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By David G. Almond

"Economies of scale and the variable characteristics of many large ore deposits have resulted in grinding installations having large SAG mills fitted with variable speed drives. Problems associated with effective operation of this equipment has created the need for better operator knowledge of mill load behavior and the need for new instrumentation capable of providing accurate information about load movement and composition. This paper discusses two performance enhancing instruments available to address these needs.INTRODUCTIONThe economic demands of new mining projects has resulted in increased equipment sizes and this is particularly evident in milling where there are now many SAG mills operating in the range of 28’ – 40’ diameter with installed powers of 8 – 22 MW. These mills are designed to process very high capacities of ore daily2 and the variable characteristics of many of the deposits mined results in the need to vary the mill speed during processing to cater for the changes in ore hardness. Mill speed changes often result in operational problems associated with high-energy impacts striking the mill liners reducing energy efficiency as well as degrading the grinding media through breakage, accelerating liner wear and in extreme cases even breaking liners. These operational problems have given rise to the need to have instruments aimed at providing operators with accurate information in real time about the load composition and movement within SAG mills.The original concepts for operating SAG mills were derived from past experience of operation of conventional fixed speed ball mills. However, the relatively stable conditions that applies with fixed speed and constant ball charge in ball mills supports theoretical approaches to the grinding process based on energy calculations such as Bond’s Work Index concept. Later a set of parameters Si and Bij were used for describing load kinetics, but all of these methods neglect the details of load movement and how the energy is actually used inside of the mill. It is now commonly accepted as a basic rule that SAG mills grind by impact whilst ball mills grind by attrition and consequently SAG mills are designed and operated to generate a large amount of impacts by the cataracting load without paying sufficient attention to where these impacts are actually occurring. Many mills are operated on the basis that as long as the mill is consuming the expected power there is little attention given to the class of bodies using this energy and this often results in the inefficient use of energy."

Jan 1, 2005

By E. C. Morton

The Western Australian School of Mines (WASM) has developed a facility for the static testing of large scale ground support elements. The samples are loaded using a mechanical screw feed jack. Instruments measure the displacement and the load being applied. Methods have been developed for testing of mesh, shotcrete and membranes. Mesh samples are restrained on all sides and loaded centrally. A new large scale punch test method has been developed to evaluate shotcrete and membranes. These products are sprayed on to a sandstone substrate containing a centrally located isolated disc. The sprayed product is loaded by displacing the disk. The different test methods are described in detail. The test program involved two types of mesh, several shotcrete mix designs with different curing times and one membrane product. The force-displacement responses are used to directly compare the products. The implications for implementation of the different ground support products are discussed.

May 1, 2009

By D. Bois, P. Keating, J. Lemieux, M. A. Vallée, M. Allard, R. S. Smith, L. Z. Cheng, M. Chouteau, K. Fountain

"A comparison of data from INPUT, GEOTEM, and MEGATEM airborne electromagnetic (AEM) systems over the Aldermac deposit, Québec, Canada, was undertaken to evaluate the relative ef¬ficiency of these systems for the detection of small, deep conductors. The Aldermac orebody was mined between 1933 and 1934 to a depth of 200 m. A deeper mineralized zone was discovered in 1987. The MEGATEMII system was used to acquire 90- and 30-Hz data over the deposit in 2003 as part of a research project. These MEGATEMII data were compared with data acquired by other AEM systems (GEOTEM and INPUT), and the results show that the MEGATEMII response is ten times larger than the other sys¬tems. For the dB/dt response, 30 Hz gives a better signal-to-noise ratio, whereas 90 Hz has a better signal-to-noise ratio on the B-field.The data were evaluated by numerical modeling. The deposit was best represented by a complex mod¬el, comprising two weakly conductive bodies to approximate the alteration halo, two deeper prisms repre¬senting known orebodies, and three bodies ranging in thickness from 3 to 6 m and in resistivity from 12 to 20 ohm·m, representing the chargeable overburden. This implies that the MEGATEMII system is able to see a small conductive body at more than 200 m depth in an area where the response is contaminated by conductive overburden."

