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By Laxman M. Amaratunga, Glen G. Hein

"Most mining operations utilize the tailings coarse fraction for backfilling while the fines have been traditionally disposed on surface in tailings ponds. Recently, paste fill has been developed which uses a fraction of the fine tailings (10-30% w/w of <45 µm) in the fill (Brummer and Moss, 1991). This development allows a mining company to use a portion of the fines for backfill, but still requires surface impoundment of the unused tailings. Reactive tailings have the added problem of generating acid after being oxidized in the presence of air and water. Treating acid mine drainage (AMO) from these ponds has proven to be a costly environmental problem and a long term liability.One practical method to reduce the environmental impacts associated with the tailings pond is to modify the physical and chemical prope1ties of the fine tailings so that they can be utilized in an environmentally safe manner. This was accomplished by agglomerating the tailings into pellets and using theses pellets as a coarse aggregate in a total tailings paste fill. This technology does not require any supplemental sand or gravel to be added to the fill.This paper presents the ongoing research into utilizing these tailings agglomerates as a coarse aggregate fraction, in order to create a total tailings paste fill which has supe1ior strength characteristics.IntroductionThe Canadian mining industry produces more than 500 million tonnes of tailings and waste rock annually. The waste rock is usually returned underground for use as rock fill in underground support with the tailings remaining to be disposed of in tailings ponds. Tailings consist of a coarse and fine fraction, and can be either reactive (acid generating) or non-reactive. In an effort to minimize tailings disposal, the coarse fraction has been used underground in backfill along with alluvial sands. This diversion of the coarse fraction, however still requires that the fine fraction be disposed of in tailings ponds. This disposal creates the potential for acid mine drainage (AMO), since in the presence of air and water, reactive sulphide tailings can generate acid. It is therefore an important goal in the mining industry to minimize the disposal of fine tailings into tailings ponds since the treatment of AMO is a costly and long-te1m environmental liability."

Jan 1, 1997

By P. Woo, N. Brodusch, R. Gauvin, H. Demers

The scanning electron microscope (SEM) was primary developed for imaging applications. With the introduction of the Si(Li) energy dispersive spectrometer (EDS), simultaneous imaging and x-ray microanalysis became possible. However, long working distance and high current were needed because the position and small solid angle of the EDS detector. The high spatial resolution is generally obtained at short working distance with a SEM. Unfortunately x-ray microanalysis is never performed in the best imaging conditions. The annular silicon drift detector system is inserted below the objective lens which gives a higher solid angle (up to 1.3 sr). In consequence, a lower working distance and probe current can be used. An improved spatial resolution becomes possible during x-ray microanalysis. With this x-ray detector installed on a cold-field emission scanning electron microscope, quantitative x-ray microanalysis with high spatial resolution at low beam energy and low current becomes possible with the possibility of using the various different type of imaging at the same time. Also, since the count rate can be as high as 1,500 kcps, which lowers significantly the detection limit of elements as well as the minimum feature sizes of different phases that can be distinguished.

Jan 1, 2014

By H. Tokuichi, H. Ohta, K. Sato

A physical separation technique of tantalum capacitors from waste print circuit boards is in a development phase in Japan. A tantalum capacitor consists of an exterior coating of resin (epoxy resin) mixed with silica fillers and a sintered tantalum compact, the latter of which is impregnated with manganese dioxide (MnO2). In this study, we developed a process to recover tantalum oxide from tantalum capacitors by roasting, fluidized-bed separation, and hydrometallurgical treatments. Our aim was to recover tantalum oxide with high purity from waste tantalum capacitors. In this process, the exterior resin was selectively decomposed by roasting without oxidation and pulverization of the sintered tantalum compact at a proper temperature. Furthermore, the decomposed resin with silica filler on the surface of the sintered tantalum compact was mostly removed by fluidized-bed separation. The surface of the sintered tantalum compact was cleaned by ultrasonic irradiation in an acid solution to remove the MnO2 paste as well as residual Ag paste and silica filler. The impregnated MnO2 in the sintered tantalum compact was leached as Mn(II) ions using acid solutions via the reducing effect of tantalum metal to recover purified tantalum oxide.

