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By S. C. Tadolini

"Roof bolts are the primary means of roof support in underground mines in the United States. Presently, post installation loads are measured on about 10% of the expansion anchor and combination bolts by using standard torque wrenches that Unit the accuracy of the measurements to about ±30%. No system has been available for measuring the loads generated on full-column resin grouted bolts. Mine roof bolts are long, slender, non-uniformly sized, roughly forged, and placed in a corrosive environment, creating obstacles for researchers trying to utilize ultrasonic reflection techniques to measure strain. A complete measurement system has now been developed by the U.S. Bureau of Mines, Denver Research Center. This system is being used successfully in determining loads measured on expansion anchor and combination bolting systems, examining the transfer mechanics of full-column, resin-grouted bolts, and measuring loads generated on bolts in an underground roadway support system.IntroductionMillions of roof bolts are used each year to provide structural support in underground U.S. metal, non-metal, and coal mines. Approximately 45% of the bolts used are expansion anchor, 45% are resin-assisted combination or full-column resin-grouted, and 10% are friction anchored. The effectiveness of these bolting types is dependent on a number of factors including rock density, bolt length, in situ stresses, rock properties, bolt pattern, installation tension, etc. When implementing expansion anchor and resin assisted combination bolts , installation and post-installation tension, or load, is a major determinant of support quality and subsequent behaviour of the reinforced roof. The method by which these systems carry roof loads is well understood and documented(1,2).The importance of these installation and long-term bolt tension measurements is well recognized and are required in the federal government regulations(3). A statistical sample of the bolts, lout of 4 during installation and lout of 10 inby the last open crosscut, is required to be measured using current torque wrench technology. The disadvantages of this system are the disturbance and weakening of the anchorage and limitations of accuracy to approximately ± 30%. The method by which resin-grouted bolting systems carry roof loads is much less understood, mainly because testing procedures used to evaluate point-anchor systems are not effective on resin-grouted bolts.Due to these problems, the U.S. Bureau of Mines designed a measurement technique that provided a highly accurate, low cost-per- bolt system that led to the development of ultrasonic pulse echo instrumentation. The original system was designed to measure the tension only on expansion anchor bolts. Recent developments and system modifications have permitted measurements on resin-assisted and full-column resin-grouted bolts. The use of these types of support systems has been increasing steadily in underground coal mines.The development of this instrumentation has provided laboratory and field results that help to explain the mechanics and relationships between the applied loads and the interaction between the grouted bolt system consisting of the bolt tendon, grout, and rock."

Jan 1, 1990

By T. R. Yu, D. B. Counter

"Kidd Creek Mines primarily employs sublevel blasthole stoping for underground mining operations. During pillar recovery in No.1 Mine and panel mining in No.2 Mine, the free standing fill exposures may reach a height of 100 m or more. Consequently, a competent consolidated rockfill has been required to fill the mined openings.Since 1976, a total of more than fifty stopes and pillars have been filled with over four million tonnes of backfill. Fill exposures, up to heights of 100 m, have stood up well after blasting adjacent to the backfill.The fill material consists of a graded aggregate, Portland cement and ground slag. Sand has been occasionally mixed in the fill material to serve as avoid filler. The aggregate is prepared in a fill crushing plant, transported underground through raises, and hauled to the filling stopes by conveyors in No. 1 Mine and by trucks in No.2 Mine.Two mixing methods are currently in use. No. I Mine utilizes the free fall of gravity; the fill materials are sprayed with cement slurry as the aggregate departs from the conveyor into a mixing culvert. The other is the use of a rotary mixer on a batch basis in No. 2 Mine.This paper describes the fill systems, filling operation s, and the backfill quality control practice, together with various testing results related to improving fill quality as well as reducing fill costs.IntroductionThe Kidd Creek minesite, located 27 km north o f the city centre of Timmins, Ontario, produces 4.5 million tonnes of copper, zinc, lead, and silver ores annually from its No. 1 and No.2 underground mining operations. Both mines employ a sublevel blasthole stoping method."

