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By G. Gabra, M. Nagamori, J. J. Panneton

"RESUMELa calcination contrôlée de 10 serpentine permet de produire de 10 magnésie active. Le degré de I'activation est mesure par 10 solubilité du magnésium dans les acides faibles tel que I'acide acétique, La serpentine contenue dans les résidus d'amiante du Québec peut donc être activée par calcination. Le teneur en magnésie active des rejets d'amiante calcinés est fonction de divers paramètres: température, temps de calcination, granulométrie, etc. Les rejets d'amiante de différentes provenances démontrent des comportements thermochimiques différents qui sont étudiés dans Ie présent travail au moyen d'ATD, de TG, de la diffraction des rayons X et de la lixivation par l'acide acétique. On constate qu'un maximum d'environ 80% de la magnésie totale peut être activé dans certains rejets d'amiante du Québec et Ie degré d'activation dépend de la provenance.vIntroductionOn sait depuis longtemps que la serpentine, Mg3Si2O5(OH)4, un silicate hydroxylé de magnésium, devient amorphe lorsqu'elle perd de I'eau d'hydratation entre 600 et 700 °C. La serpentine ainsi calcinée est alors dissociée en magnésie et en silice. Cette magnésie, dite active, est soluble dans un acide faible ou même dans les solutions de sels d'ammonium.Plusieurs precédés chimiques ont été proposés afin d'extraire sous forme de composés magnésiens la magnésie de la serpentine. La plupart de ces procédés comportent une activation par calcination de la serpentine préalablement à une Iixiviation par un acide ou par une solution d'un sel d'ammonium.Malgré la contamination par Ie fer et par d'autres éléments, les compositions chimiques des rejets de différentes mines d 'amiante du Québec, présentées au tableau 1, sont très proches de la serpentine typique. Puisque Ie degré d'activation de la magnésie varie grandement d'une serpentine à I'autre lors de la calcination, la première étape de la recherche sur la transformation chimique des rejets d'amiante du Québec se résume à une étude de leurs comportements thermochimiques."

Jan 1, 1982

By Mauro Torres, Gabriel Zarate

"AngloAmerican used the extractant M5640 at Mantos Blancos and Mantoverde and uses PT5050MD at Los Bronces Division, which are aldoxime and modifier blends. In the first case, the modifier is the esther TXIB, while in the second case is the tridecanol TDA. It is a well-known fact that the extractant degrades while the modifier does not and with time the oxime/modifier ratio can change significantly, depending on the operating conditions of each plant.In order to evaluate the effect of this varying modifier content in the plant organic, especially in the organic losses, a complete metallurgical test work program was conducted at laboratory and pilot plant scale by Flomin SNF, a new extractant supplier. Results showed that the addition of an aldoxime without the modifier did not affect the chemical performance of the extractant, but there was a significant decrease in the viscosity and density of the organic.Based on the results obtained in the metallurgical program, it was decided to run plant scale tests at the SX plants of Mantos Blancos and Mantoverde. In these tests, instead of adding extractant to maintain the organic inventory, CuPROMEX 3406 which is an unmodified aldoxime was added until the original oxime/modifier ratio was reached. This was obtained after 6 months of operation and since then both plants are running with the extractant CuPROMEX 3302 both supplied by Flomin SNF.In the case of Los Bronces, the pilot plant work was completed and it is programmed to run similar plant scale tests. In this case, the oxime/modifier ratio increased and TDA will be added until reaching the original ratio.The plant results obtained at Mantoverde and Mantos Blancos, together with the pilot plant test results obtained at Los Bronces, are discussed in this paper."

