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By A. Gronningsater

Introduction As is well known, the fight between the blast furnace and the reverberatory furnace in copper smelting during the second decade of the century ended with a victory for the latter, a result which must largely be attributed to the introduction of flotation methods. To my knowledge, no new blast furnace has been blown-in at any American copper smelter since the War. As the metallurgy of nickel, up to and including matte smelting, is generally considered a close parallel to that of copper, the introduction of reverberatory smelting in the big new smelter of the International Nickel Company at Copper Cliff, built in 1928-30, seemed a logical adherence to a well established trend. In view of these facts, considerable comment was made when Falconbridge Nickel Mines decided to retain the blast furnace in their smelter built about the same time, in 1929. A presentation of the reasons governing this choice should therefore be of interest. The basic principles underlying the Falconbridge practice, and the considerations guiding the successive steps in the operations, will be discussed in the following. To understand the problems involved, the character and mode of occurrence of Sudbury ore must be kept in mind. Large amounts of disseminated ore are generally present. Although the gangue usually runs only about 50 per cent Si02, the ore hoisted as a rule contains more silica than needed in the smelting operation; until recently this excess was removed entirely by hand picking (in later years supplemented by magnetic sorting methods), a difficult and not very satisfactory method, in which 'the border line between ore and waste is entirely arbitrary, and therefore too much depends on the judgment of the picker.

Jan 1, 1934

By E. Krause

The oxidation of Fe(II) with S(IV) and oxygen can be used advantageously in leaching systems requiring both Fe(III) and sulfuric acid; e.g., heap leaching of sulfide-containing ores or concentrates. The oxidation of Fe(II) with SO2/air is several orders of magnitude faster than with air alone; rates of up to ?20 g/L'h were observed at 25°C. The reaction proceeds via an intermediate orange-red mixed-valence Fe(II)/Fe(III) sulfite complex, which is present at an essentially constant concentration throughout most of the oxidation process. The oxidation rate of Fe(II) is approximated by the formula -d[Fe(II)]/dt = k [S(IV)] [02]0'3. The concentration of the mixed-valence complex, and thus the iron oxidation rate, decreases with increasing acidity. Certain impurities, such as SCN- and thiourea, negatively affect the rate. The oxidation reaction 2Fe2 + SO2 + 02 4 2Fe3+ + S042- is accompanied by the side reaction H20 + SO2 + 1/202 4 H2SO4. The side reaction becomes particularly important at low temperatures and whenever the oxygen mass transfer is insufficient with respect to the rate of addition of S(IV). Thus, effective oxygen mass transfer is essential for achieving high oxidation rates with minimal acid formation. The activation energy was estimated as 141 kJ/mol over the temperature range 10-35°C.

Jan 1, 2006

By L. Frost

"Since the late 1990s, increased prevalence of severe lung disease among US coal miners has been observed, particularly in central Appalachia where thin-seam mining is prevalent. While respirable silica exposures have been implicated in many cases, the available dust monitoring data from mines shows no explanatory trends. One possible factor could be differences in the relative toxicity of silica particles based on their surface condition. Previous research has indicated that particles with occluded surfaces (e.g., by a thin layer of clay) may be less toxic than those with non-occluded surfaces. However, this dust characteristic has not been investigated before now in central Appalachian mines. In the present study, 18 respirable dust samples from seven central Appalachian mines and two northern Appalachian mines were analyzed to assess the relative occurrence of surface-occluded silica. Results showed that the dust from central Appalachian mines actually had higher fractions of occluded silica, which may be due to extraction of large amounts of rock strata along with the coal. However, the central Appalachian mines also more silica particles overall (i.e., occluded + non-occluded) in terms of total number percentage of respirable particles.INTRODUCTIONResurgence of occupational lung disease among US coal miners, especially in the central Appalachian region, has been observed since the late 1990’s (Gibson, 2016; Zosky et al., 2016; Suarthana et al., 2011; Gamble et al., 2011; Antao et al., 2005; Attfield & Sexias, 1995). In addition to increased prevalence of simple Coal Workers’ Pneumoconiosis (CWP, or “black lung”), the prevalence of severe and rapidly progressive disease has risen dramatically (Almberg et al., 2018; Blackley et al., 2018a, 2018b, 2016, 2014; Graber et al., 2017; Cohen et al., 2016). While exposures to respirable silica have long been known to cause or play a role in diseases such as silicosis or progressive massive fibrosis (PMF), which is conventionally considered as the most advanced form of CWP, it is unclear whether silica is the only culprit in the recent outbreak of severe disease. Indeed, as disease rates have risen, both total respirable dust and silica concentrations generally appear to have declined since the late 1980s, based on regulatory monitoring data in the most affected mining districts (i.e., MSHA coal Districts 4, 5 and 12, which cover central WV, eastern KY and southwestern VA, and southern WV, respectively) (Figure 1). Of course, the latency of disease can vary and may be or more 10 years, but many of the most recent cases are likely to have resulted from exposures that occurred since the late 1980s."

