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By G. A. Vissac

METHODS of coal preparation have been greatly developed and improved during the past few years. Under the pressing competition from substitutes that have been displacing coal as a source of heat or power, the coal industry has been able, in many cases, not only to hold its own but to regain much of the lost ground, by offering better products of a higher value. The purpose of this paper is to present, in a brief and simple form, the principles of the latest methods and machinery used in coal preparation, in order to enable the coal operator t0 select methods and equipment capable of solving his own problems in the most efficient and economical way. The paper will deal, in the main, .with coal cleaning. It must be emphasized, however, that a complete coal preparation study covers many other operations, such as handling, screening, de-dusting, dust-proofing, and ash-conditioning. Handling Coal From the producing face to the railway car, coal must be handled with great care, t0 prevent degradation, because the coarser sizes are not only easier to clean than fines but, for our markets, they are often the more valuable. Generally speaking, such care is less important or is unnecessary in the Old Country, where conditions are different. The best way to minimize breakage is to avoid gravity methods in coal transportation. All chutes in a preparation plant should be shaken mechanically, and cleaning processes requiring large surge bins must be avoided. De-Dusting De-dusting of coal is accomplished by the use of vibrating screens, fans, or classifiers, or by a combination of these. Fines below 1/32-in. (or coarser than this, with some coals) can very seldom be cleaned efficiently. Common practice is to discard them immediately, if the moisture content of the coal allows this to be done. Otherwise, they are liable to give unnecessary trouble, either in the form of dust or sludge, which are difficult to collect or to settle. Flotation is the only process which will effectively clean fines.

Jan 1, 1937

By W. T. Irvine

"The H.B. Mine, a zinc-lead operation of the Consolidated Mining and Smelting Company of Canada Limited (Cominco), is in south central British Columbia, on Sheep Creek, a west-flowing branch of the Salmon River. It is 3 miles by road from B.C. highway No. 6, 30 miles south of Nelson, B.C., with an excellent highway connection to the smelter city of Trail, 35 miles from the mine. The nearest permanent settlement is the village of Salmo, 7 miles distant by road. The mine is situated along a belt containing other important zinc-lead deposits, including those of the Canadian Exploration Company, 3 miles to the south, and the Reeves MacDonald mine, 9 miles distant again to the southwest.Deeply oxidized outcrops were located in 1907 on the north side of Sheep Creek valley by two prospectors named Horton and Benson. The property was given the name H.B., after the two discoverers, and there commenced a long and intermittent series of mining operations designed mainly to develop and extract oxidized lead-zinc ore. In 1925 a determined attempt was made by a British syndicate to develop a large tonnage of sulphide ore, but this ended in failure after some 3,000 feet of lateral development had been done. In 1927 Cominco acquired the mine, but it lay idle until 1946 when a geological study was followed by a program of surface drilling which encountered extensive sulphide mineralization. An underground development program revealed more details of the deposit, and sufficient ore was found to warrant construction of a 1,000-ton mill. Mine preparation and mill construction were completed by early 1953, but production was deferred due to low metal prices. The mine was finally brought into production early in 1955."

Jan 1, 1949

By A. Trusiak

"Increasing energy costs in mineral processing is making it necessary for tighter controls over various parts of the concentrating process. Continuous monitoring and control of critical process streams can be easily performed by today's many computer systems. Indeed, the list of successful applications of computers to grinding and flotation control continues to grow. The day is rapidly approaching when the use of computers within concentrators for either process control or simply metallurgical accounting will no longer be a luxury, but a necessity. The benefits from such applications far out weight the initial investment.This paper describes the current application of a process control system at Brunswick Mining in north-eastern New Brunswick. Emphasis is placed on the control of a CuPb flotation circuit, and the benefits derived-from this control.No. 12 Cu-Pb Flotation: The complex nature of the Brunswick flotation process has been described else-where. 1,2,3 Four final concentrates are eventually produced, these being a Zn concentrate assaying 53% Zn, a 30% Pb concentrate, a 20% Cu concentrate, and a bulk Pb Zn concentrate assaying 35% Zn and 17% Pb. The Cu concentate, Pb concentrate, and bulk Pb Zn concentrate are all derived from a primary CuPb concentrate assaying 26% Pb, 11% Zn and 1.8% Cu.The No. 12 CuPb circuit produces approximately two thirds of the total primary CuPb concentrate. No. 12 ore, assaying 3.5% Pb, 9% Zn, 0.25% Cu and 100 g/tne silver, is ground in two parallel grinding circuits. Each line uses a 3.2 x 4.2 meter rod mill in closed circuit with two parallel 3.2 x 3.9 meter ball mills. Ball mill and rod mill discharges are classified in two stages of cycloning. The secondary cyclone overflow, or flotation feed, is controlled at 65% passing 37 microns."