Jan 1, 2016

By N. J. Bunce, B. Vollick, A. F. Souza, D. Bejan, P. Maharaj, C. Shamshoom

"Copper electrodeposition from synthetic and authentic acid mine drainage (AMD) is inhibited by ferric iron (Fe(III)), through re-oxidation of copper. Polarization to –2.5 V versus standard hydrogen electrode prevents the dissolution of predeposited copper from the cathode. Above pH 1, copper deposition is independent of Fe(III) concentration to approximately 25 g/L. Fe(III) is reduced to ferrous iron (Fe(II)), with a high combined current efficiency for iron and copper reduction. In a divided cell, the conversion of Fe(III) to Fe(II) is almost quantitative, with efficient transfer of acidity from catholyte to anolyte. Prospects for developing electrochemical technology for complete remediation of AMD are discussed.RÉSUMÉ En raison de la réoxydation du cuivre, le fer ferrique (Fe(III)) nuit à l’électrodéposition du cuivre à partir de drainage minier acide (DMA) véritable et synthétique. Une polarisation à -2,5 V par rapport à l’électrode standard d’hydrogène empêche la dissolution du cuivre déjà déposé de la cathode. À un pH supérieur à 1, la déposition du cuivre est indépendante de la concentration en Fe(III) jusqu’à environ 25 g/L. Le fer ferrique est réduit en fer ferreux (Fe(II)) avec une efficacité combinée élevée pour la réduction du fer et du cuivre. Dans une cellule divisée, la conversion de Fe(III) en Fe(II) est près de la quantité théorique, avec un transfert efficient de l’acidité de catholyte à anolyte. Les perspectives de développement d’une technologie électrochimique pour la remédiation complète du DMA sont discutées."

Jan 1, 2016

By Pritii Tam Wai Yin, Tony Leong, Neetu Bansal, James Vaughan, Amiel Boullemant

Mercury is a toxic element that is associated with many industrial activities including the Bayer process. Mercury is known to enter the Bayer process with the bauxite feed material and has the potential to accumulate as the process solution is recycled. This paper explores the speciation of mercury in Bayer process streams. The process parameters considered are the oxidising-reducing potential, basicity and temperature. The results of a refinery circuit survey are presented and compared with the thermodynamic predictions. Electrochemical measurements on relevant synthetic systems are also included.

Jan 1, 2014

The Daly-Judge Tunnel, begun in 1887 to drain mine workings, is now a source of potable water for the Municipality of Park City. The water is exceptionally clean; however, when roof falls occur, suspended solids levels become too high and water from the tunnel cannot be used in the treatment plant. In 2001, the tunnel was inspected and rehabilitated to ensure that the water supply for the 2002 Winter Olympic Games would be uninterrupted. Twenty-five timber tunnel sets were removed, and 86 steel sets were installed. Timber lagging was installed or replaced over about 230 m (750 ft) of tunnel length.

Jan 1, 2006

By J. Ji

Oxygen requirement during acidic pressure oxidation of sulphide concentrates can he as high as 800 kg/(m3.h) in the first compartment of an autoclave when the total oxidation residence time is shortened to 30 minutes. To achieve such a high oxygen transfer rate while maintaining satisfactory oxygen utilization, a pilot plant campaign was launched at the Placer Dome Research Centre in 1998 to measure the oxygen mass transfer coefficient in a 5 m3 pressure vessel at a total pressure close to 500 psig with a 30 HP gassing agitator. Maximum oxygen flowrate to the pressure vessel was designed at 3 600 kg/h. In this paper, a great deal of information is provided on the gassing agitator, pressure vessel. measurement procedure and results. Without oxygen sparging, the oxygen mass transfer coefficient is 2 600 h-1 with a mixing power input of 5.0 kW/m3 in solution. With oxygen sparging, the oxygen mass transfer coefficient increases to 3 500 h-1 with a mixing power input of 4.0 kW/m3 in solution.

Jan 1, 2004

By G. De Gregoriis, G. E. da Mota, H. P. Dias, J. E. M. Blasques

In recent years, like many other smelters around the world, Albras has been increasing line current. The necessary adjustments to the work practices and process variables need to be made on all potlines, which in this case means on 960 cells at the same time. It was a challenging task to adjust the process parameters to the new amperage. In July 2003 potlines 2, 3 and 4 were running at 1 73 kA and the goal at that time was to reach 180 kA. Preliminary trials with 180 kA were performed on a group of 10 boosted cells on potline 3 in order to investigate possible operational problems. It was decided to increase the amperage in three steps accompanied by continual evaluation of the state of the process. In July 2004 the target of 180 kA was achieved with good operational results. This paper presents the overall results, the parameter adjustments made during the amperage increase in the booster group and how the increase was implemented on the other potlines.

Jan 1, 2006

By C. Edwards

ABSTRACT The versatility and specificity of rare earth elements (REEs) have led to their use in an everincreasing variety of applications in new technologies; consequently, demand for REEs has increased significantly. However, separating REEs safely and effectively is a complex and expensive process. This paper presents a review of REE mineral processing technologies, providing an update on current capabilities in REE mineral processing and REE extraction from the major commercially valuable REEbearing minerals such as bastnaesite, monazite, and xenotime, as well as the ion-absorption type of REE deposits.

