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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 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 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 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 Minji Jun

Leaching of the surface-coated metals from waste acrylonitrile butadiene styrene plastics (WABSP) has been investigated in the ammoniacal solution. The ammoniacal leaching of WABSP (typically containing 4.1 wt% Cu, 1.3 wt% Ni, 0.03 wt% Cr) efficiently dissolved copper and nickel, leaving unleached chromium as fine particles. The results indicate that the anions of different buffer media for metals leaching follow the order: CO3 2− > Cl− > SO4 2−. A declined leaching with high NH4OH–(NH4)2CO3 ratio (8:1) and a steep rise in extraction with high pulp in solution revealed the influence of anion (CO3 2−) and cation (Cu2+) present in the system. The leaching performed in carbonate solution by maintaining 5.0 M total NH3 concentration as NH4OH–(NH4)2CO3 ratio, 4:1; pulp density, 200 g/L; agitation speed, 400 rpm; temperature, 20 °C and time, 120 min yielded the maximum extraction (>99%) of copper and nickel. The study reveals a plausible recovery of the surface-coated metals from the WABSP without altering the properties of polymer material that can be recycled separatel

By A. E. Raddatz

The U.S. Department of the Interior, Bureau of Mines, is investigating methods for recovering tungsten from domestic resources. A calcine-quench-leach (CQL) technique was developed for the extraction of tungsten from the Golconda, Nevada, hot springs ores. The deposit contains disseminated tungsten, which cannot be beneficiated by conventional techniques and must be treated by hydrometallurgical methods. The CQL procedure consists of calcining the ore at 900° C for 3 hr, quenching in a solution containing 1.5 parts of NazCO3 and 1.5 parts of NaOH per part of WO3 in the ore and at a solution- to-ore ratio of 3 to 1, leaching at 200° C and at 250 psig for 3 hr, and filtering and washing the residue. A series of closed-circuit batch tests showed that 92 pct of the tungsten was extracted from a sample containing 2.32 pct WO3. The pregnant solution contained 11 q of W03/1.

Jan 1, 1982

By Urs Peuker, Thomas Mütze, Thomas Leißner, Inga Osbahr

"Mining in the Ore Mountains (Germany) has taken place for hundreds of years and in several stages. Each stage produced its typical tailing. According to the former state of the art of the processing much valuables could not be extracted. This was caused by complex interlocking of the minerals, low grades, or by upgrading processes not sensitive enough for small particle sizes. Thus the processing of these fractions was uneconomically. Modern technical possibilities and rising metal prices bring these tailings disposals into focus for the processing of strategic metals. Mining disposals can be classified in heaps with coarse, low grade rock, which did not undergo a processing, slags and fly ash from smelting, and tailing disposals from mineral processing. The latter is under investigation to integrate them into a geographic information system. The mineral processing of the material will be part of the database besides information about geographic location, structure, amount of valuables, and details of ownership of the heap. The tailings disposals are composed of coarse, low grade, interlocked particles from density separation processes, medium size particles, which are tailings from flotation, and fines with liberated valuables, which did not undergo flotation. Therefore mineral processing of these disposals comprises the separation of liberated gangue by classification, beneficiation of locked valuables with density separation, liberation of valuables by grinding, and beneficiation of liberated valuables with flotation. The presented work shows the results concerning the composition of one exemplary heap and results of the classification as well as of the preconcentration tests achieved with material from this heap."

Jan 1, 2014

By Yun Guo, Min-Min Lin, Bing Xie, Hong-Yi Li

The extraction of vanadium from vanadium-containing APV-precipitated wastewater by W/O microemulsion system was studied. In the N263/n-heptane/isoamyl alcohol/NaCl microemulsion system, N263 played double functions of cationic surfactant and extractant to form microemulsion in n-heptane, and isoamyl alcohol was added in the microemulsion as a cosurfactant. The extraction of vanadium by W/O microemulsion systems effectively accelerated the extraction and improved the extractatability because of the enormous increase in the micro-interfacial surface area. Due to ion association with extractant, vanadium was extracted into the microemulsion phase. The influence of different parameters on the extraction ratio (E%) were investigated, including the volume ratio of aqueous phase to microemulsion, surfactant concentration, pH value of the feed solutions, cosurfactant concentration, extraction time as well as temperature. In optimum conditions, the extraction ratio (E%) can reach 97.2%.

Mar 1, 2018

By Hassan Saalman, Mahmoud Mahmoud, Omneya El Hussaini

The Ferruginous sandstone ore was obtained from Ramlet Hemeyir area, Southwestern Sinai, Egypt, was found to contain 10% rare earth elements (REE). A representative sample has been subjected to sulfuric acid agitation leaching. About 93% of the REE content was released by mixing the ground sample to -100 mesh with concentrated H2SO4 at solid/liquid (S/L) ratio of 1/2 for 5h at a temperature ranged from 250 to 300o C. The REE was precipitated from the prepared sulfate leach liquor of pH 1.0 by adding 35% oxalic acid at 25o C. The precipitation yielded 98% REE. For selective separation of yttrium, the rare earth cake was leached with 31.5% nitric acid. Next, the aqueous solution was contacted with di-(2-ethylhexyl) phosphoric acid (D2EHPA) where 98% yttrium extraction was achieved. Stripping was done with hot H2O (80o C). Finally, a precipitation with oxalic acid was done followed by ignition at 950o C. A yield of 95% yttrium oxide was produced.

