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By S. Soren, K. Deo Mehta, G. Jha

The paper aims to find a sustainable sintering process through cleaner means, in order to achieve the desired outputs with zero environmental impact. At present, sintering has no control over its environmental impacts. These impacts occur when fossil fuels such as coal and coke are used in metallurgical sintering. These fossil fuels are undoubtedly effective in their assigned reducing task, but simultaneously emit about 220 kg of GHGs (greenhouse gases) per tonne of sinter produced. The current attempt has focused on reducing these emission levels by switching from coke to a renewable fuel source such as biomass. Biomass is a cleaner fuel source with a significantly lower emission footprint compared to coke. In the experimental campaign, sawdust and wood-charcoal were used in substitution for coke. Experiments were designed in a progressing substitution to obtain the maximum substitution ability without compromising the sinter properties. Sawdust as raw biomass source optimally substituted 10% of coke. The substitution levels were remarkably higher (30%) for wood charcoal due to its low moisture and high calorific values. A fuel blend with 30% biomass and remaining coke succeeded in producing a sinter with acceptable quality indices. Sinters prepared with or without biomass contribution shared an identical phase transformation and microstructures. In order to to maintain consistency in the process, every effort was made to maintain all operational and process parameters identically, except for the fuelling framework. Keywords: Sintering, sustainable, biomass, GHGs, fuel substitution

Jan 1, 2020

By J. A. C. Diering

All underground openings are three-dimensional, and in many cases it is not valid to analyse two-dimensional sections. This has been a problem in the past owing to the magnitude and complexity of three-dimensional stress analysis. In this paper a three-dimensional boundary element approach, using non-conforming quadratic elements, which is specifically applicable to underground excavations, is described. In addition, the philosophy of application of stress analyses is discussed. Three case histories are presented to demonstrate the practical nature of the approach as a tool in solving mining problems.

Jan 1, 1987

Palabora Mining Company is operating a copper block cave mine in the Limpopo Province of South Africa. The mine was developed in the late nineties with a target ore reserve of 220 million tonnes of carbonatite at 0.69% copper. At 30,000 tonnes per day the operation had an estimated life of mine of approximately 23 years. Cave induced open pit side wall failure occurred in late 2004 with 130 million tonnes of waste reporting to the bottom of the open pit. This failure resulted in potential sterilization and dilution of the ore reserve and subsequent reduction in life of mine. Physical and numerical modelling indicated potential losses of up to 30% of the original ore reserves. Cave Management tools based on throughput principles are utilized to manage the failed waste ingress and ensure an optimum life of mine. Extracting the maximum rate possible, as soon as draw points are commissioned, with the view of achieving the highest production in the short term, is any mining operation?s goal as more money is required sooner. This approach, of course, improves the cash flow but presents a risk in the cave propagation profile as overall draw control is compromised. At Palabora it is evident from a Mineral Resource Throughput Management (MRTM) point of view that maximum production in the short term is not the solution. If high production rates are achieved and the profile of the cave is not manage correctly, this results in failed waste reporting to draw points earlier with adverse effects on grade. This would effectively and significantly shorten Palabora?s Life of Mine. The applications of good MRTM principles are required to assist the value chain for sustained profitability. The process commenced with change management paving the way for a better understanding and a holistic view for all role players.