Jan 1, 2019

The control of AG mills fed by ores with harder and softer components can be challenging due to the sensitivity of the mill to the ratio of hard to soft components. The mills at LKAB?s iron ore operations are generally well controlled, but there are periods when the throughput falls off dramatically, due to a loss of mill filling. As the ore is upgraded in the sorting plant through magnetic separation prior to milling, this presents an opportunity to manipulate the feed blend in advance. In order to assess this opportunity, a fullscale mill charge was dumped and assayed by size to quantify the build-up of competent waste in the mills. This was supplemented by pilot and locked-cycle laboratory tests at a wide range of feed blends in order to facilitate the development of an appropriate multi-component AG mill model. This paper reports the experimental campaign requirements and methodologies, as well as the major outcomes from analysis and modelling exercises using the industrial, pilot and laboratory multi-component data.

Jan 1, 2012

By K. Ebrahimi-Nasrabadi

The dependence between the extent of conversion of cryptocrystalline magnesite to caustic magnesia was studied experimentally in the temperature range of 400?1200 °C and for calcination times up to 8 h. A TGA?DTA analysis showed that decomposition begins at temperatures as low as 340 °C and is completed at 600 °C. Measurements of specific surface area (SSA), iodine adsorption number (IN), crystallite size, and loss on ignition were carried out to characterize the product quality. It was found that while all the parameters are dependent on the extent of calcination, densification of the product caused by grain coarsening affects their final values. As a consequence, SSA and IN went through a maximum with time. The study showed that high quality caustic calcined magnesia can be produced in a narrow range of temperature (600?650 °C) and time (1?4 h). Medium quality caustic calcined magnesia can be made within a wider range of parameters.

Aug 1, 2013

By Michael Harley, E. Guy Charlesworth

"Abstract - Bedding-parallel thrust-hosted gold veins are developed in the Sabie-Pilgrim's Rest goldfield, eastern Transvaal, South Africa. The reefs display a complex history characterized by temporally and spatially overlapping events of similar mineralization and deformation. Recognition of the episodic reactivation of these low angle (5° to 10°) thrust faults implicates high fluid pressures (equalling or exceeding lithostatic pressures). Geochemical constraints limit the amount of fluid-rock interaction and the occurrence of transgressive mineralization within the Archaean granitoid basement suggests that the high pressure fluids originated at depth beneath the Transvaal Basin.Similarities exist between gold deposits of Sabie-Pilgrim's Rest and Telfer, Western Australia as well as Passagem de Mariana, Minas Gerais, Brazil. All deposits are sub-parallel to the enclosing bedding which is shallowly inclined. Furthermore, they are also hosted by contractional deformational structures and are clearly epigenetic in origin."

Jan 1, 1994

By H. E. Pawson

Giant Yellowknife is located 600 air miles north of Edmonton on the north shore of Great Slave Lake in the Northwest Territories. In its 24 years of operation some 7,730,000 tons of ore have been treated to yield $175,000,000 worth of gold.The gold ore occurs at irregular intervals throughout a fold-like sericite schist zone cutting Precambrian rocks. The ore is characterized by 40 to 80% quartz - carbonate and about 10% pyrite and arsenopyrite. Considerable amounts of stibnite and antimony sulpho-salts are also present. Gold is finely divided in the native state existing in complex association with the sulphides. In most instances stibnite is barren but minor values are evident on occasion. Visible native Fold, quite prevalent in the early operating years, is, practically non-existent today. It was readily apparent from mineral dressing investigations conducted in .1944 that Giant ore would be difficult to treat. By 1948 a flow sheet had been established and operations began. In the ensuing years, changing ore characteristics necessitated modifications to some process steps and complete abandonment of others. By 1962 the major metallurgical problems were solved and all that remained was the refining of operating performance to its highest level.With one or two exceptions today's operation follows the practice established by 1962 (Fig. 1 - Flowsheet).

