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By Patrick F. R. Murphy

An ultra-low profile LHD has been developed by EJC Mining Equipment of Burlington, Ontario for narrow-reef mining applications. This paper discusses the challenges encountered during the evolution of the machine's design and subsequent field testing. Continuing research and development focuses on optimizing machine performance and controllability, reducing emissions, and enhancing maintainability. Details will be given on field trial experience and ongoing test work being carried out at the manufacturing facility.

May 1, 2004

By J. Dean Warner

From 1984 through 1987, the Bureau of Mines investigated numerous prospects and new discoveries of columbium, rare-earth elements (REE), uranium, zirconium, and other metals associated with the multi-phased peralkaline granite at Bokan Mountain, southeastern Alaska. This report summarizes the mineral evaluations and estimates indicated and inferred resources for those prospects containing significant tonnage and grade. A total of approximately 96.2 MM lb Cb205 (43 pct of the tonnage at a grade exceeding 0.125 pct Cb205), and 11.8 MM lb U308, 27.9 MM lb Th02, 133.0 MM lb Y203, 637.6 MM lb ZrO2, and 241 MM lb REO are estimated to be contained within nine deposits totalling 37.8 million tons (at least 80 pct of resource tonnage also exceeds 0.5 pct Y203 + REO). Generally, half of the REO content is composed of the heavy yttrium subgroup. Several deposits also contain beryllium and tantalum; 2.1 MM lb of Ta205 and 8.9 MM lbs of BeO were estimated. These estimates would likely be substantially increased with additional exploration. Four deposit types were investigated; shear zones and fracture-related deposits, pegmatites, mineralized dikes, and placers. Mineralized dikes contain most of the beryllium, columbium, tantalum, yttrium, zirconium, and REE resources, whereas shear zones and fracture-related deposits (Ross-Adams mine) contain the principal of the uranium resource. Pegmatites are similarly mineralized as are dike deposits, and contain low levels of gold. Placer potential is suggested, but no deposits were found.

Jan 1, 2011

By M. W. Hildyard, M. O. Kataka

Seismic integration schemes aim to incorporate the rock mass response provided by the historical seismic activity to improve the ability of numerical models to reproduce future response to mining. Up to now, integration schemes considered the events using quasi-static models. This paper presents case studies that consider some of the seismic and modelling issues needed to extend these schemes to include the dynamic characteristics of seismic events. The ultimate aim of an automated integration scheme seems impractical for dynamic modelling. The predictions of the WAVE elastodynamic finite difference program are, however, suitable for modelling limited numbers of events to determine either the near-field velocities and damage or the effect of fractures and excavations on the far-field low frequency response. Les projets d’intégration sismique ont le but d’incorporer la réponse de la masse rocheuse à l’activité sismique historique pour améliorer la capacité des modèles numériques à reproduire la réponse future à l’exploitation minière. Jusqu’à présent, les projets d’intégration étudiaient les évènements avec des modèles quasi-statiques. L’étude présente révèle des études de cas qui considèrent certaines des questions concernant modelage et séismes qui sont nécessaires pour porter plus loin ces projets en incluant les caractéristiques dynamiques des évènements sismiques. Le but final d’un projet d’intégration automatisé semble impraticable pour la modélisation dynamique. En revanche, les prédictions du programme élastodynamique par la méthode des différences finies WAVE sont appropriées à la modélisation d’un nombre limité d’évènements de façon à déterminer soit les vitesses à petite échelle et les dommages, soit les effets des fractures et excavations sur la réponse de basse fréquence à grande échelle. Seismische Integrationsschemata zielen darauf ab, Reaktionen der Gesteinsmasse, die von den Berichten seismischer Aktivitäten stammt, zu inkorporieren um numerische Modelle in die Lage zu versetzen, künftige Antwort zum Bergbau zu reproduzieren. Bis dato betrachteten Integrationsschemata die Fälle mit quasi-statischen Modellen. Diese Arbeit stellt Fallstudien dar, die einige der seismischen und modellierenden Fragen betrachten, die benötigt werden, um diese Schemata auszuweiten, um die dynamischen Charakeristika seismischer Fälle einzuschließen. Das Endziel eines automatisierten Integrationschemas erscheint für das dynamisches Berechnen undurchführbar. Die Vorhersagen des WAVE elastodynamisch begrenzten Unterschiedprogramms sind jedoch verwendbar für das Berechnen von begrenzten Anzahlen von Fällen, um entweder die Nahfeldgeschwindigkeiten und Schaden oder den Effekt von Brüchen und Extraktionen auf der Fernfeld Niederfrequenzantwort festzustellen.