Jan 1, 1981

By K. Kryst, K. Elkayar

The increasing need to process lower grade contaminated copper concentrates and simultaneously produce higher purity copper cathodes necessitates a more diligent control in the electrorefining process. Ion exchange processes can be a powerful tool in controlling electrolyte impurities to ensure trouble free operation within the tankhouse and to help maintain cathode purity. In addition, some impurities can be recovered as a value added product. The use of a short bed ion exchange process provides a unique opportunity for impurity management in copper refineries. Several metals are discussed in this paper, including Sb, Bi, As and Ni and their methods of control and recovery within the copper electrolyte circuit. INTRODUCTION To produce a high purity copper product by electro-refining a copper electrolyte, it is important to control the concentration of impurities such as arsenic (As), bismuth (Bi), antimony (Sb), and nickel (Ni). These impurities are introduced into the electrolyte from the copper anode and can affect the copper product and the production operation. The presence of As in the copper product results in a lower electric conductivity and reduced workability of the copper. High levels of Bi in the electrolyte will result in a copper cathode product containing Bi. High Sb concentration will cause undesired floating slimes. A variety of methods have been developed to control the level of these impurities in the copper electrolyte including precipitation, liquid extraction, and ion exchange. These impurities can be recovered and used. This paper discusses novel methods that utilize short bed ion exchange technology to control these impurities. Also, the use of ion exchange to recover acid from the liberator cells bleed for reuse in the tankhouse is reviewed. SHORT BED ION EXCHANGE Short bed ion exchange technology has been employed extensively in the metal finishing industries for over forty years. Unlike conventional ion exchange, short beds are fixed beds without freeboard space. The application of short bed technology results in significant benefits when compared to conventional ion exchange. The resin used in this technology has a significantly smaller diameter than conventional ion exchange resin. As a result of the smaller resin size, exchange kinetics improves dramatically. This facilitates operation at higher flow rates and the use of smaller beds with a small mass transfer zone length. Smaller beds are especially important for applications that use chelating resins due to the high cost of the resin.

Jan 1, 2019

By G. Granata, C. Tokoro, M. Matsuoka, K. Horiuchi

"Given the high demand of copper and the lack of new high-grade deposits, low grade copper ores containing copper oxides and refractory sulfides need to be processed as well. However, in order to recover copper from refractory copper ores, the leaching performances should be improved. In this study, we investigated the effect of the mechanical activation in the leaching of copper from covellite. The influence of a mechanical treatment by planetary ball milling on leaching rate in sulfuric leaching media was evaluated by using X-ray diffraction (XRD), mineral liberation analysis (MLA) and X-ray absorption fine structure (XAFS). Results showed a significant increase in copper leaching from around 50%up to 100% along with an improvement of the leaching kinetics when using planetary ball milling. XRD spectra showed the disappearance of the covellite peak from (101) face, thus suggesting the amorphization of covellite associated with the use of planetary ball milling. XAFS results showed that the coordination number of Cu changed from 6 to 3, thereby confirming the effectiveness of the mechanochemical effect.INTRODUCTIONDuring the last decades, while the copper grade in copper ores has been decreasing, the concentration of impurities such as arsenic tends to increase, thus requiring for the processing of low grade copper ores. For a suitable processing of copper minerals, hydrometallurgical processes are preferable due to the lower costs, the lower environmental impact and the easier implementation (Watling, 2006). However, for an economical recovery of copper, the leaching performances must be improved.Among the recent advancements in leaching technologies, the mechanochemical activation allows for an improvement of the leaching performances by means of high-intensity grinding(Mohammadabad et al., 2016). In the first place, the improvement is the result of the dramatic decrease of particle size, which results in an increase of the mineral surface area and the consequent enhancement of leaching kinetics. Furthermore, the intense mechanical stress due to the high-intensity grinding could induce the rupture of the crystal structure producing a decrease of crystallinity. Such changes of the mineralogical structure, also known as mechanochemical effect, can lead to more leachable mineral, thereby resulting in a greater dissolution of copper. Last but not least, whereas the mineral surface might undergo passivation during leaching, the high-intensity grinding can continuously refresh the surface of the mineral, thus exposing portions of unleached bulk mineral."

