Search Documents

By R. M. S. Häggblad

The article deals with stabilisation and regeneration of Ce02 and Ce02/Zr02 based Pt water gas shift catalysts, subject to high initial deactivation. The reaction gas species effect on the catalyst deactivation was investigated by H2-TPR. Activity measurements enabled the effect of different promoters, some added to the Ce02 based catalysts and some to the Ce02/Zr02 based Pt catalysts, to be investigated. The catalysts were also characterised by BET and CO-TPR. Deactivated catalysts activity was restored by using various regeneration methods. Of the two selected carriers the Ce02/Zr02 based Pt catalyst showed the highest resilience to deactivation. For the two different carriers, Ce02 and Ce02/Zr02, Wand Re were the best promoters when the catalyst was subject to deactivation. Experiments with Hr TPR indicate a fast initial change in the platinum oxides concentration and composition. The CO-TPR was used to make conclusions about the various regeneration effects of water and oxygen on the catalyst. Finally it is suggested that not one deactivation mechanism is possible and which mechanism that dominates is dependant on the catalyst and the reaction gas composition.

Jan 1, 2005

By H. C. L. Reemeyer

"The mining industry faces increasing economic, environmental and social challenges, as easier deposits deplete and ores become more difficult to mine and process; head grades decline; projects migrate to increasingly remote and difficult geographies; and the industry faces more stringent environmental and social standards. While these factors increase costs and risks, they also present opportunities for a more integrated approach to mining, processing, materials handling and waste management. Two key areas of risk are in tailings and water management. As mining and processing tonnage rates continue to increase, the disturbance footprints, water and energy consumption and scale of tailings storage facilities also rise. These factors can be mitigated by more selective mining, ore sorting and pre-concentration, and more intensive tailings dewatering. Such alternatives must be considered in conjunction with the selection of mining method, metallurgical flowsheets, major equipment selection, and key process design criteria. For example, increasing the grind size of a concentrator reduces energy consumption and eases tailings dewatering, but may cause challenges with metal recovery. Emerging technologies will be reviewed, and economic and environmental trade-offs for eco-efficient alternatives will be presented for metal sulfide mines.INTRODUCTION The metals mining industry faces significant challenges that can affect its financial viability and social license to operate. There has been a consistent decline in ore grades and a relentless push towards larger operations to capture perceived economies of scale. With these larger operations, the environmental footprint of mines has increased. Water and energy consumption, and resulting GHG emissions have also risen. In recent years, two issues have raised special concerns for stakeholders and companies: • Water scarcity, particularly in drier regions of the world such as Northern Chile, SW USA, Southern Africa and Australia • Tailings risks, following the high-profile failures of the Fundao tailings dam in Minas Gerais, Brazil in 2015 and the Mount Polley tailings dam in BC, Canada in 2014"

Jan 1, 2017

By E. Uzun

"In this study, H2SO4 and HCl leaching behaviour of an ore sample containing 1.87% of Ni, obtained from the Karaçam (Eskisehir-Turkey) lateritic deposit that started operation recently was investigated. Acid concentration, temperature and reaction time were investigated as variables by comparative leaching tests and the optimum leaching conditions were selected on teh basis of Ni, Fe, Co and Mg behaviour. For H2SO4 leaching, optimum leach parameters were 95°C, a 12 normal H2SO4 concentration and 90 minutes of reaction time, and under these conditions Ni and Fe dissolutions were 97.6% and 95.4% respectively. On the other hand, 70°C, an 8 normal HCl concentration and a reaction time of 90 minutes were found to be the optimum parameters for HCl leaching, and Ni and Fe dissolutions found were 97.1% and 97.8%, respectively.INTRODUCTION Today, nickel and various nickel compounds are obtained from sulphide and lateritic nickel ores (Moskalyk & Alfantazi, 2002). Even though about 72% of the world's nickel reserves constitute lateritic ores, only about 42% of nickel is produced from these ores. The decrease in production from sulphide deposits, on the other hand, and increases in the use of nickel in the iron and steel, and battery production sectors, and especially the production of nickel cast irons in recent years, have accelerated the importance of making use of lateritic deposits (Luo et. al., 2010; Chang et. al., 2010). In the beneficiation of sulphide nickel ores, mostly pyrometallurgical methods are used. On the other hand, about 39% of lateritic ores are beneficiated by pyrometallurgical, and 61% are beneficiated by hydrometallurgical methods (USGS, 2009). In laterite beneficiation, the ore properties are of significance; that is to say, limonitic ores are eneficiated by high pressure acid leaching or via the Caron process, nontronitic ores are beneficiated by high pressure acid leaching or heat treatment (smelting), and ores containing serpentine by pyrometallurgical methods (Berezowsky, 2004; McDonald & Whittington, 2008a; Dreisinger et. al., 2015)."

