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By Tony Robles

"A new process is disclosed for the recovery of gold and/or silver values from activated carbon wherein they are adsorbed as cyanide complexes. A unique chemistry allows the new process to use mill process water containing an inexpensive, non-hazardous compound as the eluant. The new process has an elution period of 5 hours at 98°C. The gold in the eluant can be recovered by electrowinning or zinc precipitation. The new elution process is simple, reliable, and provides an eluted carbon that may not require thermal regeneration before reuse. Retrofitting the new process to existing processes such as Zadra, Anglo, and others will not require significant capital expenditures.IntroductionThe efficient and reliable elution process for gold/silver from activated carbon is a very important process in the gold mining industry. The most common techniques for elution of gold cyanide from activated carbon are the Zadra and Anglo procedures.Specialized elution techniques like the Micron and Acetonitrile processes have not been commercialized to significant extent probably due to relatively complex unit operations and expensive reagents.Important functions of the gold recovery circuit are:Efficiently recover gold/silver from the loaded carbon.Carbon is in good condition for reuse for optimum recovery.The pregnant solution is amenable to electrowinning or zinc precipitation to recover gold/silver from the solution.The whole process to be simple and short to reduce gold inventories.Present Elution ProcessesTable 1 shows some carbon elution processes that are available today. The two main elution processes used in the industry are: (1) the Zadra process, in which hot cyanide solution is circulated through an elution column and electrowinning cells; and (2) the Anglo American Research Laboratory (Anglo) process, consisting of a pretreatment step with caustic cyanide solution, followed by elution with deionized water."

Jan 1, 2000

By M. Cai

A practical simple method is proposed to estimate the tensile strength (st) of strong brittle rocks from the strength ratio of R= sc/st = 8 sc/sci, where sc is the uniaxial compressive strength and sci is the crack initiation stress in an uniaxial compression test. 8 is the Griffith strength ratio and the term sc/sci accounts for the difference of crack growth or propagation in tension and compression in uniaxial compression tests. sci can be reliably obtained from volumetric strain measurement or acoustic emission (AE) monitoring. With the strength ratio R determined, the tensile strength can be indirectly obtained from the result of uniaxial compression tests. In addition, a practical estimate of the Hoek-Brown strength parameter mi is presented and it is suggested that mi = 12 sc/sci can be applied to strong, brittle rocks in high confinement zone and mi = 8 sc/sci can be applied to low confinement to tension zones. It is found that the predicted tensile strengths and Hoek-Brown strength parameter mi using this method are in good agreement with test data. The rock strength parameters like st and mi, which require specialty tests such as direct tensile (or Brazilian) and triaxial compression tests for their determination, can be reasonably estimated from uniaxial compression tests.

May 1, 2009

By Karla Kaboth, Brad Garraway

"The surface of many concentrate thickeners has traditionally looked like that of a flotation cell. Over the last three years Normandy Golden Grove Concentrator, in partnership with Outokumpu Technology have developed and implemented an innovative breakthrough deaeration technology to address this problem.When installed in the thickener feedline, the system successfully detached air bubbles adhering to the solid particles and removed air bubbles from the feed stream, subsequently presenting a fully deaerated feed to the thickener feedwell.Tests to date have shown an 80% reduction in overflow solids. As a result, costs associated with concentrate loss and measures to recover concentrate that settles in process tanks and ponds have been minimized.MOTIVATIONIt is recognised in mineral processing operations that the entrainment of finely divided air in thickener feeds can cause significant problems in the operation of a thickener. This is particularly the case with thickening of flotation concentrates, in which air bubbles are strongly attached to the mineral particles and are difficult to remove. Such particles attached to fine air bubbles rise to the thickener surface, and hence over the weir and into the thickener overflow. High solids carryover can lead to a direct loss of mineral product, and if the product settles in process water tanks and ponds, the recovery of this product by means of intermittent dredging operations is usually expensive and inconvenient.Many concentrate thickeners are fitted with devices which attempt to alleviate this problem. Froth rings or baffles, skimmers and froth sprays, all play a part in reducing the amount of solids carried over into the overflow. However, many froths are very stable and persist despite these measures. In the worst cases, a thick froth layer forms which builds up to a depth greater than the froth ring, and hence may overflow or pass beneath the ring . Any concentrate that finds its way past the froth ring can also cause blockage of the overflow weir and launder, and has the potential to upset clarification performance by cutting off sections of the overflow handling system. The formation of a froth layer on the thickener also makes it more difficult for the operator to monitor the clarification performance."