Jan 1, 2013

By W. S. Hegler

Introduction The San Antonio mine is on the north shore of Rice lake, thirty miles east of lake Winnipeg and 120 miles northeast of the city of Winnipeg. Since production was started in 1932, about 2,000,000 tons of ore have been mined. Currently in progress is an extensive development programme which involves 11,500 feet of crosscuts, 3,800 feet of raises, 1,753 feet of shaft sinking, and 160,000 cubic feet of slashing, all in footwall waste rock. It is the purpose of this .paper to indicate the necessity for this work and to describe the methods employed to carry it out. Where unknown factors are as prevalent as they are in the average Canadian gold mine, a definite answer cannot be given as to what is the most economical and efficient system for mining dipping deposits. At San Antonio, the host rock for the orebodies, locally termed basalt, occurs as a sill some hundreds of feet in width and dipping at approximately 45 degrees from surface to an unknown depth. It may be assumed that it continues well below the lowest depth established to date by diamond drilling. At present, a vertical winze (No. 3 winze) hoists the ore from the 2,360-foot level to the 1,600-foot level, on which it is hauled about 900 feet to No. 3 shaft, also vertical, and there hoisted to the mill bins on surface. To maintain ore reserves at a five-year figure, at full production, additional deep exploration was required immediately. It was decided to sink No. 4 winze, a vertical opening of the same size as the other operating shafts, to provide ten additional levels (17 to 26) at 150-foot intervals, with additional depth to allow for an underground crusher installation and usual loading-pocket facilities for three-ton skip-hoisting. The winze was bottomed at 1,753 feet below the collar, at a depth of 4,117 feet below surface, and is the deepest man-sized opening in Manitoba.

Jan 1, 1949

By R. Blosser

Phelps Dodge introduced a process improvement model known as Quest for Zero in 2001. In 2004, the focus of Quest for Zero shifted from corporate-led initiatives to projects led by site personnel. The Hydrometallurgical Division at Morenci focused on two (2) significant process improvements: reducing electrolyte bleed and organic losses. In September 2004, Morenci kicked off a bleed reduction project to reduce electrolyte bleed by 45%. By the end 2005, electrolyte bleed was reduced by 55% through understanding and reduction of variance. Significant process changes made during this time period include changes in acid-addition strategy, reduction of uncontrolled bleed, and optimization of bleed placement. In April 2005, the organic reduction project began with the goal of reducing losses by 15%. Operational improvements during this project included improved organic recovery systems, mixer speed optimization, and picket fence installation. By the end 2005, this project reduced organic losses by 35% and current losses are 48% lower than at the start of the project. These projects and smaller Questfor Zero projects have resulted in sustainable annualized savings of $13 million in Morenci hydrometallurgical operations. Through the use of Six Sigma tools, process variances can be controlled and reduced.

Jan 1, 2007

By Steven Hart

In September 2009, Newmont began the commissioning of a 3 stage crushing and milling circuit at its Boddington Gold Operation in Western Australia. The Boddington ore has been recognized as one of the hardest and most abrasive ores treated to date and this played a pivotal role in the selection of crushing and HPGR technology for this operation as opposed to the more conventional SABC circuit. This paper will examine the scale up from laboratory to pilot plant and full scale operation of the crushing circuit. There have been many challenges to overcome in the first 18 months of operation; however, the HPGRs have in general met design expectations and performance is expected to improve further with ongoing design changes

Jan 1, 2012

By J. E. Hawley

Abstract Spectrographic methods for the analysis of minor constituents such as platinum-group metals, gold, silver, and silica in palladium and platinum refinery sponge are described. These involve the conversion of the sponge metals to the black form by reduction with high purity aluminum filings in either acid or alkaline media. Standards of known composition are prepared in the same manner from either refinery or spectrographically standardized sponge metals. Samples of the black metals are mounted in pellet form and a multisource spark discharge used for excitation. The range of concentrations of impurities studied are those common to refinery samples. Excellent reproducibility of intensity ratios using major constituents as internal standards, is demonstrated and two methods have noted by which the accuracy of the procedures are shown to be of the order of 3 per cent, on the basis of synthetic preparations. Introduction During the past two years an investigation has been carried out in the Spectrographic Laboratory, Queen's University, to determine the suitability of modern spectrographic techniques for the precise analysis of the platinum-group metals palladium, platinum, rhodium, iridium, and ruthenium in the form of refinery sponge analyzing about 99 per cent of the major constituent.

