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By Nikhil Dhawan, Veeranjaneyulu Rayapudi

This study investigates upgrading of low-grade banded hematite quartzite iron ore (Fe ~31%). Conventional beneficiation was found to be futile. The susceptibility of iron phases to microwave exposure and their selective absorption assists in the liberation of iron values. Microwave exposure of coarse particles at 540 W for 10 min yielded a concentrate with Fe 56.30% and recovery of 50.68%. Conventional carbothermic reduction at 500 °C, 60 min and 9% charcoal yielded a concentrate with Fe 57.6% and iron recovery of 68%. The presence of sufficiently bonded silica leads to an easy formation of a fayalite phase. The microwave reduction design yielded FeG of 57.6%, FeR of 47% and a yield of 24% at an optimal condition of 540 W, 8 min and 6% charcoal. It was found that a small fraction of microwave-irradiated ore-charcoal mixture melted rapidly, and pure ferrite balls were observed within 8 min. An optical micrograph of a ferrite ball reveals the retained austenite and martensite phase.

By Nikhil Dhawan, Veeranjaneyulu Rayapudi

This study investigates upgrading of low-grade banded hematite quartzite iron ore (Fe ~31%). Conventional beneficiation was found to be futile. The susceptibility of iron phases to microwave exposure and their selective absorption assists in the liberation of iron values. Microwave exposure of coarse particles at 540 W for 10 min yielded a concentrate with Fe 56.30% and recovery of 50.68%. Conventional carbothermic reduction at 500 °C, 60 min and 9% charcoal yielded a concentrate with Fe 57.6% and iron recovery of 68%. The presence of sufficiently bonded silica leads to an easy formation of a fayalite phase. The microwave reduction design yielded FeG of 57.6%, FeR of 47% and a yield of 24% at an optimal condition of 540 W, 8 min and 6% charcoal. It was found that a small fraction of microwave-irradiated ore-charcoal mixture melted rapidly, and pure ferrite balls were observed within 8 min. An optical micrograph of a ferrite ball reveals the retained austenite and martensite phase.

By David R. Skuse, Jacek A. Kostuch, Steven M. Fortier

"Particulate ground calcium carbonate (GCC) is used as a filler and coating pigment in a wide range of applications including papers, paints, plastics and films. In this paper we will overview the requirements for GCC used in both filling and coating applications in the paper industry and highlight the processes used in the production of these materials.INTRODUCTIONCalcium carbonate is both abundant and very widely distributed occurring most commonly as limestone, marble and chalk (Harben 2002). Given, the high purity, whiteness and relative inertness of these materials it is not surprising that they are heavily utilised in a range of industrial applications. Worldwide ground calcium carbonate (GCC) capacity is estimated at 50.5 million tonnes per year (Roskill, 2002). This usage is divided between relatively coarse materials that are used in applications such as landscaping, cultured marble and carpet backing and finer grades that find application in plastics, rubbers, inks, paints and paper. Given the usage of calcium carbonate products across a very wide range of consumer products, the outlook for GCC is very promising with growth expected in many markets. Growth is anticipated across many applications in the emerging nations and more specifically in paper applications on a global basis.Over the last several decades the GCC processing industry has evolved to be well adapted to both the cost effective production and delivery of standard products and the production of high performance speciality products. Many of the standard unit operations found in the metaliferous raw material processing industry are also used, albeit in some cases in modified form, in the processing of GCC. Coarse products (diameter > 3 - 10 µm) are typically produced by dry processing routes. Whereas, the finer products are typically produced by a combination of wet and dry processing."

Jan 1, 2004

By T. Guler

Turkey has important celestite reserves in Sivas. These reserves have gypsum and/or calcite as gangue minerals. Celestite-calcite ore cannot be concentrated by gravity methods due to the low efficiency of gravity methods in fine sizes. Research was tried to concentrate celestite to a market grade by flotation. Parameters such as pH, particle size, dosage and types of collector and depressant were examined. Cyanamid A-845 and Na-oleate combination as collector, sodium silicate as depressant, and Aerofroth-65 as frother were found as the optimum flotation reagents. By applying 3-stage cleaning, a concentrate assaying 95,34 % SrS04 was obtained with 84,93 % recovery.