Jan 1, 1973

By J. Esson, R. A. D. Pattrick, M. K. Nimfopoulos, D. A. Polya, K. M. Michailidis

"Economic Mn-oxide ore deposits of commercial grade occur in the Rhodope massif near Kato Nevrokopi in the Drama region, Northern Greece. The Mn-oxide mineralization has developed by weathering of continental hypogene rhodochrosite-sulphide veins. The vein mineralization is con-?ned by tectonic shear zones between marble and metapelites, extending laterally into the marble astabular, pod or lenticular oreshoots (up to 50 m x 20 m x 5-10 m). Supergene oxidation of the hypogene mineralization led to the formation of in-situ residual Mn-oxide ore deposits, and secondary infills of Mn-oxide ore in embryonic and well developed karst cavities.Whole rock geochemical pro?les across mineralized zones con?rm the role of thrusts and faults as solution passageways and stress the importance of these structures in the development of hydrothermal and supergene mineralization at Kato Nevrokopi. Three zones are recognized in the in-situ supergene veins: (A) a stable zone of oxidation, where immobile elements form (or substitute in) stable oxide mineral phases, and mobile elements are leached; (B) a transitional (active) zone in which element behavior is strongly in?uenced by seasonal ?uctuations of the groundwater table and variations in pH-Eh conditions; and (C) a zone of permanent ?ooding, where variations in pH-Eh conditions are minimal. Zone (B) is considered as the source zone for the karst cavity mineralization. During weathering, meteoric waters, which were CO2-rich (PCO2~10-3.8 to 10-1.4) and oxygenated (fO2~10-17 for malachite), percolated downward within the veins, causing breakdown and dissolution of sul?des and marble, and oxidation of rhodochrosite to Mn-oxides. Karst cavity formation was favored by the high permeability along thrust zones. Dissolved Mn2+was transportedinto karst cavities in reduced meteoric waters at the beginning of weathering (pH~4-5), and as Mn(HCO3)2in slightly alkaline groundwaters during advanced weathering (pH~6-8). Mn4+-oxide precipitation took place by fO2increase in ground waters, or pH increase by continuous hydrolysis and carbonate dissolution. In the well developed karst setting, some mobility of elements occurred during and after karst ore formation in the order Na>K>Mg>Sr>Mn>As>Zn>Ba>Al>Fe>Cu>Cd>Pb."

Jan 1, 1997

By J. Esson, R. A. D. Pattrick, M. K. Nimfopoulos, D. A. Polya, K. M. Michailidis

"Abstract-Economic Mn-oxide ore deposits of commercial grade occur in the Rhodope massif near Kato Nevrokopi in the Drama region, Northern Greece. The Mn-oxide mineralization has developed by weathering of continental hypogene rhodochrosite-sulphide veins. The vein mineralization is confined by tectonic shear zones between marble and metapelites, extending laterally into the marble as tabular, pod or lenticular oreshoots (up to 50 m ?? 20 m ?? 5-10 m). Supergene oxidation of the hypogene mineralization led to the formation of in-situ residual Mn-oxide ore deposits, and secondary infills of Mn-oxide ore in embryonic and well developed karst cavities.Whole rock geochemical profiles across mineralized zones confirm the role of thrusts and faults as solution passageways and stress the importance of these structures in the development of hydrothermal and supergene mineralization at Kato Nevrokopi. Three zones are recognized in the insitu supergene veins: (A) a stable zone of oxidation, where immobile elements form (or substitute in) stable oxide mineral phases, and mobile elements are leached; (B) a transitional (active) zone in which element behavior is strongly influenced by seasonal fluctuations of the groundwater table and variations in pH-Eh conditions; and (C) a zone of permanent flooding, where variations in pH-Eh conditions are minimal. Zone (B) is considered as the source zone for the karst cavity mineralization. During weathering, meteoric waters, which were CO2-rich (PCO2 ~10-3.8 to 10-1.4) and oxygenated (fO2 ~10-17 for malachite), percolated downward within the veins, causing breakdown and dissolution of sulfides and marble, and oxidation of rhodochrosite to Mn-oxides. Karst cavity formation was favored by the high permeability along thrust zones. Dissolved Mn2+ was transported into karst cavities in reduced meteoric waters at the beginning of weathering (pH~4-5), and as Mn(HCO3)2 in slightly alkaline groundwaters during advanced weathering (pH~6-8). Mn4+-oxide precipitation took place by fO2 increase in ground waters, or pH increase by continuous hydrolysis and carbonate dissolution. In the well developed karst setting, some mobility of elements occurred during and after karst ore formation in the order Na>K>Mg>Sr>Mn>As>Zn>Ba>Al>Fe>Cu>Cd>Pb."