Jan 1, 2001

By K. Honkivaara, S. Junnikkala, J. Aho, K. Föhr, P. Lampinen

"Bubble area flux has been claimed to be a major tool to dimension flotation equipment. The paper describes laboratory scale (batch and 301 continuous cell using plant feed) and small-scale (1 ,8m3) industrial experiments with coarse-grained massive sulphide ore.Samples were taken with different froth thickness and air addition rates while the bubble size was measured with an UTC capillary device. Problems of averaging the bubble size measurements and the bubble area flux measurements are also discussed.Material balances were calculated using the commercial program BILCO. The flotation rate constants were calculated for pulp and for froth assuming that we can extrapolate the flotation rate constant of the pulp phase from tests using very thin froth thicknesses.The results indicate that the bubble area flux is not alone an adequate dimensioning tool. With fine sizes there was a nearly linear relationship between the bubble area flux and rate constant for constant impeller revolutions. For coarse sizes the form of the relationship is non-linear and seems to go through a maximum value. Thus the final flotation outcome is not a simple function of bubble area flux but depends on the particle size distribution and other more traditional flotation cell dimensioning variables. The effect of the froth residence time in the total recovery of coarse particles is large, but will not be treated in this paper."

Jan 1, 2000

By G. Warren

"In the past, smelting and refining of the precious metals recovered from ore has been generally looked upon as a small and relatively minor component of the business of gold mining. This is not surprising when one considers that the manpower, space and capital components of a gold refinery at an operating mine can be several orders of magnitude smaller than other unit operations at the same location. This philosophy is changing however, as economic and environmental pressure are forcing a re-evaluation of all aspects of gold mining and milling.This paper describes and evaluates three furnace technologies available and used by precious metal operators. The operation of arc, induction, and reverberatory furnaces are discussed and their strengths and weaknesses with respect to capital cost, impurity removal, slag quality, and environmental concerns highlighted.INTRODUCTIONIn the past, smelting and refining of precious metals recovered from ore has been generally looked upon as a small and relatively minor component of the business of gold mining. This is not surprising when one considers that the manpower, space, and capital components of a gold refinery at an operating mine can be several orders of magnitude smaller than other unit operations at the same location. Generally, gold producers will spend a great deal of effort to squeeze the last drop of efficiency out of a leach or grinding circuit via simulation or testwork but will design and build the refinery based upon past practices or industry standards. That past practices or industry standards will provide a functional refinery is undisputed; that this philosophy achieves the efficiencies that are possible or demanded in other plant operations is debatable.Environmental and worker hygiene requirements are considerations as well. Pollution control equipment can become a substantial contributor to both the capital and operating costs of a refinery and furnace selection has a marked influence upon the type and cost of this equipment. Slag characteristics and subsequent processing options available also play a role in furnace selection as precious metal slag losses and the recovery effort can greatly influence the cost / benefit analysis of a particular technology."