Jan 1, 2009

By Randy Clarkson

"Free gold particles occur in many base metal and precious metal mines throughout the world. Occasionally mine operators are unaware of the presence of free gold particles in their ores because they have not seen free gold hand specimens or at the working face of the mine.Due to the high density and malleability of gold, large circulating loads of free gold particles can build up over time in grinding circuits. This gold is often deposited in unintended traps throughout the mineral processing system.These unidentified free gold deposits represent a serious security risk as well as inaccessible inventory of gold. Deposits of free gold particles often remain undetected by mine management until they are accidentally discovered while repairing mill equipment. The trapping and subsequent release of gold particles can frustrate attempts to calculate accurate metallurgical balances.Nuclear tracers have been used to determine the free gold recovery efficiency of gravity recovery equipment, drills, and sampling equipment (Clarkson, 1994, 1997, 1998, 1999 and Walsh, 1986, 1992). In 1995, the author used radiotracers to locate and eliminate gold traps at La Mine Doyon (Clarkson, 1997). In 1998 and 2001, the author also salted mineral processing circuits at Royal Oak's Pamour Mine in Timmins (Clarkson, 1999) and at Gold Corp’s Red Lake Mine in Ontario with radioactive gold particles.The gamma and X-ray radiation from gold's nuclear isotope (Au198) were detected through steel, in pipes, and in moving equipment with scintillometers. The radiotracers proved to be a safe, rapid and accurate method of identifying gold traps, improving security, determining the recovery efficiency of Knelson gold concentrators and estimating retention times of gold particles in various processes."

Jan 1, 2003

By J. A. McLennan

Nickel laterite deposits are typically formed from tropically weathered mafic-ultramafic complexes. The resulting nickel concentration is found within soil horizons and is mineable with regular dozer units. At the time of mining, digging elevations are needed at selective mining units (SMUs). Moreover, for mine planning, the costs due to dilution and lost ore require mining elevations to be specified. A conditional simulation approach is used to calculate ?optimum? SMU mining elevations. Multiple realizations of the ore base surface are simulated. For a single bottom surface realization, dilution and lost ore costs are calculated over the entire vertical extent of each SMU. The optimum mining elevation is the elevation where the lowest summed cost of ?accepting dilution? and ?wasting ore? occurs. The reported SMU dilution is the cost of mining the volume of waste located below the bottom surface and above the optimum mining elevation and the reported SMU lost ore is the cost of not mining the volume of ore above the bottom surface and below the optimum mining elevation. For all ore bottom surface realizations, optimum SMU mining elevations and their associated dilution and lost ore costs are calculated and compiled to represent production uncertainty.

Apr 1, 2005

By Ian D. Pirie

Presentation outline ?Project location ?Geology ?Resources and reserves ?Results of the FEED study ?Social and environmental considerations ?Project path forward