Jan 1, 1999

By Michael J. Mankosa

The increased emphasis placed on recycling of consumer and industrial waste products has generated a need for advanced separation techniques for recovery of valuable by-products. As a result, the metals recycling industry has made great strides in the recovery of used wire products through the application of chopping and screening processes. Unfortunately, in many instances, the final plastic reject still contains a substantial portion of valuable metal. Typically, this material has been discarded, since an appropriate separation technique has not been available. Recent work conducted by Carpco, Inc. in Jacksonville, Florida indicates that a high-grade metal and/or plastic product can be recovered from this waste material using a novel density separation process. Extensive test work has been conducted using this procedure to validate separation performance as well as equipment capacity. This paper will present the basic operating principle of the separator, as well as results from tests conducted using various feed materials.

Jan 1, 1995

By J. M. Figueiredo, A. R. Silva, M. C. Coelho, R. M. Guedes

"With the continuous depletion of conventional sources of non ferrous metals, low grade complex sulphide deposits have been regarded as an alternative source of supply of zinc, copper, lead and also silver and gold. In spite of very important efforts in research in this domain it is still a long way to go in order to achieve reliable technological solutions both in ore dressing and metallurgical treatments which enable these complex materials to be regarded as commodities. The RLE process to treat bulk concentrates such as those obtained from portuguese ores, under development at LNETI and QUIMIGAL premises, combines experienced unit operations and innovative configurations and is an effective contribution for the future use of these materials. The results of three years of experimental work and main conclusions on the overall feasibility of the process are presented.IntroductionIn spite of the very well-known problems associated to the development of complex sulphide deposits, the economical importance of the enormous reserves of these materials, widespread all over the world, is so evident that its future use as source of supply of basic and precious metals and other minor elements is almost unavoidable.Critical steps in that direction are mainly concerned with favourable economical conditions and sound metallurgical practice for the treatment of nonferrous metals concentrates, which are nowadays still considered as non-conventional raw materials.Since Portugal is one of the countries where these orebodies occur more extensively, Portuguese authorities have dedicated since the beginning of the seventies, a big effort on finding economical and reliable processes for the recovery of copper, lead, zinc and also gold and silver, and for this purpose, several alternatives on ore dressing and metallurgical processing were studied, however without full sucess.Five years ago LNETI, a research laboratory, and QUIMIGAL, a chemical company, both state owned, decided to launch a research programme for the development of a metallurgical process for the treatment of NFM concentrates, which received financial support from NATO under its programme ""Science for Stability""."