Jan 1, 2001

By K. D. Hester

"1.0 INTRODUCTIONThe uranium industry started in Canada with the reopening of the gravity concentrator at Eldorado's Port Radium mine in the North West Territories in 1942. In 1948, uranium exploration and processing was opened to the private sector, and the Federal government offered special pricing incentives to the industry. Eldorado explored the Lake Athabaska deposits in northern Saskatchewan, and in 1953 opened the Beaverlodge mine. Other mines in the area soon followed. In 1953 also, the deposits in the Blind River - Elliot Lake area and in the Bancroft area of Ontario were discovered. Mines and mills were quickly developed in these areas.Canadian uranium production peaked at 15,900 short tons of U3O8 in 1959. In that year, uranium exports at $300,000,000 exceeded those of all other minerals in value. On October 1, 1959, Canadian uranium operations employed 13,900 people.Up to 1956, it was a seller's market. By 1958, world production of uranium was far ahead of requirements. Accordingly, in November of 1959, the USAEC announced it would not renew it's options for continued delivery of uranium from Canadian mines. This led to the ""stretchout"" of existing uranium contracts, and many mines were shut down (see Table 1) .In the early 1970's, a commercial uranium market seemed to be developing. In Ontario, at Elliot Lake and Bancroft, mines reopened. But competition was coming from the Athabaska basin of Saskatchewan. Currently it is these high grade, open pit deposits of northern Saskatchewan which are in the forefront of Canadian uranium developments. Canadian developments since 1942 reflect the political and market forces, as well as metallurgical progress and changes in environmental standards. This paper attempts to show these process changes in broad perspective."

Jan 1, 1987

By Sanjeeva Latchireddi

"Deteriorating ore grade and high demand of metals has tremendously influenced the unit capacity of SAG mills. Pebble circuits have been added to the existing SAG circuits to increase their unit capacity. To make the pebble circuits justifiable, the grate openings are increased almost to 4.5 inches to draw large quantity of pebbles. The inherent transportation problems associated with SAG mills in closed circuit with pebble fraction are more complicated than the single stage SAG mill operation, because the coarse pebbles behave significantly different to that of slurry in pulp lifters. In addition, operation of mills at higher speed, to take advantage of precise design of high release angled shell lifters significantly reduces the discharge efficiency of the conventional pulp dischargers. The presence of a sizable quantity of pebbles in the pulp lifter limits the space available and reduces the flow gradient through the grate thus increases the load inside the mill and hence the mill power-draw. This paper analyses all the problems associated in discharge of slurry and pebbles in SAG mills with pebble circuits. The Outokumpu patented Turbo Pulp Lifter (TPLTM) aims at eliminating the inherent material transport problems to ensure good grinding conditions. Recent installation of TPLTM in a 26-ft SAG mill has proven these claims by reducing the specific power consumption while increasing the mill capacity.INTRODUCTIONThe conventional wisdom is to look at the semi-autogenous grinding (SAG) mill as just another grinding mill in which ball charge, mill filling, and mill speed are varied to attain maximum throughput. Liner design for the SAG was also approached more from a wear perspective rather than a metallurgical and hydraulic-flow perspective. The extensive research work carried out by Latchireddi and Morrell (1997,2003a and 2003b) has demonstrated to think of the SAG as more of a pump in which the grates and pulp discharger end (“impeller”) were designed to efficiently discharge the product sized particles smaller than the grate size with minimal recirculation out of the mill.The rock load in the mill is essentially depends on ore characteristics and the discharge rate of broken particles. The discharge rate of the product material depends on how efficiently the discharge pump (grate and pulp lifters) is operating. Similar to the impeller design affecting the pump capacity, the pulp lifter design affects the discharge capacity (or mill throughput) of the SAG/AG grinding mills. The general arrangement of AG/SAG mills discharge system is shown in Figure 1."

Jan 1, 2007

By R. Sharon, P. M. Hawley, X. Ochoa

"Development of Barrick Gold’s Pierina gold mine in north-central Peru requires construction of a waste dump for disposal of nonore grade materials stripped from the open pit to expose the ore. The waste dump site is located near the open pit in a U-shaped valley that is underlain by thin colluvial and morainal soils with locally thick peats and glaciolacustrine deposits. Bedrock underlying the site consists of variably weathered volcanics. The site is subject to moderately high seasonal rainfall and significant seismicity. The geotechnical quality of the waste rock is highly variable, ranging from low-quality, degradable, soil-like materials containing appreciable quantities of silt and clay, to high-quality, hard, durable, blocky waste rock.A hybrid dump design concept was developed that requires in-pit segregation of highand low-quality waste, and selective placement of materials in the dump to maximize stability and enhance internal and subsurface drainage. The design includes a rock drain beneath the dump along the axis of the valley and a series of connected finger drains. Extensive peripheral diversions and stormwater control measures to inhibit infiltration of precipitation are also included.This paper discusses the unique challenges presented by the site and waste materials, the design methodology and stability analyses, and planned monitoring systems."