Jan 1, 2017

By Jean Scott, G. A. Collins, G. W. Hodgson

WHILE it is well known that trace metals in crude •oils and distillates have been a matter of deep concern to refining engineers because of the damage done .by the metals to cracking catalysts and refractories (8, 5), the trace metals in crude oils may be of great benefit in correlation studies and in origin studies. During the geological processes in which the oil is formed from •animal and vegetable life, the hydrocarbons are subjected to a host of inter-related reactions, both physical and chemical. It has been well demonstrated that the lighter materials, certainly the short-chain paraffin hydrocarbons, are degradation products of the heavier molecules, and it might be concluded that for correlation studies it would be advisable to direct attention as much as possible not to the degradation products of the large molecules but to the large molecules themselves. Be this as it may, the methods at •hand to conduct the study are not particularly precise and, in the past, attention has centred on the simpler small molecules, as evidenced by the great mass of work involving distillation and chromatography with painstaking separation of the simpler 'front-end' hydrocarbons.

Jan 1, 1954

By J. -P. Lempainen, K. Haavanlammi, K. Valkama

The Outotec nickel matte chloride leaching process is especially suitable for small capacity standalone nickel refineries. The matte is leached with hydrochloric acid and oxygen in multiple steps in atmospheric OKTOP® reactors. Oxygen feed prevents reductive conditions and hydrogen formation in the leaching process. A sufficiently high acid concentration keeps the iron dissolved in the solution. The iron is removed after leaching by precipitation and the pregnant leach solution continues to metals recovery. The metals recovery from the chloride solution is done by solvent extractions, where ammonia is used for pH control. Copper and nickel are stripped from the organic phases into sulphate solutions. Metals can be recovered from the purified solution by conventional electrowinning. Other alternatives are precipitation or hydrogen reduction processes. The ammonium chloride in the solvent extraction raffinate is regenerated into ammonia. The calcium chloride solution from ammonia regeneration is routed to acid regeneration where sulphuric acid is added and ultra-pure gypsum is precipitated. The hydrochloric acid solution from this step is recycled back to leaching. The process yields good overall metal recoveries and expensive

Jan 1, 2014

By S. A. Scott

In 1913, potash mineralization was recognized at a depth of over 7,000 feet in an oil-well hole (Norcanol's Radville No. 1 well) drilled near the town of Raddlle, Saskatchewan. Although, at this depth, potash is not considered mineable by conventional underground mining methods, this discovery shed the first light on the potash deposits of Saskatchewan in much the same manner as oil-well drilling was responsible for the first discovery of potash in the Carlsbad Basin of New Mexico in 19:25. The discovery near Radville alerted the oil-well geologists and, as a result potash mineralization was recognized in the Verbata No. 2 well, near Unity, Saskatchewan, in 1946. The mineralization, which was at a depth ?of 3,466 feet, graded more than 20 per cent K20 over a thickness of eleven feet. This discovery, as its depth, thickness and grade were all amenable to mining,