Jan 1, 1983

By Yufeng Yao, Robert P. Lowell

Abstract - Conductive heat flow anomalies associated with hydrothermal sulfide mounds on the seafloor have been studied. Sulfide mounds tend to focus conductive heat flow because sulfide minerals have a greater thermal conductivity than country rock. The topography of the mound above the seafloor, however, leads to a reduction of heat flow, particularly near the center of the mound. Because the focusing effect is constant, whereas the topographic effect decreases with distance away from the center, the focusing effect dominates toward the edges of the mound. The magnitude of these effects depends upon the conductivity contrast and the height/radius aspect ratio of the mound. Conductive heat flow anomalies at sulfide mounds also result from fluid circulation within the edifice. For the active and relict mounds at the TAG hydrothermal field, which have an aspect ratio of approximately 0.3 and a conductivity contrast of 3, the heat flux through the center of the mound should be roughly 70% of the background heat flux. The maximum effect of heat flow focusing is to increase the heat flux by about 30% above background. By comparing these effects with conductive heat flow data at TAG, we find that advective processes dominate conductive heat flow at both the active and relict mounds. Advective circulation can also give rise to much larger lateral and vertical temperature gradients than would result from conduction alone.

Jan 1, 1999

By T. Lamb, T. Aston

"The Gays River Mine is located within a complex hydrological regime and underground mining environment. Early in the mine history, the operation was' affected by an inability to control mine water inflows. More recently, the reopened mine operation had been unable to maintain consistent drainage of groundwater to the mine workings and sustain low groundwater levels.In 1992, a series of investigations were completed, which in part Were funded under the Canada-Nova Scotia Mineral Development Agreement. The studies undertaken included: hydrochemical and drainhole well screen assessments, an air photo study of sinkhole features, numerical modelling studies and a major piezometer installation and overburden soil study. The results indicate that the water draining to the underground workings is comprised of two distinct components: precipitation over the drawdown cone catchment area and seepage from the nearby Gays River. In addition, a third ""balance"" component was' also found, which represents uncertainties in the areal extent of the drawdown cone and river seepage components.After outlining the history of the groundwater problems experienced at the Gays River Mine, this paper examines the various studies undertaken during 1991-1992 and concludes that control of groundwater levels in the mine area is achievable. A number of recommendations are also proposed for future study ."

Jan 1, 1993

By K. L. Ford, D. R. Lentz, J. H. Rickard, R. P. Taylor

Abstract - Coupled gamma-ray spectrometric and lithogeochemical studies throughout the Heath Steele Mines area, Bathurst Camp, New Brunswick, indicate that primary and secondary geochemical differences exist between the stratigraphic footwall and hangingwall crystal tuff units which host the massive sulfide deposits. Of particular interest is Th, and to a lesser extent, Zr and Y, which are more concentrated in the hangingwall than the footwall crystal tuffs, whereas the opposite is true for TiO2. This expands on earlier lithogeochemical-chemostratigraphic results concentrated in the eastern part of the Heath Steele belt that led to revision of the local stratigraphy, and necessitated a structural re-interpretation of the area with implications for exploration. Although K depletion (decreased K and K/Th) is typical of chloritized footwall zones of chloritization, these features were not fully apparent here due to the recessive nature of these altered zones. In situ radiometric analysis using a portable gamma-ray spectrometer is a rapid, inexpensive analytical technique for chemostratigraphic characterization of rock units within volcanogenic massive sulfide deposit settings, provided there is a contrast in the distribution of natural radioelements, particularly Th and K.

