Search Documents

By K. Liddell

Current practice for the recovery of platinum group metals (PGM) from flotation concentrates is by energy-intensive smelting processes followed by separation and refining of the base and precious metals. Energy costs are rapidly increasing in South Africa and this trend is forecast to continue in future years. Moreover, the smelting operations are highly susceptible to deleterious constituents, particularly chromite; and the increasing tonnages of UG2 chromititic ore being mined results in an increasing chromite content of smelter feeds. Hence there is a substantial incentive to assess and implement alternative processes to smelting for the treatment of PGM concentrates. The Kell process is a patented hydrometallurgical alternative to smelting that is totally insensitive to the chromite content of the feed concentrate and comprises three commercially proven unit operations (pressure oxidation leaching of base metals and sulfur, roasting and chlorination of precious metals). The Kell process provides the opportunity for existing and new PGM producers to significantly reduce their energy consumption, and in particular their consumption of electricity, by removing the need to smelt PGM concentrates. Energy costs, greenhouse gas emissions and installed power requirements are also much reduced. Compared with smelting and refining these reductions are calculated as: ? Total energy consumption-50% reduction ? Electrical energy consumption-84% reduction ? Energy consumption costs-76% reduction ? C02 emissions-70% reduction ? Installed power requirement 92% reduction. The data obtained from simple laboratory batch amenability testing can be readily applied to the Simulus mass and energy balance simulation procedure to determine energy and reagent consumptions, equipment selection and process optimization. The applicability of Kell to treating low-grade 'dirty' concentrates with high UG2 content as well as to polymetallic concentrates containing base metals as the major payable metals with subordinate PGM means that these savings can also be expected when the process is used for treating a range of concentrate types and grades. Keywords PGM,UG2, Merensky, smelting, chromite, electricity, hydrometallurgy, Kell Process, Eskom.

Jan 1, 2011

By Derek H. C. Wilton, David T. W. Evans

"Abstract - The Midas Pond gold prospect, central Newfoundland, is hosted by sheared and altered felsic and mafic pyroclastic rocks of the Tulks Hill volcanics, Victoria Lake Group, and has been trenched and tested by 19 diamond-drill holes. Gold values are sporadic, with the best diamond-drill intersection assaying 7.3 g/t over 0.9 m, and the highest grade channel sample containing 14.7 g/t over 1.15 m. Alteration and mineralization are confined to a 200 m wide brittle-ductile shear zone. This shear zone formed in response to regional D1 deformation, and the shear zone fabric parallels the regional S1 foliation. D2 deformation resulted in broad Z-shaped flexuring of the shear zone. The deformation was probably related to the ca. 396 Ma to 420 Ma Salinic Orogeny.Advanced argillic alteration and an extensive halo of Fe-carbonate and pyrite surround the gold mineralization. Alteration minerals include pyrophyllite, paragonite, quartz, plagioclase, chlorite, fluorite, Fe-carbonate, and pyrite. As such, the alteration seems to reflect combined epithermal (argillic) and mesothermal (carbonate) mineralizing environments. The argillic alteration and fluorite represent a regional hydrothermal effect on the felsic volcanic rocks; whereas, the Fe-carbonate and pyrite represent a mesothermal alteration associated with auriferous quartz veins. The Au, which is spatially associated with pyrite, occurs in three structurally-controlled vein sets. These veins are confined to the contact between a highly deformed breccia (the banded mafic tuff) and structurally overlying, altered felsic tuffaceous rocks. The vein sets include: V1 boudinage veins that parallel the shear-zone fabric (C-shear veins), and V2 and V3 extensional fracture veins which are controlled by an S2 fracture cleavage. V1 veins are the earliest and contain the lowest concentrations of Au. V2 and V3 veins are concentrated within the hinges of the D2 flexures.Pb isotope data for galena separates indicate that the Pb in the vein systems was probably of local derivation. Sulfur, C, and O isotope data for mineral separates are similar to data for typical mesothermal lode gold deposits. Fluid inclusion data suggest the presence of at least three hydrothermal fluid types in the region, viz., (1) an early saline fluid which produced the argillic alteration, (2) a CO2-bearing fluid directly related to the auriferous mineralization, and (3) H2O-NaCl fluids which flowed through the Midas Pond shear zone at a lateral distance from the zone of mineralization. The mineralizing fluids are interpreted to have been CO3 -2-rich types which originated through metamorphic dehydration and decarbonation of mixed island-arc and continental crustal rocks. Gold precipitation resulted from the reaction of these fluids with the Fe-rich breccia unit. Pre-existing and simultaneously crystallizing pyrite served as loci for Au precipitation."