Jan 1, 1951

By The Mill Staff

Introduction The Hollinger cyanidation mill commenced operating in June, 1912, and 49,779,978 tons of solids had been discarded by the end of 1950. The present tailing disposal plant has been in operation since 1924 and holds 39,762,250 tons. The tailing disposal plant is on the property of Kayorum Gold Mines, Limited, about a mile and a quarter south of the mill. The area available for tailing covers 640 acres and is roughly l ¼ miles long by ¾ mile wide. Tailing was first deposited on the McKay veteran claims, l mile long by 1/2 mile wide, which were later included in the Kayorum property. This site sloped from northeast to southwest, the slope amounting to 106 feet between the northeast corner and the low land to the southwest. The low land was wet and a considerable area was covered with muskeg up to 17 feet deep. It was drained by a brook running across the property in a southwesterly direction. Later, when more ground was needed, the low-lying mining claims adjacent to the south and west boundaries were acquired to allow for the extension of the plant. This entailed some extra difficulties and expense that would not have been encountered had the whole of the present site been acquired with the original purchase. At the present time, the top of the tailing deposit stands at two elevations, but eventually the tailing deposited on the extension to the original plant will be raised to the same height as the older deposit. Previous to 1924, all of the tailing was deposited on Low land and in lakes near the mil1. Between 1924 and 1926 an increasing tonnage was sent to the McKay claims until, in 1926, all of it was being sent there. In 1936, the area of the deposit was extended and shortly afterwards all of the tailing was deposited on the new site.

Jan 1, 1951

By A. Anderko, R. D. Springer, R. D. Young

A thermodynamic model has been developed for calculating phase equilibria and other properties of multicomponent electrolyte systems. The model has been designed to reproduce the properties of both aqueous and mixed-solvent electrolyte systems ranging from infinite dilution to solid saturation or pure solute limit. The model incorporates formulations for the excess Gibbs energy and standard-state properties coupled with an algorithm for detailed speciation calculations. The excess Gibbs energy model consists of a long-range interaction contribution represented by the Pitzer-Debye-Hiickel expression, a second virial coefficient-type term for specific ionic interactions and a short-range interaction term expressed by the UNIQUAC equation. The accuracy of the model has been demonstrated for common acids and bases and for multicomponent systems containing aluminum species in various environments.

Jan 1, 2005

By J. E. Nesset, F. Rosenblum, J. A. Finch, R. Payant

ABSTRACT Under certain conditions of moisture and oxygen, sulphides can spontaneously self-heat. We tested a hypothesis that galvanic interaction between some sulphides can promote self-heating. Galvanic interaction is controlled by the rest potential difference between sulphides and particle size. A series of binary sulphide mixtures was tested. Individual sulphides and mixtures with low rest potential difference did not self-heat, but mixtures of high rest potential difference did. When testing pyrite-sphalerite mixtures, self-heating increased inversely with particle size, with the fineness of the pyrite governing self-heating, indicating the rate-limiting step is the reduction reaction. The results support the hypothesis that galvanic interaction contributes to sulphide self-heating. A possible mechanism based on hydrogen sulphide as an intermediate product is proposed. The understanding gained will be of interest to those involved in storage, shipping, and disposal of sulphides.

Jan 1, 2012

By R. K. Collings

The annual production of mineral and mineral-based wastes in Canada is conservatively estimated to be 400 million tons. Over 90 per cent of this amount is waste rock or mill tailings from the mining and mineral processing industry. The remainder are slags and ashes from the metallurgical industry, and sludges from the chemical industry. Because of impurity content, remote location, or both, most wastes have limited usefulness and, indeed, are a liability in that they must be transported to waste disposal sites and contained therein. Certain mineral wastes do, however, have potential value. Some contain small percentages of metals that may be economically recoverable with price increases; others consist of mineral material that may be used for such purposes as road base, construction aggregate, railroad ballast, and in the manufacture of bricks or blocks. Increased interest has been shown in mineral wastes in recent years. This paper reviews the types of mineral wastes available in Canada, briefly describes a number that have potential value, and notes possible uses and related research by CANMET and other organizations.