Jan 1, 2017

By T. M. Waterland

A GREAT deal of work has been done at Britannia during the past three years in improving and standardizing rounds for development headings, and the results obtained have been very gratifying. The round which has recently been adopted as a standard for all drifts and cross-cuts has, in some instances, doubled the average length of break formerly obtained. It has also increased the footage of advance per labour shift and reduced explosives' consumption. At the present time, with a shortage of skilled labour, more control is necessar7 in directing the drilling of rounds. The use of a standard round has been o great assistance to the foremen and shiftbosses. NATURE OF GROUND Most of the ground encountered consists of a moderately silicified chlorite schist, known locally as G.M.S (green-mottled schist). It cannot be termed hard ground, because a three and one-half inch Leyner will drill at an average rate of twelve inches per minute. Due to the abrasive nature of the ground, the gauge of the steel wears rapidly and in most cases one set of steel is needed for each hole. The ground is not 'free breaking and holes must be spaced correctly to obtain a maximum break. DEVELOPMENT OF STANDARD ROUND FOR THREE-SHIFT OPERATION Late in 1940, a development programme was started involving 7,000 feet of drifting, 6 ft. By 7 1/2 ft. In cross-section, in two headings. Plans were drawn up to advance the headings on three shifts. Equipment used in the headings was as follows: D.A.35 Leyner?machines, Eimco Model 12B mucking machines, nine one-ton end-dump cars, one Mancha Little Trammer on the longest heading, 20-pound rail, 4-in. 15 compressed-air line, 15-in. Fan pipe connected to 5 h. P. Coppus fan. Diamond-drill cross-cuts, 16 feet in length, were cut every 250 feet. Switches were installed at these cross-cuts and they served as passing points for the cars. One siding seventy feet in length was cut in each heading at approximately the half-way point. The cycle of operations planned was to muck out, drill, and blast each shift. Previous performance in this type of ground indicated that a daily advance of 12 feet to 14 feet might be made. However, if a better type of round could be developed, an increased daily advance would be possible. A study of the time available for each operation in a net working shift of six and one-half hours indicated that the maximum depth of round that could be mucked, drilled, and blasted each shift would be six and one-half feet. When this maximum depth was decided upon, the next problem was to develop a satisfactory round. Due to ventilation difficulties and to the short net working shift, the whole round had to be blasted at one time. In other words, no blasting of the cut-holes could be done.

Jan 1, 1942

By Dennis Martin

What is Borehole Hoisting Borehole Hoisting is the use of a hoisting plant to hoist rock, workers and materials or both, vertically through a borehole, between a mine level and surface or between two levels of an underground mine. Screen and bolt ground support on walls of 5.5 m diameter borehole hoisting shaft