Jan 1, 1998

By M. A. Bérubé

Le but principal de cette étude était de vérifier certaines hypothèses de nature minéralogique susceptibles d'expliquer en tout ou en partie les variations importantes observées au concentrateur no 1 de Mines Gaspé quant aux performances du circuit B de broyage et de flottation. A cette fin, nous avons examiné six séries d'échantillons prélevés lors d'une campagne de mesures effectuée au cours de l'été 1980. Ainsi, la capacité du circuit de broyage semble dépendre principalement devariation s dans les caractéristiques minéralogiqu es et texturales du minerai, par exemple du rapport grenat/quartz, le grenat ayant une influence néfaste sur le rendement de cette étape, contrairement au quartz . Quant au processus de concentration, la fraction du cuivre sous f orme de minéraux difficilement récupérés par flottation, en particulier des carbonates de cuivre, varie de façon significative dans le minerai, et cela peut certes expliquer en partie des récupérations parfois plus basses. Malgré ces variations important es de la composition minéralogique globale du minerai et de sa broyabilité, le degréde libération des sulfures de cuivre dans le produit de broyage dépend presque seulement de la granulométrie de ce produit. Afin de garantir des performances égales à l'étape de concentration, il est fortement suggéré de contrôler cette granulométrie à un niveau constant.

Jan 1, 1984

By Frank S. Palkovits

Over the past 30 years, underground paste backfill systems have been implemented across the globe in many configurations, each with a primary focus of providing high quality backfill to enhance mine operations. However, a number of key design drivers are not always well understood within design engineering firms that occasionally undertake this work, leaving opportunities on the table. Perhaps not always obvious to those newly embracing paste technology, identifying key drivers and opportunities attainable through a paste backfill system is generally not straightforward, requiring innovative approaches to securing higher productivity, availability, performance, maintainability, and operator health and safety. This paper describes a number of applications and audits of paste backfill systems for several challenging mine operations. Key challenges include iterative design aspects of availability of good backfill materials, plant location and configuration, general arrangement, and binder receiving/storage systems.

Oct 1, 2010

By N. Bartie

Realising the circular economy (CE) is faced with some significant challenges. Process metallurgy and its infrastructure play key roles at the heart of making the CE work. Therefore, the enabling role of process metallurgy within the CE will be central to the discussion in this paper, touching among others on product and system design as well as the key metallurgical and other process fundamentals that need to be investigated and understood to make the CE a reality. The central role of materials and its processing will be discussed in an integrated circular cities perspective. A key focus will be a discussion on designing a resilient “Smart Materials Grid” using and innovating metallurgical process engineering tools, which will manage the flows through Sustainable Circular Cities. The discussion will be using copper as leitmotiv of the discussion i.e. from copper ore, to metal, to complex products, recycling, product design and simulation and its impact. INTRODUCTION The transformation of our economies from linear to circular models in which waste materials become resources is the focus of the Circular Economy (CE) paradigm. Its important intentions: to take control of the damage we have been doing to the planet to ensure its livability and the welfare of future generations. This can only be achieved through careful utilisation of human, natural and economic resources. The concept is derived from several others (Beaulieu et al., 2015), of which Cradle-to-Cradle, Extended Producer Responsibility (EPR), Industrial Ecology, and Life Cycle Thinking are just a few. Other relevant concepts include resource efficiency (RE), design-for-recycling (DfR), and various others discussed by Blomsma and Brennan (2017) and Homrich et al. (2018), as well as various R-hierarchies e.g. Reduce, Reuse, Recycle (3R). However, does society comprehend these concepts, the actors and actions that influence them and how they interact? Do we know the limits of CE? Does society understand the key role of Copper and other metals to rendering the CE operable? True life cycle and systems thinking, focussed on simulation-based, process and metallurgical design are required to answer these questions. Lieder and Rashid (2016) identify nine widely-varying fields that study the concept ranging from sustainability sciences and engineering to economics, business, law, and social sciences. Overlap between questions addressed highlights the need for interdisciplinary collaboration (Nanz et al., 2017) in the assessment of the current status, and further steps needed to transform to CE. This doesn’t apply just in academia, but also in industry where concrete implementation of CE needs to occur at multiple levels, thus necessitating cooperation between government, civil society and other stakeholders (Beaulieu et al., 2015).