Jan 1, 1997

By O. A. Carson

Introduction Although the equilibrium conditions of the copper-tin alloys have been the subject of a ?great amount of research, the equilibrium diagram is still in doubt. The first systematic investigation of the series was made in 1894 by Le Chatelier, who published a freezing-point curve. In 1895, Roberts-Austen and Stansfield published the complete liquidus and partial solidus curves, which they modified somewhat in 1897. Their liquidus as given in 1897 is essentially that accepted today (Figure 3). Haycock and Neville also published a liquid us curve in 1897. In 1898, William Campbell published a series of micrographs showing the changes in microstructure which accompanied the thermal changes of Stansfield? s diagram. The classic work on the bronze series was published by Heycock and Neville in 1903. Their diagram was based on both thermal and microscopic evidence, and is the first for this series of alloys which can be considered nearly complete. Further, it was the first bronze diagram in which any attempt was made to apply the phase rule in the interpretation of the thermal and microscopic results. That portion of Heycock and Neville's diagram covered by the present work is reproduced in Figure 7.

Jan 1, 1929

By H. Wotruba, L. von Ketelhodt, M. Kowalcyk, C. Kleine, M. Robben

Sensor-based sorting is a financially and technically viable technique that is operating with little or no water and relatively little power consumption. Sensor-based sorting has already proven its operational stability at two gold applications in South Africa. Sensor-based sorting cannot only turn waste dumps and diluted mining blocks into reserves. The implementation can also cause a decrease in specific operating costs of mill and mine when for example replacing selective mining by bulk mining methods with subsequent sorting. Successful installations, operational experience, test work data and flow sheet integration from mine-to-mill underline the impact and potential that sensor-based sorting can have on the gold mining and processing industry.

Jan 1, 2011

By S. Mostaghel, Fariba Azgomi, Steven Goodman, M. Oliazadeh

"It is known that copper loss to the slag phase can be in the forms of chemical dissolution and mechanical entrainment. The contribution of each mechanism to the total Cu-loss depends on different parameters and may vary significantly. To reduce emissions to the atmosphere and to shorten the converting cycle, most industrial processes are aiming at production of high matte grades. Increased oxygen-to-sulfur potential, required by these processes, results in a highly oxidized slag, which in turn leads to an increased concentration of the chemically dissolved copper in the slag phase. Hence, slag cleaning processes are becoming increasingly important for economic viability of the smelting operations. Hydrometallurgical, pyrometallurgical, and mineral processing routes have been used for copper recovery from slag. The current article reviews a trade-off study evaluating two of the commonly used methods (namely flotation and electric slag cleaning furnace) by comparing the technical pros and cons, operability/maintenance requirements, and capital/operating costs of the operations.INTRODUCTIONLarge volumes of smelter and converter slags are produced by copper smelters around the world. It is estimated that for production of each ton of metal 2.2 tons of slag is produced (Gorai, Jana, & Premchand, 2003). Considering the current pressure on the mining and metals industry for “zero-waste, sustainable and environmentally friendly” extraction of metals, it is absolutely essential for metal producers to improve their efficiencies in order to remain competitive. To meet these objectives, almost all metal producers, from steelmaking plants, ferroalloys, base and precious metals producers, are trying to reduce their metal losses to the slag phase. In particular, copper producers are aiming to produce higher matte grades in their smelting furnaces to reduce the converting time and improve their overall throughput. Copper smelters need to minimize their copper/precious metal losses to the slag phase before final landfilling/use of slag. The current article, therefore, briefly reviews the copper loss to the slag phase, methods of metal recovery from the slag and then presents a trade-off study conducted for a green field flash smelter of the sulfide concentrates. A comparison was made between two of the major process routes for metal recovery from slag, namely slag cleaning furnace and flotation."

