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By Ting An Zhang, Chengzhi Jiang, Xiujing Zhai, Feng Xie

The extraction and concentration of nickel from aqueous solutions by emulsion liquid membranes (ELMs) using tributylphosphate (TBP) as extractant has been investigated in this research. The emulsion liquid membrane consists of a diluent (kerosene), a surfactant (Span 80), a mobile carrier (TBP), and a stripping solution (ammonia solutions). Important factors influencing the separation process including membrane composition, feed composition, and mixing speed have been examined. The results show that more than 90% of nickel in the feed solution could be extracted by ELMs under the experimental conditions, indicating the feasibility to use the designated ELMs to detoxify nickel-containing industrial waste solutions.

Jan 1, 2014

By B. R. Reddy, Y. V. Swamy, B. V. R. Murthy

Chromitiferous overburden containing nickel in oxide form is reduced with CO-CO2-N2 gas mixtures. This converts the nickel oxide to metallic nickel and restricts the iron oxide to the magnetite state to extract the nickel. Reduction experiments were carried out on 500-g samples in a bench-scale vertical reactor. The tests were run to establish the process parameters, including the optimum partial pressure of CO, the reducing gas flow rate, granule size and roasting time. The reduced calcine is then subjected to an ammonia-ammoniacal carbonate leach to extract nickel selectively as its amine complex. More than 90% of the nickel could be extracted with a volumetric CO/(CO+CO2) ratio of 0.57 at 750°C and a roasting time of 90 min. The addition of 2.5% (by weight) charcoal to the granules improved nickel extraction by 11 % to 19%. Nickel leaching efficiency with ammoniacal solution is compared to the bromine-methanol method.

Jan 1, 2000

By Z. H. Khan, A. K. Saha, D. D. Akerkar

The possibility of extracting nickel from a low-iron, highly siliceous nickel ore of Indian origin using roast reduction followed by an ammonia leaching process has been examined. The ore used in the investigation contained 0.82 wt% Ni, 6.20 wt% Fe3O4, 12.04 wt% MgO and 60 wt% SiO2. Phase changes in the constituent minerals on heating were studied by differential thermal analysis which revealed endothermic peaks at 110 and 695 degrees C and an exothermic peak at 810 degrees C. Significant process variables in studies of roast reduction include the amount of reductant, the temperature of reduction and chemical kinetics. Maximum nickel extraction was achieved by reducing the ore with 10 wt% fuel oil at 750 degrees C. The reduced ore was given a preconditioning treatment in concentrated ammonium hydroxide solution to oxidise most of the metallic iron and was then leached in aqueous ammoniacal solution at ambient temperature to dissolve nickel. This gives 88.5% extraction of nickel as soluble ammine complexes.

Jan 4, 1992

By M. A. Cawley, J. W. Jordan

"IntroductionTHE ORE PILLARS in the lower ""A"" ore body of Waite Amulet Mines, Limited, can be divided into two groups, those mined in the open by sub-level blast-hole methods, and those to be mined .against coarse rock fill. Of the three pillars in the last group one has been mined and is dealt with in this paper. The method of mining pillars against coarse rock fill was suggested by the late Dr. Oliver Hall, who was Consulting Engineer for the Company when recovery of No. 2 East pillar was being planned.HistoryThe Waite Amulet mine is in northwestern Quebec, nine miles north of the town of Noranda. The property is approximately three miles long by one mile wide. The present Company, which was formed by the amalgamation of Amulet Mines, Limited, and Waite Ackerman-Montgomery Mines, Limited, commenced operations in the summer of 1937. In the fall of 1939 a new .flotation mill was erected with a capacity of 1,200 tons per day, which was increased to 1,800 tons per day in 1943. The mill is presently operated at 1,400 tons per day.To date, seven orebodies have been found, the major ones being Lower A, East Waite, and Waite. Three separate concentrates are produced, copper, zinc, and pyrite. Total tonnage mined to the end of 1954 from all orebodies has been 7,340,000 tons, which averaged 4.6 per cent copper, 5.3 per cent zinc, and 0.037 oz. gold and 1.27 oz. silver per ton."