Jan 1, 2003

By M. Keen

The backfill system at Cleveland Potash?s Boulby Mine places waste filter cake underground in the form of a high concentration slurry in worked out areas of the mine reducing the mine?s environmental impact. The backfill system delivers slurry approximately 1100 m down the mine and up to 11 000 m horizontally underground. The Boulby backfill system is unique in that the placed backfill waste is not used for the usual purposes such as support, pillar extraction or reducing ventilation requirements. It was specifically implemented for environmental reasons. This influenced the design and resulted in some unique features of the system. A previous paper (Wilkins et al., 2004) described the design and commissioning of the backfill system. This paper describes the operational experience with the backfill system and backfill placement since commissioning in May 2003 and looks at lessons that can be applied in other backfill designs.

May 1, 2007

By John Lindsay

"Although diamond mining is a relative newcomer to the Canadian mining industry, with the first diamond mine coming on stream in 1998, Canada is already a significant producer of high quality diamonds. This paper will review diamond production trends and highlight some of the projects currently under development, which have the potential to enhance Canada’s role as a significant diamond producer.Diamond recovery plants generally incorporate well-established mineral processing and recovery techniques, with issues such as product security, cut-off sizes, and diamond preservation becoming important considerations when developing process flowsheets. A diamond processing plant must achieve three key objectives, namely:• Liberation of the diamonds from the host rock;• Separation of the heavy minerals, which include the diamonds, from the host rock;• Recovery of the diamonds from the heavy mineral concentrates.Processing technologies and equipment commonly utilized in diamond processing plants will be reviewed and some of the significant issues associated with each option briefly discussed.Typical process flowsheets will be presented.Finally, an overview of De Beers’ Victor diamond project, located in northern Ontario, will be provided. This potentially significant diamond producer is currently at a pre-feasibility stage of development.DIAMONDS IN CANADAWorld Diamond ProductionCommercial diamond mining commenced in earnest with the discovery of the Kimberley diamondfields in South Africa in the late 1800s, and this resulted in a dramatic increase in the supply of diamonds. As can be seen from figure 1, world diamond production, mainly from South Africa, Angola, Congo, and Namibia, remained relatively constant through the early 1900s until the development of the large Russian kimberlite pipes in the 1950s. Production increased again in the 1960s and 70s with the development of the Botswana mines, but the most significant increase in global production came from the Argyle mine in Australia, which came on stream in 1983 and produced a peak of almost 43 million carats in 1994, representing approximately 40% of the world market. Production from Argyle has since fallen-off as grades have declined and the strip ratio has increased with the deepening of the open pit."