Jan 1, 1988

By Paul M. Jr. Musgrove, Donald C. Moore

Conventional Smelting Practice Conventional copper smelting practice varies from smelter to smel¬ter, but generally consists of some or all of the following unit processes: roasting, smelting, converting, and fire refining. Roasting. Copper sulfide concentrates can be smelted directly or after an initial roasting step. Roasting is used in some smelters because roasting prior to smelting increases smelting capacity, less energy is required to melt hot roaster calcines than wet sulfide concen¬trates, roaster off gases are high in Sot concentration, 5-15% SO2, and some volatile impurities are removed from the concentrate prior to smelting. However, many smelters do not use roasters, because the problems associated with handling hot dry calcines outweigh the advantages mentioned. Concentrate roasting is performed in multiple hearth or fluid bed roasters. If the moisture is low, roasting can be performed autogenously, usually at 500-600°C. High roasting temperatures are avoided because excess oxidation of the iron compounds may lead to magnetite formation. Magnetite is detrimental to reverb operation because mag¬netite can combine with refractory minerals to form a highly viscous slag. This slag prohibits efficient matte-slag separation and leads to excessive copper losses. Also, magnetite can settle through the matte layer, deposit on the furnace bottom, and consequently reduce furnace capacity. Roasting is carried out only on sulfide concentrates prior to smelt¬ing in reverb or electric furnaces. For smelting processes, such as the flash and continuous that rely on the exothermic heat of oxidation of the sulfur minerals, roasting is not practiced. Reverberatory Smelting. The predominate copper smelting fur¬nace for the past 50 years has been the reverb. These furnaces are typically 100-120 ft long, 30-35 ft wide, and 12-15 ft high. A typical furnace layout is shown in Fig. 2. Refractory brick linings cover all internal surfaces of the furnace. Originally the flame was directed to reverberate or reflect off the furnace ceiling and melt the feed material. Current practice is to direct the flame down the furnace length to melt the concentrate. A method of charging the concentrates or calcines, generally along the side walls to minimize refractory erosion, is incorporated in the furnace design. The copper concentrates, calcines, and fluxes charged into the reverb undergo a series of complicated reactions as the temperature of the mixture increases. The reaction of the iron and copper sulfides with the oxygen in the furnace produces a molten Cu25-FeS mixture called matte. Copper smelting metallurgy is based on the fact that sulfur has a greater affinity for copper than for iron and most other common metals. Therefore, in a system containing copper, the copper will preferentially remain as a sulfide compound until all of the other metals have been oxidized. The oxidized metals combine with silica to form a silicate slag that floats on the matte and is removed from the system. Reverberatory furnace smelting chemistry can be approximated by the following chemical equations: FeS2 + O2 - FeS+ SO2 (1) The formation of FeS ensures that any copper present other than as sulfides will be reduced by the relationship: CuO2 + 2FeS + O2 - CuS + 2FeO+SO2 (2) or 2Cu +FeS - Cu2S + Fe (3) As the molten charge travels down the furnace, continued oxida¬tion of the iron minerals and sulfurization of the copper minerals occurs. When all of the copper has been converted to sulfides, the iron sulfides can then be further oxidized as: FeS + (3)2 O2 FeO + SiO2 (4) The FeO reacts with the silica added as flux in the furnace charge. A simplified equation is: FeO + SiO2 -FeO SiO2 (5) The iron silicate slag formed is skimmed from the surface at the end opposite the burners. The copper content of reverb slag is usually less than 0.6% Cu and is discarded. Matte is removed along the side wall and is taken to the converter for oxidation of the remaining sulfur and iron. The main objectives in reverberatory smelting are to produce a molten Cu2S-FeS matte containing 30-60% Cu and a throwaway slag. Production of matte permits complete conversion of all copper minerals into copper sulfides, which can migrate because of specific gravity differences, through the lighter slag layer. Also, the molten matte droplets collect the noble metals, gold and silver, as the matte settles in the furnace. The large settling area of the furnace provides enough separation time to produce a low grade slag, which can be discarded without further processing. High heat losses are associated with reverberatory smelting be¬cause of the large volume of gases sweeping through the furnace. Therefore, an outside source of heat is required to keep the smelting reaction going. Natural gas, fuel oil, or pulverized coal are used as this heat source.

Jan 1, 1985

By V. F. Kozyrev, L. B. Tsymbulov, L. Sh. Tsemekhman, A. N. Golov

Gipronickel Institute, JS has developed a new process and equipment for melting copper raw material in a two-zone Vaniukov furnace resulting in the production of white matte and low-grade slag in terms of copper content. Tests on processing copper nickelcontaining ore from Norilsk (Cu - 17 %), Ni - 3 %) were conducted in Vaniukov pilot furnace (total area 11.4m2) of Severonickel Combine. The amount of the processed material was about 11 thousand tons. Principal results of pilot testing are presented.

Jan 1, 2007

By K. Takahashi, G. Dodbiba, T. Fujita, N. Sato, S. Matsuo, K. Okaya

Different materials are used for manufacturing various parts of a cellular phone. Considering that a great number of mobile phones are being discarded every year, it is important to recover these materials and then recycle them in order to save resources. The primary objective of this work is therefore to find and suggest ways on how to recover various materials from discarded cellular phones for recycling. Much effort, for example, has been put in recovering and recycling tin oxide (ITO), a mixture of indium (III) oxide, (UI2O3) and tin(IV) oxide from the discarded liquid crystal displays (LCDs). Here we report a simple and cost-effective process, which includes a HC1 treatment process for altering the structure of the indium (III) compound into a chlorideinduced indium compound, in order to facilitate the vaporization of indium at a relatively low temperature. The experimental results indicated that 96.2 % of indium can be recovered from ITO, when the sample was heated at 973 K for 90 minutes. In addition, some thermodynamic calculations were also carried out in order to better understand the mechanism for chlorideinduced vaporization of indium.

