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By Randolph R. Dreher, Michael L. Schreder

"In recent years, FFE Minerals has designed, installed and commissioned several large grinding mill circuits. This paper discusses the issues related to the erection and commissioning of large grinding mill circuits. The three circuits discussed are Collahuasi in Chile, Minera Los Pelambres in Chile, and Century Zinc in Australia. These circuits include large SAG and ball mills with both gear and gearless drive arrangements, as well as, both trunnion and shell supported mills. Aspects of these features are discussed with respect to their impact on installation and commissioning. While the normal challenges of mill erection and start-up were experienced, there were also a few unexpected challenges encountered partly due to the large scale of the equipment. An overview of the equipment and design ratings of these circuits is also provided.COLLAHUASIOverviewThe Collahuasi mills were started-up in November of 1998. The site is located in the Andes Mountains at an elevation of 4300 meters. The grinding circuit includes two lines featuring one SAG mill and one ball mill for each line. The circuit features a recycle conveyor around the SAG mills with provision for adding a pebble crusher in the future. The design tonnage of the grinding plant is 60,000 MTPD with an ore BWI of 10.5 kWh/MT.The SAG mills are 32’ x 15’ (1 3’ EGL) with 10,800 HP drives. The drive arrangements are dual pinion through pneumatic clutches. The mills are driven by low speed synchronous motors with CCV variable speed drives. The SAG mills are trunnion supported and use hydrostatic lubrication. Feed to the SAG mills is via a rock-box type feed chute on a cart. The mills discharge on to trommel screens with a bucket wheel to remove the oversize.The ball mills are 22’ x 36.5’ (36’ EGL) with 12,500 HP drives. The drive arrangements are dual pinion through pneumatic clutches. The fixed speed mills are driven by low speed synchronous motors with GE Quadramatic load sharing drives. The ball mills are shell-supported and feature multi-pad bearings with hydrodynamic lubrication. The mills are fed via a rubber lined sump box with feed spout on a cart. The ball mills also discharge on to trommel screens. Twenty-six inch (26”) diameter hydrocyclones are in closed circuit with the ball mills."

Jan 1, 2001

By X. Tang, N. Ripepi

The Langmuir model has been widely used for the coal-methane sorption process through an isothermal test. However, even though the temperature-dependent Langmuir model was recognized theoretically and experimentally, the quantitative relationships between the Langmuir constants a and b and temperature have not been clarified and confirmed by laboratory test data. This paper collects 158 methane-coal sorption data under different temperatures that apply the Langmuir model to the relationships between the Langmuir constants a and b and temperature. Even though the Langmuir model cannot be used to show the physical state of methane adsorption on coal, it is widely used because of its simplicity and the reasonable explanation of its parameters a and b. The results from these published isothermal tests indicate that (1) there is no consistent relationship between Langmuir constant a and temperature, (2) the Langmuir model is a temperature-dependent model whose temperature-dependent feature is mainly shown by the relationship between the Langmuir constant b and temperature, and (3) the relationship between the Langmuir constant b and temperature can be described by both exponential and linear relationships with dried coal. This paper provides a database for the temperature-dependent Langmuir model that can be used for the reservoir simulation of coalbed methane reservoirs and for the coal seam degasification modeling of underground coal mines, considering the heating effects induced by geothermal and heat injection through factors such as hot water, hot steam and ultrasonics.