Jan 1, 2003

By Jianwen Yang, Yucai Zhou, Yanhong Dong, Daixiong Chen

"The study of combined depressants for flotation separation of lead from copper has been carried out to solve the copper-lead separation problem. Test results show that the depressant of CMC (Carboxyl Methyl Cellulose) or sodium sulfite plays an important role on the selective depression of galena flotation. When the two reagents above are combined, the selective depression effect on galena is further enhanced. When the mass ratio of CMC to sodium sulfite is 1:4, galena was best depressed, with the lowest lead content in the copper concentrateClosed-circuit tests indicate that under a certain pH value of the slurry, with the best ratio of CMC to sodium sulfite as depressants, the copper concentrate with the grade of 28.23% copper and 4.12% lead is obtained, and the copper recovery is 84.87%; the lead concentrate with 68.12% lead and 0.86% copper is obtained, and the lead recovery reaches to 92.20%. An efficient separation of copper and lead is achieved with the combined depressants. INTRODUCTIONComplex copper-lead-zinc sulfide ore is abundant in China. With the rapid increase of mineral consumption, it has become one of the important sources to produce copper-lead-zinc metals. However, it is difficult to achieve effective separation of copper and lead, due to the complicated ore properties, such as fine dissemination, low copper grade, complex symbiotic relationship and similar floatability of copper and lead.Poisonous reagents of bichromate and cyanide are often used as depressants for copper lead separation. They have good separation efficiency, but pose serious pollution issues. As the environmental protection is becoming more and more important, it is a trend to replace traditional toxic reagents with new nontoxic efficient chemicals. Organic depressant FCLS studied by Liu has strong selectivity against galena, and does not depress chalcopyrite. A new type of galena depressant of CLS-01 discovered by Jin-lin Yang shows good results in the separation a mixed copper and lead concentrate. In addition, dextrin, chitosan and some other natural organic compounds (Drzymala, 2003; Huang, 2012; Qin, 2013) have also been studied to separate copper and lead."

Jan 1, 2016

By H. Perez, A. Lavoie, V. Prévost, P. Létourneau, P. Lind

The Home smelter is a custom copper smelter located in Rouyn-Noranda, Quebec. In 2005, the smelter produced 145,000 t of copper containing 807,000 oz of gold, 20,388,000 oz of silver in addition to 515,000 t of sulphuric acid. Recently, a flexible flow sheet has been developed incorporating the intermittent use of the Noranda Continuous Converter allowing for the increasing or decreasing of production based on market conditions, while attaining 90% or greater sulphur fixation regardless of operating mode. For 2006, the smelter is operating in full production mode and production is expected to be approximately 185,000 tm of copper. Mechanical modifications, process changes, metallurgical concepts, minor element treatments, and operational techniques that have allowed the smelter to develop this flexibility are presented. Current efforts to increase the productivity of the converter aisle and anode furnaces as well as improvements being made to address fugitive emissions are also discussed.

Jan 1, 2007

By Emilie S. Williams

The Doyon Mine?s Zone 1 (Altered Zone), found along the southern boundary of the mine, encompasses approximately 20% of the mining reserves. The mineralisation is found in a series of steeply dipping narrow veinlets, which are contained in an extremely weak host rock: a sericitic schist with RQD values ranging from 0 to 20. Longhole mining methods are used, with either longitudinal or transverse approaches, depending on the width of the ore zone. Dilution is of critical importance in the mining of this zone, particularly in longitudinal stopes, where dilution levels have averaged 55% since 2000. Major factors influencing dilution in this zone include the stope geometry and standup time. A two-phase approach has been developed to mine and backfill longitudinal stopes: the top of the stope is pastefilled after approximately half of the broken ore has been extracted; the remainder of the ore is extracted once the plug has cured; and the remaining void is filled. This method has been used in six stopes to date, resulting in an average dilution of 17%. This paper details the application of this technique and its effect on dilution.

Oct 1, 2004

By Mark Fernald, Christian O’Keefe, John Viega

"In this presentation, the SONARtrac™ technology platform will be described. CiDRA's proprietary clamp-on, passive sonar array-based flow meter technology performs two independent measurements – flow rate and fluid characterization. Firstly, the meter provides the volumetric flow rate of the mixture by measuring the speed at which naturally occurring turbulent structures convect with the flow past an axial array of sensors. Secondly, the meter uses similar sonar-based processing techniques and naturally occurring sound in the process fluid to measure entrained air levels. The result is a unique ability to measure the flow rate and entrained air level of most fluids – clean liquids, high solids content slurries, and liquids and slurries with entrained air.Also to be presented is the application of the sonar-based technology platform in a variety of oil sands processes, hydrotransport, and minerals beneficiation applications. In particular, difficult slurry flow measurement and control in the areas of comminution and flotation such as mill discharge, hydrocyclone feed/overflow, final concentrate, thickener discharge, and tailings will be discussed. In addition new real time maintenance free process measurement capability will be presented in the areas of entrained air in slurry, collection zone gas holdup (?g), and slurry velocity profiling. The operational advantages and value of these measurements, even in the presence of scale buildup, will be discussed.INTRODUCTIONFlow measurements in the mineral processing and oil sands industry suffer from the limitations placed by previously available flowmeters including the commonly used instruments such as ultrasonic meters, magmeters, turbine meters, orifice plate meters, vortex flow meters, Coriolis flow meters, and venturi meters. The desire or requirement to make a non-contact measurement that is accurate and robust, and can be performed on practically any type of fluid within practically any type of pipe has driven the creation of a new class of flowmeters. This new class of flowmeter technology utilizes sonar-based processing algorithms and an array of passive sensors to measure not only flow, but also fluid composition."