Jan 1, 2011

By A. M. Muhamedilova, S. M. Narbekova, L. V. Semushkina, D. K. Turysbekov, N. K. Tussupbayev

The research and development of new, more selective collectors in order to improve flotation efficiency represent one of the main priorities in the development of innovative technologies pertaining to flotation separation of various substances and minerals. The use of collectors’ mixtures is considered as the universal instrument for ensuring necessary proportion of chemically and physically sorbed collecting reagents in every individual case. Currently hard-dressed, refractory ores and man-made raw materials, which are characterized by low content of valuable components, fine dissemination of mineral assemblages and similar technological properties of constituent minerals, are widely involved in the recycling processes. The possibility of processing tailings, which are left after flotation beneficiation, has been considered in the case study of tailings of the Zhezkazgan beneficiation plant and the Tishinsk copper ore deposit with utilization of the modified multifunctional flotation reagents. Modified poly-functional collector is the mixture of composite Aerofloat, TC-1000 and butyl xanthate. Reactants ratio is 1:1:3. The advantage of the suggested flotation reagents is that they are composed of two polar groups and have a long hydrocarbon radical. This structure, being in the water during flotation process, plays a dual role: first, as a collector adsorbing on the mineral surface it produces metal complexes with polar groups in the form of bridges; second, apolar radicals flocculate valuable slimed components, thereby intensifying flotation process. It has been shown that flotation of Zhezkazgan tailings with utilization of the lesser modified reagent amount compared to butyl xanthate has enabled to produce crude copper concentrate with 13.0 % copper grade and 80.22 %recovery. Compared to the baseline technology, copper content of crude concentrate has been increased by 5.1 % and the recovery by 31.4 %. It is shown that tailings flotation of the Tishinsky field by applying modified collector allows for increase of the of useful components extraction in a collective concentrate with: copper 9.63 %, lead for 8.41%, zinc 9.2 %, iron 2.73 %, gold 3.57 %.

Jan 1, 2018

By Nikhil Dhawan, Débora Moreira, Yogendra Pratap, Himanshu Tanvar

A glauconitic siltstone sample with a potash content of 10% is investigated as a potential potash source. The fine grain structure comprises potash bearing phases of glauconite, muscovite, and microcline. The as-received samples showed poor dissolution ~ 5%. Planetary ball milling was attempted for the dissociation of the underlying silicate structure for the potash release, and ~ 60% potash was obtained after 12-h milling. Mechanical activation (8 h, 60% NaOH) prior to heat treatment (600 °C, 30 min) was found successful in recovering more than 95%K values from glauconite. The final residue consists mainly of quartz, indicating a complete breakdown of the silicate matrix. The mechanical activation process followed is simple, adaptable, and less energyintensive compared to existing methods.

May 4, 2020

By E. I. Gedgagov, A. V. Tarasov, V. M. Paretsky, A. D. Besser

"Hydrometallurgical technology and equipment have been developed in the Gintsvetmet Institute (Moscow, Russia) for processing of rhenium-containing alloys waste to produce commercial- grade metal products with recovery of all valuable components. The technology is based on waste treatment with hydrofluoric acid with subsequent decomposition with nitric acid at pH = 1 to 1.5 to ensure maximum degree of rhenium conversion to the liquid phase of pulp. Some experimental lots of ammonium rhenate and ammonium molybdate were produced at a pilot plant. The following main techno-economic performance values have been achieved when processing grinding residues (balance-based tests were conducted with a lot of SP-2 alloy waste) to produce ammonium rhenate of high purity, as well as tantalum-niobium, nickel-cobalt and tungsten-molybdenum concentrates: rhenium recovery of 92.3% into a commercial product from grinding waste; production cost of 170-210 US Dollars per 1 kg of rhenium in commercial-grade ammonium perrhenate taking into account its recovery; tantalum and niobium recovery of 75- 77% into bulk concentrate; nickel and cobalt recovery of 88-89% into a mixed concentrate; tungsten and molybdenum recovery of 88% to 91% into concentrate. Introduction During the recent years the use of rhenium has been increasing at a fast pace worldwide [1]. The growing prices for this refractory metal are an essential factor encouraging collection and processing of rhenium-containing wastes. Double-component waste, such as molybdenum- rhenium or tungsten-rhenium alloys can be readily decomposed and the process of rhenium recovery does not cause, therefore, any difficulties. The main separation of metals takes place at the stage of waste digestion with acid, because molybdenum and especially tungsten remain in precipitate in the form of hydrated oxides and rhenium reports in the liquid phase, from which it can be easily extracted by solid ion-exchange or solvent extraction techniques. However, most of rhenium is contained in super-alloy waste [2]. Such waste has different contents of valuable components, such as tantalum, niobium, molybdenum, tungsten, cobalt, nickel, chromium and other metals. The basis of all such super-alloy waste is nickel. Refractory metals in the super- alloy composition are in the form of metallized target-oriented eutectics, and due to this reason hydrometallurgical decomposition of such alloys is a difficult technological task."