Jan 1, 2007

By K. Basaure, S. Palma, R. L. Castro

"Mud rushes are sudden inflows of fine wet ore within underground mines; they can harm people and equipment as well as cause production delays. Block and panel caving mines are prone to mud rushes in their production levels, with mud flowing from the drawpoints. Different mines have developed methods to control drawpoint conditions and thus perform safe ore extraction. The aim of this study is to develop a decision matrix and apply it to samples from drawpoints for the Diablo Regimiento mine, Codelco, El Teniente Division (Chile). The Abrams’s Cone was used to perform slump tests on mud samples and characterize their consistency by setting the saturation degree and packing. The unconfined strength for different saturation degrees and packing was then evaluated. The results show that conditions for fluid response in slump tests correspond to conditions for low strength in unconfined compression tests. We also found the water content value at which the flowing consistency changes from plastic to soft behaviour according to the slump test classification for each sample. This value varied between 12.2% and 16.9%, depending on the sample tested. Finally, a fluid security factor was defined in order to design an extraction pattern for the samples tested, based on the water content. We conclude that a very densely packed ore is not prone to flowing. The flow properties depend strongly on the specific properties of the ore tested, and consequently a specific criterion should be developed for different types of ore at a particular mine. The criterion developed for Diablo Regimiento ore provides critical water content values from 11% to 15%, depending on the ore type. IntroductionBlock/panel caving has the lowest cost and the highest production rates of all underground methods (Heslop, 2000; Hubert et al., 2000). Good management of the risks is important to ensure the security of personnel and infrastructure. Operational risks for the method include rockbursts, air blasts, and mud rushes (Hubert et al. 2000). A mud rush is defined as a sudden inflow of mud from a drawpoint or other underground opening into the mine (Butcher, Joughin, and Stacey, 2000; Butcher, Stacey, and Joughin, 2005). Mud rushes can cause dilution, production delays, and damage to equipment as well as injuries and fatalities. The elements necessary for mud rushes are mud-forming material, disturbances, a discharge point, and water (Butcher, Joughin, and Stacey, 2000). All these elements are present in block/panel caving mines. Mud-forming material comes from ore milled by the caving process; water comes from underground and surface sources. Drawpoints in the production level are discharge points for the mud. Finally, several types of disturbance could trigger the mud rush phenomena: blasting, earthquakes, operation of equipment (Call & Nicholas, Freeport McMoRan, and Hydrologic Consultants, 1998; Jakubec, Clayton, and Guest, 2012), arch collapse above drawbells (Butcher, Joughin, and Stacey, 2000; Butcher, Stacey, and Joughin, 2005; Jakubec, Clayton, and Guest, 2012) and ore drawing (Butcher, Joughin, and Stacey, 2000; Butcher, Stacey, and Joughin, 2005; Call & Nicholas, Freeport McMoRan, and Hydrologic Consultants, 1998; Jakubec, Clayton, and Guest, 2012)."

Jan 1, 2017

By F. H. Deist, H. D. Small, J. D. Austin, L. F. Duvel

The development and application of a computer system to aid in the planning of production in mines by F. H. DEIST, L. F. DUVEL, J. D. AUSTIN, and H. D. SMALL Published in the Journal, December 1972. Discussion by J. W. WILSON and R. W. HAZELL Although the Mine Planning Simulation Programme was originally developed as an aid to coal-mine planning, its extreme flexibility has enabled us to simulate gold-mine planning with equal facility. The initial application of the Mine Planning Simulation Model was in a small area of a large Free State gold mine where the primary aim was to assess the applicability of the programme to the gold-mining situation.

Jan 12, 1972

By M. J. Nicol

The results are reported of experiments on the influence of a number of variables on the rate of loading of gold cyanide onto activated carbon and the equilibrium capacity of the carbon. These variables include the concentrations of gold and free cyanide in solution, the pH value and ionic strength of the solution, the concentration of organic compounds in the solution, the temperature, the particle size of the carbon, and the mixing efficiency. The experiments were carried out under approximately steady-state conditions to simulate the conditions in a countercurrent carbon-in-pulp absorption plant. In addition, the results are reported of experiments to establish whether the absorption of gold cyanide onto activated carbon is a reversible process.