Jan 1, 2003

By G. B. Harris, A. R. Laplante

Fine tungsten was used to mimic gold and fine and coarse silica and magnetite to mimic gangue. The optimum flow rate of fluidisation water varied from 1.7 l/min for fine silica to 5.6 l/min for coarse magnetite. Overloading at optimum flow rate was then investigated. With fine silica gangue no overload occurred under any of the conditions studied, and there was very little overload with fine magnetite. Coarse silica caused appreciable overload, but only after a significant bed of tungsten had been formed; with fine magnetite gangue the onset of overloading occurred earlier. Finally, with coarse magnetite severe overloading occurred almost immediately. Overloading was modelled using a first-order kinetic rate constant for erosion of the concentrate bed and a critical mass of constant above which it takes place. The results clearly demonstrate the importance of removing oversize to maximise recovery, particularly with a high-density gangue

Jun 18, 1905

By Jr. Boyle

This report analyzes the availability of chromium in the form of ferroalloy products (high-carbon and low-carbon ferrochromium and ferrochromium-silicon) and chromite products (metallurgical, chemical, refractory, and foundry sands) in market economy countries (MEC's). A total of 873.5 Mmt of in situ material, containing 202.5 Mmt of chromium, in 10 MEC countries was analyzed for availability as of the January 1989 study date. Extraction and beneficiation of this in situ material would result in 475.4 Mmt of marketable chromite products. As analyzed, 288.5 Mmt of these chromite products would be available for export or for internal chemical and refractory use and 186.9 Mmt would be smelted in-country to produce 74.3 Mmt of high-carbon ferrochromium, 3.9 Mmt of low-carbon ferrochromium, and 2.0 Mmt of ferrochromium-silicon, all available for export. A network flow model, which includes centrally planned economy countries (CPEC's), was also constructed to analyze world chromium supply and demand interactions for the period 1986 through 1994 and to address a hypothetical North American embargo on direct imports of all South African chromium products beginning in mid-1990. The model shows capacity constraints developing in the MEC high-carbon ferro chromium industry in 1987 and a possible subsequent overreaction by the industry in terms of new capacity additions. The results suggest that sufficient chromium is available to North America should an embargo on direct imports from South Africa be imposed by its governments. However, the effect on export revenues in South Africa is relatively small, and timely increases to world ferrochromium, chemical chromite, and foundry sand chromite capacities would be required.

Jan 1, 1993

"INTRODUCTION These safety standards for anthracite mines have been prepared for use as a guide for Federal inspectors. They are a revision of ""Tentative Inspection Standards for Anthracite Mines,"" published as Bureau of Mines Information Circular 7282 issued in May 1944 and used as a guide in Federal inspections. No provision made in these safety standards should be construed or operate to nullify any existing State statutes; they are intended to supplement the State statutes in the interest of increased mine safety. Mines are inspected in accordance with the Federal Coal-Mine Inspection and Investigation Act of 1941, Public Law 49, 77th Congress, H. R. 2082. Any unusual hazard noted by an inspector and not covered by these standards will be discussed in the inspection report, and suitable remedial procedure will be recommended.ARTICLE I. - SURFACE STRUCTURESSection 1. Breaker and cleaning plant.Electric switches and controls in breakers and cleaning plants should be enclosed.Where coal is dumped at or near air-intake openings, reasonable provisions should be made to prevent the dust from entering the mine.Section 2. Lamp houseNaphtha or other flammable liquids in lamp houses should be kept in approved containers or other suitable dispensers.Flame safety lamps should be permissible and maintained in permissible condition.Flame safety lamps should be properly cleaned, examined, and assembled by a qualified lamp attendant or other competent person before each period of use. Lamp assembly should again be checked by the person using it immediately before he enters the mine.When not in service, methane indicators, flame safety lamps, and electric cap lamps should be in the custody of competent persons responsible for maintenance and testing."