Jan 1, 1963

By Richard W. Hutchinson, Christopher Osterman

Abstract -The Selebi-Phikwe deposits are hosted by amphibolite to granulite facies gneisses of the Limpopo belt. The sulfide orebodies are generally stratiform and conformable to the foliation of the wallrock gneiss. Sulfide remobilization has occurred during metamorphism and has produced discordant sulfide veins, called piercement structures, which project into the hanging wall and foot­ wall gneisses. These piercement structures are compositionally similar to the stratiform sulfides, consisting mainly of pyrrhotite, pentlandite and chalcopyrite. However, texturally the piercement sulfides are coarse grained and inclusion-free, whereas the stratiform sulfides are fine grained and packed with tectonically produced wallrock inclusions. The piercement structure sulfides and adjacent stratabound sulfides were sampled and analyzed for major base metals and PGE's for comparative purposes. Most of the piercement structures were enriched in Cu, Ag and Au, whereas they were generally depleted in Ni, Co, Fe and PGE's. Changing elemental associations were observed from the stratabound sulfides and along the length of the piercement structures. This and the degree of depletion or enrichment of various major and minor elements in the sulfide piercement structures may reflect the relative mobilities of metals under metamorphic conditions. The similarities between sulfide piercement structures and the footwall Cu-POE veins in some Sudbury deposits are reviewed.

Jan 1, 1994

By A. F. Brock

Introduction During the past ten years, shaft sinking has been going on continuously at Inco and new shafts and shaft extensions have been sunk under a broad diversity of conditions. The Murray mine and the Stobie section of the Frood-Stobie mine have?been opened and a main hoisting and supply shaft has been sunk at each of these mines; No. 2 shaft at Murray was sunk initially from surface to 1,960 feet and is now being extended; No. 7 shaft at Stobie was sunk from surface to 3,105 feet. No. 2 shaft at Garson has been extended from 2,200 feet to 3,200 feet in depth. At Levack, No. 2 shaft is being extended from 1,860 feet to 2,850 feet for ore hoisting from the 2,650 lev el. At Creighton, No. 6 shaft, collared on 52 level, was sunk 1,740 feet to a total depth of 5,563 feet below surface. It is the deepest Inco working. At Creighton, also, No. 7 shaft is being sunk from surface to the 30 level to handle the ore from caving operations. Ail Inco main hoisting and supply shafts sunk during the past twenty years have been vertical rectangular shafts equipped with two cages and two ore skips. The cages are standard and interchangeable for all shafts except Garson No. 2; they are single-deck, 13 ft. 4 in. by 5 ft. 2 in. inside and are designed to carry a maximum material load of 16,000 lb. The man load is 48 men. Skip capacities at the several shafts vary from 8 to 15 tons. Forty-two hundred feet bas been established as the standard vertical lift, and the permanent hoisting equipment at each shaft is designed for eventual hoisting from that depth.

Jan 1, 1950

By J. Jin, J. D. Miller

"Bubble attachment at hydrophobic surfaces is of critical importance in the flotation recovery of mineral resources and fundamental experimental results have been reported in the literature describing the details of film thinning, rupture, and displacement. Now, for the first time, bubble attachment phenomena are examined using Molecular Dynamics Simulations (MDS). The simulation involves a nitrogen bubble containing 906 nitrogen molecules in a water phase with 74,000 water molecules at the hydrophobic molybdenite (001) surface. During the simulation period of 1 ns, film rupture occurs and the nitrogen bubble attaches to the molybdenite (001) surface, resulting in a contact angle of about 90o. In contrast, the film is stable at the hydrophilic quartz (001) surface and the bubble does not attach. Contact angles, determined from MD simulation after stabilization of the three phase line of contact, are reported and these results agree well with experimental and MDS sessile drop results. In this way, film stability and bubble attachment are described with respect to interfacial water structure for surfaces of different polarity. For example, the interfacial water molecules for the hydrophobic molybdenite (001) surface have relatively weak interactions with the surface when compared to the hydrophilic quartz (001) surface, as revealed bythe presence of a 3 A water exclusion zone at the molybdenite/water interface. INTRODUCTIONAttachment of air bubbles to mineral particles is of fundamental importance in understanding flotation separation phenomena. It is well known that attachment of an air bubble at a hydrophobic surface includes film thinning, film rupture, and film displacement (Somasundaran, 2006; Wilson, Adlard, Cooke, & Poole, 2000). Experimental techniques, such as high speed video, can catch the process of bubble attachment at a mineral surface (Drelich & Miller, 2012; Niecikowska, Krasowska, Ralston, & Malysa, 2012). However, molecular scale information to examine the phenomena has not yet been reported."