Jan 1, 2019

By R. F. Favreau, R. A. Morin, G. W. Kuzyk, P. J. Babulic, N. J. Tienkamp

"Atomic Energy of Canada Limited is constructing an Underground Research Laboratory (URL) as part of a comprehensive program to evaluate the concept of nuclear fuel waste disposal deep in crystalline rock formations. Careful blasting methods have been used to minimize damage to the excavation surfaces. Good wall quality is desirable in any excavation. In excavations required for nuclear waste disposal, the objective will be to minimize blast-induced fractures which may complicate the sealing requirements necessary to control subsequent movement of groundwater around a sealed disposal vault. The construction of the URL has provided an opportunity for the development of controlled blasting methods, especially for drilling accuracy and optimization of explosive loads in the perimeter and cushion holes. The work has been assisted by the use of blast simulations with the mathematical model Blaspa.This paper reviews the results of a recent project to develop a controlled method of full-face blasting, and compares the observed field results with the results of a blast simulator called Blaspa. Good agreement is found between the two, and the Blaspa results indicate quantitatively how the blasting may induce damage in the final excavation surface. In particular, the rock in the final wall may be stressed more severely by the cushion holes than by the perimeter holes. Bootleg of the rock between the perimeter and cushion rows occurs when the burst out velocity imparted to it by the explosive loads in the perimeter holes is inadequate. In practice, these findings indicate that quantitative rock stress and rock burst-out velocity criteria can be established to minimize wall damage and bootleg. Thus, blast simulations become an efficient way to design controlled blasting and to optimize quality of the excavation surf ace."

Jan 1, 1989

By G. W. Poling, G. W. Hawthorn

"""Unassayable gold and platinum group metals"" have come into vogue in the 70sand 80s in certain jurisdictions as a means of perpetrating fraud. Usual arguments are that a particular ore is not amenable to ""conventional fire assaying"". Explanations for unassayable gold usually revolve around: evaporation of micron-size gold; vaporization of organic gold complexes; volatization of gold halides; alloying of gold with PGM's which prevents fusion or alloying which prevents collection .This paper reviews several of the myths and truths of gold and PGM assaying with the knowledge that not a single mine operatesin thefree world producing gold from unassayable ore.INTRODUCTIONAccurate analyses of exploration data and estimation of ore reserves are two of the most important functions of professional geologists and engineers in mine development. The problems encountered become particularly acute when dealing with precious metal values where very low grade deposits (in the range of 0.0 I Troy oz per ton or - 0.2 ppm (g/t) by weight) sometimes are economic or constitute ""ore"". To emphasize this ""scarcity effect"", a gold assay of 0.4 g/t (0.01 T oz/st) would result from the presence of a single 420 J.tm (35 mesh) gold flake (which weighs I milligram) in a 5 kilogram siliceous sample.The current high gold price and the prospects of high profit margins awaiting exploitation of even very low grade gold deposits, has spawned a sizeable group of incompetent or fraudulent gold assayers. These ""assayers"" will often report significant precious metal values where no such values occur. Failures of legitimate assay labs to confirm these ""significant"" assays are often attributed, by the charlatan, to their capability of detecting ""unassayable gold"" while the check assayer cannot. Geologists and engineers lacking detailed knowledge of sampling and assaying techniques are often hard-pressed to convince a gold-fever aroused company executive or investor-client to abandon a property on which attention-grabbling values have been reported, often by more than one laboratory. Professionals engaged in such evaluations must acquire an understanding of contemporary precious metal sampling and assaying techniques . Only then can they determine whether misrepresentation of values have been perpetrated"