Jan 1, 1983

By G. van Aswegen, J. D. Bosman

Gold bearing deposits in the Witwatersrand consist predominantly of narrow tabular conglomerates with quartzite host rock. Extensive mining of these deposits at depth results in significant stresses being induced in the periphery of the tabular excavations. Where remnant pillars are created excessive stresses have resulted in the dynamic failure of the abutments. At the Hartebeestfontein Mine four brittle shear zones have recently been exposed and investigated. For most part, they represent abutment failure into the weaker hangingwall strata, striking parallel to the abutment, dipping towards the stope, with a sense of displacement consistent with closure of the stope from above. These shear zones are similar to those described in various publications by Ortlepp2: they are essentially brittle shear zones characterised by the presence of very finely comminuted ‘rock flour’ on an obviously freshly sheared surface. The angle between the shear zone wall and extension fractures indicate the sense of shear and the intersection of the extension fractures with the shear zone walls define a strong lineation normal to the direction of slip. It is suggested that these structures are referred to as ‘mining induced dynamic brittle shears’. Les gisements d’or du Witwatersrand consistent principalement de conglomérats tabulaires peu épais contenus dans une quartzite mère. Une exploitation poussée de ces gisements en profondeur a pour résultat l’apparition d’importantes contraintes induites à la périphérie des excavations tabulaires. Là où des piliers restants sont créés, des contraintes excessives ont eu pour résultat une rupture dynamique des appuis. Dans la Mine d’Hartebeestfontein, quatre zones fragiles de cisaillement ont été récemment exposées et étudiées. Dans la plupart des cas, elles représentent une rupture des appuis dans les strates moins solides de la paroi suspendue, orientée parallèlement à l’appui, inclinée en direction du chantier d’abattage avec une direction de mouvement consistante avec l’affaissement du toit de la galerie d’abattage. Ces zones de cisaillement sont similaires à celles décrites par Ortlepp2 dans diverses publications: Ce sont essentiellement des zones cisaillement fragiles caractérisées par la présence de « farine de roche » très finement réduite sur une surface évidemment fraîchement cisaillée. L’angle entre la paroi de la zone de cisaillement et les fractures de tension indiquent le sens du cisaillement et l’intersection des fractures de tension avec les parois de la zone de cisaillement définit une forte délinéation perpendiculaire à la direction du glissement. Il est suggéré que l’on se réfère à de telles structures comme des « cisaillements dynamiques fragiles induits par l‘exploitation ». Goldablagerungen im Witwatersrand bestehen überwiegend aus schmalen tabularen Konglomeraten mit Quarzitwirtgestein. Der umfangreiche Abbau dieser Ablagerungen in der Tiefe resultiert in signifikanten Drücken, die auf die Peripherie der tabularen Aushöhlungen wirken. Wo Restpfeiler hergestellt wurden, resultierten exzessive Drücke im dynamischen Versagen der Auflager. In der Hartebeestfontein Grube wurden kürzlich vier spröde Scherzonen freigelegt und untersucht. Zum größten Teil representieren sie Auflagerversagen im Stratum der schwächeren Stollenfirste, parallel zum Auflager verlaufend, in Richtung Abbaustelle abfallend, mit einem gewissen Grad an Verschiebung konsistent mit der Schliessung der Strosse von oben. Diese Scherzonen sind jenen ähnlich, die in den verschiedenen Publikationen von Ortlepp2 beschrieben werden: sie sind im Wesentlichen spröde Scherzonen, charakterisiert durch das Vorhandensein des sehr fein pulverisierten „Gesteinsmehls“ auf einer offensichtlich frisch abgescherten Oberfläche. Der Winkel zwischen der Scherzonenwand und den Bruchverlängerungen deutet die Richtung der Scherung an und die Kreuzung der Bruchverlängerungen mit der Abscherzone grenzt sich gegen eine starke Lineation normal zur Richtung der Lehne ab. Es wird vorgeschlagen, daß diese Strukturen als `vom Abbau verursachte dynamisch spröde Scherungen’ bezeichnet werden.

Jan 1, 2003

By A. P. Van Wyk, G. Pirnaar

The effect of vanadium on the susceptibilities of chromium and chromium-nickel steels to reversible temper embritlement was evaluated. The steels investigated were approximately equivalent to steels of as 970 Grade 708A42 and as 970 Grade 817M40, the major difference being that molybdenum had been replaced by vanadium in various concentrations. It was found that the vanadium had a detrimental effect and, in al the concentrations tested, enhanced the susceptibility to reversible temper embritlement of both the chromium and the chromium-nickel steel. The detrimental role played by vanadium in reversible temper embritlement is explained by a consideration of the effect of vanadium carbide, as well as of free vanadium, in solid solution in ferrite.