Jan 1, 2016

By J. R. Porter, C. Panaou

"A significant effort is expended on preparation of the capital estimate for feasibility studies. The development of the operating cost estimate is usually allocated a significantly smaller budget and less emphasis is placed on ensuring the accuracy of the operating cost. This paper presents a comparison of capital and operating costs of hydrometallurgy processing plants in order to demonstrate the importance that variations in operating costs have on the overall life-of-mine costs. The analysis is supported by a case study on the operating costs of a water treatment plant at an existing Australian hydrometallurgy facility at the feasibility design level. INTRODUCTIONDetermination of the economic validity of a minerals processing project is a critical outcome of a feasibility study. In many cases, the financial results of the feasibility study will determine if the project should proceed to execution. Ensuring that these results are reliable and built upon sound engineering practices allows companies to make informed decisions on resource opportunities and the selected technology.The methodology for developing a capital cost estimate at the feasibility level is well established. Specialist engineering estimators use budget quotations, first principle estimation and rules-of-thumb to develop an estimate with an accuracy of typically -10% to +15%. Cost reporting software packages are used to methodically build up the estimate from design information supplied by the engineering disciplines. Risk analysis and capital expense workshops are carried out to identify potential deficiencies in the estimate. It is widely recognized that in order to successfully produce a feasibility level capital estimate with an appropriate level of accuracy, a rigorous development process is necessary.The development of the operating cost estimate is typically less rigorous than the capital cost estimate. The operating cost is normally derived by the process engineering discipline and rarely are specialist estimators used to support the development. The allocated budget for producing an operating cost is usually smaller and less emphasis is placed on ensuring the accuracy and contingency reflect the most likely operating cost scenario."

Jan 1, 2012

By K. Janghorban

Multilayered ceramic-metal composites were produced by solid state diffusion bonding of polycrystalline alumina and three Al alloys containing 1 to 2%Sr, 2% Mg and Mg + Si (5057 Al alloy), The multilayered structures were fabricated at 610°C under a compressive stress of 3 MPa for 2 hours in He-5%H atmosphere, The ceramic-metal interfaces were studied by SEM and TEM equipped with energy dispersive spectrometer (EDS). Addition of 1 wt % Sr to Al promoted good adhesion, without producing any interfacial products, while the Al-2%Sr was poor in this regard due to the presence of dispersed A14Sr which is an intrinsic phase in the Al-Sr phase system, The Al-2%Mg alloy produced a continuous film at the interface, which was characterized to be MgAl2O4. Al alloys containing Mg and Si (5057 commercial alloy) produced not only the spinel structure (MgAl2O4), but another phase rich in Mg and Si (close to Mg2Si), These latter interfacial reaction products were not the intrinsic phases in Al-Mg alloys, and were reaction products at the Al2O3 ?Al alloy interfaces.