Jan 1, 2016

By N. D. L. Burger

An effective prototype face support system to minimize rockfalls in rockfall-prone mines was developed and patented during a project sponsored by the Mine Health and Safety Council (MHSC). An experimental development model (XDM) roof support unit was developed and evaluated. Mines with rockfall problems in the face areas can use the roof support system to reduce fatalities and injuries in the face area of stopes. The roof support system consists of two similar support units connected to each other via two crank mechanisms. Each unit consists of a headboard supported by a ?wishbone? structure (top and bottom leg). In the first design a threaded bar with struts similar to a scissors jack keeps the legs apart. (In the final version that was field trialed, the threaded bar mechanism was replaced with a hydraulic cylinder). All the components are manufactured from steel. To move the system forward the first support unit is collapsed, the second remaining in the loaded position. The first unit then hangs from the second unit to which it is connected. The first collapsed unit is then manually cranked forward and prestressed. If the required position is still not achieved, the other unit is released and moved forward. The specification for the system was determined and presented at a workshop with industry participants. Different concepts were developed and evaluated against the system specifications. A technology demonstrator was then developed and tested on surface. The technology demonstrator development process included detail design, building and testing of components and subsystems, design reviews and the building and commissioning of the technology demonstrator. The testing of the technology demonstrator was done in a 500-ton hydraulic press, in a mock-up stope and underground. A risk analysis, in which technical, logistical and economic aspects were assessed, was done to determine the critical areas of the system. During the next phase of the project working prototypes were developed for underground tests and evaluation. This process indicated that it was necessary to adjust the system specifications and a redesign was called for. In an iterative process test units were built for purposes of evaluation, specification verification and field trialing. After the successful conclusion of the trials the equipping of a complete production stope started for purposes of integrating the system into the current mining process.

Jan 1, 2008

By Howard J. Pincus

The purpose of the study is to investigate the effects of fracture filling and associated alteration on the mechanical behavior of rocks. Results are presented for specimens from the Climax and Henderson mines in Colorado and the San Manuel mine in Arizona. The rock types studied from each location are from 1) Climax, medium-grained - granite, fine-grained granite, granite porphyry, schist, and gneiss; 2. Henderson, rhyolite porphyry; 3. San Manuel, quartz monzonite and quartz monzonite porphyry. A description of the rocks was made including: color, texture, mineral composition, composition and fabric of fracture fillings, density and porosity. Unconfined and triaxial compression tests were run on the Climax and San Manuel specimens. Sonic relocities were measured during the compression tests; Shore sceleroscope hardness was measured before the compression tests. Tentative conclusions are that some relation appears to exist between compressive strength and the following properties and features for all the specimens: dynamic Young's modulus, sonic velocity, density, rock composition, and composition and fabric of fracture fillings - veinlets. There appears to be some relation between compressive strength and Shore scleroscope hardness and porosity for quartz monzonite (San Manuel). A relation appears to exist between compressive strength and texture for schist and between com¬pressive strength and extent of failure on fracture fillings for fine-grained granite, schist, and gneiss (Climax). We plan to perform additional analyses of data collected so far and data yet to be collected in the next few months, with emphasis placed on devising a practical index of alter¬ation that can be used in the field. Limitations on interpretations arise from such factors as heterogeneity of the rocks studied, the small range of confining pressures used, and other factors presented in the text.

Jan 1, 1973

Mr G H Park was born and brought up in Zimbabwe, in mining from the University of the Witwatersrand. Starting as a Learner Official at Rand Mines, he went Mr D J van Niekerk was educated at Helpmekaar Boys' High School in Johannesburg and the University of Pretoria. Mr C J C Janse van Vuuren was educated at Klerksdorp High School, Pretoria Technical Colege, the University of Pretoria, and Unisa. He holds the folowing degrees: BSc (I&S), BSc Hons (Met Pta), ND (Met Tech), AEP (Unisa). Prof. R F Sandenbergh was educated at the Afrikaanse Hoerskool Germiston and the University of Pretoria. He holds the degrees of B Ing (Chem), M Ing (Chem), and D Ing (Metal). President's Award Professor Jannie van Deventer, Head of the Department of Metalurgical Engineering at the University of Stelenbosch, has been honoured with one of the five President's Awards for 1990. These awards are made by the CSIR's Foundation for Research Development. G R Bozzoli Awards* Prof. Alan Kemp, Dean of the Faculty of Engineering at the University of the Witwatersrand, has appealed to industry not to cut bursary funding during recessionary times.