Jan 1, 1987

By D. K. Weatherley, T. P. Oladele, L. B. Bbosa

Comminution is a critical stage of mineral processing which aims to reduce the size of ore particles through breakage, consequently increasing the likelihood of liberation of valuable minerals. Several factors, such as pre-existing cracks, mineralogical composition and particle size, are known to affect ore breakage behaviour. While investigating the contributing role of each factor is essential to improve understanding of breakage, isolating individual factors by experiment is typically impractical. Numerical techniques such as the Bonded Particle Model-Discrete Element Method (BPM-DEM) have been developed as a means of studying closely controlled breakage conditions. This study evaluates this technique in modelling the effect of pre-existing cracks breakage behaviour for synthetic ore specimens during impact breakage in a short impact load cell (SILC). Synthetic ore was modelled as a collection of discrete entities connected by brittle-elastic beams. Model parameters were calibrated against macroscopic material properties of a conventional homogenous synthetic ore. The extent of pre-existing cracks within specimens was prescribed by varying discontinuities between the connecting entities. An approach with the insertion of seed points was used to model the mineralogy of the specimens. Results showed that pre-existing cracks resulted in a decrease in the fracture force as the crack density increased. Additionally, the irregularity of fracture patterns within rock specimens after impact became more apparent as the crack density increased. The findings were used to emphasise some of the attributable causes of the variation observed between ore hardness characterisation tests and in situ ore behaviour. Keywords: Element Method; Bonded Particle Model, synthetic rock sample; rock fracture, preexisting cracks

Jan 1, 2020

By S. M. Spottiswoode

Rock mechanics benefits of stope support in deep-level mining originate principally from the ability of the support to resist relative movements between the hangingwall and the footwall. While the role of stabilizing pillars and backfill on reducing the energy release rate (ERR) at the face is well understood on the basis of elastic behaviour of the rock mass, the effect of support on controlling movement of fractured ground around deep-level stopes has not yet been thoroughly studied. As part of a programme of studying the behaviour of this fractured ground, geophones were placed 10 m from the face on the roof and floor of a longwall stope in which packs were used as supports. It was found that differential movement between the roof and floor generally exceeds average stope movement during a seismic event. Conventional support therefore transmits very little seismic energy; this is expected to remain true until complete convergence occurs. In backfilled stopes, backfill can become at least as stiff as the fractured wall rock. This will have the effect of reducing the length of the hanging- wall beam. Amongst other rock mechanics considerations, a shorter hanging- wall beam should: (i) increase the resonance frequency of the hangingwall and thereby decrease the levels of vibration for seismic events with M > 1,0; (ii) reduce absolute vibration levels by increasing the mass of rock immediately adjacent to the stope which is shaken by seismic waves; and (iii) reduce differential movement between the roof and floor and thereby reduce cyclic strain on the backfill and on gully support.

Jan 1, 1988

By G. A. Georgalli

A control philosophy to be applied by Lonmin to maintain constant electrode immersion in their no. 1 circular smelting furnace is put forward. This control philosophy is facilitated though a combination of thermodynamic models used to predict the slag composition and empirical correlations to estimate the furnace geometric factor as well as the slag conductivity. The model can determine the operating resistance set point required in order to maintain a desired electrode immersion as a function of the feeds to the furnace, slag level and temperature. If modifying the resistance set point cannot maintain the desired immersion without moving out of the furnace P-V-I operating envelope, the model can be used to determine the amount of slag modifiers (silica or lime) required to shift the slag chemistry back into a region where the immersion can be controlled though resistance set point manipulation.