Jan 1, 2005

By Jon A. Carlson, W. Scott Williams

The EKATI"" Diamond Mine commenced operation in October 1998. It is operated by BHP Diamonds Inc., which is a part of the Non-Ferrous and Industrial Materials Division of BHP Minerals, a business unit of BHP Company Ltd. of Melbourne, Australia. The EKATI"" Diamond Mine is a joint venture of BHP Diamonds Inc. (51%), Dia Met Minerals Ltd. (29%), Charles E. Fipke (10%) and Dr. Stewart L. Blusson (10%). In December 1993, an office was opened in Yellowknife, Northwest Territories specifically for the then Northwest Territories Diamonds Project (EKATI"" Diamond Mine) development. EKATI"" Diamond Mine is located 300 kilometres northeast of Yellowknife and 200 kilometres south of the Arctic Circle in the Northwest Territories of Canada (Figure 1 ). The site is located in a remote tundra area of continuous permafrost, north of the tree line, which is part of the Canadian Shield. Wekweti (Snare Lake), a Dogrib community approximately 180 kilometres southwest of the project, is the closest neighbour. The mining area and facilities lie within the 10 960 hectares land leases held by the joint venture. The total claim block (core zone and buffer zone joint ventures) covers 344 000 hectares. To date, more than one hundred kimberlite pipes have been identified within the claims. Five of these provide the basis for the present mine plan. Total capital investment in the project is over C$700 million. At full production employment will be about 600, including employees and long-term contractors, with over 33% being Aboriginal from the Northwest Territories.

Jan 1, 2001

By Gilles Tremblay

The legacy of orphaned/abandoned mines, with their associated environmental liability, human health concerns and the financial costs of clean up is a serious issue facing Canada. In 1999 and 2000, a number of stakeholders put forth requests to Mines Ministers to establish a joint industry-government working group, assisted by other stakeholders to review the issue of abandoned mines. The Ministers supported this initiative and asked that a multi-stakeholder workshop be organized to identify key issues and priorities. The Workshop on Abandoned Mines, held June 2001 in Winnipeg, developed consensus, guiding principles and recommendations that were presented at the Mines Ministers? Conference, September 2001. Ministers agreed on the importance of a large-scale program for the rehabilitation of orphaned/abandoned mines sites. A multi-stakeholder Orphaned/Abandoned (O/A) Mines Advisory Committee was formed and a plan was developed to study various issues relating to the implementation of remediation programs across Canada. Three task groups were formed to address the following key areas: Information Gathering; Community Involvement; and Barriers to Collaboration. The presentation will include a summary of the results of the three projects and the future work plan for the program.

Nov 1, 2002

By G. Barr

"The Polymetallic Twin Metals Project deposits in Northern Minnesota are being developed by Twin Metals Minnesota. Amongst the flowsheet alternatives being examined at a prefeasibility study level is one involving flotation of a bulk concentrate followed by high temperature autoclaving of the concentrate using the PlatsolTM Process to solubilize all the valuable metals in the concentrate (Cu, Ni, Pt, Pd and Au). Autoclave operating conditions have been established in which recoveries of all the pay metals are in the 92 to 99% range, and the autoclave residence time has been reduced to an hour. The stability of the precious metals in solution after autoclaving has been assessed, as well as the impact of internally recycling process liquors to the feed to the autoclave.INTRODUCTIONPlatinum group metals (PGMs; platinum, palladium, rhodium, ruthenium, iridium and osmium) are traditionally recovered from their ores by production of a high-grade concentrate by flotation and/or gravity concentration followed by smelting of the concentrate. Concentrates are smelted to remove nonsulphide minerals and converted to remove iron and sulphur. The resulting high PGM grade matte is processed hydrometallurgically to leach the base metals, leaving a residue with a very high grade (20-40%) PGM concentrate for final refining to individual PGMs. Due to the complexity of this process and its high capital cost, this process can only be applied economically to high grade PGM ores. Lower grade PGM concentrates might be accepted as copper and nickel concentrates but may not yield attractive payment terms for the PGMs. The mineralized Duluth Gabbro region of Minnesota hosts several low grade base and PGM deposits and Twin Metals Minnesota is working to develop the Twin Metals Project deposits. Earlier studies at both pre-feasibility and feasibility level, by other mining companies, have failed to yield an economically viable project based on the traditional process. It is possible to produce a saleable copper concentrate from these ores, but economic recovery of nickel and PGMs into saleable concentrates proved elusive.In the late 1990’s and early 2000’s the Platsol™ process was developed at SGS in Lakefield, Ontario (at the time Lakefield Research Limited) for the NorthMet deposit (also within the Duluth Gabbro) owned by PolyMet Mining. The process developed was a high temperature chloride-assisted pressure oxidation process capable of dissolving simultaneously the base metals (Cu, Ni, Co, Zn), gold and the PGMs."