Jan 1, 2014

By Jianhong Xu, Hailong Hou, Yu Jing, Yundong Wang

Microcapsules containing extractants have been applied to the extraction of metal ions, organic acid, pesticide, and herbicide. The difficulties of conventional liquid-liquid extraction in phase mixing and phase separation (e.g. high energy consumption, emulsification phenomenon, and large equipment requiring housing) could be overcome in a liquid-solid system by using microcapsules. Even some extractants with high viscosity and poison can be encapsulated conveniently into microcapsules. In the work described in this paper, polysulfone microcapsules containing 2-ethylhexylphosphonic acid mono-2-ethylhexyl ester (P507) were successfully prepared. Then, the extraction rate of lanthanum, samarium, and erbium in dilute solution into the microcapsules containing P507 was studied. An extraction kinetic model is proposed based on the interfacial reaction model accompanied by intraparticle diffusion expressed by Fick’s law. The extraction efficiency was obtained by using the kinetic model with the separately determined complex formation rate constant and the intraparticle diffusion coefficient. From the results, it was confirmed that the kinetic model could predict the experiment in a microcapsule system over a wide concentration range of lanthanide solution. In the future, the model could provide good guidance for the design and application in extraction of lanthanides.

Jan 1, 2014

By Aimee Williamson

Bio-mining is a cost-effective, energy efficient and potentially environmentally protective method of exploiting a variety of ore deposits, from high-grade base metal deposits of nickel to lower-grade uranium deposits such as are found in the uraniferous quartz-pebble conglomerate beds of the Quirke Syncline, Elliot Lake, Ontario, using bacterial metabolic processes are used to enhance the solubilization of the metal(s) of interest. The cost effectiveness of bio-mining in Elliot Lake depends on the bacterially-influenced dissolution metals of economic importance, as well as on the ability to suppress microbiological activity following the release to solution of the economically important elements. This poster provides an overview of elemental recovery results and release kinetics for more than 50 elements. The data is obtained using a combination of advanced spectroscopic tools (ICPAES and ICP-MS) to analyze solutions obtained from a series of environmentally controlled biologically mediated uranium ore extraction experiments. The focus will be on rare earth elements, potential environmentally toxic elements, and metals of economic importance. Furthermore, the importance of these findings will be put into an economic and geologic context for the re-development of the mining industry of the Elliot Lake Camp.

Aug 1, 2010

By John D. Carriere, Jean Leung, Daniel. Laguitton, Michael .. J. Burrows

"The objective of this expert system application in a copper-zinc concentrator, is to increase the availability of assay data through substitution for erroneous data. The assay data is used by the Operations personnel and automatic process control strategies to control flotation circuit performance. Two online X-ray fluorescence analysis systems analyze 24 streams across four divisions to provide the required data. Data availability of individual streams average 90% while the resulting circuit availabilities are 70%. The unavailability of correct data results from mechanical, X-ray or process problems. The erroneous data can be identified by process experts using knowledge relating stream interdependencies and historical process behaviour. An expert system was devised to provide real time identification of erroneous data and substitute for that data with the ultimate aim of improving metallurgical response by both the operator and automatic control systems.The paper discusses the early development work and architecture of the knowledge base designed using Neuron Data's expert shell NEXPERT.INTRODUCTIONThis paper describes the early development of an expert system (ES) prototype at the Kidd Creek concentrator. The project is a joint undertaking of the Kidd Creek Division of Falconbridge Limited and CANMET. Kidd Creek entered into the project to apply heuristic rules to the XRF assay data to improve the metallurgical control of the flotation process by increasing availability of analysis data. This project provided a problem of suitable size and future expandability while making use of a wide base of in-house expertise and experience. Another important benefit is the introduction of expert system technology to the site providing tools and experience for future system selection and technology application. CANMET has an on-going interest in assessing various aspects of ES technology (hardware, software, knowledge engineering, integration) while addressing problems in process operations that could most readily benefit from the technique. CANMET's role in the project was to provide project assistance, financial support, substitution algorithm assistance and software for data substitution. The knowledge bases (rules) have now been completed and are to be tested in the online environment."