Feb 1, 1948

By D. G. Willard

The U.S. Department of Interior, Bureau of Mines has evaluated the potential resource and annual availability of cobalt from 27 deposits or districts in 11 market-economy countries. More than 95 pct of primary cobalt production from market-economy countries, mainly byproduct output from copper and nickel operations, was analyzed. Approximately 1,200,000 mt of cobalt is potentially recoverable at the demonstrated resource level. Very large additional resources exist, especially in laterite deposits, but are at the inferred level. The producing operations evaluated in this study are expected to produce about 19,000 mt of cobalt in 1989, significantly less than the available capacity of about 31,000 mt. More than 14,000 mt per year of cobalt is produced from copper-cobalt deposits in Zaire and Zambia, of which 11,000 mt can be produced at a copper price of $O.60jlb or less with a O-pct discounted cash-flow rate of return (DCFROR). Approximately 4,400 mt of cobalt is produced annually from nickel-cobalt deposits in six countries, of which about 1,800 mt (in Canada and Brazil) could be produced at a nickel price of $1.50jlb or less at a D-pct DCFROR. Approximately 200 mt per year of cobalt is produced from a cobalt mine in Morocco. Most of the evaluated nonproducing operations and prospects that could potentially produce cobalt as a principal product or byproduct would require a total cost (including a 15 pct-DCFROR) significantly exceeding January, 1989 market prices of the principal commodities. Principal-commodity dependence, transportation routes, cobalt toxicity, and other issues that could affect the availability of cobalt are addressed herein.

Jan 1, 1991

By A. Pareek, V. S. Masampally, V. Runkana

Mineral processing plants, in general, consist of comminution and separation circuits. Rougher flotation is considered as the heart of the separation circuit with concentrate grade and mineral recovery being the key productivity and quality parameters that affect the downstream operations and profitability. Online prediction of these parameters can help improve process performance in real-time. Past efforts to predict flotation cell performance have focused on building physics-based or machinelearning (ML) based models. Although physics-based models are popular in process design applications, ML models are more appropriate for online monitoring, control and optimisation due to their lower computational requirements. Deep learning (DL) is gaining prominence nowadays for application in industrial operations where a large amount of data is available. We have developed a DL model based on long short-term memory (LSTM) neural networks to forecast the two key performance indicators (KPIs), namely, mineral recovery and concentrate grade in a mineral processing plant. Rougher feed characteristics such as mineral grade and particle size distribution, upstream measurable disturbances such as throughput of the comminution circuit and pulp density, and control variables such as aeration rate, pH, and collector addition rate are considered as explanatory variables to train the model. The model was tested with data from an industrial mineral processing plant. For real-time dynamic predictions, the trained model appropriately accounts for mismatch in measurement frequency of real-time process data and off-line laboratory analysis data. Such a time series model can be used as a state observer in model predictive control techniques for optimal control of unit operations in a mineral processing plant. Keywords: Mineral processing, flotation, time-series, dynamic modelling, deep learning

Jan 1, 2020

By S Gillies, H W. Wu

A study has been undertaken into mine ventilation systems currently in use within Australian modern coal longwall (LW) extraction mines. It reviews systems and discusses evolving changes being adopted to address the more complex challenges. There is a strong move to longer panels, wider faces, greater extraction heights, increased production rates, more efficient ventilation and decreased personnel. Each of these changes is motivated by improvement in productivity and generally leads to cost reduction. In addition, mine workings are moving deeper, which results in increased ventilation control issues such as higher total and respirable dust levels, greater seam gas contents in parallel with lower in situ permeabilities, spontaneous combustion and heat management issues. A comprehensive database of Australian LW ventilation practice is analysed. Currently there are a variety of LW panel ventilation circuits used in Australian underground coal mines due to various combinations of seam characteristics, gas emission rates, spontaneous combustion, geological features and surface constraints. The main issues usually addressed in the designing and planning of ventilation circuits for LW panels are airway velocity, gas concentrations, LW cutting methods (eg Bi-di, Uni-di or half web), ventilation of control devices, pressure differentials and leakage paths and understanding gas concentrations across the length and width of the goaf. If the ventilation circuit can manage the applied contaminant load (gases, heat and dust) at an acceptable cost and circuit duty, then supplementary controls, such as gas drainage, refrigeration and dust sprays and scrubbers, may not be required. CITATION:Gillies, S and Wu, H W, 2013. Review of Australian longwall panel ventilation practices, in Proceedings The Australian Mine Ventilation Conference , pp 87-94 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By C. Bergmann, V. Ross