Jan 1, 2016

By P. Amelunxen

It is well-known that the particle breakage mode that occurs in a semi-autogenous mill is significantly different from that which occurs in a ball mill, enough so to warrant very different lab scale breakage tests and scale-up models. This phenomenon, at the plant scale, leads to changing process constraints depending on the particular properties of the ore. For example, hard or coarse ores might bottleneck the SAG/AG mill throughput, while the ball mills might constrain throughput when milling soft or fine ores. Because the power draw of a ball mill is fairly constant compared to that of a SAG/AG mill, periods of SAG/AG mill bottlenecks lead to increased ball mill circuit specific energy consumption and a finer grinding circuit product size, which together constitute a loss of efficiency of the overall process. Because HPGR circuit throughput is less sensitive to changes in ore hardness, this situation is less common or severe in those circuits. Furthermore, the ball mills are somewhat decoupled from the upstream crushing circuits through the use of surge bins, which also helps to mitigate this effect. As a result, ore hardness variability constitutes a significant cost penalty that should be applied to SAG mill/ball mill/pebble crusher SABC circuits when evaluating the relative economics of different grinding circuit configurations. Nevertheless, the issue is by-and-large ignored, because most current HPGR trade-off studies are based on averaged ore hardness parameters. This paper provides an estimate of the magnitude of this penalty for a typical operation by using Monte Carlo simulation methods and plant data from Freeport-McMoRan?s autogenous, semi-autogenous, and HPGR-based comminution circuits. A case study from an in-house 96K t/d sulphide project is presented to illustrate the concept. It was found that by including variability effects, the trade-off NPV of the HPGR circuit over that of the SAG mill circuit increases to $84 million from negative $33 million, effectively reversing the original decision of selecting a SAG mill circuit for the fixed hardness case.

Jan 1, 2012

By S. Nikolic

"Calcium ferrite slags are currently used in a number of copper-converting processes. Despite the industrial importance of this system the phase equilibria have not been fully investigated. Characterization of this slag system is necessary to improve the control of process parameters, including fluxing and operating temperatures. Recently, a modified experimental method has been developed; this involves rapid quenching of slag samples equilibrated at high temperatures with the resulting phase assemblages analysed using electron probe microanalysis (EPMA). In the present study, experiments were completed under controlled temperatures and oxygen partial pressures using CO/C02 or H2/C02 gas mixtures. The investigated slags were supported during equilibration by a substrate of the primary phase. This technique removes the limitations arising from the use of crucibles and facilitates rapid quenching of the melt. Experiments were carried out at 1200°C and 1250°C (1473 Kand 1523 K) and in the composition range 0-40 wt.% Cu20, 0-25 wt. % CaO and 45-80 wt. % Fe203. Liquidus and solidus data are reported for the primary phase field of spinel between the oxygen partial pressures of 10-4.5 to 1 o- 6-5 atm. The analyzed compositions of liquids and solids are used to construct the phase diagram of ""Cu20"" -CaO-""F ez03"" system in equilibrium with metallic copper at fixed oxygen partial pressures."

Jan 1, 2007

By X. Gu

Mining operations, especially surface coal mines, have great impacts on the local ecosystem and the environment. However, it has been a common practice in mine design that the economic evaluation considers only the operating and capital costs, while the costs incurred due to ecological damage are left out. Such practice underestimates the total cost of mining and ignores the difference in ecological impacts between different mine design alternatives. With growing interest in sustainable development, the mining industry is under increasing pressure to operate not only based on economic and engineering principles but also on principles of sustainability. Therefore, estimating ecological costs of mining and considering such costs in the mine design process is an important step towards reducing ecological impacts at the design stage. The objective of this paper is to quantify the ecological costs of open pit coal mining and to demonstrate the influence of such costs on the ultimate pit design. The ecological costs of mining are estimated based on the areas and types of land damaged by mining activities and the carbon emission from energy consumption. Four major components of ecological cost are identified, namely, lost value of direct eco-services, lost value of indirect eco-services, prevention and restoration costs, and the cost of carbon emission from energy consumption. Necessary equations are given for estimating these costs. An algorithm is described for ultimate pit optimization in coal deposits with near-horizontal coal seams. A case study is presented in which the ultimate pits with and without ecological costs are optimized and compared to show the influence of ecological costs on the pit design outcome.