May 1, 2010

By S. Yalcin, S. Chang, M. Oliazadeh

"Present day ferrochrome demand is on the rise, as more steel products require this material as an alloying element. It is projected that the stainless steel demands will grow by about 6% per year to 2020; as such, the supply and demand for ferrochrome appears promising. Chromite is the only economical source of chromium, which is commonly upgraded by gravity techniques and used to produce ferrochrome. The gravity separation methods are not able to treat particles finer than 0.1 mm efficiently. Therefore, in processing chromite ore, a large quantity of fine tailings; some 25% of the mineral values have been reported as lost in slimes during processing. The chromite content of these tailings is typically 10-20% Cr2O3 and accumulates into a large stockpile/pond. A chromite concentrator project in Canada, considers a cut-off grade of 20% Cr2O3. This indicates further processing might be required to recover chromite from low-grade stockpile when the high-grade deposits are depleted. Low-grade ores normally indicates disseminated chromite ores with low liberation and consequently fine comminution. This paper reviews the techniques for treating chromite from fine material and provides some examples of successful operations in fine chromite recovery. Beneficiation techniques for fine particles include gravity separation, magnetic separation, and flotation. Several case studies are presented here that used one of the aforementioned methods or a combination of them.INTRODUCTIONChromium is one of the essential element and important material for production of ferrochrome alloys and special steel. Chromium occurs as chromium spinel, a complex mineral with chemical composition of (Fe, Mg) O (Cr, Al, Fe)2 . Chromite ore is exclusively found in ultramafic igneous rocks. Major associated gangue minerals are talc, quartz, serpentine and iron oxides.Chromite is a heavy mineral, with a specific gravity in the range of 4-5, relative to associated gangue minerals in a chromite ore, which makes gravity concentration a suitable process. This is generally true for coarse particles, ranging from 1 mm to 100 mm. Typically, the ore is prepared by crushing (and grinding if necessary) to a suitable size and the material is sized for the appropriate beneficiation techniques. It is common to use heavy medium separation (e.g. dense media drums, dense media cyclones, or Dyna Whirlpool) to treat coarse size and intermediate size particles (1 – 100 mm). Intermediate and fine particles (20 – 0.1 mm) could be treated by jigs. Spirals, shaking table, sluices, and Reichert cones can upgrade fine particles. Figure 1 shows a typical chromite beneficiation process flowsheet. As can be seen, it consists of three major sections: feed preparation (crushing, screening), de-sliming, and concentration. Beneficiation recoveries range from 80 to 85%, although the ratio of upgrading is only 2 to 3."

Jan 1, 2013

By Y. Özçelik

Rocks show a variety of engineering and physical properties that affect typical quarry operation such as diamond wire cutting in the marble quarry. The physical and mechanical properties of rocks depend on their mineralogical and petrographical characteristics. The purpose of this study is to investigate the effects of mineralogical and petrographical properties of a rock, such as grain size and mineral content, and how these affect the wearing of diamond bead and cutting rate in a diamond wire cutting operation in a marble quarry. Statistical analysis is used to carry out the study. Three different rock types were examined, namely, hard rock (andesite) and softer rocks like limestone and marble. The study revealed that the mineralogical and petrographical properties of the rocks, such as percentage of quartz, biotite, plagioclase, matrix minerals, and opaque minerals, are important. In addition, the grain size of quartz and plagioclase in andesitic rocks and calcite grain size for carbonate rocks significantly influenced the wear rate and cutting rate in diamond wire cutting operations.

Jan 1, 2005

By Jim Wickens, Vince Lofftus, Mark Kravetz

"This paper discusses how Bald Mountain Gold Mine (Ely, NV) dramatically reduced mercury production by treating process solutions with a liquid reagent based on organic sulfide chemistry. Cherokee Chemical Co. Inc. (Vernon, CA) markets this reagent as UNR 81 1A.A summary of the bench tests comparing various reagents is given. The ensuing plant trial is discussed and successful mercury removal results are presented. Data is presented to show that early removal of mercury in the process reduces the amount of mercury collected at retort. Data is also presented to show that use of this reagent does not reduce gold recovery. Plant trial treatment costs are discussed. Finally, the potential of this technology to substantially reduce closure costs as compared to traditional mercury removal options is discussed.INTRODUCTIONBald Mountain Mine produces gold. Mercury is a costly by-product of gold production because no revenue is generated from mercury. Traditionally, mercury leaches out with gold and follows the gold throughout the processing circuit until it is eventually condensed and collected at Retort. Minimizing or eliminating mercury production minimizes worker health and safety risks, minimizes environmental liability, and reduces costs. In the long term, minimizing mercury production will make mine closure easier and less costly.Bald Mountain Mine dramatically reduced mercury production by treating process solutions with liquid reagents called UNR 811 and UNR 811A. Several reagents were bench tested before proceeding with a plant trial. The reagent complexes the mercury to form an insoluble organic sulfide precipitate. In independent tests, the precipitate passed acid mobilization tests and is considered stable."