Jan 1, 2008

By O. Tshifungat, A. Luzanga, N. Goel, S. Sandoval

"Tenke Fungurume Mining Company (TFM), located in Katanga Province, Democratic Republic of Congo (DRC), began production of copper cathode and cobalt hydroxide in March 2009. Copper production capacity was 115 000 t/a utilizing a semi-autogenous grinding, agitated leach, countercurrent decantation process followed by solvent extraction and electrowinning (SX-EW). Cobalt production capacity (as cobalt hydroxide) was 8000 t/a utilizing a three-stage purification/precipitation process treating raffinate from the copper SX-EW process. Recently a phase II expansion was completed, in which copper electrowinning capacity was doubled and four stages were added to the SX plant. Corresponding enhancements were made to mining, milling, and processing to bring production to 195 kt/a copper and 15 kt/a cobalt. An additional sulphuric acid plant is currently being constructed, adding 1400 t/d sulphuric acid to bring total acid production to 2250 t/d for use in leaching at TFM.Initial copper electrowinning results were satisfactory, but cathode quality and current efficiency decreased as copper production increased. The Freeport-McMoRan (FCX) Technology Centre and TFM personnel began operation of two bench-scale copper electrowinning cells in the tankhouse to diagnose chemistry effects on the cathode and anode. The results indicated an excellent electrolyte chemistry, producing finely crystalline copper at 430 A/m2 current density with 98% current efficiency. Attention next turned to the physical factors of the electrowinning cells, including electrode insulator geometry and performance. A demonstration cell was selected in the tankhouse and was operated with varying cathode and anode insulator geometries. Operation of the demonstration cell with a three-side cathode edge strip, A-style anode insulators, and an improved cell furniture design increased current efficiency from 77% to 89% at 400 A/m2 current density. Cathode quality improved correspondingly. TFM is now operating at 92% current efficiency."

Jan 1, 2016

By B. P. Watson

In South Africa, the Bushveld platinum group metal deposits are unique in that they are the largest platinum group metal deposits in the world. These deposits occur as two distinct stratiform tabular orebodies and strike for many hundreds of kilometres. Mining is extensive, with depths ranging from close-to-surface to 2 000 m. The mining method is a variation of planar open stoping. Pillars are widely used to support the open stopes. Crush pillars are commonly used in this role, where the residual pillar strength provides the required support resistance to stabilize the stoping excavations. This paper describes the direct measurement of stress within these crush pillars, and describes the stress?strain behaviour of these pillars. These findings indicate that the measured pillar failure stress is too high. The implications of these findings are that roof damage and pillar bursting could occur, and some examples of this type of damage are shown. The paper concludes that this problem could be mitigated if the three dimensional pillar geometry were modified and the pillars were cut smaller.

Jan 1, 2007

By N Morales, G Nelis

Short-term plans must meet the long-term goals while accounting for shovel selectivity, operational space, and scheduling restrictions. There are several approaches that deal with these issues to deliver a feasible short-term schedule. This paper compares two approaches based on mathematical programming to address the short-term mining cut definition and scheduling in an integrated way. They work by scheduling extraction to comply with long-term production goals, but they differ in how mining cuts are modelled. The first approach schedules Selective Mining Units (SMUs) but aggregates them around predefined locations. The second approach does not work on SMUs; it decides directly what mining cuts are to be used from a pre-defined set of shapes. It assigns a destination and schedules the extraction for the selected mining cuts. We applied both approaches in a real case study to compare their results in terms of profit, mining cut design, expertise needed, profit, and optimisation performance.