Jan 1, 2000

By J. Ekengård, H. Srinath, M. Andersson, P. Jönsson, D. S. Sarma

The present paper deals with mixing between the metal and slag phase during the ladle refining process, including vacuum degassing. Dispersion of metal drops in ladle slags has been studied. The results show a larger quantity of metal droplets in ladle slag during the alloying and stirring phase than during the subsequent vacuum degassing phase. The thermodynamic and kinetic aspects of the interaction between metal drops and the slag matrix have been considered and probable reasons for the observed distribution discussed.

Jan 1, 2004

By P. J. Rider

Diamond Value Management?s objective is to achieve $/ton revenue improvement across De Beers group mines. The respective %improvement contributions are measured against targets set by each of the operations. Degree of target achievement is measured via a maturity model at each operation. DVM comprises seven elements. It spans several value chain facets and various collaborative projects are conducted under its umbrella. New technologies in De Beers are also evaluated from a DVM perspective.

Jan 1, 2003

By A N. Kasnanto, D A. Sims, J Pocoe, B Pulungan

In this case study we compare the sampling quality of diamond drilling (DD) and reverse circulation (RC) drilling at the high sulfidation Purnama gold–silver (Au-Ag) deposit located in northern Sumatra, Indonesia.There was a systematic bias of 23 per cent gold metal under-call in the diamond drilling based Ore Reserve estimate compared to the mill reconciliation over three years of production. The RC based Grade Control data and mill production data were in much closer accord. Investigation led to quantification of the significantly reduced sampling imprecision (or improved precision) of RC drilling samples relative to diamond drilling samples and the determination of a grade bias between the drill types caused by sample volume and metal distribution in the mineralisation. An infill RC drilling program was undertaken to improve data density across the pit for inclusion in the Mineral Resource estimate. This subsequently increased Au metal in the estimate by 16 per cent.The sampling imprecision of primary diamond drill sampling is rarely tested. Duplicate sampling can be used to quantify sampling imprecision at each stage of subsampling. In diamond drilling it is tested at the primary sampling stage via second half core sampling, while RC sampling is tested with field duplicate sampling at the rig. Duplicate data can be compared in a number of ways to determine imprecision including Thompson-Howarth plots which quantify imprecision to allow comparison between different methods and protocols. Comparing sampling imprecision between data sets and data types allows determination of optimum drilling and sampling protocols and can reflect upon the underlying validity of Mineral Resource input data and estimation outputs.CITATION:Sims, D A, Pocoe, J, Kasnanto, A N and Pulungan, B, 2017. Why combine reverse circulation and diamond drilling in Mineral Resource estimation at Purnama?, in Proceedings Tenth International Mining Geology Conference 2017, pp 413–422 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 20, 2017

With the employment of a system of ore pillars to support the hanging wall during the mining operations at the North Mine, it was found that, when the stopes were worked out and the pillars remained intact, the main footwall drives developed an undulating longitudinal section due to the shrinkage of the filling in the worked out stopes.These depressions collected water, which made the footwall drives very wet under foot, and everitually worked its way into the pillar workings and the levels below.Water was also finding its way into the shafts because of the silting-up of the level drainage systems and, as both the shafts are down-cast, this tended to increase the humidity of the air entering the mine.To overcome these troubles it was decided to try a system of siphons as a method of subsidiary drainage.SYSTEM AT PRESENT IN USE.All drainage water is collected as near to its source as possible, run into a sump in an adjacent low lying part of the level, and continuously removed from these sumps by a siphon system.Each level has its own system. Water is collected from the sumps by a 1-in. pipe and delivered to a pipe running the length of the level along the back of the footwall drive. At the shaft the pipe is increased to...