Jan 1, 1978

By Qin-yong Ma, Jing-shuang Zhang, Bo Han, Pu Yuan

In order to improve the utilization rate of blasthole and blasting effect in vertical shafts, 5 kinds of millisecond blasting model tests with three circle vertical blastholes have been conducted in various delay times with concrete as model material. The blasting effects of millisecond blasting model tests are investigated from cutting volume, cutting depth, and blasting fragmentation, then a reasonable delay time is obtained. In the middle of concrete model, there are 4 blastholes uniformly distributed in the first circle, 6 blastholes uniformly distributed in the second circle, and 10 blastholes uniformly distributed in the third circle. Each blasthole is charged with one millisecond delay electric detonator, and the rest is blocked with stemming. The depth of blastholes in the first and second circle is 180mm, which is 20mm deeper than those in the third circle. After initiation, the blasting cavity is cleaned up, then the cutting volume and cutting depth are measured. Meanwhile, the blasting fragments are gathered and have analysed by function Rosin-Rammler. The test results show that when the inner two circle blastholes initiated simultaneously and the third circle blastholes initiated with millisecond delay, the reasonable delay time is 50 ms. At this time, the cutting volume and cutting depth of blasting model are 3920 cm3 and 68 mm respectively. The fragmentation indexes of blasting fragments, X50 and X80, are 53.2 mm and 74.6 mm respectively. While when all three circle blastholes initiated with millisecond delay, the reasonable delay time is also 50 ms. At this time, the cutting volume and cutting depth of blasting model are 10705 cm3 and 118 mm respectively. The fragmentation indexes of blasting fragments, X50 and X80, are 51.0 mm and 81.7 mm respectively. The researches indicate that for millisecond blasting with multiturn blastholes, the reasonable blasting scheme is multiturn blastholes initiated in millisecond delay and the reasonable delay time is 50 ms.

Jan 1, 2015

By R. S. Douglas

Introduction The operations of The Granby Consolidated Mining, Smelting and Power Company, L1mited, have always been concerned with the mining of low-grade copper ores. In the fifty years of this Company's existence, nearly 60,000,000 tons of such material has been successfully mined and treated. During these years, many changes in mining practice took place. However, at Copper Mountain, due to the large tonnages of very low grade material, it was realized that something more than changing mining methods would have to take place if these low-grade orebodies were to be mined at a profit and the Company and. its employees enjoy many years of successful operation and employment. Production costs were mounting commensurately with rising taxes and the increasing cost of supplies, labour, freight, and smelting, while the value of the product on the market was not increasing in price to compensate for the higher costs of production. A survey of the mineable ore in 1944 indicated that, of the 12,000,000 tons in the reserves, only 3,500,000 tons was now profitable ore. This reduction in the commercial ore reserves was a serious matter, as it meant that the mine had only two years of life at the profitable shipp1ng rate of 4,800 tons per day. Further, the location in the mine of the orebodies making up the 3,500,000 tons was not suitable for such high production.

Jan 1, 1950

By Luke Evans

What are Best Practices? What are Bad Practices? Who Uses Best Practices? Why Bother? Acceptability Guidelines Published Best Practices Published Bad Practices Additional Best Practices Some Best Practices for Low Grade Deposits Interpretation Guidelines Big Bear Gold Deposit Example Conclusions Questions

May 1, 2013

By C L. Briend, O&apos

This paper presents the main features of coal mining in Japan, as determined by the author while attending the International Coal Conference in October of 1963. The general economy of the country is considered, and the requirements for coking coal are outlined. Statistics and comments on production and consumption are then given, with accompanying cost figures, and the characteristics of Japanese coal mining are compared with those of Canada. Mining methods and underground transportation are covered, with emphasis on undersea mining. Finally, a description is given of a visit to the Miike colliery, the largest in Japan. I T was my privilege to attend the International Coal Conference held in Tokyo in October, 1963, to meet many prominent members of the Japan Coal Association and to visit one of the major coal fields. This is probably the reason why I was asked to present this review of the coal situation in Japan. In doing so, it must be pointed out that I am not an expert in coal mining, that my

Jan 1, 1965