Sep 1, 2021

By Franz E. Pawlek

Synopsis This is a summary of investigations in the field of pressure leaching carried out at the Institute for Extractive Metallurgy of the Technical University of Berlin-West during the last seven years. The efficient dispersion of gas in a liquid and its dissolutions as well as the events at the phase boundary when leaching ZnS are reported. At the end two examples are given for the use of pressure leaching in the processing of smelter by-products. INTRODUCTION About 20 years ago, experiments were independently started in Canada and in Berlin on the subject of dissolving sulfidic ores without preliminary treatment by means of leaching under increased oxygen pressure. In Canada ammoniacal solutions were mainly employed, and F. A. Forward and his co-workers were successful in applying this process on a large industrial scale1, 2. This process is used in Fort Saskatchewan to produce about 15,000 tons of Ni per year. In Berlin, investigations were started on leaching with acid solutions3, 4. This process had also been used on a large industrial scale in the D.S.A., but the choice of unfavourable working conditions led to great difficulties, and to giving up this process for industrial use. Later on in Berlin work was carried out especially on the kinetics of pressure leaching 5, 6, 7, 8, 9 and its applications to the processing of smelter by-products, in particular of mattes and speisses10, 11, 12, 13, 14. The dissolution of ores and by-products can be subdivided into four steps: 1. the efficient dispersion of the gas, so that its dissolution occurs over as great a surface area as possible, since only the dissolved oxygen reacts with the solid, 2. the suspension of the solid, so that the entire surface can take part in the reaction, 3. the phase boundary reaction, 4. the removal of the reaction products. The first two steps belong to the field of process engineering and will be discussed from this point of view. The aim of these experiments15 was to find out first of all the limits of efficiency of self-sucking stirrers in solutions of low viscosity with regard to the mass transfer. In addition, the effect of ultrasonics in the same reaction system was studied. The experiments were carried out in a lucite container as shown in Fig. 1. Four baffles of 0.1 d width were in the stirring vessel. The height of these baffles was adjustable. For every experiment one litre of liquid was used. An infinitely variable direct current motor, which was fixed with a hollow shaft, served as the driving power for the stirrer.

By Steve Konieczki

[ ] Great Basin Gold Ltd. (GBG) owns the Hollister Mine and operates it through the wholly owned subsidiary Rodeo Creek Gold, Inc. The property is located in northern Nevada on the Carlin Trend gold belt within the historic Ivanhoe Mining District. The mine is located on the western edge of the Northern Nevada Rift, a major north-northwest trending structural zone that has localized middle Miocene volcanic rocks and numerous epithermal gold deposits. The precious metal deposits are low-sulfidation epithermal systems characterized by banded quartz veins with electrum and silver selenides. The Hollister Mine will employ the selective underground mining method of cut and fill. This will be combined with split shooting (resuing mining method) to optimize recovery and reduce dilution. The focus of mining will be on extraction of ?ounces? from the narrow-veins rather than trying to maximize tons. The initial drifts will be excavated along the strike of the vein and then capped with a substantial concrete sill pillar. Back stoping with hand held drills will be used to selectively mine the vein. The stopes will be backfilled with waste rock material. With every lift, an 18 inch concrete cap will be poured onto the backfill waste rock to prevent gold values from being lost into the fill. Targeted ore production rate will be over 400 tons per day. With current ore reserves, project life is estimated to be around eight years; however, Hollister geologists view the site as one of the most prospective systems for bonanza vein deposits in Nevada. Approximately $14