May 1, 2013

By W. G. McBride

THE present demand for men with engineering training exceeds anything in the history of the profession. Recent estimates indicate that mechanization of war has reached such a stage that at least twelve horsepower is required now for every one horsepower needed to run the machines of the first Great War. The scale of operations exceeds anything in the past, while more territory and people are involved than ever before. Many of our allies have inadequate or no facilities for producing the thousand and one articles absolutely necessary to the conduct of modern warfare. Lack of preparedness has made it necessary to produce at an unparalleled rate to gain parity with the enemy, while our losses have been so heavy that replacement, in itself, presents a task which might well stagger less courageous peoples. This means that not only is there an unprecedented demand for engineers to produce the machines, munitions, and supplies of war, but there is a correspondingly large need for men with engineering skill in the uniformed forces to operate and maintain the vast numbers of ships, 'planes, tanks, trucks, radios, and a myriad of other devices to carry on the varied activities of war. This abnormal demand came upon us after almost ten years of stagnation in the engineering industries. During this time many trained engineers drifted into other callings, while the lack of jobs for graduates caused a falling off in the number of men training for this profession. Consequently, the supply of well trained engineers was below normal when the madman of Germany precipitated us into the present turmoil. Added to all this, we in Canada are still suffering from governmental inefficiency in the utilization of the available supply of engineers. In the armed forces, engineers have been assigned to office and other work where their engineering training is not only wasted but for which men with more suitable training could be found. Mining engineers are debarred from service as aeronautical engineers, and students in mining engineering are denied the bursaries provided for aiding students in engineering. Mining and metallurgical engineers are both barred from the Ordnance department of the army. This arises from a misunderstanding in relation to engineering training. In Canadian engineering schools, all engineers are given a sound training in the branches of human knowledge which are fundamental to engineering: mathematics, physics, and chemistry. The basic difference between the several categories of engineers lies in the fact that training in the application of this fundamental knowledge to the solution of engineering problems is given in different subjects.

Jan 1, 1942

"The QEMSCAN system is the third generation of automated mineral analysis systems that began with the QEM*SEM at CSIRO 30 years ago (Butcher et al. 2000). It is now a successful commercial instrument with 39 instruments in operation around the world. Back Scattered Electrons (BSE) and Energy Dispersive (EDS) X-Ray spectra are used to create digital mineral images with mineral identification occurring online. Low-count EDS spectra are used preferentially to BSE brightness, allowing minerals to be accurately identified by chemical composition. Individual minerals or groups of similar compositions are identified by comparison to a comprehensive mineral database incorporated into the QEMSCAN software. Optical mineralogy and Scanning Electron Microscopy is a metallurgists’ tool for determining the ore process mineralogy and delivering a better flowsheet. These techniques have limitations, provide limited information and sometimes fail to provide representative information. In contrast, QEMSCAN provides information on the nature and occurrence of all mineral species present plus the liberation characteristics, looking at thousands of particles in the sample presented. Some of the data presented includes:• Particle and grain sizes• Dominant mineral species• Particle shape• Calculated metal content• Density• Deportment of elements of interest between the mineral phases• Quantitative mineralogy• Mineral associations, locking and liberation• Textural relationships (particle images)It can also be used on concentrates, polished sections or drill core. This is a significant improvement in ore characterisation providing lithology, mineralogy, liberation, ore grade and porosity from a sample. Theoretical grade-recovery curves can be created to be mineralogically limited from QEMSCAN analysis which, when compared with metallurgical results, are valuable in evaluating whether liberation or chemical selectivity is driving flotation performance. A number of case study examples are given where the QEMSCAN had a significant bearing on future project development at an early stage of exploration and mine development."

Jan 1, 2012

By D. J. MacKinnon

"The Canadian Copper Company, formed in 1885, began mining the nickel-copper ores of Sudbury in 1887 and producing nickel-copper matte from its smelter in 1888(1). Although it was quickly appreciated that the separation of nickel and copper (i.e. refining) was critical to the production of a marketable nickel product, refined nickel was not produced in Canada until about 1918. From the beginning of the industry, there was considerable political agitation to refine nickel in the Province of Ontario(2). The companies, however, claimed that the existing processes, viz. the sodium sulphide ""tops and bottoms"" process operated by the Orford Copper Company of New Jersey, U.S.A.(1,3,4) and the carbonyl process(5,6) developed by the Mond Company at Clydach, Wales, were too costly to operate in Canada. Nickel was first refined in Ontario about 1918 when the International Nickel Company opened a nickel refinery at Port Colborne, Ontario(6). Initially, however, the refinery at Port Colborne employed the Orford sodium sulphide process, and the first electrorefined nickel in Canada was produced in 1920 by the British America Nickel Corporation whose mine and smelter were located in Sudbury, Ontario, but whose electrorefinery was located at Deschenes, Quebec, near Ottawa, Ontario. The purpose of this article is to outline the early history of nickel electrorefining and to trace the significant developments which led to the first successful commercial electrolytic nickel refinery in Canada. The development of an electrolytic process for refining nickel is due largely to the efforts of four individuals, viz. Carl Hoepfner, Hans A. Frasch, D.H. Browne and ultimately N.V. Hybinette, and the contribution of each is discussed separately below."

