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By Frédéric Bélanger, George P. Demopoulos, Pascal Coursol, Thomas Feldmann

"The use of scandium in industrial applications is expected to rise in the coming years, in large part due to the expected adoption of new, clean energy technologies such as solid oxide fuel cells or Sccontaining high strength aluminium alloys. Given that the supply of Sc is currently dominated by only a few countries (Russia and China) other sources for scandium extraction need to be investigated. This paper looks at different options for the recovery of scandium from waste solutions generated during the production of TiO2 pigments via the chloride process route. In particular, a direct solvent extraction (SX) approach and possible options for concentrating scandium before the SX step were investigated. It was found that the presence of Ti during solvent extraction causes severe formation of stable emulsions, which can complicate the successful application. This was particularly a problem after multiple contacts of the organic phase with fresh aqueous phase. On the other hand it was noticed that the presence of certain elements in the solution, e.g., iron, could possibly be utilized to concentrate scandium in a solid precipitate.INTRODUCTIONMany newly developed high-tech products depend on the use of chemical elements that are only produced in very small quantities. This can result in high and fluctuating prices, as can be seen from Figure 1, which may cause substitution with inferior alternatives, if possible at all (Reck & Rotter, 2012; US Dept. of Defense, 2013; Whittaker, 2011). One example is scandium (Sc) a member of the extended rare earth element family. There are only a few scandium ore deposits, but most Sc is produced from the processing of by-products of other metallurgical processes (Melcher & Wilken, 2013). Increasing demand for scandium metal and oxide (Sc2O3, scandia) is coming from its increased use in applications such as ceramic lasers (Li, Ikegami, & Mori, 2005; Rainer et al., 1982), high strength aluminium alloys (Lathabai & Lloyd, 2002) as well as solid oxide fuel cells (SOFC) (Thijssen, 2011). Especially, the spread of climate friendly SOFC’s, which are a commercial product today (Bloom Energy, 2014), is expected to be a main driver for future Sc2O3 demand. However, the rate of market penetration of this green technology depends on the availability of high purity Sc2O3 (>99.999%), since it specifically contributes to longer life-time and better safety of the SOFC’s (Bloom Energy, 2014; Irvine, Politova, & Kruth, 2005)."

Jan 1, 2016

By M. Kinnunen

Handling of municipal solid waste (MSW) and control of its potential resources is becoming increasingly important in modern society. The OECD (Organization for Economic Co-operation and Development) average on municipal waste produced in 2003 was that of around 570 kg per inhabitant. In European Union countries around half amount is of MSW landfilled. In Finland still 88% of MSW was landfilled in 2004. Since incineration reduces MSW volume significantly while producing heat energy, it has become one of the most common waste handling techniques. The shortcoming of incineration is however, that incinerated MSW is often not pre-sorted and it produces slags and ashes as side streams from which recovery of metals is difficult. This paper looks into possibilities of scrap metal recovery for recycling from municipal solid waste incineration (MSWI) bottom ash. In order to do that, a flow sheet of physical separation methods was developed and tested with Finnish rotary furnace MSWI bottom ash. In addition, an empirical model for estimation of the recoveries of scrap metals with physical separation methods for treatment of MSWI bottom ash was developed. According to the results of this study, for 1 Mt of incinerated MSW with 30% mass reduction, a minimum resource potential of 100 kt for magnetic metal scrap of roughly 24 % grade and about 19 kt of non-magnetic metal scrap of roughly 65 % grade. In other words, 57 kg magnetic scrap and 12 kg non-magnetic scrap per inhabitant at the mentioned grades, respectively. Although wide variations of parameters affect the MSWI bottom ash quality, it can be said that a considerable amount of metals is drained into MSWI bottom ash. Results of this study show that from MSWI bottom ash significant material streams metal scrap can be reintroduced into the resource cycle which would otherwise be lost.

Jan 1, 2014

By Frederic Benard

RECOVERY of selenium and tellurium at Copper Cliff by the Ontario Refining Co. has been previously described by the writer.1 During 1935 a new building was erected to house this operation and description of the new plant and process form the subject of this paper. The building, of modern steel and tile construction, covers an area of 5900 sq. ft. with a volume of approximately 165,000 cu. ft. The large monitored center bay is flanked by two smaller outside bays, one of which is devoted to the tellurium section. All equipment, ceilings and upper walls, are treated with aluminum paint, giving good illumination and facilitating good housekeeping, essential from a health standpoint in handling material of this nature. EQUIPMENT Layout and type of equipment conforms to the requirements of the process as described below. In the selenium section, storage tanks, neutralizing tanks and precipitators are of steel construction, lead-lined and. arranged for gravity flow (Fig. 1). The neutralizing tanks are equipped with water-cooled lead coils for controlling solution temperature, agitation being accomplished with compressed air. The precipitators are provided with mechanical agitators, and sulphur dioxide is introduced into the solution through 3-in. lead pipes leading from a manifold attached to the cover of the precipitator. Each precipitator has internal cooling coils and an air-jet exhauster to carry off fumes. Neutralized solutions are filtered in a 29-plate wooden filter press using heavy cotton duck as a filter medium, forced feed being maintained by a Welmet centrifugal pump. Dilute solutions are concentrated by evaporation in a vacuum evaporator, constructed of cast hard lead. Heat is supplied by an outside steam chest operating under 15 lb. steam pressure. The solution is circulated through the steam chest by a pump of antimonial lead at a rate approximating 400 cu. ft. per min. The evaporator is evacuated by a water-jet condenser that creates a vacuum of 26 in. of mercury under normal operating conditions. The capacity of the vessel is 60 cu. ft. and the rate of evaporation 25 cu. ft. per hour.