Jan 1, 1955

By R. W. Bartlett

High salinity geothermal brines typically contain up to several hundred parts per million of valuable heavy metals, and because of the large potential flow of geothermal brines from future power plants, metal production as a by-product could be significant A-planned 55-MW demonstration plant near Niland, California should generate enough spent brine to contain approximately $670,000 per year in gross metal values. Lawrence Livermore Laboratory estimates (1) a Salton Sea Geothermal potential of 1300 to 8700 MW, which would translate to possible yearly heavy metal values between $5,800,000 and .$l05,000,000. Since these estimates were made on a conservative basis, this geothermal reservoir is potentially a large ore deposit. The heavy metals in order of decreasing abundance in high salinity brines from the Salton Sea geothermal field are usually manganese, zinc, iron, lead, and silver. All but manganese and iron are valuable enough to be of commercial interest. Minor amounts of copper and arsenic are also present. Although heavy metal precipitates, particularly lead, have been found in scales within pipelines and obtained from steam flashing, most of the metals remain soluble during steam production; hence removal of the valuable metals from the spent brine before reinjection in ground reservoirs is indicated. Because the heavy metal concentrations are low, a very simple and inexpensive process is required. The heavy metal value in Niland brine is only ' about $1 per 1000 gallons; see Table 1. The brine composition shown in Table 1 was taken from one of the brine samples used in our study. Projected post-flash brine from a steam plant has somewhat higher Zn and Pb values, but the gross heavy metal value is still less than $2 per 1000 gallons.

Jan 1, 1980

By Luis W. Pommier

The developed processes, the subject of this paper, extract the hydrocarbons, halogenated organic compounds and metal contaminants from hazardous soils, rendering them non-hazardous and ready to be placed back on the ground. The processes first segregate a non-contaminated fraction from the contaminated soil. This initial task is performed using available minerals processing technology: sizing, gravity separation, magnetic separation, froth flotation, etc. The remaining contaminated fraction, is submitted to two leaching operations for the removal of organic and metal contaminants, whenever both are present or to a single leaching operation if only one type of contaminant is present. The decontaminated soil fraction is mixed with the non-contaminated fraction and is placed back on the ground. The leaching liquors, used at neutral pH, are non- hazardous. After decontamination the liquids are processed to extract the contaminants, now valuable materials, for commercialization. Organic contaminants are removed by distillation of the solvents. The metal contaminants are precipitated in their oxide form. In both cases the regenerated solvents are recycled. The decontamination processes are economically competitive with transportation and disposal of the soils to permitted facilities for hazardous materials, and most importantly the developed processes remove the liabilities from the Responsible Parties. The law establishes that these liabilities do not disappear with either fixation nor with transportation and disposal since the environmental liabilities are of the ? cradle to grave? type.

Jan 1, 1994

By G. M. Wilemon, B. J. Scheiner

The Bureau of Mines conducted investigations on extracting phosphate values from fine-particle wastes generated during the mining and beneficiation of phosphate ore by leaching dried wastes using a H2SO4-methanol mixture. Although phosphatic clay waste contains up to a third of the phosphate originally present in the ore, recovery of these values represents a significant challenge due to the fine grain size of the material, the particles averaging less than one micrometer in diameter. Waste samples obtained from two Florida producers yielded phosphate extractions of 71 and 84 pct, respectively, when leached under ambient conditions and at a stoichiometric acid ratio of 1.4:l. Elevated temperatures did not improve the amount of phosphate extracted. Undesirable A12O3, Fe2O3, and PlgO impurities were largely rejected; higher grade wastes gave acids that meet established industry impurity standards and lower grade samples yielded acids that marginally fell short of the standard. Separation of the crude acids from the insoluble residues was accomplished via vacuum filtration through either filter paper or commercial filter cloths. Both filtration rates and solids retention were improved by the addition of 0.5 lb/ton of poly(viny1 acetate) (PVAc) or poly(ethy1ene oxide) (PEO) to the methanol-acid slurries. These results show potential for substantially increasing the reserves of an important domestic resource.