Aug 1, 2013

By Lauri Pesonen

"Although electrical energy use at copper smelters is typically about 15-20% of operational costs, it is something which is not very often analyzed and discussed at a detail level in technical papers and presentations. This paper aims to list different equipment configurations per smelter area and estimate their electrical energy consumptions versus unit of material processed. In addition chosen smelting technology and oxygen enrichment rates affect the overall electricity consumption and therefore some overall guidance is provided for efficiency gains.INTRODUCTION In principle, processing plants, such as copper smelters, require electrical and chemical energy together with some degree of human input to process input materials into desirable output form like metallic copper. From the standpoint of metallurgical plant designers, the thermodynamics of smelting, when it comes to releasing and conserving thermal energy, is sufficiently understood and frequently discussed in the literature in detail, but judging from the quantity of publications, the way how we exactly utilize electrical energy is not. This article is meant to illustrate how electrical energy consumption is distributed among different process units and equipment types at copper smelters using various smelting technologies. In general, typically about 15-20% of operational costs are related to electricity and at locations with high electricity prices the potential savings can be quite significant for the overall profitability of the plant operations. In today’s tight profit margin environment any potential savings are more than welcome. The purpose here is not to promote any certain copper smelting technology, but to open a discussion on how and where we use electrical energy, and how the overall energy use could be lowered with the implementation of newest electric power saving designs and solutions. Using the principles outlined in this article, it should be fairly easy and straightforward to replicate the calculations and even improve them wherever it is seen necessary for the greater benefit for us all in this industry"

Jan 1, 2017

By Victor R. Eyzaguirre

Cerro Verde is one of several large porphyry-type copper deposits in Southern Peru. The copper occurs as primary sulphides, with a significant oxide capping. The exploitation of the deposit has been correspondingly divided into two stages, the first consisting of copper production from oxides, and the second from sulphides. Minero Peru successfully started the Stage I plant in June of 1977 with the first production of electrowon copper cathodes. This paper is essentially devoted to a description of the Stage I process, which can be summarized as follows: Ore is trucked to the primary crusher, where it is crushed to 225 mm and conveyed to a coarse ore stockpile. Ore withdrawn from the stockpile is secondary crushed to 65 mm, conveyed to a truck loadout bin, loaded into trucks and driven to the leaching pads. Ore is placed on the pads in lifts and sprayed with sulphuric acid solution to dissolve the copper. Copper-bearing solution forms a pond behind a dam at the lower end of each pad and is pumped from there to the solvent extraction plant. Copper is extracted from the leach solution into an organic phase and is then stripped from the organic phase to form a purified concentrated copper solution suitable for feed to electrowinning. In electrowinning, the copper is finally recovered as cathodes suitable for sale as refined copper.

Jan 1, 1978

By J. R. Donald

"Project development has been increasingly formalized in the past decade, with most organizations having extensive and descriptive deliverables documents and processes to govern the stage gate progress of capital projects. Although this has improved the general governance of capital project development and execution, it has perhaps not led to improved business success. This paper identifies specific challenges associated with developing complex projects with multiple stakeholder interests over the long timeframe of the project life cycle. The paper outlines these challenges and suggests management approaches and aspects of project development where increased focus is required to achieve improved business outcomes."

Jan 1, 2015

The major factors limiting the establishing of vegetation on the shale and sandstone wastes at Minto are high exchange acidity and low pH (shale pH 4-5, sandstone pH < 4). In the first revegetation program (1967-74), 3.8 million trees were planted without grading on 4,000 acres (red, jack and scots pine and black locust). White dutch and sweet clover, birds-foot trefoil and red top were established in selected areas, and ponds were stocked with brook trout. In the second phase of reclamation, beginning in 1979, a comprehensive plan will be initiated which will allow for forest production, wildlife development and recreational activities.