Jan 1, 1989

By M. Nowak

Extreme high grades often exert too much influence on surrounding block estimates, leading to a smearing of high grades that can cause overestimation of mineral resources. This concern underlies the decision to cap high grades, i.e., to lower the highest grades in a drill hole data set to some level deemed to be reasonable for producing reliable resource estimates. The choice of a capping level can, by itself, have a large influence on the analysis of the economic viability of a project. Despite this, it is often a choice that is made arbitrarily, guided more by art and tradition than by sound technical analysis. This paper makes suggestions for good capping practices, using methods and procedures that have been presented elsewhere, in the technical literature and in conference papers. The authors propose to use capping only if absolutely necessary, and review examples of cases where capping may not be necessary, or where capping does not make the estimates more reliable. It addresses two particular instances where capping is recommended: (1) where there is good reason to believe that estimation with uncapped grades will overstate the true underlying average grade, and (2) where there is potential for overly optimistic assessments of local block grades in the vicinity of some very high grade assays. The proposed capping practice is based on adjustments to only a very few high grade assays. The level of this capping can be assessed by bootstrapping and through analysis of the effect of extreme values on summary statistics like the average and the coefficient of variation. The paper recommends that good practice include an examination of the potential for local overestimation in higher grade areas.

Aug 1, 2013

By M. Remillard, M. Garon, M. Thérien, D. Hodouin

"Nineteen streams in the copper and zinc flotation units of the concentrator of the Matagami division of Noranda Inc. have been sampled for particle size distributions chemical assays, and gold and silver contents in each particle size interval. The data are upgraded by the mass balance package BILMAT. These results, in conjunction with microprobe analyses, are used to quantify the silver losses in the copper clean¬ing flotation stages and to determine how overall silver and three silver-bearing minerals are recovered in the plant final products.RESUME Dix-neuf produits des circuits de flottation du cuivre et du zinc de la division Matagami de Noranda Inc. ont été échantillonnés pour leur distribution granulométrique, analyses chimiques et le contenu d'argent et d'or des fractions granulométriques. Le bilan métallurgique a été obtenu à l'aide du programme BILMAT. Ces résultats couples avec une etude minéralogique à la microsonde sont utilises pour quantifier les pertes argentifères aux diverses étapes de nettoyage du circuit de cuivre et pour de¬terminer la repartition de l'argent et des trois minéraux argentifères dans les produitsfinals de l'usine.INTRODUCTION The Matagami division of NORANDA INC. produces copper and zinc concentrates from a sulphide ore which contains a small amount of gold (0.34 g/t) and silver (20 g/t). Only a small part of the silver (~-30%) is presently recovered in the copper concentrate. The objective of this study is to evaluate the metallurgical behaviour of silver, and secondarily of gold, in the flotation circuit for a possible increase of its recovery in the copper concentrate to improve its economic value. For that purpose it is worthwhile to study the minerals which bear silver and to determine the critical flotation stages where they are lost in the tails. The methodology proposed here is based on a coupling of a mineralogical study and a detailed mass balance calculation in the flotation circuit."

Jan 1, 1986

By Dale A. Sketchley

"Abstract - Large differences in gold content between rock fragments or small volumes of unbroken rock, commonly referred to as the nugget effect, are related to sample and subsample weights, crushing and pulverizing sizes, and density, shape and size distribution of gold grains. Gy’s sampling theory, which relates gold grain characteristics to sample weight, fragment size, and error, provides a method of estimating a sampling constant that quantifies deposit characteristics that contribute to the nugget effect.Sampling control tests provide methods for determining the sampling constant. Sampling control charts, comprising nomographs that relate sample weight, error and particle size, use the sampling constant to optimize sample weights and comminution sizes, and to minimize the variability introduced by incorrect or inadequate sampling and sub-sampling procedures. The sampling constant is useful in classifying ores into arbitrary categories according to expected sampling difficulties. Spatial variations in the sampling constant can be used to aid in defining ore types for modeling geological domains, which should have unique sampling protocols. Examples that illustrate the relationship between the sampling constant and expected sampling problems, and the spatial variation of the sampling constant, are given for various deposits. Case histories of sampling programs are presented.Quality control procedures consist of monitoring sample contamination amd accuracy and precision of analytical results. Accuracy, precision and contamination are monitored by including standards, duplicates and blanks into batches of samples being analyzed. Analytical results of standards and blanks are plotted on a batch-and-time basis with control lines at plus or minus two/three standard deviations as confidence limits of the accepted round-robin value. Examples are given that illustrate a laboratory system in statistical control and the various types of problems encountered when interpreting quality control data. These include sample transcription errors, degradation, contamination, sampling errors, detection limit problems, instrumental drift, lack of statistical control, laboratory bias, procedural problems and data tampering.Where bias is negligible, precision, as a function of concentration, can be obtained from duplicates using the Thompson-Howarth precision algorithm, which also allows calculation of the practical detection limit. Examples are given that illustrate variations in precision, with grain size, for different types of duplicates, as well as mineralogical changes and sampling errors."