Jan 1, 1989

By Vladimir Lobanov, Vladimir Skorohodov, Andrew Evdokimov, Ural State Technical University 19, Boris Radionov

"The bulk of raw materials in metallurgy of secondary platinum metals are the wastes such as multilayered ceramic-metal systems, e.g. the spent catalysts of the chemical and petrochemical industry, automobile catalysts, etc. Due high chemical resistance of a ceramic basis the extraction of platinum metals by leaching even at very fine crushing is poor. The study of sorption leaching of platinum and a palladium from the spent automobile catalysts are carried out. The average content of PGM is 0.5%. A few methods of raw materials sampling were compared: 1) direct hydrometallurgical leaching of platinum metals from the fine-grained catalyst and atomic absorption analysis of the obtained solutions; 2) fusion to a metal or sulphidic collector of precious metals with the alloy analysis by direct spectrometry; 3) fusion with dissolution of a precious metals collector and atomic absorption analysis of PGM in solutions. Leaching allows detecting out no more than 90% of a palladium and 80% of platinum. Fusion to a collector gives the best results at the chemical ending. Researches on sorption leaching are executed. Under optimum conditions the extraction of platinum metals to concentrates is comparable with similar indicators of collector fusion.The large part of raw materials in metallurgy of secondary platinum metals are the multilayered or tridimensional ceramic-metal systems, in particular the spent catalysts of the chemical and petrochemical industry, automobile catalysts, ceramic-metal condensers, etc. The content of the sum of PGMs can be from hundreds grams per ton to several percent. For processing of such materials the hydrometallurgical methods of platinum metals leaching by various reagents are used commonly. High chemical resistance of the ceramic basis covering the metal particles causes low recovery even at very fine grinding. Other fundamentally different method of catalysts opening includes dissolution of the ceramic basis in the solutions of hydrofluoric acid and the subsequent recovery of platinum metals from insoluble residue. Application of this method is associated to use of corrosive medias and appropriate equipment. Fusion of the specified materials with collecting of PGM by copper, nickel or lead at the specialized enterprises provides higher extraction in semiproduct. At the same time, complexity and duration of the subsequent extraction of platinum metals concentrates causes high total cost per unit. Thereby this method is economically acceptable only for rather poor raw materials (no more than 0,1%). Undoubtedly, that the raw materials with the high content of PGM are reasonable to process by individual technology. In the present work the processing of some kinds of ceramic-metal raw materials, particularly automobile catalysts (neutralizers) with various collectors at the conditions of a target fusion are investigated."

Jan 1, 2008

By R. J. Traill

Introduction Following the discovery of high-grade deposits of pitchblende in the Great Bear Lake area of the Northwest Territories by Mr. Gilbert LaBine in May, 1930, samples were submitted to the Mines Branch for examination and investigation as to suitable methods of treatment (1). The purpose of this paper is to present, briefly, the results of the investigative work conducted on the ores and an outline of the methods developed for the extraction of the radium content and recovery of the by-products. The paper is presented in two parts. Part I gives a review of the history of radium production and of the principal processes used in other countries. Part II describes the work and gives the results of the Mines Branch investigation. Part 1-Radium Production In Other Countries Historical Outline Uranium was discovered in 1789, and more than a century later Madame Curie, while studying the physical properties of this element, observed that the ore of uranium exhibited greater activity than the pure uranium com-pounds she had prepared and traced the cause to the presence of a new element which she named radium. Eventually, research proved radium to be a disintegration product of uranium, and always present in uranium ores.