Jan 1, 2001

By Timo Niitti, Antti Rinne, Janne Turunen, Sami Grönstrand

"This paper describes novel methods to enhance the flow dynamics of a mechanical flotation cell that can be applied to an existing installation. There are several reasons to alter the flow dynamics. The most common motivation is to improve the metallurgical performance of the cell, for instance, to improve the flotation recovery in particle size ranges that are normally difficult to recover.The flow dynamics enhancements are associated with the geometry of the rotor-stator area and additional features in the tank. The rotor-stator area has been modified to provide more concentrated turbulent kinetic energy or respectively more pumping efficiency with less attrition to the slurry. The tank can be fitted with various geometrical features, aimed to create a desired particle suspension profile. Depending on the targeted particle size range and application, badly selected or designed features can also have a negative effect on the metallurgical results.Besides new flotation equipment, the enhancements can be applied to existing flotation cell installations, thus enabling improvements in the mixing profiles, solids suspension, tank sanding and metallurgical performance of operating flotation circuits. Results of several tests carried out in existing plant operations are presented, concerning both flow variables and metallurgy.The common factor to above-mentioned enhancements is the use of modern computational fluid dynamics tools in the design phase, as well as the validation of the obtained CFD results in laboratory and production scale. Validated variables include mixing profile, power draw and local air hold-up.The paper demonstrates the effects of various flow geometries on the mixing profiles and metallurgical results of existing mechanical flotation cells. It shows how modern computational fluid dynamics analysis can be used to screen and select different approaches before carrying out costly trial-and-error experiments in plant scale."

Jan 1, 2006

By John Frostiak, Lloyd Buckingham, Kenneth G. Thomas, Luis Wilson

"The El Indio roasting facility associated with El Indio Mine in Central Chile has operated since 1981. It is a facility developed to remove arsenic and antimony from gold bearing enargite/pyrite flotation concentrates.The three installed roasters are Nichols Herreschoff units each with 14 hearths and measuring 4.85 metres diameter. Hearths are rabbled by central shaft driven rabble arms. Product is progressively heated to 750°C under neutral to reducing atmospheric conditions to drive off labile sulphur, arsenic and antimony.Ancillary equipment includes skip hoist feed arrangements, water cooled screw conveyors for transporting and cooling calcine, gas cyclones, post combustion chambers, heat exchangers, electrostatic precipitators, bag houses, gas scrubbers and packaging facilities .Operating practices, history, mineralogy and results are reviewed. The improvements to arsenic and antimony extraction, ambiental conditions and environmental achievements are discussed.INTRODUCTIONHistoryThe El Indio property lies in the El Indio belt, a copper, gold and silver area approximately 175 km long by 10 km wide. The property is located in the Andes Mountains of the central part of Chile at latitude 30"" South and 160 km from the sea-side city of La Serena. Plant location is at 3850 meters above sea level.The property currently consists of two underground mines feeding the El Indio plant and an open pit feeding the newer Tambo leach plant. The property has been controlled by a succession of owners (St. Joe Minerals, Fluor, Bond Gold International and Lac Minerals Ltd.) before acquisition by Barrick Gold Corporation in 1994."

Jan 1, 1999

By W. J. Tough

THIS paper offers examples of some of the typical calculations which occur in the ventilation of metal mines. The fi.rst section of the paper discusses the problems associated with natural-draft pressure; the second section presents the more common of the calculations concerned with the operation of mine fans; the third section illustrates the integration of the individual problems covered by the preceding sections into the larger problem of determining the operating characteristics of a fan for a particular mine duty. There is a very definite limit to the extent to which pure arithmetic can be successfully applied to the solution of min<; ventilation problems. As mines advance from the exploration, development, and initial production stage to the rime when underground workings become exceedingly complex, there is a considerable increase in the relative importance o theory and judgment in the solution of ventilation problems. These calculations illustrate the application of mathematics to ventilation, but they also in-elude examples of instances where arithmetic cannot be conveniently applied, and important estimates, based on judgment compounded of experience and a knowledge of the theoretical consideration involved, must be made. The inclusion of such estimates is intended as a guide and a warning to the reader; a guide in assisting him to recognize and estimate his own incalculables; and a warning that the solutions must not be blindly applied to his apparently similar problems. These sample calculations are intended to assist the reader in applying the theory contained in the standard texts on metal-mine ventilation. They have been selected from the ventilation experience of a British Columbia mine.