Jan 1, 1998

By Masaharu Moteki

At present, the Unmanned Construction System (UCS) is adapted on initial operation against the disaster. For instance, it has been used in Unzen. When the scale of the disaster is big, or the danger to a person is big, the UCS is expected as a quick and safe construction technology. The Remote Controlled UCS is introduced into construction site to repair earth-fill embankment or to remove rock, as urgent measures after the disaster outbreak. For improvement of the operability, shortening of the skill time, reduction of the fatigue, examination of the most suitable system (placement of a monitor and the joy stick) of the remote system is a main theme in the Public Works Research Institute (PWRI). In the report of this year, when the operator controls construction machine directly or he controls it using remote system (he can check the situation of site only via monitor), the degree of achievement, construction efficiency and construction precision were inspected. The remote system which is based on the system used at Unzen was developed at PWRI.

Aug 1, 2013

By R. E. Whitehead

"Abstract - Biotite from aplite that appears to have crystallized simultaneously with disseminated Cu-Mo mineralization at MacLeod Lake in northern Quebec yields K-Ar dates in the range 2032 ± 44 Ma to 2097 ± 47 Ma. The mean value, 2065 Ma, is some 400 Ma younger than the 2485 ± 64 Ma date obtained from a hornblende concentrate separated from a granodiorite sill that is spatially, but apparently not genetically, associated with the disseminated Cu-Mo deposit. The mean biotite date of 2065 Ma at MacLeod Lake, a K-Ar date of 2055 Ma reported from an area 50 km east of MacLeod Lake and published dates of about 2000 Ma for galenas from the Chibougamau Cu-Au ores, suggest a widespread thermal and mineralization event, in that part of the Superior Province east of James Bay, that differs from the common late Archean mineralization that is related to volcanism and metamorphic-intrusive activity.Sommaire - Les datations au K-Ar sur des biotites qui auraient cristallise en même temps que la minéralisation disséminée en Cu et Mo dans une aplite mise en place au Lac MacLeod, au nord du Québec, s&apos;échelonnent entre 2032 ± 44 Ma et 2097 ± 47 Ma. La datation moyenne, soit 2065 Ma, est d&apos;environ 400 Ma plus récente que la datation obtenue pour un concentre de hornblende extrait d&apos;un filon-couche de granodiorite qui est spatialement mais ne serait pas génétiquement associée à la minéralisation disséminée en Cu et Mo. La datation moyenne de la minéralisation au Lac MacLeod, soit 2065 Ma, la dataion au K-Ar à 2055 Ma citée pour une région à 50 km à l&apos;est du Lac MacLeod, de même que les datations d&apos;environ 2000 Ma établies pour des galènes qui se trouvent dans le minerai aurifère et cuprifère à Chibougamau, suggèrent un phénomène thermique répandu qui aurait produit la minéralisation dans cette région à l&apos;est de la Baie James, dans la province du lac Supérieur. Ce phénomène serait diffèrent de celui qui aurait produit la minéralisation répandue d’âge archéen tardif qui elle serait associée à l’activité volcanique et à l‘activité intrusive accompagnée de métamorphisme."