Jan 1, 1933

By Desmond M. Levins, Reginald K. Ryan

"Removal of radium from uranium tailings would substantially lessen their environmental impact because it would eliminate, or greatly reduce, the hazards arising from radon emanation, leaching of radium, dust dispersal and gamma radiation. A process is described to extract radium from tailings by leaching in 3M Nael solution. The radium can be rapidly leached at ambient temperatures by either multi-stage contact in agitated vessels or washing on a tailings filter cake. Extraction efficiencies ranging from 80 to 94% have been measured for tailings derived from a number of Australian ores.Methods for radium recovery from the leach solution by precipitation, ion exchange and adsorption have been examined. Precipitation as barium/radium sulphate appears to be the most suitable approach. A possible flow diagram for the process is proposed which involves recycle of the saline leach liquor.IntroductionDisposal of mill tailings is the most contentious environmental issue facing the uranium mining industry. The conventional approach to managing tailings is to store them in a dam above ground. Attempts have been made to stabilize some tailings piles by physical or chemical methods. In a few cases, the tailings have been covered with earth and revegetated. However, most tailings piles from abandoned mills are in a neglected state and subject to the erosive and dispersive forces of nature. The cost of remedial action is high. The United States Department of Energy (DOE) has estimated that it will cost up to $130 million to rehabilitate abandoned mill sites in the U.S .A..Environmental authorities are now insisting that uranium mining companies improve their methods of tailings disposal. In the United States, the preferred approach is burial, either in mines or in specially excavated pits. Tailings retention systems as operated in Canada are regarded as storage facilities until an acceptable method of disposal is found. In Australia, the Fox Commission recommended that tailings from the Ranger project should be returned to the open-cut mine. Returning tailings to the mine has considerable merit, but it is not always a viable option. In many underground operations, it is not physically possible to fit all the tailings in the worked-out mine. In other cases, the groundwater hydrology of the mine may be unfavourable."

Jan 1, 1980

By R. A. Wildenboer, R. F. Sandenbergh

Rare earth elements (REE) are present at concentrations of approximately 0.36 % in phosphogypsum stacks located in Phalaborwa, South Africa. The REE are present in the phosphogypsum in solid solution in the gypsum and as fluorine and aluminium rich precipitates. Recovery of the REE offers opportunities for valorization and reducing the environmental impact of the phosphogypsum waste. This paper presents the results of investigations into the leaching behaviour of rare earth elements and other impurities from phosphogypsum using aqueous sulfuric acid. The effects of acid concentration, temperature, residence time, and the influence of impurities were evaluated for their impact on the solubility of rare earth elements in a sulfate medium. It was found that a significant fraction of the REE can be leached from the phosphogypsum in a range of conditions much less severe than traditional hard rock rare earth processing conditions. The REE were leached at a sulfuric acid concentration of 110 g/l, temperature of 30°C – 40°C and a residence time of approximately 8 hours.

Nov 6, 2024

By Zhazira Baltabekova, Nina Lokhova, Arailym Dukembayeva, Madali Naimanbayev, Zhantore Dzhurkanov

"Recycling of industrial waste provided they have enough valuable elements requires significantly less investment compared with ore processing. Promising man-made raw material of rare earth elements (REE) is phosphogypsum. Works performed on dismantlement of phosphogypsum only concern phosphogypsum processing from apatit. In this case, none of the developments was tested for phosphogypsum produced on the basis of phosphorus, chemical and mineralogical composition of which is markedly different from apatites. We studied phosphorites of the Karatau deposits (Kazakhstan). Mineralogically all the layers are composed of three main ingredients: phosphate, silica and carbonate, which account for 92-96% of the weight composition of ores and rocks. Of impurities there are pyrite, iron hydroxides, gypsum, fluorite. In addition to the actual composition of phosphate in phosphate ores there are apatite. Silica is presented by chalcedony and quartz, and carbonate by calcite and dolomite. The results of X-ray microanalysis showed point cluster of grains containing REE in samples of phosphate ore. Analysis of the grain showed that REE are in the form of phosphates and silicates, which according to numerous studies in apatites is not established. In this regard, phosphogypsum, waste of production, is different in composition too. Therefore, transfer of techniques proposed in the methods for extracting REE from phosphogypsum after processing apatite, did not give the stated results and required a new approach to solving this problem. It should be noted that the phosphogypsum produced by processing the Karatau phosphorite contains medium group of REE by 2.6, and heavy group by 7.8 times more than phosphogypsum from processing the Kola apatite. Given the shortage of medium and heavy REE research to develop ways to extract the lanthanides from phosphogypsum after processing phosphate rock is promising. The study contains results of studies on the extraction of REE from phosphogypsum after Karatau phosphorite processing."