Jan 1, 1938

By Mark Lawshe, Shosie Serata

The conventional method of mining underground, using narrow rooms and wide pillars, has proven to be poorly adapted to the powerful earth forces that act on underground openings. The high costs of roof bolting, timbering, steel arching, and other means of artificial roof support for both room-and--pillar operations and longwall gate roadways attest to the limitations of traditional mine design. Too often, productivity halts when corners buckle under "cutter roof' stresses, floors heave, and roof strata fall. In the 1960s, an entirely different approach to strata control was developed. The Stress Control Method (SCM) stabilizes failing roofs and heaving floors by removing the cause of failure - highly concentrated stresses around the openings. This is achieved by regrouping three to 10 rooms into a single entry where the rooms are separated by very narrow yielding but stable pillars. SCM employs a specified excavation sequence to redirect earth pressures away from the mine openings. The yield pillars act as scaffolding until a permanently protective secondary stress envelope forms around the entry (lower part of Fig. 1). The result is a dramatic improvement in roof stability and a significant reduction in artificial support costs. Some roof bolting must still be done, mainly to meet Mine Safety and Health Administration (MSHA) requirements. Potash, salt, coal mining The stress control method was first applied to the deep potash mines of Saskatchewan. The new SCM design actually increased room widths to enlarge the stress envelope, redistributing the stress concentration away from the entries (Serata and Schultz, 1972). This resulted in a turnaround from pervasive roof fall/floor heave conditions that threatened operations in the eight new mines opened in the Saskatchewan potash field during the 1960s. In addition to eliminating all roof bolting, SCM successfully stabilized ground conditions and substantially increased productivity over the original design (Serata, 1976, 1982). Similarly, the deep Boulby potash mine of Yorkshire, England, was about to close after various conventional means of combating roof failure had failed. An SCM test area was excavated that resulted in the ground quickly stabilizing without need o artificial support (Serata, 1982; Chilton and Laird, 1982). Another example is the Diamond Crystal salt mine in Louisiana. Here, multiple level mining was being extended from the third to the fourth level downward in the Jefferson Island salt dome. Extensive roof failure was occurring at 396 m (1300 ft) ii the third level. SCM design principles in creased the room size from 24 to 49 m (80 t( 160 ft) on the fourth level at a depth of 457 m (1500 ft). This unprecedented expansion of room width was first questioned and later ap¬proved by MS HA, which expressed satisfaction in seeing the improvement documented by field validation data in 1973. The most recent SUM application to salt mining has been in the Michigan Basin a: Goderich, Ontario. The anticipated abandonment of Domtar Chemicals Group/Sifto Salt Division's deep operation in bedded sale was prevented by a multi-year SCM implementation. The success of SCM in dry mining[lecc] to its application to solution mining of salt This adaptation of the method is now being implemented in bedded salt at Fort Saskatchewan, Alberta by Dow Chemical Co., and in St. Clair, MI by International Salt Co. Application of SCM to solution mining in a salt dome is proceeding at Napoleonville. LA by Dow Chemical. The same technology has also been applied to cavern design for the Louisiana Offshore Oil Port (LOOP), and the compressed air energy storage (CAES)

Jan 1, 1989

By S. D. D. Welsby

"A pilot-scale flotation column was used to evaluate four polyglycol frothers at Teck’s Highland Valley Copper Mine. The column was fed with flotation plant feed from the grinding circuit cyclone overflow. Frother dosage was changed by controlled addition to the column feed. Pulp, froth and metallurgical performance were assessed using gas dispersion measurements, water recovery and copper, molybdenum and gangue grades and recoveries. Ranking the frothers based on pulp and froth phase performance using gas dispersion measurements and water recovery yielded contradictory rankings. All of the frothers changed the froth structure in a similar manner such that, at the same structural conditions, they all gave the same degree of separation between water and entrained solids. The only difference between the frothers was what dosage was required to achieve a certain structure. Changes in pulp structure appeared to be related to the critical coalescence concentration (CCC) for each frother. Changes in froth structure were not completely described by the CCC and involved other factors, such as the hydrophobicity of particles in the froth. All of the frothers were capable of achieving the same metallurgical result, just at different dosages. Optimal dosages for the frothers were anywhere from the CCC, where copper and molybdenum grades were maximised, to 2 to 3 or 4 to 7 times the CCC, where copper and molybdenum recoveries, respectively, were maximised.INTRODUCTION Background Due to supply issues, Teck’s Highland Valley Copper (HVC) Mine replaced Dowfroth 250C, their auxiliary frother, with Huntsman Polyfroth W31. The primary criterion for selection was that the new frother gave the same performance as Dowfroth 250C, minimising disruption to the circuit. Frothers play a dual role in flotation, altering the bubble size in the pulp zone and the stability of the froth. A stronger or weaker frother, or a blend of both, will alter the balance between pulp and froth effects and, ultimately, metallurgy. Dowfroth 250C and Polyfroth W31 may not necessarily be the best frothers for HVC from a metallurgical standpoint. There may be a different frother or frother blend that could deliver higher recovery in the bulk roughers."