The concept of underground ore processing, in which valuable ore and waste are separated at a coarse particle size to improve the economics of operations and minimise the environmental impact, seems attractive from a sustainability perspective. Perceived benefits of such an operation include energy and cost savings due to a reduced hoisting load, a reduced footprint of surface operations, and using material for backfill that has not been contaminated by process chemicals. Progress in the South African mining industry in this area has however been below expectation, and opportunities for underground processing in deep, narrow reef operations were therefore explored by an Excel-based simulation with sensor-based sorting (SBS) as concentration technology. Various PGM mining scenarios incorporating factors such as the mining rate, depth of operation, and effects of dilution and particle size on the metallurgical efficiency, were compared. Whilst these initial results can be regarded as indicative only, they have shown that the economics of such an operation are driven by selective mining at significantly higher rates than achievable with conventional methods. Also, the cost of preparing, conveying and placing rock backfill will need to be lower than that of hoisting waste and depositing it on a surface dump. Another aspect to be addressed is whether ore sorting machines would be able to operate under harsh conditions underground and what provision would need to be made to ensure their efficient operation. Further investigations should focus on refining costs, incorporating environmental considerations in the modelling of economics and sustainability, and compare greenfields with brownfield scenarios. Keywords: Ore sorting, hoisting, backfill, mining rate, economics simulation

Jan 1, 2020

By N S. Weerasekara, S Town, E Lessing, J A. De Torres

Philsaga Mining Corporation (Philsaga) had three main aims in converting their lining system from steel to metal cap. They wanted to increase mill availability, minimise localised high wear and optimise particle breakage to deliver increased efficiency within the milling circuit. Combining Discrete Element Method (DEM) modelling and empirical interpretation of liner wear resulted in a rigorous approach to understanding the impact of mill liners on overall mill performance. This new understanding was subsequently used to not only derive a new lining design to optimise performance and thereby improve overall plant economics, but also to guide further lining development.In this work, detailed DEM modelling, developed from actual plant data, together with optimising liner design for wear paved the way towards the installation of a Metal Cap lining system at Philsaga’s Ø6100 mm SAG Mill. The results were a reported increase in SAG mill throughput of 16 – 20 per cent due to improved breakage through more efficient use of installed power, a reduction in localised high liner wear with overall increase in liner life and reduction in liner replacement time. Site also reported an increased gold recovery of 2 per cent which the improved breakage might have contributed to. This paper will present the design challenges and compare the Metal Cap lining systems performance against Steel lining system. CITATION:Weerasekara, N S, Lessing, E, Town, S and De Torres, J A, 2018. Metal cap versus steel lining system at Philsaga Mining Corporation’s Ø6100 mm SAG mill, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 653–664 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By Jonas Addai-Mensah

Despite the fact that nickel (Ni) laterites accounts for about 60 % of the world?s nickel mineralization, commercial processing of complex, low grade Ni laterite ores by conventional beneficiation (e.g. magnetic, electrostatic and flotation separation) still presents intractable challenges, particularly in heap leaching. Invariably, low grade laterite ores require aggressive hydrometallurgical techniques (e.g., highly acidic lixiviant leaching at ambient or elevated temperatures) for Ni and cobalt (Co) extraction. As a precursor to heap leaching, agglomeration of the multi-mineral phase feed particles into robust and porous granules of 4-10 m permeable bed, is essential. In the present work, several basic and applied studies of the agglomeration behavior of real Ni laterite ores and selected oxides and clay minerals (hematite, quartz smectite and kaolinite) which constitute the predominant host gangue phases of typical laterite ores are performed. Isothermal, batch agglomeration tests involving sulphuric acid solution as a binder were carried out to produce coarse, narrow-sized (5 ? 40 mm) granules which were well characterized. The influential role of feed ore characteristics (e.g., mineralogy/chemistry and primary particle size distribution) and binder composition (30 vs. 44% w/w H2SO4) and dosage (15-25 wt.%) in agglomeration behavior and granule properties were examined through several characterization techniques including granule size, compressive strength, bed percolation and re-wetting stability and 3-dimensional micro-structure analyses. Links between feed mineralogy/chemistry and size, binder formulation, agglomeration behavior, curing conditions and final granule properties and structural characteristics are established. The implications of the findings to Ni laterite plant agglomeration for enhanced heap leaching are discussed. Keywords: nickel laterite, complex low-grade ores, agglomeration, heap leaching, micro-structure