Jan 1, 1990

By John Chipman, Peter J. Koros

The distribution of copper between liquid silver and liquid iron, Fe-C, and Fe-C-Si alloys was studied at 1600°C. From the data and the activity of copper in silver obtained from the phase diagrams, the activity coefficients of copper in iron up to 3 atomic pct are log y, = 0.90 — 2.4 N, and y": = 8.0. The effect of carbon up to saturation is expressed as log y = 1.8 Ni. The effect of silicon up to 11 atomic pct is insignificant. SLOWLY rising copper content of steel being produced is focusing attention on the lack of information on the behavior of copper in liquid steel. Chou' measured the activity coefficient of copper in liquid iron through its distribution between liquid silver and iron. He obtained a value of 9.12 at infinite dilution. Chipman' calculated the activity coefficient of copper in liquid iron at 1600°C from the phase diagram and obtained a value of 12. Ameiln and Pfeiffer- investigated the possibility of removing copper from liquid steel through the use of lead, silver, and bismuth as solvents. They also calculated that the activity coefficient of copper in liquid iron would be increased five or sixfold by addition of 4 pct C. Langenberg and Lindsay' are investigating the effectiveness of lead and sodium sulfide in removing copper from liquid iron. The experimental method was similar to that used by Chou' and Chipman and Floridis." Silver, copper, and iron were melted together, the copper distributing itself between the two layers so that at equilibrium a,.,, (in Ag) - a,.,, (inFe) and N,.,, (in Ag) yc,, (inFe) (in Ag). [21 N,.,, (inFe) Thus, from the concentration of copper in each layer and the activity coefficient of copper in liquid silver, the activity coefficient of copper in liquid iron can be calculated. The effect of an alloying element such as carbon on the activity coefficient of copper can also be determined from its effect on the distribution, provided the new element is essentially insoluble in silver. The one serious limitation of this method is the increasing mutual solubility of silver and iron in the presence of a third component. The mutual solubilities of pure iron and silver are negligible, but increase with increasing copper concentration beyond the range investigated. Ag-Cu System—The activity coefficient of copper in silver at 1600°C had to be estimated, since direct experimental data are not available. Kawakami" measured the heat of mixing of Ag-Cu alloys at 1200°C by a calorimetric technique. Scheil' calculated the heat of mixing in this system by use of the phase diagram, assuming that the solution is random and that the heat of mixing is independent of temperature. The terminal solid solution may be assumed to obey Raoult's law within the requirements

Jan 1, 1957

By Eugene O. Binyon

"Belt and shaker conveyors, used singly or as combined units, have for many years been an integral part of many coal-mine haulage systems. The belt conveyor has also been used extensively as a main haulage unit in open-pit iron-ore mines of the Lake Superior district, particularly on the Mesabi Range. Few underground metal mines, however, either in the ferrous or non-ferrous branches of the mining industry, use the conveyors exclusively or even extensively for underground haulage. At present, only five iron-ore mines in the entire Lake Superior district have belt- or shaker-conveyor installations of any significant size. Of these mines, three are in the Crystal Falls-Iron River district of Michigan on the Menominee Range, one is on the Marquette Range near Ishpeming, Mich., and one is on the Mesabi Range near Keewatin, Minn.Results of the studies of tackless haulage in the Lake Superior district will not permit the establishment of any formula to prove or disprove the advantages of belt or shaker systems at all underground mines. The studies do show that trackless haulage can be used to advantage where the proper conditions exist. Advantages are particularly apparent when development can be planned sufficiently in advance of exploitation to permit installation of belts or shakers as part of the development program and where the systems are extended to include haulage in stopes as well as main entries.During the winter of 1947-48, the Bureau undertook a study of all of the underground ore conveying systems known to be in operation in the Lake Superior district for the purpose of making available to the entire mining industry the technologic data developed in the area. Because conditions that led to the installation of the belt and shaker conveying systems were often a function of the character of the deposit or the method of mining, a brief description of those conditions is included in this report.After a preliminary investigation into the extent to which trackless conveyors were employed in the district, it was found that a detailed study at five bites would result in fair appraisal of that method of haulage in the underground iron-ore mines.Each of the five significant installations is described separately in this report. In one instance, however, company policy does not permit identification of the mine wherein the conveying is done. In that instance the operation is designated as mine ""A"".Unless otherwise specified, all tonnage figures quoted herein are gross.The cooperation of officials of all of the companies operating underground mines in the Lake Superior district permitted successful completion of preliminary studies of mining practices and conveying systems.Special acknowledgment is made to R. D. Satterley, general superintendent, and to operating staffs of the Sherwood and Morris mines of the Inland Steel Co. The assistance given by W. D. Whitman, assistant superintendent, and Douglas Tippett, mine captain at the Morris mine, and P. D. Pearson, superintendent, R. O. Marsten, engineer, and W. Johnson, mine captain, at the Sherwood mine, is acknowledged."