Jan 1, 2007

By Sami Hindstrom, Travis Orser, John Starkey

"A great deal of effort has been expended in the last 11 years by the authors and others to identify the correct power required to grind ore in a SAG mill, by developing ore hardness variability functions for the ore body to be mined. This paper describes the next step in the SAG design process, which is choosing a SAG mill to accomplish the required specified design performance.At the test work stage, each ore hardness data point is a unique and discreet bit of information that represents the energy required to grind the ore represented by that sample at optimum SAG conditions. When the design criteria are set, it is now considered necessary that the required power draw in the mill chosen be provided at optimum conditions as well.Based on actual 3-ft diameter pilot plant results, we can now show that the grinding efficiency and throughput drop when the optimum load limit is exceeded. Future work is expected to show a similar reduction in efficiency when exceeding optimum speed. Providing specified power by exceeding these optimum conditions is not adequate and leads to a lower cost but underperforming SAG mill.It is possible in some cases to modify a SAG circuit to force it to perform better, but this often takes time, much effort and costly modifications. In most cases the capital spent on a properly sized SAG mill may prevent production losses that would be many times greater over the extended start up period. This is the direct result of choosing the proper SAG mill at the start.INTRODUCTIONThe sequential steps required for the design of any grinding circuit, have changed dramatically in the last seven years. Previously, it was considered necessary to do six foot diameter SAG mill pilot plant tests on representative samples of minus six inch ore (F80), in order to understand design requirements. It can now be shown that this method is costly and not likely to produce the desired result because it is difficult to recover six-inch material from deep within a new mine. The representative nature of the sample is often compromised when the sample is selected."

Jan 1, 2003

By Eiichiro Mizushima

Catalyst systems comprising (Ph3P)AuCH3 and acid co-catalysts were found to be highly efficient for hydration of alkynes to give ketones. Methanol was the best choice of solvents. Introduction Hydration of carbon-carbon unsaturated bonds is one of the environmentally benign routes to the carbon-oxygen bond formation. Hydration of alkynes forming carbonyl compounds, an important variation in this category, is a well known reaction which has been extensively studied. The reaction by acid-catalysis is the most conventional method, while it is applicable only to electron rich acetylenes to obtain satisfactory yields (1,2). The hydration of simple alkynes usually requires another catalyst (co-catalyst), typically a mercury(II) salt (3), which is inconsistent with the guiding principle of Green Chemistry (4).

Oct 1, 2003

By A. C. Wilson

"In 1991, local prospector C. “Mickey” Clement recovered at least three alluvial diamonds from the Michipicoten River. Two of these stones were sent to the Department of Mineralogy, Royal Ontario Museum, where they were identified as industrial-grade diamonds with weights of 1.05 and 1.13 carats. Both stones were described as frosted and graphiteinclusion riddled.In 1995, prospector Sandor Surmacz and geologist Marcelle Hauseux discovered diamonds in bedrock approximately 20 km north of the Michipicoten River. An 18.1 kg bulk sample of a xenolith-rich lamprophyre, taken from an outcrop at the side of the Trans Canada Highway north of Wawa, yielded 1 macro diamond and 5 micro diamonds. All but one was gem quality (Sage, 2000).Since then, more than 36 occurrences of diamonds in bedrock have been reported in an area approximately 30 km2 in size, centred 20 km north of the town of Wawa. The occurrences are hosted within a sequence of unusual, Archean-aged, heterolithic breccias previously considered to have no economic significance. Ten kimberlite occurrences also have been confirmed in the Michipicoten greenstone belt in the last five years. Two of these kimberlites are diamondiferous.To date, expenditures on property acquisition and diamond exploration in the Wawa are estimated at CDN$120 M.This paper summarizes the geological observations made on these occurrences and describes the relationship between the occurrences and the regional geology."