Jan 1, 1989

By Yves Prévost, Guy St-Amant, Guillaume Dion

In 2010, the foreseen increase of copper grade in feed concentrate at the Horne Smelter was a significant challenge to maintain a steady copper flow throughout the process. The Horne Smelter process includes a Noranda Reactor, a Noranda Converter, pyro refining vessels and anode furnaces. Of the many challenges involved, improving the single casting wheel performance was one of the key topics. Its reliability had to be improved to meet the new production objectives, as anode copper production was to increase from 170 000 to 225 000 tonnes/year. A multidisciplinary team was dedicated to improve the casting wheel availability. The improvement project allowed pinpointing failure modes and reviewing operational and maintenance practices. The common understanding developed by the team allowed improving small components with a high gain-to-effort ratio. The project allowed reaching operational targets without any significant investment. INTRODUCTION In operation since 1927, the Horne Smelter, located in Rouyn-Noranda, in Quebec, has undergone several technological changes since its original construction. Built to smelt the Horne mine concentrates, the smelter also accepted some local custom feeds from the first years of operation. When the Horne mines stopped production in 1976, the plant became a 100% custom smelter. With the development of the Noranda Reactor (1973) and Noranda Converter (1997), the Horne smelter increased its productivity and improved its environmental performance. Continuous improvement of the Horne process allowed maintaining a high level of flexibility in the operation of the smelter, while improving its overall capacity.

Jan 1, 2019

By Jose R. Hernandez-Aguilar

"This paper revisits data from previous studies (Hernandez-Aguilar, 2010, 2011) to better understand the link between metallurgical performance and bubble size (Db) in the collection (pulp) zone of full-scale flotation systems. To obtain conclusive evidence exclusively related to the collection zone, the experiments were designed in those studies to control both the bubble generation method and the influence of the froth zone on the overall metallurgical performance in industrial and pilot-scale flotation columns. The results showed that generating small (1 mm) bubbles had a significantly higher beneficial effect on the recovery of coarse (> 75 µm) particles compared to the effect on fine (< 16 µm) particles. It was observed that, under certain practical conditions, the recovery of liberated coarse particles was close to zero, but could be recovered almost completely by reducing bubble size. The evidence revealed an apparent inconsistency of the so-called “k–PSb” model: at constant froth recovery, increasing Sb did not always result in higher recovery, as the model predicts. Overall, the evidence showed that it was not the Sb but the bubble size that proved to be a more reliable link to metallurgical performance. Only under highly suboptimal bubble size conditions (Db = 3 mm) did the froth zone seem to play a role, possibly reflecting an effect of pulp-zone bubble size on froth stability. However, for the most part, both the bubble generation method and the froth zone played inconsequential roles. An empirical model that better describes the kinetics of particle collection is proposed.INTRODUCTIONStudies showing the link between metallurgical performance and the interactions of bubbles and particles in the collection (pulp) zone of industrial flotation systems are scarce. A common belief among a number of mineral processing practitioners considers that “small” bubbles are favourable to recover “small” particles and that, conversely, “larger” bubbles are more efficient at floating “coarse” particles. However, perceptions like this are often substantiated by anecdotal evidence rather than by rigorous data analysis. What is the optimum bubble size for a particular case, how operating conditions can be adjusted to control a target bubble size, and whether the bubble size needs to be changed in the first place to optimize the operation of a flotation plant are questions that usually remain without solid answers.The studies reported by Hernandez-Aguilar (2010, 2011) were, in part, performed to test the effect of bubble size on flotation response and, thus, help address some of the above questions. The focus of the studies was to optimize the metallurgical grade-recovery performance of the Cu-Mo separation plant at Teck’s Highland Valley Copper mine (HVC). The outcome was compelling and significantly valuable: consolidated production data showed that in 2008 – the year when the recommendations from the studies started to be implemented – the Mo plant recovery was 89%, and gradually increased to 92% and 94% in 2009 and 2010, respectively, along with an increase in Mo concentrate grade. (Note: these production figures were not included in the publications quoted above; however, they were shown during an oral presentation at the 43rd Annual Meeting of the Canadian Mineral Processors in 2011.)"