Oct 1, 1949

By Z Chang, S A. Mrozek

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. Antamina, Peru is the largest Cu-Zn skarn deposit in the world. Skarn mineralisation is focused around a multiphase Miocene intrusive complex comprising at least eight distinct phases identified by cross-cutting relationships, including truncated veins, xenoliths and chilled margins. Skarn alteration and mineralisation, as observed in drill core, remain open at depths exceeding 2.2 vertical kilometres.Porphyry intrusions occur as four major phases with subphases (from oldest to youngest): P1 (a, b), P2 (a, b), P3 (a, b) and two dykes of uncertain timing named Oscarina-1 and Oscarina-2. P1a is a hornblende-biotite quartz diorite porphyry with pervasive K-alteration, local endoskarn and 30–40 per cent sheeted quartz veins, all of which obscure original igneous textures. P1b dacite porphyry is a minor unit with <10 per cent phenocrysts and <3 per cent quartz veins. P1b cuts P1a; P2 dykes cut both units. P2a porphyritic biotite quartz monzonite contains K-feldspar megacrysts up to 4 cm long, local K-alteration and 2–80 per cent quartz stockwork veins. P2b biotite monzonite porphyry contains approximately 50 per cent phenocrysts and typically <10 per cent quartz veins. P3a (brown groundmass porphyry) and P3b (coarse phenocryst porphyry) are minor in volume and extent, and both cut P2 bodies. Oscarina dykes comprise two separate but parallel bodies. Oscarina-1 is a feldspar-phyric rhyolite porphyry, while Oscarina-2 is an intensely chloritised andesite porphyry. Oscarina-1 has a U-Pb (laser ablation inductively coupled mass spectrometry (LA-ICP-MS)) age of 10.1 ± 0.1 Ma, whereas P2 U-Pb zircon ages (LA-ICP-MS) range from 10.0 ± 0.2 Ma to 9.9 ± 0.3 Ma.All the intrusive phases contain quartz veins, with P1a and P2a having the most intensive K-alteration and most abundant quartz veins ± endoskarns. Each phase may have contributed to mineralisation at the right depths and within favourable wall rocks (ie marble) during uplift of the Antamina structural block. Therefore, a model of multiple intrusions during continuous uplift explains the vertically extensive (>2.2 km) skarn and mineralisation at Antamina.CITATION:Mrozek, S A, Chang, Z and Meinert, L D, 2015. A model for the intrusive sequence and Cu-Zn skarn formation at the Antamina deposit, Peru , in Proceedings PACRIM 2015 Congress, pp 423–430 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By Jiulong Cheng

Underground coal mining causes overburden deformation and failure, one major form of which is bed separation in the overburden. Surface subsidence can be controlled by injecting grout into the bed separations. For a long time, the grouting results of bed separation can not be validated. This is the problem in the application of this technique. This paper analyzes the conditions leading to bed separation, develops mathematical models for seismic wave fields generated before and after bed separation, and bed separations grouted with different materials, and computes the seismic wave field characteristics using the finite difference method. The results show that the existence of bed separation causes changes in the characteristics of seismic wave field, and that different grouting materials generate different seismic wave field characteristics. Intact strata, bed separation, and bed separation filled with air or water all display distinctive seismic wave field in amplitude and frequency spectra. After mining, the original location that provides strong seismic reflection wave no long exists. It is replaced by a weak and syncline-shaped reflection wave. The area, where the seam wave disappears, is larger than that of the mined area. The location, where a bed separation occurs, generates a set of reflection waves stronger than the normal one. The energy of the reflection waves varies with the materials filling the separations. Bed separations filled with air generate stronger reflection waves than those filled with water. The major frequencies for those filled with air are lower than those filled with water. Therefore, it is possible to determine the property of filling materials by analyzing and comparing the amplitude and frequency spectra of the reflection waves. Case studies of application of seismic method for exploring bed separations and grouting were analyzed. The results show that it is feasible to use seismic technique for exploring bed separations and grouting.