Jan 1, 2014

By P. Zhang, S. Al-Thyabat

"Phosphate rock contains traces of rare earth elements (REEs) and can be a secondary source of these critical materials as large tonnages of phosphate rock are mined annually. Attention has mostly focused on the extraction of REEs from phosphogypsum, which contains more than 70 percent of the REEs reporting to phosphate concentrate, with only limited work conducted on REE extraction from sand tailings and slime even though they account for 16 percent and 15 percent, respectively, of REEs mined with phosphate matrix.In this work, phosphate flotation tailings were upgraded by gravity separation and froth flotation. Gravity separation was conducted using a laboratory shaking table, while flotation was conducted in a Denver D-12 flotation cell. The concentrated tailings were then leached by nitric acid followed by REE extraction with solvent and ion-exchange resin.The sand tailings were assayed as having 2.6 percent phosphate (P2O5 ) and 198.1 µg/g REEs. It was found that the shaking table could produce tailing concentrate assayed as having 8.6 percent P2O5 and 616 µg/g REEs but with only 20 percent REE recovery, while the froth flotation produced froth concentrate assayed as having 8.1 percent P2O5 and 368.2 µg/g REEs with 63.5 percent REE recovery. Leaching the flotation concentrate with 5.2 M (25 percent) nitric acid followed by extraction with solvent and ion-exchange resin yielded precipitates with REE contents of 0.926 and 0.314 percent, respectively, compared with 0.716 and 0.213 percent when table concentrate was used.IntroductionRare earth elements (REEs) consist of a group of 17 elements: the 15 lanthanides along with yttrium (Y) and scandium (Sc). They are rare not because of their scarcity in the Earth’s crust but because they are rarely found in economic and mineable deposits. REEs are called “seeds of technology” because of their importance to some critical technological industries such as renewable energy, hybrid cars, electronics and batteries. On average, an electric car or a hybrid car contains about 10 kg of REEs (Grasso, 2012; Radhika et al., 2011).REEs are usually categorized into light (LREE) and heavy (HREE). LREEs consist of lanthanum (La), cerium (Ce), praseodymium (Pr), neodymium (Nd), promethium (Pr), samarium (Sm), europium (Eu), gadolinium (Gd) and scandium (Sc), while HREEs consist of terbium (Tb), dysprosium (Dy), holmium (Ho), erbium (Er), thulium (Tm), ytterbium (Yb), lutetium (Lu) and yttrium (Y). About 200 minerals contain traces of REEs, but only three are considered economically significant: bastnasite, with the formula La,Ce(CO3)F; monazite, with the formula(Ce,La,Nd,Th)PO4; and xenotime, with the formulaYPO4 (Jordens, Cheng and Waters, 2013)."

Jan 1, 2016

By F. Cheuk, E. Krause, V. G. Papangelakis, G. A. Moldoveanu, S. Abendschein

Systematic studies have been conducted at the University of Toronto in order to evaluate lanthanide extraction from clay minerals from different geographical regions. It was found that rare earth elements (REE) are easily recovered into an aqueous solution via an ion exchange mechanism by leaching with electrolyte solutions containing monovalent cations under ambient conditions. In this study we assessed extraction levels of individual REEs as a function of initial clay composition and leaching conditions. It was found that, in spite of variations in clay origin, all clays consistently reached peak REE extraction levels with fast kinetics. However, terminal extractions were clay-specific. The feasibility of using seawater as a leaching solution and potential synergies with ammonium sulphate as a lixiviant was also investigated. Seawater was found to be inferior to using ammonium sulphate, and provided no observable synergy in leaching performance when used together. The aqueous phase trapped in leached clays was shown to contain significant amounts of REE, as demonstrated in clay washing tests. Techniques to improve the REE concentration in leachate through process variations were investigated. However, it was found that most techniques capable of increasing REE concentrations in the resultant leachate also resulted in reduced REE extraction.