Jan 1, 2016

By A. R. MacPherson

"This paper considers four aspects of autogenous grinding. Part 1 discusses statements on power efficiency of autogenous grinding compared to conventional crushing and grinding systems. It also demonstrates how high-power efficiency can be obtained from autogenous grinding systems. Part 2 considers a simple method of testing ores to determine the type of autogenous plant which can treat these ores on a commercial basis. A comparison of testing for simple small plants vs large plants or complex ores is included. Part 3 considers important aspects which have to be considered in proper design of an autogenous grinding system. Part 4 discusses problem ores and how they can be treated in an autogenous system. INTRODUCTIONAfter about 35 years of using autogenous grinding plants in the mining industry, it is perhaps time to review where we are with this, by now, relatively old development. Autogenous grinding in this paper means grinding run-of-mine ore by fully autogenous means or by using a ball charge to assist the grinding (semi-autogenous). Also, autogenous grinding can be eitherby dry or wet.The use of autogenous grinding in the fifties was making its slow start, and it was not until the sixties that more of the mining industry would consider its use . By the seventies, most builders of new plants would have considered autogenous grinding, whether they used it or not, in the final design.In the eighties, most large plants use autogenous mills and to reduce over-all operating costs, many older conventional plants are now looking and actually retrofitting these old plants with autogenous grinding mills.It is evident that many misconceptions concerning autogenous grinding are still current. Some come from improperly designed plants which result from poor test work. There are many considerations in the proper design of autogenous plants. Some ores have been found to give problems even in a well designed plant. These ore types are beginning to be classified and recognition in the test stage allows steps to be taken in the design so that these ores can be efficiently handled in autogenous grinding plants.Therefore, this paper will address the following aspects of autogenous grinding:1. Misconceptions regarding autogenous grinding.2. Testing of ores to ensure a well designed commercial plant.3. Important considerations in mill design.4. Difficult ores, their recognition and types of design to handle such ores."

Jan 1, 1989

By T. J. Bruce, G. E. Agar, R. Stratton-Crawley

"A technique by which the design of a flotation circuit may be optimized is described and examples of the application of this methodology are given.Three interrelated criteria are defined and the application of these criteria to the selection of the optimum residence time for each stage of a flotation circuit is demonstrated. The criteria are:(1) Add no material to the concentrate that is lower in grade than the feed to that separation stage.(2) Maximize the difference in recovery between the desired mineral and the gangue.(3) Maximize the separation efficiency.Three independent systems were examined experimentally and it was shown that application of any of the criteria yielded similar results.It was demonstrated how the application of rate measurements (Criterion #2) could lead to the selection of a set of optimized times for several stages in a flotation circuit and that these could in turn be used to simulate the results of a locked cycle test. The predictions were checked against the experimental results from a locked cycle test and there was substantial agreement between the predicted and observed results."

Jan 1, 1980

By Sadan Kelebek, Gerhard W. Heinrich, Peter F. Wells, Peter J. Whittaker

"Based on plant mineralogical data & circuit modelling, a pilot plant campaign was conducted at Strathcona Mill to identify options for improving the performance and operating flexibility of the copper-nickel separation circuits. Emphasis was given to options for combined treatment of the two feed streams (primary nickel ore rougher concentrate and copper ore circuit concentrate) to obtain more flexibility in the use of existing column flotation capacities. The pilot test results demonstrated that circuits which incorporated preconcentration prior to column flotation gave the best metallurgical results. Maximum column carrying rates were also related to the preconcentration. Predictions that increased circulating loads from the column scavengers would improve copper-nickel separation performance were confirmed.IntroductionThe copper-nickel separation circuits at Falconbridge's Strathcona Mill produce a high grade (>30% Cu) copper concentrate most of which is treated at the Falconbridge Kidd Creek smelter near Timmins. The remaining nickel concentrate which still contains approximately 20% of the feed copper is combined with other lower grade nickel/copper concentrates and shipped to the Falconbridge nickel smelter. The separation is required because of the limited capacity for copper processing in and downstream from the nickel smelter. Since no nickel is recovered at present in the Kidd smelter, the challenge is to minimize the nickel content of the copper concentrate.At present there are two independent separation circuits treating the rougher concentrates from the copper and nickel circuits as shown in Figure I. The minerals treated are mainly chalcopyrite and pentlandite but there is up to 25% of the copper present as cubanite in the copper ore. The circuits were kept separate because the flotation rates and froth characteristics were very different, probably because of the presence of cubanite and the large difference in the concentrate grades feeding the circuits."