Jan 1, 2006

By Brian W. Stump, David P. Anderson, John Wiegand

Mining explosions are designed for a variety of purposes including the fragmentation and movement of materials. The blast design is dependent on the particular application intended and the material properties of the rock. The range of mining applications from hard rock quarrying to coal exposure to mineral recovery leads to a great variety of blasting practices.

Jan 1, 1997

By M. R. Saharan

A brief overview of rockburst control measures is presented. The paper then elaborates on one key rockburst control measure, which is known as destress blasting. A historical background of the technique is presented in order to understand different aspects of its field application. The publication enumerates key aspects of the technique as being used in practice in South Africa and North America. The paper also presents an analysis of three case studies in order to demonstrate difficulties associated with the field application of distress blasting in hard rock mines.

Jan 1, 2006

The acid plant consists of the following units: 1. Glover tower. 6. Mills-Packard type chambers. 3. Reaction towers. 6. Gay-Lussac towers.The sulphur dioxide bearing gases, used in the manufacture of the acid, are obtained from the roasting of zinc concentrate. Six Herreshoff type furnaces are available, each with an effective hearth area of 520 sq. ft.The concentrate contains approximately 30% sulphur, of which about 72 is ultimately recovered as acid. The oxides of nitrogen are added to the gases in the requisite proportions partly from nitrous vitriol and from nitric acid in the Glover tower, and partly from nitrous vitriol in some of the chambers and connecting flues.The acid leaving the Glover tower is cooled by the ordinary water-cooling coils, while the cooling of the chamber gases is effected by circulating water over the outer surface of the chambers, and by spraying nitrous vitriol and weak acid into them.Flow of gases throughout the plant is regulated and controlled by means of regulus metal fans, fitted with variable speed gears.

Jan 1, 1934

CENTRAL mine ore consists mainly of galena and marmatite in a siliceous and rhodonitic gangue, the average assay values being approximately 11.5 oz. per ton Ag, 14.5% Pb, 14.5% Zn, 1.3% Ox Pb, 0.3% Ox Zn.Owing to the 1923 fire in the underground workings, much of the ore now being mined has been subjected to a certain amount of oxidation, as shown by the oxidised values above.This factor is of great importance from a flotation point of view, as a sulphide surface which has been subjected to the influence of oxidation is not as amenable to flotation concentration as a fresh sulphide surface. The importance of this factor is increased when a differential flotation is to be made, and, as in the case of Central mine ore, the sulphide, which is to be floated first, has been oxidised...

Jan 1, 1931

By R. E. Moore

Refractories are unique ceramic materials, which are used in large quantities in steel making furnaces, metal smelting vessels, and glass melting tanks. They are required to possess high thermal shock properties and good slag corrosion resistance as well as creep resistance at high temperatures. Therefore, they are made from very refractory minerals such as lime, periclase, corundum, spinel, and zirconia. Slags are vitreous materials produced as a residue in steel refining processes. They cover a wide range of compositions and may contain pure oxides, silicates, and sulfides as well as fluoride phases depending on the melting and smelting processes. These minerals exhibit spectacular cathodoluminescence color under electron bombardment; therefore, cathodoluminescence (CL) microscopy is found to be a superior technique for the characterization of refractory corrosion by slags. Examples included in this study are (1) reaction of fluorine containing mold slags with ZrO2-C nozzle refractories promoting crystallization of the Cuspidine, Ca4Si2O7(F,OH)2, (2) corrosion of fusion cast refractories in TV panel glassmaking furnaces and formation of glass defects, and (3) reaction of spinel based castable refractories with copper smelting slags.