Jan 1, 1942

By The Staff

FOR forty years cyanide men have been intensively engaged on the treatment of sulpho-telluride ores. So far as we know, the processes used today were-with the important exception of the concentration plus high-lime process at the Victor mill (1915)-in use, or had been tried, in the earl y years of the century. Roasting, bromocyanide, ordinary cyanide, table and vanner concentration, floating and treating the concentrate, roasting con-centrates prior to treatment, etc., all were used at Kalgoorlie in 1904 or earlier. Flotation, followed by shipment of the concentrate to the Golden Cycle mill, was the practice at the Vindicator mine, Cripple Creek, for some years prior to 1917. Similar practice was employed in the Portland Independence mill at Victor, Colorado, and was lacer superseded by the table concentration, high-lime cyanide treatment of rails of the older Portland Victor mill. During all this time, an immense amount of research was being clone on these ores by the staffs of numberless companies. Conclusions were drawn and mills built. But if the results of one per cent of this research were ever published in sufficient detail for intelligent use by others, the Lake Shore staff has been unable to discover them, or does not recall having clone so. Government laboratories-as for instance chose of the Mines Branch, Ottawa, in or about 1927-have touched upon certain methods in their publications, but they simply have not had the time to finish the job. The paper by W. E. Johnston, Tellurides-Problem or Alibi? (E. & M.]., Aug., 1933), aroused wide interest and comment. It published conclusions, some right and some wrong, and gave the evidence adduced by the author i:n support thereof. Conclusions published without supporting evidence are to most of us worse than no conclusions at all. We suggest chat, if even one-tenth of the operators had published full, detailed results of their laboratory work, ad vance in the treatment of sulpho-telluride ores would have been very much accelerated. Therefore, we venture to inflict upon the Institute a compendium of the results of laboratory work performed during the past few years at the Lake Shore mill. It is really a summary of the more interesting results to date, prepared by the mill staff for their own use and benefit. But we hope chat the discussion under any heading is so detailed chat it will be intelligible to anyone interested in that phase of the subject.

Jan 1, 1936

By T. A. Grammatikopoulos, B. Vasily, A. H. Clark

"The St. Lawrence deposit is a large wollastonite skarn (ca. 9 Mt @ 41.3% wollastonite) adjoining the gabbroic-to-syenitic Leo Lake pluton in the granulite facies Frontenac Terrane of the Grenvillian Central Metasedimentary Belt. It occurs within a horseshoeshaped horizon of quartzite, open to the east. The skarn extends over 1.2 km2, and comprises wollastonite-dominant, wollastonite-clinopyroxene, and diverse quartz-feldspar-sulphide (-titanite) layers, interbanded with, and strongly folded within, the quartzite. As in the world-class wollastonite deposits of the Willsboro district, New York, no carbonate protolith is preserved, and calcite is a negligible constituent of the mineralized zones. Wollastonite development is nonetheless ascribed to metasomatism of a metacarbonate unit by magmatogene, silica-rich, CO2-poor (XCO2 < 0.05) fluids, at P = 350 MPa and ca. 560°C, or, alternatively, 550 MPa and ca. 630°C.Several high-grade zones, averaging >40% but locally exceeding 80%, of translucent, white, tabular wollastonite grains, with lengths of 0.2 cm to 5 cm, have been delimited by drilling. Preliminary beneficiation studies yield high aspect-ratio and high-purity (low Fe and Mn) concentrates.St. Lawrence compares favourably to other known prospects and operating wollastonite deposits: having good chemical and physical properties, low iron content, and high aspect ratios (>20:1). In addition, the deposit has a favourable size and location, and will benefit from a low Canadian dollar. Canada currently has no wollastonite production."