Jan 1, 2008

By P. R. Jochens, D. D. Howat, C. W. Wolhuter

Prof D. D. Howat (Member): I am a man with a grievance this afternoon. The other authors of the paper entered into a conspiracy to ensure that I presented the paper, and finally even had the audacity to say that this was the least that could be expected from me as I had done so little of the other work. Looking at the distinguished audience here that includes some figures who have forgotten more about cement that some of us have ever learned, I realise how right the other authors were. I feel very much cast in the role of the fool who was pushed in where angels fear to tread. I hesitate to say, Sir, that I would not use the word angels in any definition of the authors of this paper. In presenting the paper, may I try to put forward a few simple points about cement that are probably as little known to the majority of the audience as they were to me-at least until very recently. Concrete, made by mixing Portland cement, sand, crushed rock and water is the most widely used constructional material in the world. Generally the production of actual Portland Cement is about two-thirds that of steel and by the time this is formed into concrete the total is about five times the world tonnage of steel produced. A very rough approximation suggests that world consumption of concrete is about one ton per person per annum. Apart from fresh water mankind consumes no other material in such quantities. What mankind originally required was a mortar to be used for joining blocks of stone or bricks. Such mortar was originally made from limestone burned in a kiln and then mixed with water. Production of mortar of this type began over 5,000 years ago. Some samples from pyramids in Egypt have shown this mortar to be in good condition although over 4,500 years old. The progress of technological development was very slow until 1824 when an English bricklayer, Joseph Aspidin, patented a new mortar material which he made by burning limestone and clay together in his kitchen stove. Aspdin called his product 'Portland Cement' because, in colour, it resembled a well-known building stone quarried in the Isle of Portland on the south coast of England. Aspid's same types of raw materials are used today in modern cement production and the product is still referred to as Portland cement. A lime-containing material such as chalk or shell and a clay-type material such as shale, slate or clay itself, are mixed in the required proportions, ground and fed to a kiln where the minimum burning temperature is about 1,500°C. The raw materials react to form hard chunks of a material known as 'clinker'. About 5 per cent of gypsum is added to the clinker and the mixture is then ground to a very fine powder, the function of the gypsum being to control the rate of setting of the concrete. The key oxides in cement production are lime, silica, alumina and iron oxide. The possible reactions occurring during and the products of the burning process are shown in Table I.

Jan 12, 1968

By Y. Tanaka, H. Miki, Y. Aoki, H. Okamoto, G. P. W. Suyantar, T. Hirajima

Due to environmental limitations and the scarcity of freshwater, flotation using seawater or desalinated seawater as process water becomes more common for mining operations. On the other hand, seawater has been reported to affect the flotation behavior of sulfide minerals. Few reports are available on influence of pyrite depression, recycled seawater and lime consumption on flotation by using seawater. In this study, the flotation tests were performed using ore sample to investigate the effect of pH with seawater and process water used in plant on the flotation behavior in a laboratory-scale flotation cell and a pilot plant-scale. The results of laboratory scale flotation indicated that copper and molybdenum flotation kinetics decreased with increasing pH. Furthermore, the iron (mainly pyrite) recovery decreased with increasing pH. The pilot plant-scale flotation test showed that water with a higher concentration of magnesium and calcium (plant processes water) consumed a higher amount of lime. Because of low lime consumption and high copper and molybdenum recovery and grade, pH 8 is suitable compared with pH 8.5 for flotation. Keywords: seawater, process water, copper-molybdenum rougher flotation, pH, pilot test

Jan 1, 2020

By L. A. Vermeulen

A simple argument shows that it is impossible for the flight of a rod in contact with a litter bar in a rotary mill to be initiated at that point on the upside of mill rotation when all the forces acting on the rod are in equilibrium. The argument was confirmed experimentally in a study showing that the lift of litter bars is a function of their height, that the rod velocity at the point of departure has a significant radial component, and that the subsequent flight of a rod is along a parabolic trajectory as predicted by elementary mechanics. A quantitative account of the observed effects, which is based on a dynamic theory of the sliding interaction between rods and litter bars, revealed interesting properties of the motions of rods down the surfaces of litter bars when the latter are in uniform circular motion during the rotation of a mill.

Jan 1, 1985

Copper cyanide is a common component of cyanide-treatable precious metal ores. The copper concentration in production heaps can be predicted from laboratory column tests, but the exact correlation is not necessarily intuitive. Generally, heap leach operators like to keep copper concentrations in solution below 300?500 ppm and may note problems with gold recovery and cyanide consumption when copper concentrations exceed this amount. There are several methods of copper removal from cyanide solutions including ion exchange; direct electrowinning; acidification, volatilization, and recovery (AVR); and sulphide precipitation such as the sulphidization, acidification, recycling, and thickening (SART) process. SART involves acidification with addition of soluble sulphide, separation of the resulting copper sulphide precipitate, and addition of lime to re-establish alkalinity prior to returning the solution to the leaching process, recovering both copper and cyanide as valuable products. In principle SART is very simple. Yet some SART plants that have been built may have been unnecessarily complex. This paper explores the basics of SART and makes the case for a simple plant design as applied to the heap leaching circuit.

Jan 1, 2012