Jan 1, 2001

By R. W. Vieth, S. A. Bradford

THE orientations of tetragonal tin crystals are commonly examined by either the transmission Laue method or the back-reflection Laue method. The predominant planes and zone axes of the pattern are plotted on a Wulff net and indexed by measuring the angles between planes and directions. The problem of identifying the Laue spots has been approached in several different ways. Nicholas 1 tabulated interplanar angles for ten planes but did not relate the angles to permutation of indices or sign. A table of angles between adjacent planes and of angles between adjacent zones was published by Chandrasekhar and veal.' These tables are very helpful, provided the film shows all the low-order zones that occur within about 20 deg of each other. Frounfelker and Hirthe's compilation3 of angles between the (001) plane and twenty-nine other planes is useful for the indexing, only if the (001) spot can be found on the film. For all but the simplest orientations, however, more information is needed to index the diffraction spots. Standard stereographic projections, such as those published by Bicelli, 4 provide part of the information required. By their nature, these projections show the positions of only a few zone axes. In the tetragonal system many low-order zone axes are not normal to the planes having the same indices. Consequently, planes and directions can have entirely different positions on a stereographic projection. For example, the [011] zone axis is 33 deg from the normal to the (011) plane. The crystallographic orientation usually has to be deduced from a combination of the predominant zone axes and a few low-order planes that appear on the film. In addition to a standard stereographic projection of the planes, a stereographic projection of the directions is needed. Such a projection can be constructed from the angles given in Table I. To provide more information about angles between planes and directions, we have calculated interplanar angles up to (332) and angles between planes and directions up to [332] on an IBM 1620 computer. The angles are differentiated according to permutations of sign and indices. Those doing research on tin single crystals are welcome to request a copy of these tables from the authors.

Jan 1, 1967

By Andrzej Walentek

This paper presents the results of numerical modeling of the fractured zone around longwall gateroads. The numerical computations have been based on the results of underground measurements of the fractured zone around longwall gateroads, which were performed for various geological and mining conditions and measured with the use of the endoscopic method with a digital camera. Having results of the underground investigations, trials of duplicating them by the Finite Element Method using software PHASE and the Hoek-Brown failure criterion were performed. In order that the fractured zone predicted by the model-based computations closely simulate that measured underground, some selected parameters in the Hoek-Brown failure criterion were modified based on the earlier determined relations used for prediction of the roof-to-floor convergence in longwall gateroads. As a result, the model-based computations produce results very close to those obtained from the underground measurements. Determination of the fractured-rock zone around the mine roads is of great importance to mining practice, especially in planning and designing supports for longwall gateroads. The influence of the front abutment pressure causes the fractured rock zone around the roads to increase and results in an increase of support load. Moreover, proper determination of the fractured rock zone is one of basic elements in the process of planning roof bolting for mine workings.

Jan 1, 2009

By K. O. Bennington

The thermodynamic properties of two manganese silicates, pyroxmangite and fowlerite (zincian rhodonite), were experimentally determined by the Bureau of Mines. The standard enthalpies of formation were determined by hydrofluoric acid (HF) solution calorimetry, the low-temperature heat capacities by adiabatic calorimetry, and the high-temperature heat capacities by differential scanning calorimetry over the range 322 to 1,372 K. Data for the standard enthalpy of formation (AH298), the derived standard entropy (S298), and the standard Gibbs energy of formation (AGf 298) are as follows: Pyroxmangite [(Mn0.9554Cao,0446)(Si03)J- AHf298 319.668±0.46 kcal/mol AH298 (from oxides)... -7.215±0.11 kcal/mol 5295 23.785±0.048 cal/(mol'K) AGf298 -301.190±7.96 kcal/mol Fowlerite [(Mno.7o3Zno.0884Ca0.2086)(Sio3)]--AHf298 ?... -329.914±0.419 kcal/mol AH298 (from oxides)... -8.400±0.104 kcal/mol 5598 21.949±0.059 cal/(mol-K) AGf298 -310.716±7.99 kcal/mol These experimentally determined values were combined with data from the literature to calculate the Gibbs energies and equilibrium constants of formation over the temperature range of the measurements. Standard enthalpies, Gibbs energies, and equilibrium constants of formation are given as a function of temperature from both the elements and the constituent oxides.