Jan 1, 2014

By M. Hakala

A computer program was developed to simulate transient strains and stresses during an overcoring in the situ stress measurement. The solution is based on superpositioning of elastic stresses. The measured strains can be compared with the calculated ones to check if the measured transient behaviour is accordant with the interpreted in situ state of stress. If not, the in situ state of stress can be calculated based on any transient or final strain values. The transient stresses can be compared with the strength envelope of intact rock and thereby estimate the core damage potential. Technically, the developed code fulfilled all the objectives, and the numerical error was found to be less than 5%. The basic idea can be applied also for different overcoring probes with minor modifications and recalculation of stress tensors. The analysed case studies showed clearly the advance of having an objective and practical quality control method to study the reliability of the in situ stress measurement data. The study was set up and funded by the nuclear waste management companies Posiva Oy ( Finland ) and Svensk Kärnbränslehantering AB ( Sweden ). Un programme informatique a été développé pour simuler des tensions et des contraintes passagères pendant la mesure in situ de contrainte de surcarottage. La solution se fonde sur une superposition de contraintes élastiques. Les contraintes mesurées peuvent être comparées avec celles calculées pour vérifier si le comportement passager mesuré correspond à l'état de contrainte in situ interprété. Sinon, l'état de contrainte in situ peut être calculé à partir de n'importe quelle valeur de contrainte passagère ou finale. Les contraintes passagères peuvent être comparées avec l'enveloppe de force de roche intacte et ainsi il est possible d'évaluer le potentiel de dégâts au noyau. Techniquement le code développé a rempli tous les objectifs, et le taux d'erreur s'est révélé être inférieur à 5 %. L'idée de base peut être appliquée aussi pour différentes expériences de surcarottage avec des modifications mineures et le recalcul de tenseurs de contrainte. Les études de cas analysées ont montré clairement l'avantage d'avoir une méthode de contrôle de qualité objective et pratique pour étudier la fiabilité de données de mesure de contrainte in situ. L'étude a été fondée et financée par les sociétés de gestion de déchets nucléaires Posiva Oy (Finlande) et Svensk Kärnbränslehantering AB (Suède). Es wurde ein Computerprogramm entwickelt, um Transient-Beanspruchungen und Spannungen während eines Überentkernens in situ-Spannung-Messung zu simulieren. Die Lösung beruht auf Superpositionierung von elastischen Spannungen. Die gemessenen Beanspruchungen können mit den berechneten Daten verglichen werden, um nachzuprüfen, ob das abgemessene Transient-Verhalten mit dem interpretierten in situ-Zustand der Spannung übereinstimmt. Wenn nicht, kann der situ-Zustand der Spannung auf irgendwelchen Transient- oder endgültigen Beanspruchungswerten berechnet werden. Die Transiet-Spannungen können mit dem Stärke-Umschlag des intakten Felsens verglichen werden und dadurch ein Kernbeschädigungspotential kalkulieren. Technisch erkannte der entwickelte Code alle Ziele und hat gefunden, dass die numerische Fehler weniger als 5 % waren. Grundsätzlich können auch verschiedene Überentkernen-Untersuchungen mit kleineren Modifizierungen und Wiederberechnung des Spannung-Tensors angewandt werden. Die analysierten Fallstudien zeigten deutlich den Fortschritt, objektive und praktische Qualitätskontrollmethode zu haben, die Zuverlässigkeit in Situ-Spannung-Messung-Daten zu studieren. Die Studie wurde aufgestellt und durch Atommüll-Management-Firmen Posiva Oy (Finnland) und Svensk Kärnbränslehantering AB (Schweden) finanziert.