Jan 1, 1995

Presentation Overview Corporate Overview Kutcho Project History Exploration Development Permitting Capstone -Pursuit of Value Building a mid-tier copper producer focused in the Americas Two operating copper mines; Cozamin& Mintomines High grade and low cost 2010 forecast: 80-85 million pounds of copper in concentrates Total Cash Costs* of ~US$1.10/lb to US$1.20/lb of payable copper produced Significant by-products of gold, silver, lead & zinc Growth potential Further expansions at existing Cozamin& Mintomines High grade copper development: Kutcho project, BC Investments in Nevada Copper & Silver Wheaton 100% North American assets Low risk ?Canada & Mexico ?Conventional operations

Oct 1, 2010

By P. C. Holloway, M. J. Collins, K. R. Buban

Historically, there has been little incentive to process copper-arsenic concentrates by leaching, rather than via established smelting operations. However, with ever tightening environmental emission standards and more arseniferous concentrates being produced, increasingly sophisticated and upgraded off- gas handling facilities are required in smelting operations, generating greater quantities of challenging metallurgical dust. In this context, oxidative pressure leaching offers distinct advantages. The current state of pressure leach technology delivers high extractions from copper concentrates, while producing an environmentally stable process residue, without atmospheric emissions of sulphur dioxide or arsenic. However, as was shown in an earlier study, copper extractions and environmental leach residue stability are impacted negatively by arsenic levels above 2.5 to 3.0 wt%. In further studies on pressure leaching of copper concentrates, Sherritt has developed process conditions that allow even highly arseniferous (> 7 wt% As) feed materials to be converted into environmentally stable residues, with excellent copper extractions. At the same time, pressure leach residues can be generated that are amenable to silver recovery via direct cyanidation, without pre-treatment by lime boiling or hot curing. The results of pilot plant test work are discussed. INTRODUCTION As the production of copper has expanded over the years, resource exploitation has extended beyond relatively clean, high grade and easily accessible ores, and copper sulphide concentrates increasingly come with higher levels of contaminants, such as As, Hg, Sb and Bi, and lower copper and sulphide sulphur concentrations. However, as concerns on the impact of the emission of such contaminants into the natural environment have mounted, import/export restrictions have made the trading of copper concentrate more difficult, and, in particular, the export of concentrates with greater than 0.5 wt% As; under ever tightening environmental emission standards, these lower quality concentrates can be a significant strain on the off-gas handling facilities in smelting operations. The processing of arsenic bearing copper concentrates is, therefore, an increasing challenge as copper producers seek to exploit more complex mineralization and to extend the life of their operations. In this environment, processing arsenical copper concentrates with high temperature pressure leaching (>200°C) must be considered as an alternative to pyrometallurgical processing. Leaching kinetics for copper minerals at these temperatures are rapid and acid is generated for potential use in heap leach operations. Pressure leaching of copper concentrates eliminates gaseous emissions of arsenic and sulphur, while generating environmentally stable residues, as measured by the Toxicity Characteristic Leach Procedure (TCLP) (Kalanchey et al., 2007). Both medium and high temperature pressure leaching of copper concentrates have also proven themselves technically on an industrial scale (Marsden et al., 2004) with high copper recoveries.

Jan 1, 2019

By Bian Xue, Wu Wenyuan

In order to recycle the iron and rare earth from the Baotou tailings, a roasting method was developed involving the addition of C, Ca(OH)2, and NaOH, designed to both decompose the rare earth minerals and convert haematite into magnetite and allow its recovery. The effect of various parameters such as roasting temperature, roasting time, C addition, Ca(OH)2 addition, and NaOH addition on the reduction ratio of haematite (iron sesquioxide) and decomposition ratio of rare earth were investigated. The results showed that the reduction ratio of haematite and the decomposition ratio of rare earth could be as high as 85% and 95% respectively when the roasting temperature was 650°C, roasting time was 60 min, Ca(OH)2 addition was 4%, NaOH addition was 3%, and C addition was 1.5%.