Jan 1, 1987

By M. Lahtinen, P. Latostenmaa, I. Kojo, E. Peuraniemi, M. Lindgren, S. Jyrkönen, M. Jåfs

The year 2019 marks the 70th anniversary of Flash Smelting. The technology was operated on industrial scale for the first time at the Outokumpu Harjavalta copper smelter in Finland in 1949. Since then Flash Smelting technology, now owned and developed by Outotec Oyj and the first plant, now owned by Boliden Harjavalta Oy, have been evolving throughout history by optimizing the process for various feeds and raw-materials, as well as through developing the equipment and operating practices. Even though sustainability was not a popular theme during the first days of Flash Smelting, it has still been the key principle in guiding the development. This is well proven by the wide adoption of Flash Smelting for the processing of copper and nickel concentrates, mostly due to its ability to fulfil the economic and resource efficiency criteria demanded by its operators. This paper describes the history and improvements of the Outotec® Flash Smelting process by looking into the early days of operation and how the development of the process and equipment has taken place over time as a joint effort between the technology provider and the users of the technology. INTRODUCTION In the 1940s, Outokumpu Oy refined the concentrate produced in the Outokumpu mine into metallic copper initially in Imatra, and later from 1945 to 1949 in Harjavalta, based on the most modern electrical melting process of the world at that time. After the war, Finland lost a significant part of its hydropower production, and due to the increasing use of electricity, the availability of electricity weakened, and the price rose. To maintain Outokumpu's profitability, it was forced to look for a melting solution that was not based on electricity.

Jan 1, 2019

By Ron Glowe

This paper shows the potential of a new computer model, using the Glowe-Tech (GT) Tonnage Analyzer program to compare drilling and blasting results. This model uses the crushers and conveyors as measuring tools in aggregate quarries in the USA and Canada. All data is collected using commercially available kilowatt transducers and an analog input/output automation module. There are two areas of measurement used to compare results. The first area is the tonnage moved by key conveyors to show the split in products produced and the actual rate of production. If belt scale readings are not available then the Glowe-Tech Tonnage Analyzer can be used to convert kilowatts from the conveyor motor to tonnage. The second is the tons of rock crushed per kilowatt hour of energy consumption by the primary and secondary (optional) crushers. This modeling process works best if blasts are mucked and crushed in sequence. There are many other variables such as crusher types, crusher settings, screen sizing and actual geology in the pit that can influence the results, however if variables can be standardized then definite trends can be seen when comparing drilling and blasting results

Jan 1, 2005

By M Dickie

During the 1980s and early 1990s low cost gold treatment plants were at the forefront of the Australian engineering and mineral processing business. This opportunity was the result of a high gold price and innovative design by emerging engineering companies in an expanding market. The low cost approach to gold plant design and project execution resulted in smaller and shorter life projects becoming viable and the multiplicity and fast track approach of the projects resulted in sharp learning curves. From the mid 1990s to the early 2000s the relatively low commodity prices and resultant reduction in minerals opportunities also resulted in innovative design in both small and large plants, albeit on a reduced scale. The innovation was associated with alleviating capital cost pressures for large capacity low-grade (eg NewcrestÆs Cadia) or small-scale (eg JubileeÆs Cosmos) projects to either improve NPV or to enable the project to be won and profitable in the competitive lump sum environment. Between 2004 and 2008 the engineering market became less competitive from an engineerÆs perspective due to the high demand for engineering services in a high price commodity market. The need for innovation remained but was driven overtly by shortages in water, power and personnel and long equipment lead times. This paper discusses some of the factors that contribute to the capital cost of a project and the scope for improving the likelihood of financing and increasing the project net present value (NPV). Case studies with recent large copper concentrator projects will be used to demonstrate how a value driven approach to project and plant layout can result in significant improvement in NPV.

Jan 1, 2009