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By D. W. Ginsberg, D. Cieplinski

"The lifetime cost of a material handling system, or any mine system, can be reduced through the application of quality engineering. Dynamic simulation is a powerful tool that can be used to enhance the experience and ingenuity of the design team. It cannot only de-bottleneck the system to improve system productivity, but also define critical design criteria to eliminate over-design and reduce capital cost. Eliminating over-design also reduces operating costs by lowering component replacement costs and by ensuring all parts of the system operate at a higher utilization and efficiency. IntroductionA mine owner’s profitability is improved when a mine is developed at the lowest lifetime cost. This cost relates not only to capital and operating costs, but also, to a great extent, on the efficiency and reliability of the operation. These factors are tied directly to the quality of the engineering which depends on the experience of the engineers involved, and their ability to apply known system information in a rational and constructive manner.It is becoming apparent that it is increasingly more difficult to deliver the quality of engineering expected and needed to provide attractive lifetime costs using traditional design methods. Historically, system information has been evaluated based on a limited amount of information to which global rules-of-thumb and the engineer’s personal experience were applied to develop design criteria. However, market forces are dictating that more system parameters be defined and that they be defined more accurately so that the most economical system can be developed. This increase in the volume of information and the need for more accurate predictions is straining, if not already exceeding, the ability of engineers to apply simple global rules of- thumb and to combine all the information in a rational, comprehensive and representative set of design criteria. Furthermore, as proposed system production rates increase beyond established industry practice, the extrapolation of the engineer’s experience becomes more difficult and increases the risk in the proposed design."

Jan 1, 1998

By P. Hajiani

"It has been shown that the economics of the hydrometallurgy of a rare earth process depend upon the amount of acid needed to extract the metal per unit mass of the ore. In this paper, we show how to save a significant amount of acid treating a low grade ferrocarbonatite and extracting +99% constituent REEs. In order to save the acid, the alkali earth metals, originated from gangue minerals were removed prior to REE leaching. Moreover, the operating conditions of leaching reactor are optimized to leave iron behind and to leach REE selectively and effectively. Consequently, the hydrochloric acid consumption per ton of ore drops significantly compared to other processes.INTRODUCTIONThe viability of an extractive metallurgy process of rare earth elements (REEs) to produce an intermediate mixed concentrate depends upon a low-cost simple flowsheet (Bogner, 2014). It has been known that regardless of the mineral, the acid consumption is intimately associated with the operational cost (OPEX) of any REE project, particularly when the initial REE grade is low. Besides acid, the other reagents consumed in the process impact the OPEX. This makes minimizing acid consumption in an REE method a critical priority. The acid consumption in different projects has been reported to be equivalent of 100 kg to 1000 kg HCl (100%) per tonne of ore (Average: 500 kg/t HCl 100%) (Bogner, 2014). This value is strongly correlated with the grade of REE in the ore and therefore in the pre-concentrate.Previously, it has been stated that the Montviel flotation concentrate requires about 300 kg HCl (100%)/t ore to attain +99% REE recoveries using direct leaching (Yu, Verbaan, Pearse and Britt 2013). Hence, for an ore feed rate of 2500 t/d ROM, a large acid production plant (744 t/d HCl 100%) would be required for acid leaching of such concentrate. The equivalent Chlor-alkali plant is estimated to cost M$ 150 (CAPEX) and may require 100 MW power. Therefore, there is a necessity to improve the extractive hydrometallurgy and to reduce the acid consumption significantly."