Sep 1, 2012

By C. G. Anderson

The treatment of concentrates for the recovery of the precious metals imposes many processing and environmental challenges. Hydrometallurgical technology for recovery of values from refractory precious metal materials has been developed at Sunshine Mining and Refining Company. Historically, conventional pyrometallurgical methods have been used for to recover values from refractory precious metal materials. However, enhanced environmental awareness has worked to constrain typical refractory material processing methods. This hinders and limits the ability of established smelting or roasting technologies to fill the needs of new sources of refractory materials. An alternative is to treat refractory precious metal bearing materials by hydro metallurgical techniques. These methods are considered more environmentally appropriate since they do not produce gaseous pollutants like sulfur dioxide. Industrially, Sunshine Mining and Refining Company employs hydrometallurgical technology to process refractory precious metal materials. The Sunshine antimony leach uses a mixture of sulfur and caustic under ambient pressure and boiling conditions. The Sunshine oxidation leach uses a mixture of sulfuric acid and a minimal amount of sodium nitrite under relatively mild oxygen pressure and temperature conditions. These conditions result in a nitrogen species catalyzed sulfuric acid pressure leach. Since hydrometallurgical technologies provide industrially proven processing options for many types of refractory precious metal containing materials, their usage should be considered. Ample work has been conducted to prove the effectiveness of Sunshine refractory processing technologies on various materials. This has included platinum group metal materials, refractory gold concentrates, silver materials and indium materials. In some cases, complete process flowsheets have been designed to recover both base and precious metals from these materials and to treat the generated waste streams in environmentally sound manners.

Jan 1, 1996

By Behzad Vaziri Hassas, Firat Karakas, Mehmet Celik, Onur Guven

Since dewatering is a costly process, removal of moisture from relatively fine size lignites is an important industrial problem. As the lignite gets finer, the presence of micropores and relatively low hydrophobic nature of lignite lead to trapping of large quantities of water in the coal matrix. Lignites used in power stations need to be dried to a certain degree of humidity before it is transferred to the combustion chamber. It is generally wise to dewater coal as much as possible by physical means before it is dried. However, dewatering cost in some cases may be prohibitive. Thus, sometimes removal of even little percent moisture may be of great help in such operations. Surfactants have been used to increase the efficiency of coal dewatering. It is crucial to select appropriate type and amount of surfactants in order to get higher efficiency. Since the hydrophobic characteristic of coals increases with increasing the coal rank, dewatering of bituminous and anthracite can be accomplished using low HLB surfactants (Oily), however, these surfactants are not efficient in the case of low rank coals like lignites and brown coals. It is suggested that adsorption of low HLB surfactants on high rank coals is governed by Acid-Base interactions rather than electrostatic and van der Waals forces. Since this characteristic in lignites is not strong enough, low HLB surfactants could not be effectively adsorbed on low rank coals to make surface more hydrophobic. In this paper, a special set-up has been designed to study the effect of high HLB surfactants on dewatering of fine lignite particles. It is shown that high HLB surfactants function more efficiently on low rank coals than low HLB ones.