Jan 1, 2003

By John Sleavin

As part of the Link Light Rail project in Seattle, WA, work was completed at the Pine Street Stub Tunnel (PSST) in early 2007 for the Central Link project. This tunnel was excavated using cut-and-cover construction within the limits of Pine Street. The stub tunnel provides a turn back via double crossover for light rail trains running in the Downtown Transit Tunnel (DTT), as well as a connection point for the next phase of the project, the University Link (Fig. 1). Designing the new University Link tunnel connection to the existing PSST was a tricky task. The on-site geotechnical conditions, buried obstructions, geometry, requirements for construction and balancing operational considerations and neighborhood stakeholder concerns associated with the recently completed PSST combined to create a uniquely challenging assignment.

Jan 1, 2010

By Júlia Guimarães Sanches

O curso de Engenharia de Minas sofreu uma expansão, juntamente com todo o ensino superior no país, por volta de 2005. O objetivo do presente estudo é avaliar as características dos cursos do país e realizar uma comparação com cursos de outros países com grande produção mineral. Foi feito um levantamento inicial a partir de informações disponíveis nos sites do MEC, INEP e das universidades. Como complemento, realizou-se um questionário com os coordenadores dos cursos de Engenharia de Minas do Brasil. Foram compilados ainda dados relativos à estrutura curricular e carga horária dos cursos de Engenharia de Minas dos EUA, Canadá, Austrália, Chile e África do Sul. Em relação à comparação das grades curriculares com os cursos de outros países, observou-se que no Brasil o curso possui uma maior cara horária de Tratamento de Minérios em detrimento da lavra. Nos demais países estudados, o foco das disciplinas é a lavra, já que nestes países a área de Tratamento de Minérios é assunto de outras Engenharias.

Oct 30, 2017

During the course of the major research in the Metallurgy Department of the University of Melbourne into the creep characteristics of lead and .its dilute alloys, it has been the practice to prepare the majority of test specimens from rolled strip. From time to time during the progress of the work, certain variables involved in. the rolling of the experimental ingots have been investigated. It is proposed to collect together in this paper the results of these experiments. Some general features relating to the type of deformation produced have been discussed in a previous paper.While it is admitted that the conditions prevailing through this work differ' in some respects from those of industrial rolling practice, it is felt that many of the factors investigated have an important practical significance. Some of the work described forms a continuation of the researches of R. S. Russell (2) , (3) , (4), into the recrystallization of various pure leads after deformation by rolling.MATERIALS USEDIn this work, a variety of "pure" leads and certain dilute alloys made from the purest of the basis leads have been studied. The compositions of the basis leads and the corres...

Jan 1, 1940

With the proving of a mineable orebody at Orlando in May 1961, Peko Mines N.L. was faced with the decision of whether to mill the ore on site or to haul it 25 miles to central facilities at Peko. Orlando was a relatively small orebody, so that it was decided to haul, although no reliable assessment of possible ore haulage costs was available.The necessity for a good haul road was recognised from the beginning and a 17 mile, 12 ft wide, single seal bitumen road was constructed, to link the mine to the Stuart Highway and Peko. The initial selection of 20 ton capacity diesel-powered end-tipping ore carts was made, with the best information available at the time.Since then some 130,000 tons of ore have been carted over a period of 3 years, and this paper presents the experiences of this haulage operation and the lessons learned.A second mine has just been brought into operation at Ivanhoe, 19 miles from Peko. For this haul it was decided to use an ore haulage vehicle of the side tipping road train type, which would cart some 70 tons, or more than three times the load of the vehicles initially selected.Only about 5,000 tons had been hauled at the date of this paper, but preliminary figures indicate a cost 01 1· 65d./ton-mile as against 2·47d. for the smaller vehicles.It would appear that the road haulage of ore from small outlying mines to a central plant can be highly economical, if certain basic rules concerning haul roads and vehicle selection are followed. This paper does not try to determine the maximum tonnages and haul distances which would finally make road hauling uneconomical...