Jan 1, 2014

By R. F. Tylecote

"The closure of the iron and steel making plant at Consett in Durham in 1980 marked the end of an era-a long era in the world&apos;s iron and steel making history. Iron had been smelted in the Tyne-Wear area of the United Kingdom since before the Roman period, but apart from evidence of smithing in the Roman period, the historical evidence for iron production on a large scale dates from the 14th century.The paper describes sites in the area that first worked the bloomery process and then, in the 16th century, as did most of Europe, graduated to the blast furnace process. The main emphasis is on the early introduction of coal to iron making at the end of the 18th century and the use of the steam engine to render integrated sites possible.IntroductionIron has been smelted in the Tyne-Wear area of the United Kingdom since before the Roman period, but apart from clear evidence of smithing in the Roman period, the historical evidence for iron production on a large scale dates from the 14th century. Like many other countries, the United Kingdom benefitted from the visits of foreign observers, such as Jars, Svedenstierna, de la Houliere and Dufrenoy, who left detailed records. In modern times, information gained from excavations of iron and steel making sites has supplemented the historical record.Natural ResourcesUp to the end of the 19th century, the industry was wholly dependent on its own resources. Iron ore, limestone, wood, coal and water power were all found a short distance away from the main valleys of the Tyne and Wear. The main sources of iron ore were the coal measures, with rider ores from the lead vein s in Alston and Weardale, and carbonate ores from shale horizons in the Carboniferous limestone. Good coking coal was available from West Durham and reasonable coal from the Tyne valley and Northumberland limestone. Limestone was plentiful."

Jan 1, 1983

By Z. B. Karshigina, Ye. G. Bochevskaya, E. A. Sargelova, Z. S. Abisheva

At the present time of depletion of reserves of high grade rare earth deposits, it has become necessary to involve low grade raw materials in the production. Among the rare-earth deposits in Kazakhstan, the most advanced one is Kundybai. The mineralogical composition of ore is represented by the minerals: a-quartz SiO2, kaolinite Al2(Si2O5)(OH)4, muscovite KAl2(AlSi3)O10(OH, F)2, albite Na(AlSi3O8), rutile TiO2, and hematite Fe2O3. It contains: 59.1–68.5% SiO2, 19.1–28.4% Al2O3, 7.2–10.3% Fe2O3, 0.51–0.84% TiO2, 0.10–2.32% CaO, 0.39–2.10% MgO, and some other elements. Rare earths are located in the clay part (kaolinite, goethite) in the ion-adsorbing form, and they are represented by the main ore mineral churchite - YPO4·2H2O. In the ore the amount of rare-earth elements (total REEs) varies from 210 to 600 ppm. Elements of the cerium group predominate (La, Nd, Ce, Sm, Pr) and the yttrium group is represented by Y, Gd and Dy. Extremely thin intergrowth and accumulation of REEs in clay minerals cause difficulties in their recovery using flotation and gravitation beneficiating methods. The purpose of this work is the development of an extraction method for the main mineral components of ore with transfer of the total REEs into the solution with high extractive efficiency. Results of the experimental work into high-temperature processing (hereinafter referred to as sulphatization) of ore by sulphuric acid with subsequent dual-mode water leaching of the sulphate product (sinter) are presented in this manuscript. The influence on the ore sulphatization process of some physical and chemical parameters: liquid-to-solid (L/S) ratio (0.36–0.86 cm3/g), temperature (100 – 500°?), concentration of sulphuric acid (3.0–18.62 mol/L) and duration (0.5–2.5 h) was studied. Optimal technological parameters of the process were determined to be: L/S ratio 0.36–0.46 cm3/g; temperature 500°?; concentration of H2SO4 9.0 mol/L; soaking time 1.5 h. Under these conditions it is possible to obtain extractions of ~ 85% REE, as well as 58% aluminium and 10% iron. XRD and IR-spectroscopy analyses were used to determine the material constitution of sulphated cakes obtained under optimal conditions. It was found that the primary minerals are quartz and muscovite; however different sodium aluminium silicates and albite minerals are present as well. In the future, it is proposed that the leach residue cakes are processed into precipitated silica (“white soot”).

Jan 1, 2016