Jan 1, 1999

By Randy Settgast, Daniel Moos, Hyunil Jo, Ghazal Izadi, Christine Baba

"Analyses of production logs of multiply fractured horizontal wells in unconventional plays show that only a small percent (~30%) of perforation clusters contribute to production. We present in this paper one possible approach to mitigating this problem, based on use of a fully 3-D numerical model of fracture propagation through a heterogeneous medium, by studying the influence both of spacing and of perforation friction on the initiation and propagation of simultaneous hydraulic fractures from a horizontal wellbore. The results indicate that while limited-entry perforating can help to equalize the lengths of fractures simultaneously growing from multiple clusters, it cannot equalize their widths. Both fracture width and fracture length can be equilibrated by combining non-even cluster spacing with limited-entry perforating. This has obvious benefits as it allows equalized placement of proppant across the entire series of perforation clusters, resulting in similar fracture widths and conductivities which enhances the likelihood that all clusters will contribute equally to production.INTRODUCTIONThere is substantial evidence from production logs and other information that during the hydraulic stimulation of horizontal wells, not all perforation clusters within a given stage will produce. One possible cause is the interaction between the hydraulic fractures emanating from each perforation cluster. Typically fractures created at the outer perforation clusters create “stress shadows”, i.e. increases in compressive stress in the minimum principal stress direction, which limit the growth of fractures growing from interior clusters; this is confirmed by modelling [1, 2, 3]. An example of this can be seen in Figure 1 which shows the final configuration of simultaneously propagating fractures from multiple, evenly spaced clusters when there is no perforation friction. The inner fractures are substantially shorter, reproducing the results by Bunger and Peirce [3]. The inner fractures also have substantially smaller widths.One approach to equalize the growth of simultaneously propagating fractures from a horizontal wellbore is by using limited-entry perforating as a method to equilibrate the flow rates into each of the different perforation clusters. Perforation friction is caused by the finite diameter of the entry points into the formation, and leads to a pressure drop across each perforation which is proportional to the flow rate squared. Fractures that take more fluid have a higher fluid velocity through the perforations than fractures receiving less fluid. This causes a larger pressure drop between the wellbore and more rapidly growing fractures, compared to the same pressure drop between the wellbore and the fractures that grow more slowly. Because pressure within the wellbore is approximately the same at each entry point, the larger pressure drop leads to less pressure being delivered to rapidly growing fractures, thereby decreasing their ability to propagate and reducing the amount of fluid they are able to receive. Because the mass rate of injection into the wellbore is constant, this forces more fluid into the fractures that were propagating more slowly, increasing their growth rate. It also increases the pressure required to drive fluid injection. The balance of these effects makes the system largely self-regulating and delivers more balanced fracture growth. While this effect is well known, it is often neglected in current numerical models."

Jan 1, 2015

By R. C. Rowe

DR. W. F. GRAY (Contributed discussion, presented at meeting by Mr. S. C. Mifilen): Mr. Rowe's suggestion that the Institute should organize itself to interpret the needs of the mining industry to federal and provincial mines departments should commend itself to our membership as sensible and practical, and as one which these public administrative bodies may welcome. One has to start with the premise that those who head the departments of mines in question have the welfare of the industry at heart. Also, it must be postulated that the technical viewpoint of our membership is necessarily free of political considerations. With these reservations, one ventures to state that both Mr. Rowe in his communication to the present session of the Institute, and Professor M. Y. Williams in his earlier communication to the Western Meeting, on The Geological Survey and Mining Development (1), voice the regret general among our membership at the abandonment of the honoured name, and to some extent of the status, of the Geological Survey of Canada. The writer, in 1932, devoted his Presidential Address to the Institute to an appreciation of the value to Canadian mining of the Geological Survey of Canada (2), affirming the desire of the Institute "that the work of the Survey shall be maintained at the highest pitch of efficiency the resources of our country will permit, because, to scrimp expenses on the Survey, is 'to save at the spigot and lose at the bunghole'." He believes that is still the opinion of the Institute, and, if contraction of the usefulness of the Survey is feared by our members at this time, it is incumbent upon the Institute, in fairness to those who direct our government, to say so in precise and explicit terms. Mr. Rowe puts the situation pithily when he writes: "We have reached a stage in our development where a stock-taking and co-ordination of existing knowledge and data is essential. This is a part of that basic function of Dominion mineral technical services which is national and not provincial in its scope".