Jan 1, 1966

7253. Arkhangelsky, A. D. Geological Results of the General Magnetometric and Gravimetric Surveys of the U.S.S.R. (in Russian). Internat. Geol. Cong., Rep. of the 17th Sess., Moscow-Leningrad, 1937, vol. 1, Moscow, 1939, pp. 241-250. Geophysical methods have been used in the U.S.S.R. to study geological problems. Such is the case of the general magnetometric and gravimetric surveys made jointly by the Main Physical Observatory, Main Office of Geodesy and Cartography, Institute of Terrestrial Magnetism, and Main Office of Geophysical Prospecting. Some institutions supplied the data, others processed them. Geological interpretation was made by the author. (1) Regional magnetic anomalies in the plains of the U.S.S.R. are traceable to the folded substratum overlain by sedimentary rocks; these anomalies promise to yield data characterizing the substratum in respect to strike and depth. Anomalies in European U.S.S.R. suggest that the pre-Cambrian -substratum of the East European plateau consists of elements of different ages. In the most ancient of them, the main strike of the folding is almost latitudinal; in the younger ones, the folding is oriented in the meridian. (2) The Faye and Bouguer gravity anomalies are caused primarily by differences in the density of rocks composing the upper crust and by unknown processes producing changes of density at depth; these processes are not connected with isosiatic movements. The anomalies in the European and adjoining Asiatic U.S.S.R. promise to indicate the volume of the pre-Cambrian massif of the east-European plateau and to clarify the interrelations between the Ural, Tian-Shan, and Donetz Basin folding systems, as well as between the Dnieper-Donetz depression and the mid-European depression. - V. S.

Jan 1, 1944

For its Hellyer project, Aberfoyle sought to establish a high degree of confidence in all aspects of the feasibility study. Accurate capital and operating cost estimates were an integral and essential part of the study. They were crucial to the evaluation of investment returns and competitive cost position. Both the capital and operating cost estimates were based on the comprehensive data available from exploratory mine development, pilot plant operation, detailed engineering design, and Aberfoyle's experience in the area. One of the principal project objectives was to maximise operating productivity and efficiencies and this created capital cost alternatives which needed thorough cost-benefit assessment. The estimated capital cost for construction and development was $90 million, and the project was completed virtually on time for $80 million. The first year's experience indicated that operating costs would settle to well within 10% of the estimates.

Jan 1, 1990

By E. O. Wagner, A. W. Koth, V. F. Parry, J. B. Goodman, D. C. Gernies

"INTRODUCTION A commercial-scale pilot plant for gasification of lignite was completed by the Bureau of Mines at Grand Forks, N. Dak., early in 1945 as an extension of work on, gasification of low-rank coals initiated in a smaller pilot plant at Golden, Colo., during 1943. The development work on this plant was part of a program to make reducing gases from lignite for beneficiation of low-grade iron ores. The background of this project has been described in a previous report. 6/ The present report describes the major features of the large pilot plant, a record of experimental operating data during five runs totaling about 1,800 hours, .and results of experimental work on the small pilot plant at Golden during 1945.This project has created considerable interest because heretofore there has been little success in several attempts to gasify American lignites.Furthermore, the methed employed is a departure from conventional processes. Natural lignite is fed continuously into an externally heated annular retort and is converted into water gas by heat conducted through alloy metal walls. Because of the high thermal conductivity of metal, the rate of gasification is several times greater than can be obtained when employing refractory walls. The process has been developed to a part that leaves one major question unanswered. - ""How long will the alloy last under practical operating conditions?"" - and this question has aroussed the interest of the Coal Division of the Bureau of Mines in this new development.Owing to the high reactivity of the lower-rank fuels, the water-gas reactions proceed rapidly at medium temperatures of 1,650°F. to 1,800°F. These temperatures can be attained in a reaction zone heated by external means when the temperature of the retort does not exceed 1,900°F. while transferring from 6,000 to 7,000 B.t.u. per hour per square foot. Such rates are several times greater than can be attained through refractory walls as employed in the Didier-Bubaig 7/ and the Ahrens processes. 7/ Therefore, the investment cost of the gas generator per unit of capacity will be considerably lower for the higher-capacity process. However, the maintenance cost of metal vessels will be higher than that of refractory retorts, and this emphasizes the importance of large-scale operation to determine the life of an alloy retort under practical operating. conditions."