Jan 1, 2016

By W. G. Hatch

Introduction Classified equipment originally installed in 1923 in the Britannia mill consisted of eighteen Model C Dorr Simplex classifiers, each in closed circuit with a pebble-mill. These machines were only three feet wide and of light construction, so that their raking capacities were necessarily limited, the maximum being about 20 tons per hour for each classifier. On conversion from pebble-mills to ball-mills, in 1924, attempts to increase circulating loads with higher rake speeds resulted only in overflowing a coarser product to the flotation cells. The elevator type of classifier was developed locally in 1935 and during that year two units, each having a sand carrying capacity of from 30 to 40 tons per hour, were installed. With such loads, a material increase in the grinding capacities of the ball-mills was noticed, and, after further development work on the elevator classifier, fifteen of the small Dorr classifiers were replaced by units of this type, capable of handling approximately 75 tons of sand per hour, with later alterations increasing this capacity to well over 100 tons per hour. In the meantime, the three remaining ball-mills had been placed in circuit with Dorr FX classifiers, one of which was six feet wide and the others five feet wide. Impartial and exhaustive tests conducted on the Dorr FX versus the elevator classifier showed the average results on these machines at Britannia to be substantially the same. A large amount of testing was carried on to determine the effect of circulating load on the output of the grinding units. This work indicated a rapid increase in the grinding rate as the ball-mill feed increased up to about 60 tons per hour, after which a relatively slow increase took place up to 100 tons per hour. No benefits were derived from higher circulating loads. A 300 to 500 per cent circulating load, depending on the nature of the ore treated, is considered to be within the economical range for the grinding of Britannia ores.

Jan 1, 1946

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 Alfred Annor

Deep mining requires innovative approaches in ground control technology and the search for new ground support materials including backfill. There is a need to look at mine waste materials that may be available in deep mines to assess their suitability for use as composite backfill. Composite materials are widely used in all areas of engineering because of their unique and often superior structural properties compared to those of the singular constituent materials. Over the past few years, there has been a growing interest towards the investigation and use of various types of composite backfill materials as special applications fill in mines. A research project is being conducted at CANMET- Mining and Mineral Sciences Laboratories, Natural Resources Canada, to investigate the properties of potential ground support materials for use in deep mines. As part of this investigation, composite backfill materials were prepared by combining waste rock aggregates and mill tailings in various ratios, to determine their optimal mix proportions for strength development. A few cemented rockfill and straight tailings paste backfill samples were also prepared and were used as controls for the study. Two sample sizes were used for the investigation in order to evaluate the scale effects on the material behaviour. A set of samples were cured at 44 EC to simulate the anticipated temperature conditions in deep mines. A second set of samples were cured at 23 EC to provide engineering reference data. The test results show that compared to either the cemented rockfill or the straight tailings pastefill material, the composite backfill samples were the strongest and stiffest of all the test materials investigated in this study. The composite backfill samples also produced the lowest void ratio and porosity values compared to the cemented rockfill and the straight tailings backfill material. This paper will present the results of the laboratory investigations to date. It will also outline the on-going work at CANMET - Mining and Mineral Sciences Laboratories to further assess the ground support potential and viability of using composite fills in deep mines.

May 1, 2003

By K. L. C. Gonçalves, V. K. Alves

"CVRD’s Department of Development and Technology – GETEK has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. This methodology has been applied in several prospective CVRD mining projects. It encompasses a procedure for variability analysis based on bench scale tests, pilot scale grinding and flotation tests and modeling and simulation for scaling up the industrial flowsheet.The bench scale tests include the determination of the Work Index, simplified DWT tests, rougher flotation and locked-cycle tests with samples that are representative of the ore body variability. After completion of these studies, continuous, pilot-scale flotation tests are carried out at a Mini Pilot Plant with drill core samples representative of different stages of the mine life. Finally, studies are carried out with large volume samples using a conventional grinding and flotation pilot plant. This sample is also used for liberation characterization, grinding kinetics, flotation kinetics and industrial circuit scale up. This methodology has been successfully employed in the development of CVRD’s new mining projectINTRODUCTIONCVRD has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. The definition of the processing route and the sizing and scale up of the processing system are some of the products generated during the pre-feasibility of a greenfield project. This paper is focused on the scale up of grinding and flotation circuits, operations known to be of the most importance in a mineral processing plant.The methodology for the definition of the flotation and grinding circuits is divided in the following stages: bench scale ore variability studies, pilot plant operation and modeling followed by simulation. The product of this work is a calibrated simulator that can be utilized for plant operation as well as mine planning, looking to achieve optimal operational conditions.The current models used for scale up of the grinding circuits are aimed at the generating information about the size distribution of the circuit product. However, it is well known that the liberation state of these products may be of critical importance for the performance of the concentration processes downstream. Nevertheless, CVRD has been developing a methodology for liberation kinetics during grinding, aimed at optimizing grinding circuit throughput together with downstream flotation recovery."