Jan 1, 2014

By G. Adachi, K. Machida, T. Ozaki, K. Murase

The chlorination rate with CCl4 was unexpectedly slower than carbochlorination using chlorine under the same reaction conditions. Chlorination with CCl4 was topochemical and the amount of monazite reacted after different periods suggested that the kinetic control step was the reaction between CCl4 and the monazite core at the interface. The kinetic order of the reaction was 0.5. The presence of a complex former, AlCl3, promoted the chemical vapour transport of rare-earth chlorides as complexes

Jun 18, 1905

By Chi Ru?an

Sollutions of ammonium chloride, nitrate and sulfate were studied in extraction of rare earths absorbed on a low-grade, kaolilnitic ore recently discovered in southern China. The results of percolation leaching tests and the recovery of the rare-earth values from these solutions are presented.

Jan 1, 1989

By M. Tanaka, H. Narita, T. Fuchikami, K. Morisaku, K. Nagao, T. Yoshida, K. Kuroda

The effect of various N-substituents on the performance of amide-containing tertiary amine compounds in the extraction of Rh(III) from HCl solutions was investigated to optimize the structure of these compounds for industrial use. It was found that N-2-ethylhexyl-bis(N-di-2-ethylhexyl-ethylamide)amine and tris(N-di-2-ethylhexyl-ethylamide)amine, which have long, branched alkyl chains, were suitable for Rh extraction, because they were soluble in a practical diluent (80 vol% n-dodecane?20 vol% 2-ethylhexanol) and stable against highly concentrated HCl/Rh solutions, and showed high selectivity for Rh(III) over Pd(II) and Pt(IV).

Jan 1, 2014

By James R. Evans

Orange County, adjacent to the populous and industrialized Los Angeles area, is the fastest growing county in California and is second only to Los Angeles County in total population (figure 1). In addition to an equable climate and a burgeoning economy, there is room in Orange County for growth and expansion. Growth over the past decade has resulted in rapid development of industrial tracts, housing areas, and shopping and commercial centers, mostly in former agricultural areas. The value of new building in 1972 was a record amount-over I billion dollars (figure 2). Problems relating to the proper direction of growth are the concern of both government and the private sector. The intelligent and efficient use of areas to be developed requires planning for the conservation and protection of the County's natural beauty and for the orderly development of natural resources. The availability of sand and gravel, which is used in many phases of construction, will be a major controlling factor in the future of the County.

Jan 1, 1974

By G. Mawire

The extraction of scandium from its oxide with a task specific ionic liquid [Hbet][Tf2N] offers potential advantages of lower energy consumption, recycling of the extractant and reduced environmental impact compared to conventional methods. In this paper, the influence of reaction temperature, solid to liquid ratio and water to the ionic liquid ratio on the extraction of scandium from a synthetic scandium oxide (Sc2O3) system was studied. The influence of investigated parameters on leaching was very significant as approximately 98% scandium was extracted from its oxide after 72 h. Furthermore, the ionic liquid also selectively extracted 47% scandium after 48 h from a mixture of pure metal oxides containing oxides of tantalum, niobium, and scandium. The research demonstrated the potential of [Hbet][Tf2N] to selectively extract scandium from columbite processing tailings and other scandium bearing ores like Ixiolite, Heftetjernite and Tantalite which predominately contain tantalum and niobium.

By T. C. Alex, K. M. Godiwalla, R. K. Jana

Manganese nodule, available in ocean beds, is a potential resource of the future for copper, nickel and cobalt. Recovering only these metals, containing less than 3% by weight, will result in the generation of large amount of residue. In addition to the material losses, several environmental and ecological problems arise out of this. The present study has been aimed at reducing the environmental and ecological problems associated with massive generation of this residue together with utilization of the resources. Presence of ~ 20% manganese and ~16% silicon in the residue has prompted us to check its suitability as a raw material for silicomanganese production. Lower manganese content and lower Mn/Fe ratio of the residue than that required for silicomanganese production has been corrected in this study by suitable blending. The efficacy of blending and its thermodynamic aspects using FACTSAGE software are discussed. The studies conducted indicate viability of bulk utilisation of the residue through silicomanganese production and thereby reducing the burden on environment to a great extent.

Jan 1, 2007