Jan 1, 1938

By C. V. G. K. Murty

Sillimanite, chemically named Aluminum silicate, 3Al2O3SiO2 is one of the three forms of alumino silicate polymorphs and commonly encountered in aluminous metamorphic, plutonic, and volcanic rocks and their weathering products. Most of the sillimanite deposits are granular and are of high grade available in coastal areas as associated mineral along with Titanium bearing minerals namely ilmenite, rutile etc. Sillimanite is one of the widely accepted raw materials for the manufacture of high alumina refractories (55-60 % alumina) or bricks. Sillimanite based refractory bricks are used for the lining of Blast furnaces, Arc furnaces, Soaking pits, Reheating furnaces of iron and steel, rotary kilns for cement manufacturing and general kilns for lime production. Considering the vast potential of unexploited sillimanite reserves in India, M/s Trimex Sands Pvt. Ltd., (TSPL)has set up a Mineral Sand Processing plant in 2010 at Andhra Pradesh with a capacity to produce about 50,000 tons of sillimanite per annum besides the production of various heavy minerals such as ilmenite, rutile, zircon and garnet. Traditionally in India, sillimanite is upgraded by froth flotation technique using Oleic acid as a collector cum frother, sodium silicate as depressant and soda ash as pH regulator. At TSPL, Conventional mechanical float cells are installed to float sillimanite. In addition to the usual problems associated with conventional cells, poor flow ability of the sillimanite concentrate is experienced in dry circuit due to the presence of sodium oleate. Fine garnet in the sillimanite concentrate contributes to high iron content in the final product, making sillimanite separation a nightmare for operators (lower recoveries coupled with inferior grades). Various efforts were made to improve the performance of the circuit such as introduction of additional frother like MIBC, optimizing the operating parameters such as pulp density, collector dosage, depressant dosage, air flow rate, pH etc., use of collector aids, stage wise addition of reagents etc. The challenges faced to produce a sillimanite product with stringent quality norms and the efforts made to improve the grade and recovery are described in this paper. Keywords: sillimanite, froth flotation, conventional mechanical cell

Sep 1, 2012

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

By C. Flores

It is well known that manganiferous silver ores containing manganese intimately associated with silver are not amenable to conventional extraction methods. The segregation process has been applied primarily for recovering copper from oxidized and oxide-sulfide ores. A new alternative process combining segregation roasting followed by flotation to recover silver from refractory manganiferous silver ores, was studied using a 23 factorial experimental design. Laboratory investigations have been conducted on a manganiferous silver ore analyzing 0.02% copper, 0.09% lead, 0.33% zinc, 12.55% manganese and 187 g/ton silver to study the effects of additives salt, coke and copper oxide on silver recovery. Results indicate that 79.3% silver recovery can be attained in the absence of copper oxide and 88.4% when the copper oxide content is 1.5% at '750°C, 2% salt and 1% coke and 30 minutes of roasting time. It was concluded that doping with copper oxide improves silver recovery.

Jan 1, 1994

By C. S. Ek

As a continuation of previous work, further tests have been carried out to recover silver from the residues obtained in a zinc jarosite process. Selective flotation has been performed with different reagents, and ?cleaning steps have been introduced to upgrade the silver concentrate. Leaching of crude residues, as well as flotation tailings, has been carried out with cyanide and thiourea. Finally, the different beneficiation methods are compared, and the industrial aspects? of their application are reviewed.

Jan 1, 1990

By W. Kunda

Silver nitrate solution is a very common product in hydrometallurgical processes involving the treatment of various silver bearing materials. Silver is recovered by electrolysis. The alternative method for the recovery of silver from aqueous solutions is the precipitation by hydrogen at elevated temperature and pressure. A previous study by the authors revealed the feasibility of this technique for the silver ammine sulfate system. Further study showed different behavior for silver in the silver ammine nitrate system and a systematic study was initiated to establish the optimum conditions for the process. The study showed that the process can be controlled by (1) composition of the feed solution, (2) temperature and (3) hydrogen pressure. The recovered silver product was in the form of very fine powder.

Jan 1, 1989

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