Jan 1, 2006

By Andrea R. Gerson, Roger St. C. Smart, Russell Schumann, Joan E. Thomas, Warwick Stewart, Stuart Miller

Sequential Net Acid Generation (NAG) accelerated oxidation testing using peroxide was carried out on single sulfide mineral samples. The calculation of the nett acid production potential (NAPP) assumes that all the sulfide is present as pyrite and that dissolution results in a 2:1 ratio of H+:SO42-. NAPP as kg H2SO4/t is calculated as 30.6 x wt.% S present. To test whether this calculation is valid a NAPP* value was defined in which instead of using the wt.% S present for the calculation the wt.% S dissolved is used. Stoichiometric analysis of the dissolution products of the sulfide minerals as they would occur at pH 7 indicates, that in all cases, the ratio of H+:SO42- of 2:1 is viable. The NAG to pH 7 and NAPP* closely coincided for pyrite, bornite, chalcocite and covellite. For the three Cu-containing minerals, the NAG to pH 4.5 value was 0 as is predicted on the basis of the speciation likely to be present at pH 4.5. Acid generation is due to hydroxylation of the dissolved Cu between pH 4.5 and 7.0. The NAG to pH 7 and NAPP* values for chalcopyrite, arsenopyrite and pyrrhotite were in reasonable agreement but were not for either sphalerite or galena. Zn is not completely hydroxylated at pH 7 as is assumed by the calculation. For galena, surface analysis identified polysulfide formation on of the leach residue polysulfide. The formation of polysulfide is an acid consuming reaction. Therefore under conditions of accelerated oxidation the NAPP value is a reasonable estimate of the possible acid production in all cases except for sphalerite and galena. However, under conditions of less intense oxidation, for instance in columns, waste rock and tailings dumps, the formation of polysulfides is likely to be more widespread thus reducing the correlation between NAPP and the maximum possible acid production.

Jan 1, 2004

By Rick Steen, Weldon Read, Morrism Tracy T.

Edging for starter sheet blanks used in conventional copper refineries has always been problematic for getting the edging to seal and adhere to the titanium material used for the blank so that copper is prevented from plating under the edging and cause undo problems in getting the copper sheets to separate from the blanks. A method was developed at the Amarillo Copper Refinery utilizing an injection molding process that better adheres the edging material to the titanium that makes the harvesting of starter sheets much easier to the point that robots are now used to strip with and which holds up very well to the unforgiving physical nature of the electroplating process. It is believed that this technology could be successfully utilized for permanent cathodes as well. INTRODUCTION ASARCO LLC (Asarco) was organized in 1899 as the American Smelting and Refining Company. Originally a consolidation of several lead-silver smelting companies, it evolved over the years into an integrated producer of primary non-ferrous metals, including copper, lead and zinc and associated co- products (Marcosson, 1949). Grupo México, S.A.B. de C.V acquired Asarco in 1999. Today, Asarco is a vertically integrated producer of primary refined copper and associated co-products such as gold, silver, molybdenum, selenium and tellurium. Asarco operates three open-pit mines and mills and two solvent extraction-electrowinning (SX-EW) plants all in Arizona plus a copper smelter in Arizona and a copper refinery in Texas. Figure 1 illustrates how the copper metal comes out of the ground at the mine and ends up as a refined product at the refinery. The copper refinery in Texas is known as the Amarillo Copper Refinery (ACR) and is located in Amarillo, Texas. It was built in 1975 to replace 3 other aging refineries and was designed to produce 550,000 tons (500,000 MT) of refined copper per year using conventional starter sheet technology. At the time it was commissioned it was the largest tankhouse in the world at just under 1/2 miles long under one roof.