Jan 1, 1977

By J. -F. Lepage, M. A. Tomlinson, K. K. Murray

"The Éléonore mine, located in the James Bay region of Northern Quebec, generates a gold bearing flotation concentrate rich in pyrrhotite. Direct leaching of the finely ground concentrate results in high cyanide consumptions and poor recoveries, so effective pre-treatment to passivate the pyrrhotite is required before leaching. Following a number of process upsets at the mine, test work was performed to better understand the requirements for passivation so a more effective operational strategy could be implemented. This paper outlines the testing and implementation of the new strategy at Éléonore.ÉLÉONORE CONCENTRATOR The Éléonore mine is located 800 km north of Montreal, in the James Bay Region of Quebec, Canada. The ore, containing gold disseminated in sulfide minerals and free fine gold in silicate gangue, is crushed in a two-stage crusher plant before being ground in a single 9,000 kW ball mill. The ball mill is in closed circuit, with about 25% of the cyclone underflow passing through a gravity plant with two, 48” Knelson concentrators. Gold recovery in the gravity circuit varies between 15 and 20%, with gravity concentrate leached daily in an Acacia intensive leach reactor. Cyclone overflow, at a target P80 of 65 microns, flows to the flotation plant where the gold bearing sulfide minerals, mainly pyrrhotite and arsenopyrite, are recovered to concentrate with a mass pull of about 5%. The flotation concentrate, typically containing between 80 and 120 g/t of gold, is thickened and finely ground using a 1500 kW Isa Mill. The finely ground concentrate is diluted, to generate a slurry with about 32% solids by weight, before feeding the leach plant. Initially the concentrate slurry is pre-treated with oxygen and lead nitrate, to passivate the sulphide minerals before cyanide is added for gold dissolution. Gold is recovered from the leached slurry using carbon-in-pulp (CIP). The flotation tailings are also thickened and gold is recovered in a dedicated leach-CIP plant."

Jan 1, 2017

By Piotr Litwa

The paper focuses on actions aiming at improvement of work safety in mines, particularly in those extracting hard coal, taking into account the characteristics of natural hazards in Polish underground mining and the reasons of mining disasters connected with those hazards. It refers to conclusions addressed to scientific and research centres, drawn during last decade by Commissions appointed by the President of the State Mining Authority in order to investigate reasons and circumstances of incidents, which till now, could not have been accomplished because of financial reasons. Furthermore, the paper presents basic assumptions of the Strategic Research Project Improvement of work safety in mines, announced by the National Centre for Research and Development at State Mining Authority?s President request.

Aug 1, 2013

By D. T. White

Pressure acid leaching of nickel laterite ores typically involves the production of an intermediate product that is releached and refined to pure nickel and cobalt products. BHP Billiton is developing the Ravensthorpe Nickel Project to produce up to 50,000 t/y of nickel via a mixed hydroxide intermediate, which will be refined at its Queensland Nickel (QNI) Yabulu Refinery near Townsville, Queensland. The hydroxide precipitation flowsheet has been adapted to process high magnesium ores through an atmospheric leach stage, eliminate iron(III) as jarosite and oxidise iron(II) during neutralisation. The neutralisation conditions are controlled to precipitate aluminium, chromium and silicon, and to ensure good solid-liquid separation. Soluble impurities such as magnesium are purged with the tailings stream and will be managed over the long term by crystallisation in evaporation ponds.

Jan 1, 2006

ABSTRACT This paper details the work of RISKGATE, the largest Australian Coal Association Research Program (ACARP) occupational health and safety initiative to date. The program is an interactive online risk management system developed to maximize health and safety performance in the coal mining industry by implementing continuous improvement in managing major unwanted events. The comprehensive knowledge base covers tires, collisions, fires, isolation, strata control underground, ground control open cut, explosions, explosives, manual tasks, and slips/trips/falls. All major Australian coal mining companies contributed topic experts to this broad industry initiative: nine coal mining companies, more than 80 experts, and more than 400 days of individual time for knowledge-capture workshops.

Jan 1, 2014