Jan 1, 2012

By Jeanie Wills, Theresa Nyabeze, Samantha Espley, Donna Beneteau

This presentation reports survey findings that aim to modernize the definition and understanding of the term 'mining.' The survey, conducted via LinkedIn in Winter 2024, garnered responses from a diverse group of individuals, including current or former employees of mining or exploration companies, mining consulting firms, service providers, supplier firms, as well as students and retirees with industry experience. The data, ethically approved by the University of Saskatchewan, offers valuable insights into satisfaction levels within the mining sector and suggests an expansion of the identifying characteristics of definitions of mining. This marks the third CIM gaining insights paper by Beneteau, Espley, and Nyabeze, following previous works on 'Career Satisfaction for Women in Mining' (2010) and 'Mentorship Initiatives in the Canadian Mining Industry' (2018*)."

Oct 1, 2024

By A. Quilodran, B. Jakovljevic, T. McCallum, P. Wong

"Rare earth separation by solvent extraction (SX) requires multiple separation plants each consisting of many extract, scrub, and strip stages to separate the individual metals. SX modeling advances have brought benefits to circuit design and process optimization by quickly narrowing test conditions without the need for exhaustive laboratory studies. Modeling can be used to optimize circuit configuration, O/A ratios, and acidities, which allows for CAPEX and OPEX savings. A comparison between pilot plant data and the results expected from modeling is presented for rare earth separations with CYANEX® 572.INTRODUCTIONRare earth separations can be difficult due to limited separation between the elements (Xie, Zhang, Dreisinger, & Doyle, 2014). This requires relatively complex solvent extraction circuits utilizing multiple extract, scrub, and strip stages as well as multiple SX configurations to achieve the desired metal purities. Designing rare earth circuits can be complex, and significant time and testing may be necessary to determine the required staging and optimal conditions.Circuit modeling for rare earth separations has been introduced to the industry using CYANEX® 572 extractant (Soderstrom, McCallum, Jakovljevic, & Quilodran, 2014). CYANEX® 572 is an acidic organophosphorus extractant formulation which has a weaker ligand-metal complex strength than that of the traditionally used phosphonic acid extractants. This allows efficient extraction while requiring a significantly lower acid concentration for stripping (Wang, Li, Zhao, Dong, & Sun, 2015). The reduced acid requirement in strip is expected to provide OPEX savings.The modeling can be utilized to quickly determine the required acidity and organic/aqueous (O/A) ratio for a given separation. It is also expected to reduce CAPEX requirements as optimal staging can be determined, and implemented.CYANEX® 572, combined with modeling, was utilized to produce a purified dysprosium (Dy) stream. To achieve this separation, multiple SX circuits were designed and operational conditions specified to separate Dy from the mixed rare earth chloride feed."

Jan 1, 2016

By F. E. Lathe

INTRODUCTION THE present study, based largely upon phase equilibrium diagrams and other theoretical considerations but nevertheless addressed primarily to the practical metallurgist, has been undertaken with a view to discovering what light can be thrown upon the role of alumina in slags. According to Liddell (23), "alumina apparently acts as .a base if much silica is present, as an acid if the silica is low, always as a nuisance." Newton and Wilson (26) state that .alumina "is never added intentionally, as it makes the slag viscous, and is generally an undesirable constituent of slags". Camp and Francis (3) sum up their views by saying, "alumina is seldom used intentionally as a flux, but it is present in nearly all raw material, hence unavoidable". One has only to read The Function of Alumina in Slags, by Henrich (I3), and the very extensive discussion that followed the presentation of his paper, to realize that, before the days of phase equilibrium diagrams, it was a very serious question indeed as to whether alumina should be calculated as a basic, acid, or neutral constituent. Opinions then differed widely, as well they might; today, even though no categorical answer can be given to the question raised, the function of alumina can be at least partly clarified by a study of the fundamental work done by the Geophysical Laboratory in Washington and by certain universities in both America and Europe. It will be a source of gratification to Canadian metallurgists and ceramists to know that work of this type is now under way in the recently organized Physical and Crystal Chemistry Section of the Mines Branch, Ottawa, and that results of practical value have in fact already been obtained.

Jan 1, 1951