Jan 1, 1939

By Jos. Kalbhenn

PART of the plan of re-allocation evolved by the engineers of the Dominion Steel and Coal Corporation in 1932 to retard the rapid advance of the submarine workings seawards was the opening of Dominion No. 20 Colliery, in the Harbour seam, by a pair of cross-measure tunnels from the workings of an older mine working the Phelan seam, which lies 430 feet below the Harbour seam. The tunnels, 8 ft. by 13 ft. in size, were started away from the old workings of Dominion No. 2 at a point 1.5 miles beyond the shore line and were driven on a rising grade of 1 per cent for a distance of 5,335 feet, tapping the Harbour seam at a point 2 ¼ miles off-shore. For transportation and ventilation, shafts and roadways already existing in the Phelan seam were made use of during the stages of development and up to the present time. One of the two tunnels served as a main haulage road and an intake airway, and the other as a return airway and travelling road. However, as the workings of this new colliery expanded, the need for additional airways arose as was anticipated when the plans for the development were formulated. To meet this need, work was started on October 1st, 1942, to establish a new opening between Nos. 2 and 20 Collieries in the form of a vertical shaft to provide a passageway for the ventilating air current of some 250,000 cubic feet per minute on its return to the upcast shafts in No. 2 Colliery, thus making both tunnels available for intake airways. In the interest of safety, speed, and economy, the shaft was raised from the Phelan to the Harbour seam. It is 20 ft. by 12 ft. in size and, with the exception of three timbered compartments across the narrow end, is untimbered. One of the three timbered compartments housed the ventilating tubing and pipe lines, the centre one served as a cage-way for hoisting men and material during the operation of raising, and the third one contained a ladder way. The rest of the shaft was filled with debris during the period of shaft raising.

Jan 1, 1944

"HPGRs operate in a wide array of applications in mining and minerals. These range from relatively coarse crushing of copper and gold ores at one end of the spectrum to the processing of fine iron ore pellet feed and fine industrial mineral products at the other. Traditionally these circuits involve the treatment of relatively moist materials in pre-grinding circuits; using single stage HPGRs in open circuit or as a closed circuit operation with screening or partial product recycle. More recently, HPGR is increasingly deployed in dry grinding in combination with air classification. The high throughput demonstrated by larger HPGRs has stretched the application towards coarser size reduction duties for a number of new projects. At the same time, the basic principle of inter-particle crushing for generating a product with a high fines proportion gives ground to investigate applying the technology further towards finished product duties. The advantages that HPGR brings are clear. Multi Stage grinding enables energy efficient size reduction across more of the comminution circuit. Primary HPGR serves for a coarse grinding duty, whereas a secondary HPGR is geared towards generating a fine product for downstream processing or even a finished product. This publication summarizes some of the features of staged HPGR circuits and discusses some examples of operating and projected applications.INTRODUCTION In most HPGR applications, the grinding units are used as a single process stage, either in open circuit, such as for pebble crushing [Rose 2002, Westermeyer 2000, Van der Meer 2015] or in closed circuit with classification [Van der Meer 2009, 2012] or product recycling [Van der Meer 2010]. Where a higher plant capacity is required, the HPGRs are commonly installed in multiple lines in a parallel arrangement [Parker 2005, Koski 201, Vanderbeek 2006]. HPGRs can take a wide range of feed size distributions, and grind this down to a product with a high fines content. In general view, there is a limit on the size reduction range that even an HPGR can provide. Grinding down from a very coarse ore (e.g. 70-100 mm) to a very fine product (e.g. 100 microns range) is stretching the ability of a single stage HPGR circuit. Fine grinding generally requires a classification stage that returns the classifier oversize to the head of the HPGR circuit. From one point of view, it can be demonstrated that when recycling a large proportion of fines and middlings back to the HPGR feed results in an effective feed with an overrepresentation of middlings and fines, which will eventually reduce the grinding efficiency and consequently result in an increased recirculation mass flow [Van der Meer 2014]. From another point of view the large recycling mass puts a strain or even a limit on the capacity of present-day conveyor capacity."