Jan 1, 2014

By T Roberts, S Talu

Steel wire weld mesh is a common industry standard surface support used in underground mining, traditionally installed with mining drills during the bolting cycle. As underground mining depth and stress increase, mining methods and their sequences at these depths place challenges on conventional surface support options. These challenges have provided a need for high capacity surface support options that improve development efficiency, reduce the need for secondary ground support installation and is easily installed with current mining equipment. A 5m2 mesh testing machine was built, suspending a 9500kg concrete slab, designed with multiple hold down points to test varying bolting patterns. Mesh load and displacement is accurately measured using a data acquisition system which pushes against the restrained mesh module with an available 200kN load and 1200mm displacement capacity. A testing program comparing conventional surface support, straps and combinations of both were tested for baseline results. New module designs incorporating existing mesh types and new design concepts were manufactured to allow for single pass installation during a bolting cycle. Nine mesh types, mesh/strap combinations and new mesh module designs were compared using the same test method. Based on these results, two high capacity mesh module types are produced. This paper will detail the test program, load and displacement results and underground mesh module performance, describing the two high capacity mesh modules function within the development cycle and manual handling. CITATION:Roberts, T, Talu, S and Wang, Y, 2018. Design and testing of high capacity surface support, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 303–322 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By O. H. Rostad

The importance of geology in ore reserve estimation has been stated by several authors. King et, a1 (1982) reported that the geological factor was the key deficiency where serious weaknesses in ore reserve estimation appeared, especially in grade. The foundation for a successful mining operation is an ore reserve estimate that neither overstates nor understates the true situation as to tonnage and grade. Overestimation of reserves can have devastating economic results. Overestimation of tonnage may lead to wasting money on an overbuilt plant or worse yet, if grade is seriously overestimated also, to undertaking an operation that is doomed to economic failure from the outset. Understated ore reserves can result in building a plant that is smaller than optimum which may result in higher unit costs. At worst, understated tonnage of ore reserves can result in a missed opportunity for a profitable operation. Failure of the ore reserve estimator to properly take into account all geologic and allied factors pertaining to the deposit can result in an ore reserve estimate that is overstated either as to tonnage or grade or both. This paper will focus on these factors and will not discuss rules about distances above or below a drift that should be given a classification of proven or probable, etc. Each deposit has its own particular characteristics which often require modification of general rules for defining limits of proven and probable ore blocks that work satisfactorily in another deposit. A number of references are included which discuss some of these problems. Definitions of ore reserve classifications and an illustration are given in the Appendix.

Jan 1, 1985

By M A. Wells

The various mechanisms by which phosphorus (P) occurs and is distributed within banded iron formations (BIF)-derived iron ore are examined. In unenriched, precursor BIF P occurs predominantly as discrete grains of primary apatite. In high-grade ores hosted mainly by the Brockman Iron Formation, P occurs in three main forms:1. as secondary apatite;2. other submicron, P-rich phases (eg Rare earth containing minerals (REE), vivianite, strengite,wavellite) present in pore networks; and3. as P occurring with iron oxides, in particular, goethite.However, the presence of secondary P-bearing phases does not contribute signifi cantly to the total P level in iron ore, and it is P associated mainly with goethite, which represents the most common occurrence of P in these ores.No direct evidence for the replacement of Fe by P within the structure of synthetic goethite has been demonstrated. Therefore, it is most likely that in iron ores P associates with goethite in occluded form as adsorbed PO4 3- species or as discrete P-bearing minerals, such as strengite, wavellite or vivianite, in ultra-fine, micro-pores or within intra-domain regions of multi-domainic goethite. Such occluded P is not easily extracted with commonly used lixiviants, such as weakly alkaline leachates. Recent development of a low temperature-alkalie/water leach extraction method has been successful in reducing not only the P content, but also amounts of Si and Al inthe initial ore. Refi nement of such a method to commercial application will have enormous benefit in realising the economic potential of large tonnages of high-P (ie >0.07 wt per cent P) bearing iron ore.

Jul 11, 2011

By H. J. Heinen

This report summarizes preliminary research studies by the Bureau of Mines in electrowinning molybdenum directly from pure or impure molybdenum trioxide (MoO3) by fused-bath electrolysis in an open graphite cell without need of a protective inert atmosphere. A formulated electrolyte of the tertiary system containing 41.6 percent sodium pyrophosphate, 16.7 percent sodium tetraborat, 33.3 percent sodium chloride, and 8,4 percent molybdenum trioxide, by weight, electrolyzed at 1,000° C,, gave excellent results. Current density studies were made over a range from 20 to 260 amperes per square decimeter. Current efficiencies of 80 to 90 percent were obtained over the entire range. Yields up to 500 grams of molybdenum per kilowatt-hour were obtained. Higher current densities, however, reduced yields and produced a finer grained metal. The two most important facets in electrowinning molybdenum from MoO3, in an open cell, are temperature and electrode spacing to minimize or eliminate the formation of the suboxide MoO2, the major source of oxygen contamination. At temperatures of 1,000° C. and higher, together with the proper spacing of the electrodes, formation of oxides are practically eliminated. Molybdenum metal electrowon from pure MoO3 was 99.9 percent pure and contained 0,056 percent oxygen and 0,007 percent carbon. This metal, when arc-melted, had a Vickers hardness number of 175 and contained 50 parts per million oxygen and less than 10 parts per million carbon. Molybdenum metal electrowon directly from an impure MoO3 containing 90 percent MoO3 resulted in a metal of 99.8 percent purity.