Jan 1, 2008

By Taylor R. K A, Berglin C. I K

The manganese ore deposits at Groote Eylandt are quite variable both in regard to grade and physical cond- ition. In order to treat a wider range of these ores and to increase production a new primary-crusher and' concentrator complex costing $8 million has been installed and commissioned to treat ore at the rate of 12,000 tonnes per day. The initial testwork carried out to determine the method of the ore treatment required, and on which the design of the concentrator was based, is detailed. The flowsheet and equipment used is discussed together with the operation of the plant. The production achieved and product qualities obtained are given.

Jan 1, 1973

Processing of Hard-to-Treat Copper Ore and Flotation Middlings Using Chemical Treatment

Sep 13, 2010

By M. Doucet

The introduction of high-nickel anodes to CCR in 2006 resulted in an increase of nickel in anodic slimes. Removal of the nickel in these slimes to below 3.0% is necessary in order to prevent slag viscosity problems in the subsequent pyrometallurgical step. A lower specification of 0.5% was chosen in order to minimize Ni losses to the slag and to recover this nickel as NiSO4. A 2-stage autoclave process, which was patented by CCR in 1992, was chosen to treat the high nickel slimes. Laboratory and plant?scale tests were performed in 2004 in order to validate the results of this process with the present slimes. These results along with corrosion tests were used to design two brick-lined autoclaves. These autoclaves were installed in October 2006 and successfully commissioned within 8 weeks. A "Design of Experiment" test program was performed in 2006 on an alternative autoclave process that reversed the cycles in the patented process. The effects of temperature and acid content were modeled using the results. The alternative process will be validated with the new autoclaves.

Jan 1, 2007

By G. V. S. Murthy

Beach Placer Ilmenite is an important source for production of titanium metal, titania slag and pigment grade titanium dioxide. Besides, these well known applications for which Ilmenite is mined and processed, there are other emerging processes that are being tried for obtaining highvalue products. Synthesis of Ilmenite based materials for varistor applications and direct electrochemical reduction of Ilmenite to produce ferrotitanium are two such attempts being discussed in this work. In this paper, recent efforts undertaken to study the electrical and magnetic characteristics are discussed. Ilmenite, FeTiO3, is one of the mixed-valence transition metalbearing minerals, inwhich Fe can be in two different oxidation states, Fe2+ and Fe3+.. Similarly, Ti can be in Ti3+ and Ti4+.Ilmenite is inherently suitable for making Varistors, which are devices, used for limiting the transient voltage surges in a circuit. Varistors produced from Ilmenite can withstand harsh environments seen in nuclear reactors and outerspace. Electrochemical reduction of Ilmenite can be used to prepare ferrotitanium directly without any reductant. The process essentially involves removal of oxygen from the mineral through electrolytic action using calcium chloride as electrolyte and graphite as anode. At a temperature of 950°C, it is shown to be possible that all the oxygen can be removed from the Ilmenite sample, which passes through the electrolyte, to form CO/CO2 at the anode. The process has immense potential for cost effective production of titanium metal as well. Results of the above developmental works are presented in this paper. Keywords: beach placer ileminite, varistor applications