Jan 1, 2019

By A. Burkert, A. E. Hajj

"An information system has been developed for the use of companies considering developing a new mine which may be liable to secondary sulphide ignitions or by existing operations. The system uses the Comdale X expert system to analyze information about a particular mine. The system will indicate whether the mine is likely to experience ignitions and what precautions can be taken to reduce or prevent their occurrence. The advantage of an information system using expert system and artificial intelligence principles is that it can recommend action when there is an incomplete knowledge of all the system parameters. By interrogation of the user it can determine the state of knowledge and react accordingly. It is considered that such systems could be of significant use in mine planning for new and existing properties, especially where operators have limited expertise regarding secondary sulphide ignitions. IntroductionSulphide ignitions occur in the underground mines of several countries, including Australia, Canada, Finland, South Africa and the United States of America (1·3) and there has been a significant increase in their number and severity over the last decade(2.4). The basic causes of such ignitions are not well understood and man incidents are not reported or documented.Sulphide ignitions can pose a serious threat to the health and safety of mine workers and to the cost and efficiency of the mining operation (1-5). Byberg(6) described the three major problems associated with sulphide ignitions as:1. the force of the initial explosion can endanger workers and equipment;2. the intense heat of the ignition's flame front can endanger workers and equipment and may also ignite combustible material resulting in mine fires; and3. the sulphur dioxide produced by the ignition may be of such quantity and concentration to endanger workers, as was tragically demonstrated at Geco. In this incident in October 1985, a worker was killed at a position 2400 metres from the ignition point of a sulphide ignition at Manitouwadge, Ontario."

Jan 1, 1994

By Tatyana Alexandrova

The problem of processing refractory (rebellious) gold ores is important for all the countries involved in recovering precious metals from ores. Technological aspects of processing precious metal ores are examined in two respects: enhancing gold recovery from double refractory ores; and using carbonaceous ores on a production scale. The dispersed carbonaceous matter (DCM) present in concentrates causes the adsorption of [Au(CN)2]- complex on their surface during cyanidation process. Depending on the degree of graphitization and the organic carbon content in DCM (including kerogens and bitumen), the technology used to recover precious metals varies considerably. The problem of processing refractory (rebellious) gold ores is important for all the countries involved in recovering precious metals from ores. Depending on the degree of graphitization and the organic carbon content in DCM (including kerogens and bitumen), the technology used to recover precious metals varies considerably. The possibility of rare metals recovery, including rhenium, from black-shale rocks, has been examined. Comparative assessment of analytical methods of rare metals detecting has been carried out. It has been proved that neutron activation analysis is a promising technique. It has been shown that colloid-salt fraction (nanofraction) with particle size less than 1000 nm makes up to 3 wt% of black shale samples and is extracted by water under special conditions. The use of water instead of chemically active reagents provides insolubility of the mineral matrix and fully reflects the colloid-salt element of the samples. To sum up, the justification of the choice of technology for processing of carbon-bearing ores is based on examining the type of carbonaceous matter and its adsorption behavior. When the content of carbon of high degree of metamorphism is considerable, it is possible to recover platinum group elements and rare elements from this type of ores. The obtained results predetermine the general guidelines for creating new methods and techniques of processing carbonaceous material for integral development of solid minerals in the mining and oilfield areas of Russia.