Jan 1, 2003

By B. Y. Park

This paper develops a series of three-dimensional simulations for the Bayou Choctaw Strategic Petroleum Reserve. The U.S. Department of Energy plans to leach two new caverns and convert of one of the existing caverns within the Bayou Choctaw salt dome to expand its petroleum reserve storage capacity. An existing finite element mesh from previous analyses is modified by adding two caverns and increasing the size of existing Cavern 102. The structural integrity of the three expansion caverns and the interaction between all the caverns in the dome are investigated. The impacts of the expansion on underground creep closure, surface subsidence, infrastructure stability, and well integrity are quantified. The results show that from a structural viewpoint, the locations of the two newly proposed expansion caverns are acceptable, and all three expansion caverns can be safely constructed and operated.

May 1, 2009

By Gaby Steinbach

A novel synthesis is described for the preparation of supported gold catalysts for low-temperature CO oxidation. The novel route is a variation of the inverse co-precipitation method and requires amphoteric metal hydroxides, with zinc and aluminium hydroxides chosen. Catalysts were prepared by inverse co-precipitation (ICPa) and by the novel route (ICPb) and tested for CO oxidation at 60ºC. Preliminary results showed the novel route to be superior to ICPa; the catalysts prepared by ICPb showed higher catalytic activities. The effect on catalytic activity of parameters such as pH, calcination temperature, washing of precursor and calcined catalysts and ageing was investigated. A preparation pH of 9 and a Zn to Al mass ratio of 3 proved to be optimal. It was possible to remove residual chloride by washing thereby enhancing catalyst performance. Ageing and Mg citrate addition were also shown to improve catalyst performance. Characterization of the catalysts revealed that better activities were obtained when the support was more crystalline and had a higher surface area. The presence of carbonate-containing species in the support decreased catalytic activity. 197Au Mössbauer Spectroscopy revealed the presence of Au3+ when catalysts were heat-treated at temperatures of 200ºC and below. Catalysts calcined at 400º had 75% of the gold present as Au0, with the remaining gold present as Au3+.

Oct 1, 2003

By Rob Stewart

Acknowledgements ? Josef Macek(Manitoba Geological Services) ? Sally Goodman (SRK) ? Inco Geologists ? Rogerio Montiero ? Thompson Mine Geologists Introduction ? Regional Geology ? Thompson Geology ? Stratigraphy ? Sulphide Characteristics ? Conceptual Model ? Structural History ? Belt ? Thompson Mine ? 1D Deposit ? Ore Controlling Structures ? Using Structural Data to Guide Exploration ? Summary

May 1, 2006

By Malcolm Powell

"The argument has long raged across the oceans as to what the best AG/SAG mill aspect ratio is. The North American practice of a high aspect ratio of two, diameter twice length, through to the South African and Scandinavian practice of an aspect ratio of as low as a 0.5. Within the collaborative P9 AMIRA project the MPRU of the University of Cape Town and the JKMRC of the University of Queensland have been collecting plant data pertinent to this issue over the past four years.The JK SAG mill model developed by the JKMRC is based upon data collected from mills having a range of aspect ratios from 1 up to 2. The maximum mill speed, mill load, and ball filling in the fitted data base were all below the operating range common in South Afr-ica. The extension of the P9 project into South Africa therefore lent an ideal opportunity to collect data in a whole new window of operation that would allow the range of applicability of the semi-empirical SAG model to be greatly extended.At the heart of the differences in aspect ratio is the completely different operating practice for high- and low- aspect mills. The mill filling, mill speed, and recirculating loads of closed-circuit single-stage operation have more influence on the differences in operation than the pure aspect ratio relationship. Relationships for load have been extended to 45% filling, with the modified model showing the measured peak in throughput as the mill filling increases. The ball load relationship now covers fillings of over 30% balls (of total mill volume), thus including so-called SAG-ball mills. The influence of closed circuit operation has been measured. It has been found that a key issue of the high slurry flows associated with closed circuit operation, especially when closed with a cyclone, is the pulp discharge limitation of the traditional grate and radial lifter configuration. It has also been observed that absolute mill length is more of an issue than aspect ratio. Slurry flow along the mill and charge segregation within the load are being investigated on pilot and full-scale mills to provide data for potentially modifylng the perfect mixing model that is used for the SAG mill.Improved understanding of these aspects of SAG mill operation and extension of the JK SAG model to better cover the entire range is allowing serious direct comparison of high- and low aspect mills for the first time. It is the conclusion of the researchers working in this area that there is no one best practice, that in fact mills across the range have their application based on ore type, desired product, available capital and on operating costs. Developing robust designs allows direct comparison of two-stage circuits using high-aspect primary SAG mills, with single-stage closed circuit low-aspect SAG mills. The new operations within South Africa are now considering both circuit types, and data is presented on selections that cover the full range of aspect ratio and operating practice."