Sep 1, 1947

By A. W. Dougall, T. M. Mmola

"SynopsisThe global resources and commodities market has become highly competitive. While southern Africa’s abundance of minerals resources is still unrivalled, it has lost considerable dominance in terms of production output. The sustainability of southern Africa’s mining industry is increasingly becoming dependent on its ability to manage the performance of its operations well. A valuable tool for monitoring and managing performance is the use of key performance areas (KPAs) – which are those areas of performance that are reflected explicitly or implicitly in the vision and strategies of an organization and reflect the critical success factors of the organization. This paper is a review of key performance areas in the southern African mining delivery environment. The KPAs discussed in this paper have been identified by comparing KPAs of the several mining houses engaged in surface mining operations in southern Africa and then extracting those KPAs that are common to most of them. Although the authors support the view that each organization should develop KPAs to specifically fit its needs, the study reveals that five KPAs – safety and health, costs, product quality, fleet management, and delivery should form a default list that covers the key areas that any organization should consider when choosing KPAs. KPIs are those subsets of KPAs that we measure in order to manage KPAs.IntroductionThe global resources and commodities market has become highly competitive. While southern Africa’s abundance of minerals resources is still unrivalled, it has lost considerable dominance in terms of production output. The sustainability of southern Africa’s mining industry is increasingly becoming dependent on its ability to manage the performance of its operations well. A valuable tool for monitoring and managing performance is the use of key performance areas (KPAs), which are ‘those areas of performance that are reflected explicitly or implicitly in the vision and strategies of the organization’ (Barker, 1997). The terms KPA and KPI (key performance indicator) are often used interchangeably – whether correctly or erroneously is debatable. KPIs are ‘quantifiable measurements, agreed to beforehand, that reflect the critical success factors of an organization’ (O’Neill, 2007). Each KPA probably has multiple KPIs associated with it."

Jan 1, 2015

By J. Ndlovu, R. Hundermark

Keywords: Anglo Platinum, Polokwane Smelter, pyrometallurgy, metallurgy, platinum, platinum group metals, PGM, base metals, nickel, copper, sulphide smelting, electric furnace, Söderberg, electrode, matte, slag, flash drying, matte casting, crushing, granulation Anglo Platinum?s Polokwane Smelter is situated outside Polokwane, in the Limpopo Province of South Africa. Wet concentrate is received from various concentrators along the Eastern Bushveld Complex, with 60% of the total concentrate received being from UG2 reef and 40% from Merensky reef. The concentrate grade differs from 35 g/t to 110 g/t, and from this a PGM-rich nickel-copper matte is produced. The furnace matte produced is sent to the ACP converter in Rustenburg. The Anglo Platinum concentrators, from which concentrate is received, are Lebowa, Potgietersrus, Amandelbult, and Union, as well as the joint venture Modikwa. The concentrate is fed to two flash dryers that utilise coal-fired, fluidized-bed hot-gas generators to produce the hot gas that will drive off the moisture, leaving a fine bone-dry concentrate to feed to the furnace. The dry concentrate is pneumatically transferred to a storage silo before being transferred to the feed bins situated above the furnace. Lime, if required as a flux, may be transferred separately to the furnace feed bin by a pneumatic system. The concentrate and lime is fed from the feed bins onto airslides and into the furnace. The feed rate is automatically controlled, and the system is set up to optimize the power to feed ratio. The electric furnace is nominally rated at 68 MW and the power is transferred into the furnace using six 1.6 m diameter Söderberg electrodes. Concentrate is melted by energy generated when electric current passes through the electrodes and resistive slag layer. On melting, two immiscible phases form: slag and matte. Furnace matte, containing the bulk of the base metal sulphides and PGMs, is denser than slag and collects naturally at the bottom of the furnace. The furnace is constructed of a combination of refractory brick and water-cooled copper coolers. The furnace sidewalls and hearth are cooled by air drawn through the area by cooling fans. The hearth and matte zone are constructed from refractory bricks. The copper coolers reside only in the slag zone of the furnace along the entire perimeter of the furnace. One staggered row of plate coolers is installed above the waffle coolers along the perimeter of the furnace. Matte is tapped periodically through one of the two matte tapholes into 35-ton refractory-lined ladles, and cast into matte ingots on a casting machine. The cooled matte is discharged onto a concrete slab for cooling and then transferred by front-end loaders to the crushing plant. The matte is first crushed in a jaw crusher and then a Rhodax cone crusher to a size of 2 mm before it is loaded into tankers and sent to the ACP. The low-grade slag is tapped from the furnace through a water-cooled copper insert into a short water-cooled copper ?hot launder? which discharges into a granulation cold launder. The granulated slag slurry flows to three rake classifiers, from which the coarse slag is discharged onto a conveyor to dewatering silos. The dewatered slag from the silos is dumped. The water overflow from the classifiers reports to two thickeners where the water is clarified, and the thickener underflow is returned to the rake classifiers. The smelter is a zero-effluent plant and all the water from the slag silos where the slag is dewatered, as well as the storm water, is pumped to the process water dam from where it is pumped back into the plant for process water use. The off-gas from the furnace is drawn through a forced draught cooler and into a high temperature baghouse. The dust collected in the baghouse is pneumatically transferred into bins above the furnace and is fed via the concentrate airslides. The cleaned off-gas is then vented through the main stack.