Jan 1, 1961

By E. D. Musin, S. S. Gudkov, A. V. Epiforov, S. V. Balikov

The standard technology for processing of gold-bearing sulfide raw material has many problems associated with the use of cyanide as a gold lixiviant. Therefore, many tests to find alternative gold recovery methods are being undertaken. This paper describes some technological solutions, which can become alternatives to cyanide technology in the future. Pressure oxidation (POX) of refractory gold-bearing sulfide concentrates from five different Russian gold deposits were investigated. Concentrate 1 was gold-copper concentrate. Concentrate 2 was gold-pyrite concentrate. Concentrates 3-5 were pyrite-arsenopyrite double refractory concentrates. The POX tests were at temperatures of 110, 200 and 220 °C and oxygen pressures of 1.5, 2.4 and 3.0 MPa were carried out. Gold lixiviants, such as chloride, thiosulfate, sulfite and thiocyanate were tested. The recovery of gold from sulfide gold-bearing concentrates by non-cyanide methods was 97- 99%, and from double refractory concentrates, 86-93%. The tested reagents are less toxic than cyanide, and their use will reduce of the cost of gold production. These methods make it possible to recover gold in acidic media, and eliminate the washing operation and POX slurry neutralisation. In addition, precious metals recovery from POX cakes using lead reduction smelting was investigated. It was shown that the “POX-pyrometallurgical” technology for double refractory concentrates processing could be used with recoveries of gold and silver of 97 and 96%, respectively. Keywords: Recovery of gold, refractory sulfide gold-bearing concentrates, double refractory concentrates, pressure oxidation, non-cyanide lixiviants of gold, chloride autoclave leaching, thiosulfate leaching, thiocyanate leaching, sulfite leaching, reduction smelting, POX-pyrometallurgical technology

Jan 1, 2020

By T. Bednarski, T. McCallum, L. Moya, B. Gay

"Molybdenum recovery by solvent extraction has gained interest partly due to advances in concentrate leach technology. Historically, both molybdenum and rhenium have been produced using amine-based extractants with the ability to co-extract both metals. Alternatively, a phosphinic acid-based extractant (CYANEX® 600) can be used depending on the flow sheet. CYANEX® 600 selectively extracts Mo leaving Re in the raffinate, which is a valuable feature in plants which require rejection of the rhenium from the molybdenum circuit. CYANEX® 600 piloting results are reviewed and compared to commercial experiences with an amine-based extractant. INTRODUCTION Molybdenum is most commonly found as molybdenite (MoS2) in primary porphyry deposits (about 59% of global reserves) and copper porphyries (about 39% of molybdenum reserves)[1]. Recovery from molybdenite traditionally involves flotation (concentrates typically containing 85 – 95% MoS2) followed by roasting to produce MoO3. Solvent extraction is also used to recover molybdenum from acidic aqueous solutions resulting from scrubbing of roaster and smelter off gasses and leaching of molybdenite (uranium ores). Molybdenum is also frequently found in low concentrations in copper leach inventories[2]. More recently, pressure oxidation has gained interest to treat molybdenite concentrates. Molybdenum Chemistry The chemistry of molybdenum is complex with a series of extractable species existing dependent on the pH and oxidation state as shown in Figure 1.Under strongly basic conditions, molybdenum primarily exists as MoO42- while under strongly acidic conditions MoO22+ becomes the predominant species. In between, a number of isopolymolybdates such as Mo7O246- and Mo8O264- are dominant. Equations (1) – (3) demonstrate some of the proposed molybdenum equilibrium reactions[1]."

Jan 1, 2017