Jan 1, 1992

By Wei Ye, Lei Li, Xianjian Guo, Feng He

The purpose of the paper is to study the smelting process of copper concentrate with high Cu/S ratio produced in Africa or other areas around the world and theoretically analyze the influence of slag type on the smelting process. The paper has compared the processes used successfully in smelters, including CAPEX and OPEX and other factors, to provide reference for the industry to select the best process under specific conditions. INTRODUCTION At present, a great amount of copper concentrates with high Cu/S ratio, with chalcocite being the major source of copper, have been produced or are about to be produced in the Copperbelt in Africa and other countries in the world. The copper concentrate is usually high in copper (~30–55wt.%) and low in sulfur (~10–15wt.%). The conventional processes of smelting-converting and roasting-leaching-electro-winning for treating this type of concentrate is high energy and materials consumption such as electric power\fuels\flux et al. The direct-to-blister (one-step) smelting process can be considered favorable due to a reduction in processing steps. Up to now, there are four copper smelters in the world operating with the direct-to-blister smelting process. There certainly are problems in the process such as high copper content in the slag, strict operation requirement of slag cleaning electric furnace, etc. Except technical issues, the impact of factors on economical aspect such as power and material cost needs to be considered comprehensively in process selection. In this paper the technical issues of direct-to-blister smelting process, including slag treatment process, were analyzed and the comparison to the conventional processes and economic analysis of direct- to-blister smelting process to treat high Cu/S ratio concentrate were done.

Jan 1, 2019

By Andre Vien, D. H. Hendriks, D. Willet, B. C. Flintoff

"For many years the training of mill operators was an informal process which involved senior operators training junior operators on an as-needed basis. Recognizing the needs and benefits of better training some companies, and even some provinces through legislation, have moved to ensure that a formalized training program exists in the plants. These programs tend to focus on basic safety and operations. Much remains to be done. With the more rapid incorporation of new processing and information technologies, and the goal to keep labour costs to a minimum by doing more with less, operator training has become a critical issue. This paper briefly examines the evolution of training programs, provides some measure of their effectiveness, and explores some recent developments which could have a very significant impact on the industry.PreambleOne can draw an analogy between operator training and the mechanical support of production equipment. When we buy a piece of process equipment it is critical that the mechanical installation be done properly to ensure that the device will operate at peak performance on start-up. Once the equipment is running, some form of preventive maintenance is required to ensure continued efficient operation. The operators and the equipment are both elements of the processing system, yet the former do not seem to get the same ""start-up"" or ""operating"" support. Are they not seen as critical to efficient operation? Why is this so? What impact does it have? How can we improve?These are all very logical and important questions, although the answers are not nearly so obvious. In the mineral processing industry, progress in operator training over the past thirty years has been limited and quite variable. To address the perceived need for a more coherent and consistent approach, some provinces have moved to incorporate basic training requirements in their legislation. Some plants have gone a few steps further by building integrated training programs into their operating culture. One conclusion which can be easily reached is that there are no standards, and that the range of training activity is great. In other words, a flotation operator is not a flotation operator is not a flotation operator. Their knowledge and skills are very much a function of training and experience, which are clearly corporate and site specific."

Jan 1, 1994

By T. Moyo

Ammoniacal solutions are attractive and effective lixiviants for the dissolution of chalcopyrite, allowing for the formation of stable ammine complexes with copper while rejecting iron. Chalcopyrite dissolution in ammoniacal solutions is reviewed from literature, and results from a recently completed study evaluating it through electrochemical and controlled leach studies are discussed. Ammoniacal leaching of copper ores dates back to 1916 in the leaching of low grade copper ores in the Kennecott leaching plant and in the Calumet and Hecla mines, followed by the commercialisation of the Sherrit Gordon process for the ammoniacal leaching of sulphide copper ores in 1948. Since then ammoniacal leaching has received intermittent focus from researchers with the development of the Anaconda’s Arbiter process and the unsuccessful Escondida process. In these leaching processes, the formation of surface deposits has historically been reported to result in the passivation of the chalcopyrite surfaces, whereas oxygen mass transfer in solutions has been the focus of process design improvements. Results of studies on high grade chalcopyrite material is presented in this paper which show that although a surface deposit was formed, it did not result in the passivation of chalcopyrite in ammoniacal solutions, and that while oxygen transfer is critical, it is not the direct oxidant in the reaction. INTRODUCTION Ammonia and its salts have been used in the hydrometallurgical treatment of non-ferrous metals such as copper, cobalt and nickel from oxide ores. Its chemistry allows for the selective dissolution of desired metals, leaving behind gangue including iron. Chalcopyrite is a copper sulphide ore known to be very stable and hence recalcitrant to many hydrometallurgical processing methods. Thus, the dissolution of chalcopyrite continues to be an attractive area of research. The dissolution of chalcopyrite in ammoniacal solutions is possible due to the stabilisation of copper(I) and copper (II) ions by ammonia in alkaline media. Ammonia has become more attractive as a lixiviant during the present socio-economic climate due to its low toxicity, relatively low cost and ease of regeneration by evaporation (Meng & Han, 1996). Ammonia processes are ideal when the gangue minerals are acid consuming e.g. with calcareous or dolomitic gangue (Paynter, 1973).