Jan 1, 2018

By S. J. Prout, J. M. Lang, D. Lascelles, K. Sarinas

"There is a growing demand for ore characterization methods to support metallurgical development programs from a true geometallurgical perspective. Today, there are multiple automated mineral analysis systems on the market, many of which are based on conventional mineral liberation data. However, liberation data can be limited when assessing the floatability of a mineral to predict its recovery as it doesn’t consider the surface area available for collector absorption and bubble attachment; whereas mineral exposure measurements can. Using the First-Order Rate equation to predict the recovery versus time of variously exposed particles, a model has been developed to predict the recovery of copper sulphides at a particular grind size. This model provides the opportunity for the prediction of copper recovery and grade from multiple, small-scale samples. These predicted values can be imported into resource modeling software, and guide the domaining of geometallurgical units within the ore body. Data has been analyzed from various global copper porphyry systems and compared with actual flotation testwork.IntroductionAs part of the development of a mining project, resource estimation primarily focuses on the definition of the quantity and grade of the ore in the deposit; as the project evolves from preliminary economic assessment through to feasibility, the geological model is usually further populated with hardness and recovery data.The CIM Definition Standards (2011) identifies measured mineral resource as the part of a mineral resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of “modifying factors” to support detailed mine planning and final evaluation of the economic viability of the deposit as an asset. The modifying factors include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.With market fluctuations, many investors in the mining industry have become more and more risk averse which has increased the demand for more reliable data to support these modifying factors. This has led to the widespread use of geometallurgy and statistical analysis to better inform the resource estimation. Geometallurgy is used to characterize ore variability and is especially useful for highly variable or strongly zoned deposits, several close deposits that will be milled together, projects requiring complex or new metallurgical approaches, brownfields or sight-of-mine exploration, expansion projects that exploit new, deeper or adjacent reserves, and projects with significant legacy drill core that are being re-evaluated due to new economic circumstances."

Jan 1, 2018

By K. Yamaguchi

"This Precious metals content copper scrap is treated with a copper smelting process, and copper and precious metals are recovered. The precious metal lost in the slag is increasing with increasing of treated scrap amount. To determine the portion of the precious metals chemically dissolved and that associated with the mechanically trapped matte in the slag will be a key factor to improve the recovery of those metals. The distribution of precious metals between copper matte and FeOx-SiO2 slag phases is determined at 1300 oC and a fixed partial pressure of SO2 of 0.1 atm for the matte grades from 40 to 70 mass% of copper. The distribution ratios show a tendency to increase when the grade of matte is increased. In addition, the solubility of platinum and rhodium in the FeOx-SiO2-CaO slag under the air and the solubility of platinum in the FeOx-SiO2 slag equilibrated with a pure platinum or the Pt-Cu alloy were determined at 1300oC in the range of oxygen partial pressure from 10-9 to 10-7. The solubility of platinum and rhodium in the slag tends to increase with increasing content of copper in the slag and oxygen partial pressure. INTRODUCTIONThe stringency of the supply and demand of platinum group metal (PGM) intensifies by localization for resources and increase of the demand for PGM, furthermore, smelting and the recycling of PGM in the nonferrous smelting process take an important role. Scrap containing a precious metal for the purpose of collection of PGM is treated by copper smelting recently. With increase of the treatment amounts, loss of PGM in slags is the serious problem.In considering of PGM loss in slags, knowledge of the distribution behavior for PGM between slag and matte phases or slag and copper phases and the PGM solubility in slags is important. The distribution ratios of silver and gold between the FeOx-SiO2 slag and the Fe-Cu matte were reported by Roghani (Roghani et al., 2000), Shinha (Shinha et al., 1985) and Takeda (Takeda et al., 1993) However, there is no data for the PGM distribution ratio. On the other hand, the PGM solubility in slags was reported by Sano (Sano et al., 1997) However, there is very few report for the solubility of PGM in the nonferrous smelting slags."