Jan 1, 2016

By J. A. Eisele, D. L. Pool

"The Bureau of Mines has developed a particle agglomeration technique to improve the flow of leaching solution through heaps of low-grade gold-silver ores which contain clays. Benchscale experiments and industrial practice have shown that the percolation rate of cyanide leaching solution is greatly enhanced by mixing the ore with Portland cement as a binder, moistening the mixture, mechanically tumbling to form agglomerates, and aging before heap building and leaching. A slight modification ofthe technique allows its application to precious-metal-bearing tailings, which cannot normally be treated by heap leaching because they are too fine. IntroductionMany low-grade gold and silver deposits and tailing materials from former mining operations in the western United States are potential sources for precious metals. These sources are too low grade or too small to support the capital cost of a conventional cyanidation plant but have stimulated the development of lowcost procedures, such as extraction of the precious metals by heap leaching with cyanide. The Bureau of Mines has investigated methods to improve the recovery ofgold and silver by heap leaching these low-grade materials.For heap leaching to be successful, the crushed and stacked ore must have good permeability so that uniform distribution of the cyanide leaching solution through the heaps can be achieved. Some of the most difficult gold and silver ores to treat by heap leaching are those containing excessive amounts of clays, or fines generated during crushing. Excessive amounts of slimes, generally minus 50-micrometre or 270-mesh Tyler sieve size particles, in the heap feed will slow the percolation of the leaching solution, cause channelling, or produce dormant or unleached areas within the heap, resulting in long leaching periods and low extractions. In extreme cases, the clays or slimes can seal the heap and cause the leaching solution to run off the sides of the heap rather than through it."

Jan 1, 1987

By I. A. Cameron, F. Ham, R. Sriram

"Planned increases in electric furnace productivity and higher matte temperatures needed for the Sudbury Smelter's S02 Abatement Plan, require an improved matte tap hole. Since startup in 1978, the matte tap holes have been incrementally improved by the addition of copper cooling blocks around the refractory tap blocks. These improvements provided a tap block life of about 100 taps per campaign.In July 1993, a prototype matte taphole was installed in Number 2 Electric Furnace to develop the technology to increase tap hole service life. The design was based on a heat transfer model of the tapping process to determine candidate refractory materials and cooler arrangements. The prototype tap hole features increased copper cooling, high conductivity refractories and extensive instrumentation. Eight trial campaigns are reviewed outlining the performance of silicon carbide and graphite refractories and typical heat loads during tapping.IntroductionFalconbridge Limited operates two 36 MYA, 6 electrode submerged arc electric furnaces in Falconbridge, Ontario. The furnaces, commissioned in 1978, smelt a partially roasted nickel-copper calcine, to produce matte for converting and slag for discard. Additional details of the smelter are available in the literature(1-4).Each furnace was equipped with two matte tap holes located in the furnace end wall. Early in the campaign, matte tap holes were identified as a weak area in the furnace which could be improved. From 1981 to 1987, incremental improvements were made to the design by the addition of water cooled copper blocks. In 1993, a prototype matte tap hole was designed and installed in Number 2 electric furnace to develop the technology to increase tap hole service life. Improved service life is required to meet planned increases in electric fWl1ace productivity and the higher matte temperature needed for the smelter's S02 Abatement Plan.This paper reviews matte tap hole development since 1978; outlines the prototype matte tap hole design, refractory selection and installation; and details the performance of the prototype tap hole during the first eight campaigns."

Jan 1, 1995

By P. K. Chatterjee

"The operations of a large-scale open-cut mine are simulated to produce a user-oriented interactive digital model. SIMULA is chosen as the programming language and a sophisticated data base system is designed to consider the complex information network of large mines. The main features of the program are the mine planning and mine evaluation modules. These modules are responsible for the computation of truck-shovel performance and other equipment details in a mine. Stochastic methods are employed to generate event times. The use of a Data Base Management System and the data base provide an efficient method of information storage and retrieval. Truck allocation techniques have been examined in detail and comparisons between random dispatching and conventional methods are made. The model is intended for mine operators and engineers and the program is well structured and very readable. It is hoped that these features will make the model readily acceptable.IntroductionThe Department of Mining and Metallurgical Engineering at the University of Queensland has been conducting research into computer simulation methods in mining since 1972. The research has been mainly applied to actual operations, with both open-cut and underground mining being equally emphasized. The results of the research to date have not only strengthened the methodology of the approach to complex mining problems, but have also increased confidence in the use of modelling and computer simulation technique s in mining.The model described in this paper translates a complex LHD situation into a user-oriented interactive digital model using SIMULA. The present methods of evaluating truck dispatching and trolley assist are tedious and laborious. This model provide s the engineer with readily acceptable means of rapidly assessing the relative merits of these techniques in a large-scale open-pit operation. Poor utilization figures in many current operations (for example, 40% for trucks, 70% for shovels) are of serious concern to mine officials and computer-controlled dynamic truck dispatching systems have been considered by some of these mines to improve equipment utilization ."