Jan 1, 2014

By Hsing Kuang Lin

Extraction of zinc from chloride solution with dibutyl butylphosphonate (DBBP) in Escaid 110 has been investigated. The distribution coefficient of zinc was found to be independent of equilibrium pH. This independence indicates that the mechanism of the extraction may be solvation. Slope analysis for the system of dilute zinc chloride solution reveals that two moles of DBBP react with one mole of zinc. A series of tests were conducted to bring DBBP to its full loading capacity and the analysis of the results confirms this stoichiometric factor of 2. Extraction isotherms were established by varying either the AIO ratio or the concentration of zinc in the aqueous feed. Both curves were found to be identical.

Jan 1, 1990

By Xudong Liu, Ganghua Fu, Yufeng Guo, Wei Chen, Yujia Tan, Tao Jiang

Willemite is usually found in high-silicon oxide zinc ore and zinc metallurgical secondary resources but without effective extraction method yet. In this study, zinc was extracted from willemite by sodium salt roasting and ammonia-leaching process based on the idea of zinc phase reconstruction. The effects of Na2CO3 dosage, roasting temperature and roasting time on phase transformation and zinc leaching rate were studied, respectively. The results show that Zn2SiO4 was transformed into ZnO or Na2ZnSiO4 during roasting process. After roasting at 800 °C for 2.5 h at ratio of n(Na2CO3)/n(Zn2SiO4) = 2.5:1, willemite was thoroughly transformed into ZnO. The zinc leaching rate of the roasted sample was 97.62% in NH4Cl–NH3–H2O leaching system, which increased by 91.57% compared with zinc leaching rate of 6.05% when willemite was directly leached in NH4Cl–NH3–H2O system.

Mar 1, 2017

By J. Bai, C. Zang, W. Yuan, L. Zhuang

"This study evaluates the efficiency of using mechanochemical reduction to assist the extraction of zinc from zinc ferrites by alkaline leaching. The transition of zinc ferrite into a metastable state after mechanochemical reduction with metallic iron contributes to the ready dissolution of zinc from the activated zinc ferrites in alkaline solution. Zinc ferrites were mechanochemically reduced under conditions of Fe:ZnFe2O4 mole ratio of 2:1, using 5 mm diameter stainless steel balls as the activation medium at a mass ratio of balls to raw materials of 25:1. Subsequent leaching in 6 mol/L NaOH solution at 90°C resulted in more than 70% Zn extraction. These results may be used to further develop a hydrometallurgical process for recovering zinc from zinc ferrites in alkaline solution.IntroductionZinc is an important nonferrous metal required for various applications in batteries, solder, dielectric materials, and piezoelectric materials. High-grade, easily concentrated ores are becoming scarcer: to meet world demands for zinc, it is necessary to develop economical and environmentally safe metallurgical technologies that use secondary materials such as leach residues of roasted zinc concentrates and electric arc furnace dusts. The zinc in the majority of these wastes occurs in the form of zinc ferrite (franklinite, ZnFe2O4), which is very stable and insoluble in most acidic and alkaline solutions (Langova and Matysek, 2009; Youcai and Stanforth, 2000).Several pyrometallurgical and hydrometallurgical processes (Jankovi et al., 2014; Langova and Matysek, 2010; Morcali et al., 2012; Shawabkeh, 2010) have been proposed for the treatment of zinc ferrites. Pyrometallurgical methods are costly owing to high energy consumption and their emissions of toxic gases are considered an environmental problem; hydrometallurgical methods are preferred for process economy and environmental reasons (Copur et al., 2004)."

Jan 1, 2016

By James A. Barclay, Ralph C. Kirby

The US minerals posture will be increasingly dominated by three interrelated factors. First, when viewed over the long range, the US supply situation can be expected to steadily deteriorate as demand increases. Domestic high-grade ores have been used up, and other nations compete with the US for world supplies. Second, as environmental awareness becomes more acute, an aroused citizenry is demanding that minerals industries conform their operations to rigorous new standards. Finally, the changing energy situation that confronts the US is forcing an appraisal of energy-use patterns on operations previously guided solely by economic principles and not by national priorities. While future US mineral production and consumption patterns are unclear at this time, it is certain that US demand for minerals will continue to outrun domestic supply. Thus, the US is faced with a challenge of producing more minerals and metals from lower grade resources. Meeting this challenge means a strengthening of US research and development efforts, coupled with their application in the remaining leadtimes available. Some specific examples of how the US Bureau of Mines (USBM) is assisting in the processing of low-grade resources follow.

Jan 6, 1975