides, and unburned carbon."

Jan 1, 2017

By Yuemin Zhao, Wenxuan Liu, Maoming Fan, Zhanyuan Long

"Coal is mainly composed of energy-producing elements carbon and hydrogen. In addition to these elements, coal usually contains many other valuable elements, including iron, aluminum, silicon, rare earth elements, lithium, gallium, germanium, vanadium, uranium, titanium, scandium and phosphorus. These coalassociated elements are significantly enriched in coal combustion byproduct fly ash because the majority of the main components, carbon and hydrogen elements, being removed. Table 1 provides the chemical composition range of the main element oxides in 12 Chinese fly ash samples collected from coal power plants located in Northeast China, Northwest China, Southwest China and Southeast China. It can be seen from Table 1 that the major mineral components of different ashes are similar, but the bulk chemistry of different fly ashes may vary due to the changes of ratios of the mineral components in different coals. The major mineral components in various Chinese coal combustion fly ash are silicates, iron oxides and unburned carbon.Chinese coal fly ash usually contains about 5 to 15 percent of unburned carbon, which are generally chars with irregular shapes and wide particle size distribution. In addition to unburned carbon, XRD analysis of four Chinese fly ash samples from Liaoning, Shanxi, Guizhou and Shandong provinces indicates that the fly ash samples are composed of minerals such as quartz (SiO2), mullite (3Al2O3.2SiO2 or 2Al2O3.SiO2), hematite (Fe2O3), maghemite (gamma form of hematite, ?-Fe2O3) and magnetite (Fe3O4). The mullite and quartz particles are usually present as spherical particles because they are formed by melting clays, feldspars, quartz, calcite and other common minerals in coal during coal combustion. The silicate particles have usually entrapped gas to yield hollow spherical particles because their lower melting point facilitates gas entrapment.The iron oxides are usually spherical also. The magnetite (Fe3O4) is derived from pyrite (FeS2), hematite (Fe2O3), siderite (FeCO3) and limonite (2Fe2O3.3H2O) in the coal. Maghemite (?-Fe2O3) is formed by dehydration of the mineral hydrohematite (?-FeO(OH)) and/or oxidation of magnetite (Fe3O4) during coal combustion. Maghemite (?-Fe2O3) and magnetite (Fe3O4) are strongly magnetic, so they are recovered from Chinese coal fly ash and widely used for coal processing in China (Fan et al., 2005, 2000)."

Jan 3, 2018