Jan 1, 2004

By G. Song

Magnesium has excellent bio-compatibility in terms of its density, strength, elastic modulus and toxicity. However, due to its poor corrosion performance, so far magnesium has not been successfully used in the human body as a permanent implant material. Nevertheless, there is a possibility to utilize the low corrosion resistance of magnesium to make magnesium into a temporary implant component that can gradually dissolve in the human body after surgical region heals. This will avoid the secondary operation to remove the temporarily implanted component and the chronic inflammatory discomfort caused by the implant. This paper is aimed at exploring the possibility of using magnesium as a degradable bio-material through investigating the corrosion behaviors of pure magnesium and surface treated magnesium in a simulated body fluid (SBF) solution. It was found that magnesium corroded rapidly in the SBF solution, resulting in vigorous hydrogen evolution and a strong alkalization effect. Hence, it is impractical to use magnesium as a permanent implanted material or even a degradable implant. However, a proper anodizing treatment can significantly improve the corrosion resistance of magnesium, and hence slower down the hydrogen evolution and alkalization processes, which will allow to control the corrosion or degradation of magnesium at a desired rate in the human body. With a controllable corrosion rate, it is believed that magnesium as a degradable implanted material will be eventually achievable.

Jan 1, 2006

By W. C. Whittaker

"This paper gives a brief resume of the changes in the market picture for Western coal since 1948. It then analyzes present markets and recent developments in Western Canada, the western United States and Japan, and forecasts future markets in these areas to 1975.INTRODUCTIONCOAL HAS BEEN MUCH IN THE NEWS OF LATE, both in the United States and in Canada. In the United States, this is because of the booming condition of the industry and because of the purchase of control of a number of America's largest coal producers by the integrated oil and metal producing companies seeking additional fossil fuel reserves and diversification. In this regard, it is interesting to note that ownership of America's largest strip coal producer was recently acquired by one of the world's largest copper mining companies. In Canada, the conclusion of several large-scale, long-term contracts for the sale of coal to the Japan Steel Group and the projected use of Canadian coal in steel mills in the western United States has focused much attention on the coal industry."

Jan 1, 1968

By I. Filzwieser, A. Filzwieser

The Peirce Smith converter is the key metallurgical vessel in the majority of copper smelters. During the two step process, oxygen enriched air is injected through a row of tuyeres into the matte received from a primary smelting unit. Additionally, in the first step the iron content is oxidized into a fayalite slag by the addition of silica. During the second step, the remaining copper sulphide is converted into blister copper with a final sulphur content of approximately 200 ppm. The overall efficiency of this process is determined by the total reaction time, the slag quality and quantity, and the chemistry of the final blister copper. To increase the reaction kinetics during both stages, the METTOP Process Control and Regulation (MPCR) system was developed, which consists of four elements: OPC - Optical product control ACR - Automatic computer regulation GCU - Gas control unit GIS - Gas injection system The OPC system from SEMTECH records the optical spectrum emitted by the off gas flame providing immediate, detailed process information. Using this online data, additional gas can be automatically injected into the converter through a gas injection system opposite the tuyere zone. An improved oxygen distribution in the melt and the accurate determination of the process end point results in a significant reduction in the process cycle time. Furthermore, by maintaining the reaction closer to the equilibrium the final sulphur and oxygen content can be reduced compared to the conventional process. In particular, the final sulphur content can be lower than 70 ppm, eliminating the need for an oxidation step in the anode furnace.

Jan 1, 2007