Jan 1, 2019

By M. Sider, D. L. Piron

"The cost of energy is a growing factor in the cost of zinc production. Electrowinning from sulphate solutions with periodical reverse current or from acidic chloride or alkaline solutions could be the way to lower costs. A comparison of the energy consumption for electrolysis shows that the alkaline solution may be considered the most promising alternative. IntroductionZinc electrowinning from sulphate solutions is a very energy intensive process (typical specific energy usage : around 3.1 kWh / kg zinc). This energy consumption represents 35% of the total cost of electrolysis and is the most important cost component.In order to lower costs, the energy efficiency of the sulphate process should be improved or alternative processes should be developed. In this laboratory, that objective has resulted in research on electro winning in sulphate solutions with the use of periodical reverse current (PRC) and in acidic chloride and alkaline solutions.Some of the results obtained in this laboratory and already published and new experimental results reported in this paper make it possible to present a more critical comparison of the energy consumption of the three processes mentioned.ExperimentalThe experiments at constant current were carried out as described previously. Specific conditions are indicated in Table I. Current efficiencies were obtained by weighing deposits.ResultsThe best energy consumptions for electrolysis obtained in this laboratory for the three media are presented in Table I. More detailed results can be found in the papers already mentioned."

Jan 1, 1988

By George J. Cymerman

Since 1978, Syncrude Canada Ltd. has been expanding steadily its oil sand mining and processing capacity while reducing unit energy requirements. At the current rate of ~180 M tonne/year oil sand, the originally adopted Karl Clark?s 80°C version of the Hot Water process would require some 53 PJ (53 x 1015 Joule) heat. If this amount of heat were to be derived from burning natural gas it would cost $369 M/y (at $7/GJ). Fortunately, effective recovery of waste heat from the upgrading operations at the Mildred Lake site obviated the need for water heating with natural gas. The oil sand hydrotransport technology, first utilized commercially in the mid-nineties at Syncrude?s North Mine, reduced slurry temperature from ~80°C to ~50°C, cutting the unit heat requirement from 293 MJ/tonne to 183 MJ/t. However, in the new 100 Mt/year Aurora Mine even a 50°C process temperature would still require 18 PJ of heat energy per year. Only limited quantity of 65°C water could be imported from Mildred Lake site and some waste heat would be available from Gas Turbine Generators. A new, Low Energy Extraction (LEE) process was developed to minimize heat requirements by further reducing process temperatures to 25°C in the slurry pipeline and 35°C in the Primary Separation Vessels (PSV). Unfortunately, when applied to certain ore facies in Aurora, the LEE process yielded disappointing bitumen recoveries, necessitating an increase in slurry temperatures from 25°C to 35°C in the pipeline and from 35°C to 42°C in Primary Separation. In 2004, an Optimized Low Energy process (OLE) was developed and piloted at Syncrude Research using a 2000 kg/hour pilot plant. In the OLE process, most available heat is applied up-front to produce 45°C pipeline slurry at a density of 1.58 kg/l. Once the pipeline conditioning process is completed, the slurry is flooded with cold (10°C) water, resulting in PSV temperature of 34°C. The overall heat and water requirements are the same, or slightly lower, compared to those of the LEE process ? but without the penalty in bitumen recovery. Processing 100 Mt oil sand at 34°C (96 MJ/t), rather than at 50°C (183 MJ/t), will save $61 M/y (@ $7/GJ) and reduce green house gas emissions by ~47 %. The OLE process is currently being implemented at the Syncrude Aurora Mine.