Jan 1, 1981

By T. Yalcin

"This paper describes the development of a process flowsheet designed to provide high dewatering rates and to bring about savings in tailing disposal costs. The flowsheet comprises a hydrocyclone, a settling cone and a thickener arranged in closed circuit. The mill tailing, after pulp density adjustment, is fed to the hydrocyclone which yields two products. The underflow product is sent to the settling cone for further dewatering. The overflow product, together with the settling cone overflow, goes to the thickener which, with the aid of flocculants, gives a clear effluent and a relatively dilute underflow. The latter combines with the original mill tailing to form the hydrocyclone feed. Solids finally leave the system in the settling cone underflow which can reach pulp densities of up to 78%, making it suitable for direct disposal or for use as mine backfill.Some discussion is included for the selection and sizing of equipment required for this process. INTRODUCTIONDewatering is an essential process for the disposal of mill tailings which forms an added burden in milling practice, often with little or no return. It is, therefore, desirable to conduct this process as economically as possible. One way of reducing costs would be to increase the rate of dewatering so that savings can be made in the size of dewatering unit s which would lead to lower capital and operating expenditures.A common method of dewatering mill tailings is gravity sedimentation which is frequently carried out with the aid of flocculation. In most cases, flocculation greatly enhances the sedimentation process, but only until a pulp density of 50% to 55% is reached in the settled material.Centrifugal sedimentation in hydrocyclones can provide high pulp densities rapidly, but does not result in a clear effluent.This paper outlines a suitable combination of gravity and centrifugal sedimentations that results in an over-all high rate of dewatering. The system developed produces a material of high pulp density while at the same time giving clear effluent."

Jan 1, 1989

By E. Kouame, M. D. Everell, D. Hodouin, T. Kasongo

"To properly operate a mineral beneficiation plant it is often necessary to have systematic evaluations of the performances of a piece of machinery, of one of the processes or of the plant as a whole. The state of a process can be described from a collection of data: flow-rate values, particle-size distributions and chemical compositions of the various streams of the circuit studied.However, the obtained raw data are generally redundant and inconsistent from a material balance viewpoint, due to natural disturbances, sampling errors, unreliable instrument readouts and laboratory analysis inaccuracies.BILMAT is a computer program which upgrades mineral processing data, adjusting them in a least-squares sense in order that they may verify material conservation constraints. As a consequence, the program is able to calculate the best values of unmeasured flow rates (such as. circulating load ratios). BILMAT is a very general program in the sense that it can accept various data from any type of mineral processing circuit, and its usefulness is demonstrated here for three processing circuits of a pyrochlore concentration plant: grinding, desliming and flotation.IntroductionComputation of the material balance around any metallurgical process, that is the determination of stream flow rates and composition. is a basic step in its evaluation. In spite of the simplicity of the physical laws of mass conservation. and of the very fundamental need to know the flow of components around a process. it appears, particularly in mineral processing. that material balancing is quite often a difficult task. The reasons for this involve the problems associated with collecting measurements in mineral dressing plants. They are generally poorly instrumented and invariably data must be collected through hand sampling. Due to the heterogeneous nature of the streams, segregation occurs and samples are often not adequately representative, a problem made worse by the lack of circuit sampling facilities, which are often completely ignored by plant designers. Furthermore, the nature of the measurements to be performed (for instance particle size distributions in the fine range) frequently gives inaccurate results."