Jan 1, 2006

By S. Kelebek, C. Gibson, M. Aghamirian

"Global niobium production is currently dominated by three operating mines, Araxá and Catalão in Brazil, and Niobec in Canada. Although over 90% of the world’s niobium is mined and processed in Brazil, a number of high grade niobium deposits exist worldwide. The advancement of these deposits depends largely on the development of operable beneficiation flowsheets. Pyrochlore, the primary niobium bearing mineral, is typically upgraded by flotation with amine collectors at acidic pH, and often includes a reverse gangue flotation stage prior to pyrochlore flotation. Challenges related to the flotation of niobium ores include complicated flowsheets which require pulp pH adjustment from alkaline to highly acidic conditions, difficulty recovering fine grained pyrochlore, producing concentrates that meet specifications, and minimizing the losses to reverse pre-flotation gangue concentrates, slimes, and tailings. The specific flotation challenges associated with different niobium ores are discussed. In some cases, the use of alternate flotation reagents and flowsheets are required to achieve reasonable gangue rejection and maximize recovery, particularly of very fine pyrochlore grains.INTRODUCTIONNiobium is a transition metal primarily found in the oxide minerals pyrochlore (Ca,Na)2Nb2O6(OH,F) and columbite (Fe,Mn)Nb2O6. After physical upgrading, pyrochlore is converted into a ferroniobium product which is largely used in the production of high strength low-alloy steel. Today’s niobium market is dominated by the Companhia Brasileira de Metalurgia e Mineração, whose Araxá operation produces over 80% of the world’s niobium. There are two smaller operations, Niobec in Canada, and Catalão in Brazil, which account for the remaining production (U.S. Geological Survey, 2013).A number of niobium deposits exist globally, their advancement dependent on the development of operable beneficiation flowsheets. Pyrochlore, the predominant niobium bearing mineral, is commonly upgraded by flotation with cationic amine collectors in acidic pulp conditions. However, many niobium ores do not respond to this typical flowsheet because they are either highly refractory or niobium deports to other minerals such as columbite, ilmenite and rutile, which are not as responsive to this reagent scheme (Bulatovic, 2010). Other reagents, such as alkyl hydroxamic acid, have been studied extensively and while promising, are not without their own inherent issues including lack of selectivity over iron oxide minerals, and high addition rates (Ni & Liu, 2012; Gorlovskii, Eropkiny, Koval, Strelstin, Khobotova, & Shtchukina, 1968; Gorlovskii, Khobatova, & Shtchukina, 1966).It is the aim of this report to highlight the challenges associated with the physical upgrading of niobium minerals through examination of the flotation behavior of two different ore systems. In order to overcome these challenges it is important to thoroughly understand the limitations of typical flowsheets/reagent schemes currently employed in operation."

Jan 1, 2015

By Silas Li, Michael Harris-Gifford, Jason Choi, William H. Hansmire

Twin-bore tunnels with a single track in each tunnel are currently the most common subway tunnel configuration. However, a large diameter single-bore tunnel with two tracks separated by a center wall could replace twin-bore tunnels and introduce potential benefits but require necessary modifications to the subway system. In particular, a nominal 40-foot-diameter single-bore tunnel can be considered in lieu of nominal 20-foot-diameter twin-bore tunnels. This paper discusses the benefits and modifications introduced by replacing twin-bore tunnels with a single-bore tunnel, with respect to various elements of the subway related to the tunnel, stations, track, and tunnel systems.

Dec 1, 2022