May 1, 2006

By John Keay, Douglas R. Shaw, Gregory J. Olson

"The results are presented herein of on-going laboratory investigations of heap bacterial oxidation as a pretreatment method for the recovery of sulfide-encapsulated gold from ore samples of the Rhodes Mining Lahoca deposit, located in northern Hungary. The ore contains just over 2 g Ault and as much as 6% sulfide, mostly as pyrite. Laboratory aerated column tests conducted at Colorado Minerals Research Institute (CMRI) showed that mild to moderate biooxidation of the finely crushed ore occurred rapidly in approximately three months, with cyanide gold recoveries of up to approximately 60%. The results contrast cyanide gold recoveries of less than 30% from non-pretreated ore. Mineralogical examination of biooxidized residues indicated that the mechanism of biooxidation of crushed ore was related to microorganisms following pathways along structural anomalies in the pyrite grains. Such selective biooxidation is an important result to minimize the biooxidation and aeration time and effluent neutralization costs in a commercial operation.IntroductionThis paper describes the development of a heap biooxidation process for the Lahoca gold project The Lahoca deposit is situated in northern Hungary, at the site of a former copper mining area. Lahoca is a large tonnage, relatively shallow, deposit amenable to open-pit mining Lahoca has the potential to produce I 00,000 ounces of gold annually for at least ten years at its minimum reserve base.The laboratory test work conducted initially by Rhodes Mining NL demonstrated that the Lahoca refractory ore is amenable to whole ore and concentrate oxidative pretreatment and cyanidation. Capital and operating costs, however, for such milled ore treatments were considered relatively high. Rhodes Mining subsequently evaluated heap leaching methods, particularly heap biooxidation pretreatment, as a process route that promised superior economics suitable for exploiting lower gold grade refractory ores.This paper focuses on the laboratory column (simulated-heap) biooxidation test work, particularly the proposed mechanism of selective microbial oxidation, wherein only a small amount of the sulfides are oxidized sufficient to allow commercially acceptable gold recoveries by conventional cyanide heap leaching. Selective or minimal biooxidation of sulfides is considered important to minimize biooxidation time and aeration requirements, as well as acid effluent neutralization costs."

Jan 1, 1998

By Xuanxue Mo, Zengqian Hou

"Yidun Island Arc in Sanjiang Region is one of the important Kuroko-type volcanogenic massive sulfide deposits (VMS) districts in southwestern China. Tectonically, the region is situated in the eastern portion of the Tethys orogenic belt, and is a late Triassic extensional island-arc. The island-arc underwent earlier arc construction, intra-arc rifting, later arc construction and backarc spreading during its evolution and formed a complete trench-are-basin system, i.e., the Ganzilitang suture zone, forearc basin, outer are, intra-arc rifting zone, inner arc and back-arc basin from east to west across the island-arc.Kuroko-type VMS deposits and occurrences discovered so far are limited to within the intra-arc rifting zone of the Yidun Island Arc and tend to occur in groups or clusters distributed in separate faultbound basins; e.g., Zengke, Gacun and Xiangcheng basins, in which submarine volcanic rocks, especially bimodal volcanic suites, were extensively developed. The VMS deposits formed in the waning stages of volcanic eruption in the late Triassic and are hosted in the uppermost section of acidic volcanic rocks (rhyolitic-dacitic) within a bimodal suite. The orebodies are associated with these host rocks: rhyolitic-dacitic pyroclastic rocks and exhalites and chemical sedimentary rocks.All VMS deposits in Sanjiang are of the Zn-Pb-Cu type. The architecture of the deposits is twofold, consisting of concordant lenses and layers of massive sulfides and an underlying discordant stockwork. All deposits show metal zoning. The hydrothermal alteration related to mineralization is strong and asymmetrical, and is found dominantly within the stockwork zone and host rocks surrounding the ores. There are two alteration patterns: (1) chloritization - sericitization - silicification from outer sides to core of the alteration pipe; and (2) silicification + sericitization - silicification + carbonate from the lower to upper part of the alteration pipe.The mineralization temperatures of ore-forming fluids vary in the range 200°C to 335°C. Thermal evolution of the fluids displays temperature profiles that first increase and then decrease with progression of mineralization processes. The majority of black-ore minerals (Groups I and informed during intensifying stages of hydrothermal activity, and the majority of semi-yellow-ore minerals (Group III) formed during a thermal maximum. There followed the formation of fine-grained black-ore minerals (Group IV), which interacted with barite layers during decreasing temperature phases.Oxygen and hydrogen isotope studies show that the sources of ore-forming fluids for the Kurokotype VMS deposits are two-fold: seawater and magmatic hydrothermal solutions, represented by the Kuroko deposits in Japan and Kuroko-type VMS deposits in Sanjiang, respectively. In the latter case the hydrothermal solutions reacted with host rocks and mixed with cold seawater percolating downward, and circulated through volcanic rocks resulting in development of alteration zones and formation of ore minerals in Kuroko-type VMS deposits."