Sep 1, 2012

By P. Bandopadhyay, T. C. Rao

Geologically the Indian coals belong to Gondwana formations. Being of drift origin, these coals contain high percentage of mineral matter distributed in finely disseminated form. Indian coal-preparation plants process prime coking and medium coking coals for use in steel plants. Limited resources of good quality coking coals (Table 1) could not keep pace with the rapid developments of the iron and steel industry in India. This resulted in the introduction of coal-preparation technology and the establishment of washeries in India in the last 30 years or so. The first four washeries were set up during the early fifties at West-Bokaro, Jamadoba, Lodna and Kargali. Next came the first non-coking coal washery at Nowrozabad set up by the Associated Cement Company. In 1961, following the recommendation of Central Coal Washeries Committee (1953-54), Hindustan Steel Ltd. had set up five central washeries at appropriate locations, viz Durgapur, Dugda (2 washeries), Patherdih and Bhojudih, so that the raw coal feed having similar washability characteristics can conveniently come from a number of collieries to a particular washery. Subsequently, to meet the increasing demand of metallurgical coal by steel plants, two more central washeries (at Durgapur and Chasnala) and three pit-head washeries (at Kathara, Swang and Gidi) were installed by 1970. The prime objective of washing Indian coking coal is essentially to reduce the ash content. All the washeries in India, taken together have an installed capacity of about 26 million tons of raw coal input per annum. Details of these washeries in respect of their annual capacity, main washing systems adopted and the nature and size of the products are given in Table 2. It can be observed from this table that (1) bulk of the coal is processed in units like jigs HM baths and HM cyclones; (2) only two of the washeries have cleaning circuits for treatment of fines.

Jan 1, 1980

By O Ostrovski

Iron ore reduction and iron cementation by CH4-H2-Ar gas was investigated in a laboratory isothermal fixed bed reactor. Iron ore was first reduced to metallic iron by hydrogen, then carburised to cementite by methane. Increasing temperature and hydrogen content accelerated the reduction process. Methane had almost no effect on the reduction process. Meanwhile, higher temperatures increased the rate of cementation and also the rate of free carbon deposition. The optimal temperature for cementite formation was found to be 750¦C. At this temperature, reduction of iron ore to cementite by gas containing 55 vol per cent hydrogen, 35 vol per cent methane and 10 vol per cent argon was completed in about 15 minutes. Addition of sulphur into the reducing/carburising gas strongly retarded the process of cementite formation and free carbon deposition. Cementite is an unstable phase which decomposes to iron and carbon. The rate of cementite decomposition increased with temperature up to 600¦C and above 750¦C . In the temperature range from 600¦C to 750¦C the rate of cementite decomposition decreased with temperature. Cementite was the most stable at 750¦C and least stable at 600¦C. The two-interface grain model was used to describe the process of iron ore reduction and the volume reaction model was employed to model the iron cementation process. The simulated results for both reduction and cementation are consistent with the experimental data.

Jan 1, 2000

By A B. Christie

Space-based synthetic aperture radar (SAR) data is becoming available over much of the world, includingNew Zealand. Radar imagery has a number of very advantageous features, including all-weather capability; it also has artifacts, such as layover, which require processing before these data may be used to best effect.   This paper gives an outline of SAR, details the image processing required, and discusses the use of the first New Zealand Japan Earth Resources-1 SAR data to aid geological mapping on the Coromandel Peninsula.

Jan 1, 1994

By R. M. Whyte

The Konkola Deep Mining Project (KDMP) seeks to exploit the deep level potential of the Konkola Deeps orebody, producing high grade copper concentrates for treatment at the Nkana Smelter and Refinery. As an alternative to smelting, Konkola Copper Mines (KCM) has examined a processing route involving pressure leaching of the Konkola concentrates at Nchanga, with the leach liquor being treated by solvent extraction and electrowinning to produce high purity copper cathode. The solvent extraction circuit would be operated to achieve a high ?delta? copper, thereby providing a strong acid raffinate which can be used in the existing Tailings Leach Plant (TLP) to treat Nchanga oxide copper materials. The main focus of the hydrometallurgical option is to capitalize upon the synergy that exists between the treatment of sulphide copper concentrates (acid producing) and oxide copper ores (acid consuming). The paper discusses the key features of a comprehensive piloting campaign carried out at Anglo American Research Laboratories, which involved the co-treatment of Konkola copper concentrates and Chingola refractory ore in a fully integrated pilot plant circuit.

Jan 1, 2001