Jan 1, 2012

The mechanism of thiosulfate decomposition and the products formed were investigated in the presence of pyrite or pyrrhotite in copper-ammonia solutions used for leaching gold. The presence of the sulfides significantly increased the decomposition of thiosulfate. The sulfide-surface-catalyzed oxidation of thiosulfate to tetrathionate by dissolved oxygen or the cupric tetra-ammine complex was proposed to be the dominant thiosulfate decomposition mechanism. The catalysis of the sulfides in this reaction could originate from their strong affinity for aqueous sulfur species and their semiconducting properties. Oxygen was found to be the initial driving force for thiosulfate decomposition in the absence of cupric ions and the driving force for the continuous decomposition of thiosulfate in the presence of cupric ions. Under an oxygen free atmosphere, the decomposition of thiosulfate was limited to reduce the cupric ions and the catalytic effect of the sulfides was also marginal. The dominant oxidative product of thiosulfate was trithionate in alkaline solutions especially in an oxygen rich or cupric ion rich environment. Tetrathionate was found in relatively smaller amounts in an oxygen and cupric ion deficient environment. Tetrathionate was not stable and would further decompose after extended periods. Trithionate concentrations increased almost linearly with time.

Jan 1, 2005

By J. H. Canterford

The nickel industry in Australia has followed the general pattern apparent in many other regions - lots of ups and downs, successes and failures, but always plenty of challenges. Some of these can be attributed to local geographical and infrastructure factors, others to a healthy mix of corporate good luck and management countered by decisions that, in retrospect, were less than optimum. The roles of several entrepreneurs have certainly added plenty of spice! In terms of processing technologies, there was an initial heavy reliance on the use of "off the shelf' commercially proven processes such as ammonia leaching of concentrates and mattes, and flash smelting. Over the past decade there has been an increasing dominance of plant modifications and process technologies that have been developed "in house" in order to enhance project economics and exploit lower grade and more complex resources. This presentation provides an overview of the past 40 years of the major developments in the Australian nickel industry and concludes with some guesses as to where we will be in another decade or so.

Jan 1, 2005

By M. Vallée, J. -P. Charette, D. P. Gold

"The Oka Carbonatite and Alkaline Complex is a Lower Cretaceous intrusive body of the Monteregian petrographic province. Carbonate rocks of various types, and associated alkaline rocks of the ijolite and okaite suites, are distributed in a double ring-dyke structure 4 Yi miles long by 1 1/2 miles across. Alnoite and alrioite breccia pipes intrude all rocks of the Complex and indicate, from inclusions, that a cover of Palaeozoic rocks existed at the time of emplacement.The results of radiometric, magnetic and gravity surveys of the Complex are compiled and discussed both with respect to the regional pattern and, locally, the bed-rock geology. The Complex is characterized by high-intensity anomalies.Exploration activities started during 1953, and mining operations began in 1961 on the property of St. Lawrence Columbium and Metals Corporation, at one time the world's largest producer of niobium concentrates. Because of the boudinaged and fractured nature of many of the rocks and the variability in grade of the ore minerals, the role of the mining geologist is both demanding and exacting. Rock types are varied and their mutual relation-ships ambiguous, the minerals are rare and complex, and close geological control of the ore is needed in the mine and in the mill. Closely spaced diamond drill holes are needed for mining control, and the ore-waste contact is defined mainly by assay results.In the exploration for 'carbonatites' in Eastern Canada, aeromagnetic surveys are judged to be the best exploration tool, followed in decreasing order by the search for circular structures on aerial photographs, regional gravity surveys and radiometric surveys."

Jan 1, 1967

By A. A. Brant

Self-potential methods are based ?on measurements on the surface of voltage distributions caused by natural sub-surface oxidation phenomena. The attempt is made to. correlate voltage minima or maxima with the location of mineral materials or geological conditions which may give rise to such naturally produced spontaneous potentials. Any orebody undergoing oxidation is a source of electric current. Ground waters carrying free oxygen attack the upper surface of the body, rendering this the positive pole of a galvanic cell, with the lower portion of the body the negative pole and the surrounding weak acidic waters the electrolyte. Current flow is from positive to negative through the orebody and negative to positive through the electrolyte, i.e., through the surrounding rock. The upper end of a steeply dipping, tabular orebody is hence indicated as lying below the point on the surface where the largest negative potential is observed(l). From its basis, it is apparent that the self-potential method was developed primarily for ore exploration.

Jan 1, 1942