Jan 1, 1993

By Daniel Hodouin, Mélanie Proulx, Claude Bazin

"Collector and frother are commonly used to induce the floatability of valuable minerals in flotation plants. These reagents are usually distributed down a flotation bank or sometimes added in a single location at the top of the bank. A review of the literature and some experimental results obtained from laboratory and full scale trials indicate that the way flotation reagents are distributed down a flotation bank may influence the flotation response of coarse particles. The paper summarizes these observations which show that the distribution of reagents is a potential tuning variable for process optimization.IntroductionFlotation is in many cases the only way to concentrate sulphide and some oxide minerals. Flotation is a concentration process that uses particle surface composition to induce the selectivity by mean of hydrophobic reagents. The influence of particle size on flotation was firstly analyzed in the 50's1·2 and intensively investigated in the 70's, early 80's and 90's3-18. Most of the observed results show the existence of an optimum flotation size range9 whose recovery is not significantly influenced by a reduction of collector dosage""18. This result brought the idea of using a small portion of the total collector dosage at the top of a rougher bank and to distribute the remaining through the end of the bank to ensure sufficient collection of the coarse particles that need more hydrophobic coverage to attach to air bubbles and be lifted to the surface.This paper discusses the idea of using the distribution of reagents down a flotation bank to tune the operation of the process. Unlike the literature on the effect of reagents (collectors, frother...) dosage that is dense, little attention has been put on the distribution of reagents in a flotation bank. Although research on this subject has not been pushed very far, some operators have recognized the potential of a strategic addition of reagents down a flotation bank19. Except for a few exceptions, the proportions of the reagents to be distributed down a flotation bank are decided according to operator's experience instead of process analysis in relationships with particle size. This paper recalls observations made by different researchers that lead to the idea of using the distribution of reagents as a tuning variable for the flotation process. The paper does not present a detailed analysis or optimization scheme for tuning the distribution of reagents, but analyzes some plant and laboratory results with a view toward the development of a testing methodology and ultimately a validation or rejection of the idea."

Jan 1, 1998

Heap leaching accounts for around 20% of the global production of copper. In heap leaching coarsely crushed ore particles are contacted with an aqueous lixiviant percolating through large packed beds (heaps) of the ore particles, with the loaded solution collected at the foot of the heap for metal recovery, solution make-up and recycle. While dominant for copper oxide acid leaching, it finds application also in copper sulphide (bio) leaching. Despite their apparent simplicity, heaps are in fact complex chemical reactors in which a number of competing sub-processes determine the overall rate of metal extraction. Therefore, meaningful prediction of copper production rates from heaps, as well as operational optimisation and control, all hinge on a comprehensive and quantitative understanding of these complexities. The paper gives a comprehensive overview of the various sub-processes in copper heap leaching and their interactions, including those relating to mineral-particle-solution-gas phase interactions at the micro-scale bulk transport and distribution of solution in unsaturated dense beds at the macro-scale. On the basis of a more fundamental understanding and modelling, the potential for optimized operation offers new opportunities for the extraction of value from both low-grade and complex polymetallic ores. INTRODUCTION Heap leaching is one of the percolation leaching technologies, which includes in situ leaching, dump leaching, heap leaching and vat leaching (Bartlett, 1998). The common phenomenon in all these technologies is the transport of leach solution through a fixed bed of ore particles, which interacts with the solid as it migrates along its path. The target minerals are released into solution and with its flow are carried out of the bed. The targeted dissolved metal species are recovered from solution through various hydrometallurgical techniques, especially solvent extraction (SX), before the barren solution is returned to the leach process. Heap leaching is practised in three main types of applications: sulphuric acid leaching of copper oxide ores, cyanide leaching of gold ores as well as the oxidative acid leaching of secondary copper sulphide ores, usually with the assistance of certain microorganisms and, more recently, in the presence of chlorides (CaCl2, sea water). While gold heap leaching dates back to the 1960s, large scale copper heap leaching commenced in Chile from the 1980s, and sulphide heap leaching emerged essentially as a consequence of acid leaching from mixed oxide/sulphide ores from the 1990s (Watling, 2006).

Jan 1, 2019