Jan 1, 1981

By Dale Scott

"The Aggregate Resources Inventory Program was initiated 10 produce reports which assess the aggregate resources of municipalities designated under the Pits and Quarries Control ACI, 1971.Each report consists of three copies accompanied by a text. The text describes the local mineral aggregate resources, extractive activity and characteristics of the ""Selected Deposits"". The first map shows the known sand and gravel resources and their quality. The second map indicates selected sand and gravel resources of primary, secondary and tertiary significance. The third map indicates both bedrock resources and bedrock areas suitable f or extraction. These maps and the text are prepared in a formal suited for use by planning personnel.Recently, the Aggregate Resources Inventory Program has undergone evolutionary changes. New sources of data are being utilized in the inventory reports; in particular, detailed field assessment by inspecting, sampling, drilling and backhoe pitting. Liaison and joint assessment with district and regional Mineral Resources staff of the Ministry of Natural Resources, and personnel of the Ministry of Transportation and Communications, are key components of the Program.Twenty-two reports were available to the public as of December 1980, with a further thirty-six awaiting publication. Our present production rate is 3.5 reports per month. The reports appear to be well accepted by the municipalities and have proven effective in conserving prime aggregate resource areas.IntroductionThe urbanization of Southern Ontario has resulted in a major increase in the production of industrial minerals, such as sand, gravel, clay, shale, gypsum and stone, for use as cru shed aggregate, building stone, lime and cement.Although mineral aggregate deposits are plentiful in Southern Ontario, they are fixed-location, non-renewable resources. The potential for extractive development is usually greatest in urban fringe areas where land-use competition is extreme. Comprehensive planning and resource management strategies are required to make the best use of available resources, especially in those areas experiencing rapid development. Such strategies must be based on a sound knowledge of the total mineral aggregate resource base at both local and regional levels."

Jan 1, 1982

By Dale Scott

"The Aggregate Resources Inventory Program was initiated 10 produce reports which assess the aggregate resources of municipalities designated under the Pits and Quarries Control ACI, 1971.Each report consists of three copies accompanied by a text. The text describes the local mineral aggregate resources, extractive activity and characteristics of the ""Selected Deposits"". The first map shows the known sand and gravel resources and their quality. The second map indicates selected sand and gravel resources of primary, secondary and tertiary significance. The third map indicates both bedrock resources and bedrock areas suitable f or extraction. These maps and the text are prepared in a formal suited for use by planning personnel.Recently, the Aggregate Resources Inventory Program has undergone evolutionary changes. New sources of data are being utilized in the inventory reports; in particular, detailed field assessment by inspecting, sampling, drilling and backhoe pitting. Liaison and joint assessment with district and regional Mineral Resources staff of the Ministry of Natural Resources, and personnel of the Ministry of Transportation and Communications, are key components of the Program.Twenty-two reports were available to the public as of December 1980, with a further thirty-six awaiting publication. Our present production rate is 3.5 reports per month. The reports appear to be well accepted by the municipalities and have proven effective in conserving prime aggregate resource areas.IntroductionThe urbanization of Southern Ontario has resulted in a major increase in the production of industrial minerals, such as sand, gravel, clay, shale, gypsum and stone, for use as cru shed aggregate, building stone, lime and cement.Although mineral aggregate deposits are plentiful in Southern Ontario, they are fixed-location, non-renewable resources. The potential for extractive development is usually greatest in urban fringe areas where land-use competition is extreme. Comprehensive planning and resource management strategies are required to make the best use of available resources, especially in those areas experiencing rapid development. Such strategies must be based on a sound knowledge of the total mineral aggregate resource base at both local and regional levels."