Jan 1, 2002

By Steve Cole

"The economic and strategic importance of PGE bearing deposits is fuelling renewed exploration and development efforts in the precious metals mining sector.This paper presents a review of the various techniques used in the processing of PGE bearing ores, from primary benificiation to processing routes for the different flowsheets that have been developed amongst current producers. Recent innovations such as preconcentration of ore using gravity separation, recent trends in flotation circuit design and the development of hydrometallurgical routes for the direct treatment of flotation concentrates without the need for smelting will be discussed. Data from recent projects will be presented to illustrate common characteristics of PGE ores and concentrates.INTRODUCTIONThe six platinum-group elements (PGE): ruthenium, rhodium, palladium, osmium, iridium and platinum, together with gold and silver, are considered to be ""precious"" metals. The recent spike in platinum and palladium prices coupled with the projected increase in demand for the metals has fuelled a spate of exploration and new projects in the PGE sector. Historically primary producers of PGE metals have controlled the flow of information relating to the mining and processing of these ores into the public domain. Very little published information relating to individual operations and processes is available in comparison to the abundant literature relating to other metals.The purpose of this paper is to provide a general overview of the issues relating to the processing of PGE ores. Each topic discussed could easily be the topic of a paper or book and so in many cases only the surface of the topic will be scratched. An attempt has therefore been made to provide as many references as possible.This paper discusses the classification and location of PGE ores, the mineralogical factors that affect recovery of PGE's, methods currently used for recovering PGE's and process trends in PGE metallurgy."

Jan 1, 2002

By W. Duyvesteyn, B. Sabacky, T. Spitler, D. Verhulst

"The Altair TiO2 pigment process is based on hydrochloric acid leaching of ilmenite at high-acid concentration. Titanium is transferred by solvent extraction into a high-purity aqueous stream with a high titanium concentration. HCl gas is recovered by pressure-swing distillation and is used to regenerate the leaching solution.A new method of TiO2 recovery in dense film with simultaneous dehydration and hydrolysis provides great flexibility in adjusting the characteristics of the TiO2 pigment product.Modifications of this method lead to a flexible, low-cost, high-tonnage process first for the production of nano-sized TiO2, and later for stabilized zirconia-based nanomaterials and nano-sized titanates.IntroductionProcesses for the production of TiO2 from ore by a hydrochloric acid route have been proposed in the past. Most of these processes dissolve impurities and recover TiO2 as a residue or as a hydrolyzed phase in the original solution (Turner, 1948; Berkovich, 1975; Morris, 1982; Judd, 1986; Chao, 1993). Solvent extraction of titanium after leaching in aqueous HCl has also been studied (Tsuchida et al., 1982; Narita et al., 1983) or at least mentioned as an alternative (Ellis, 1963). A number of small-scale studies on leaching and solvent extraction in both sulphuric and hydrochloric systems have appeared recently (John et al., 1999; Lanyon et al., 1999; Sole et al., 1999. As far as we know, only Kerr-McGee (Morris, 1982) and a plant in India and one in Malaysia use a hydrochloric acid route (20 wt% HCl) to commercially produce synthetic rutile.The Altair process originated in a search for an economical method to treat high-magnesium ilmenite deposits, while avoiding the production of a waste metal salt by-product. The first attempts aimed at producing pure TiO2 by leaching out the impurities. Further work centred on completely dissolving the ilmenite in concentrated hydrochloric acid and reprecipitating TiO2. To decrease the impurity level in the product, solvent extraction of titanium into a purified stream was introduced. It also became clear that to take full advantage of the increase in leaching efficiency provided by concentrated acid, it was necessary to regenerate HCl gas."

Jan 1, 2002