Jan 1, 1982

By W. T. Robson

ONE of the problems encountered in any mine is the effect of increasing pressure manifested by increasing depth. Although the depth at which excessive pressure assumes serious proportion, and the manner in which it appears in underground workings, are influenced by a variety of factors, nevertheless the difficulty is present in all mines and its gravity increases with increasing depth. In the Kirkland Lake camp, as far back as seven years ago, mining had proceeded to depths sufficient to result in the formation of pressure great enough to cause violent disruption of masses of rock. It is proposed in this paper to deal with the effect of pressure caused by mining as observed at Lake Shore Mines, and to describe some of the methods which have been adopted to counteract it. There are many contributing factors in a study of the problem, and results vary widely in individual mines and in different localities. One feature common to all, however, is that there is ground subsidence, evidenced either by a gradual settling accompanied by crushing and shearing of the ore and wall-rocks, or the disruption of the ground surrounding the periphery of an excavation may take place suddenly, attended by more or less violent shattering of the rocks. Such violent failure of a mass of rock has been designated by the term rock-burst. The history of rock-bursts extends over a great number of years. They have occurred in some form or other in practically ali places where mining operations, at all but very shallow depths, have been carried out. Since the theoretical aspects of the subject have been dealt 'with at great length by capable and experienced writers (1), the present paper will give only a brief résumé of the theories propounded by those engaged in deep mining operations in various parts of the world.

Jan 1, 1940

By T. Möller

Grinding mills are the heart of the mineral processing plant, and are production-critical machines which are required to achieve high availability. Liner wear life and a liner design that promotes ease of installation are important variables to achieve a high level of mill availability. Increased mill size results in more mill internal surface to protect, and more liner components to install and remove. This drives designers toward an increase in liner size to minimize the number of components to be installed by the liner handler. The liner attachment method can also have a large impact on mill availability as well as safety during installation. This paper outlines the experience from the development, installation and follow-up of liners installed in two 38' x 45' AG Mills at Boliden Aitik in Sweden. A new Poly-Met shell liner concept has been developed to minimize the number of liner components. The new Poly-Met shell liner components are made much larger than conventional Poly-Met or steel liners. This reduces the number of components to be installed/removed by the liner handler and helps to minimize installation time and consequently down- time. A new attachment system has also been developed to eliminate the need for the installation crew to be in the ?drop-zone? of the shell liners while installing the liner bolts. One interesting aspect of this project was the requirement to predict liner life for these new large format AG mills. The existing 22' diameter mills at Aitik provided a base-line reference for this purpose, and DEM studies were used in the project development stage to predict the wear life of shell liners for the new 38' diameter AG mills. The results from this study are included.

Jan 1, 2012

By R. A. McGuire

THE recorded production of the Sheep Creek camp from 1900, when the Yellowstone Mining Company first put its mill into operation, to the present day is in excess of $17,500,000. Of this total, $2,500,000 can be attributed to the early period from 1900 to 1916, at which time increased costs of labour and material due to the Great War forced abandonment of any operations on a large scale for a number of years. Since the camp was revived in 1929, at which time the Reno mine started large-scale operations, and the early 'thirties', when the increased price of gold induced the commencement of operations on other properties in the district, total gross production has been in excess of $15,000,000. In addition to the above totals, intermittent shipments were made by lessees and small operators, principally during the period of inactivity from 1916 to 1929, but no records of this production are available. The Sheep Creek camp would rank fifth in the Province-behind Rossland, Portland Canal, Bridge River, and Hedley-in gross gold production, but, because of relatively inexpensive mining and milling costs, would occupy a more prominent position in regard to total dividends paid. Operating mines at the present time include Reno Gold Mines, Gold Belt Mining Company, Kootenay Belle Gold Mines, and Sheep Creek Gold Mines. Each one of these companies, with the exception of Kootenay Belle, represents a consolidation of several smaller properties, all of which were in existence for a number of years prior to 1929. These four companies, at the present time, own, or hold options on, practically all the important claims in the district with the exception of a very few claims and fractions still held by private owners.

Jan 1, 1942