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By G. Harper

"AbstractThe Suetut. Copper deposit, 230 miles northwest of Prince George, B.C., was discovered in August of 1971. Subsequent exploration has included 56,417 feet of surface diamond drilling in 139 holes. The deposit occur in gently dipping volcanic rocks of the Upper Triassic Takla Group, which out crops as rugged mountains.The general Mesozoic sequence comprises some 20,000 feet of green and red basaltic to andesitic rocks. The Lower Formation is characterized by subaqueous basalt, andesite and augite porphyry flows with interbedded breccias. They were rapidly deposited in tectonic depressions of limited extent and steep gradients. Facies changes are abundant. The Middle Formation. consists of volcaniclastic rocks made up of fragments from the underlying formation. Two cycles are discernible , both with basal turbiditic, tuffaceous sandstone and shale units overlain by massive, volcanic conglomerates exhibiting graded and cross bedding. A broader, shallower-water basin is indicated. The top of the uppermost unit was subaerially deposited. The Upper Formation is a variable sequence of volcanic sedimentary rocks. Detritus was supplied from a wide source area and included multiple reworked fragments of the lower formations and material from other rock groups known in the region. Zeolite facies metamorphism is recognized throughout: the entire sequence. Epidote is particularly abundant in the Middle Formation, demonstrating the results of metasomatism.The mineralized zone is approximately concordant with the strata over 100-200 feet of the Middle Formation succession . It lies just below the zone of transition from subaqueous to subaerial deposition in the top unit. Hematite, chalcocite, pyrite, bornite, chalcopyrite and native copper are disseminated throughout all sizes of lithic fragments and matrix. Concentration into the specific zone is believed to be due to upward leaching of copper during lowgrade metamorphic and metasomatic reaction. Precipitation occurred in the zone of hydration characterized by prehnite-pumpellyite below the oxidized subaerial rocks."

Jan 1, 1977

By J. E. Nesset, S. R. Rao, J. A. Finch

"For base metal sulphides, iron rejection starts in mineral processing. This review focuses on changes in plant practice specifically to improve iron sulphide rejection by control of contaminant ion effects and discusses selected general innovations in mineral processing that can be expected to play a role. One group of changes to control contaminant ions is in size reduction (the use of stage grinding and inert stirred milling) and another is in system chemistry (order of reagent addition and exploitation of the flexibility of sulphoxy reagents). Plant examples illustrate the successes, indicating that there are several options available. The general innovations start by considering the new tools for measuring gas dispersion properties and quantifying froth characteristics by imaging. Examples of their application to improve concentrate quality illustrate the potential. Other innovations stress adapting geostatistics to model variation in milling and flotation response to provide feed forward control and rigorous application of statistical experimental methodology in plant test work. An example from one concentrator shows improvements in concentrate quality from application of basic principles can be impressive.INTRODUCTIONThe more iron control (rejection) that can be accomplished in mineral processing the lower the subsequent metal extraction costs and environmental implications of iron disposal, particularly in hydrometallurgical processes. Iron rejection has been a contributing target to periodic attempts to apply mineral processing to nickel laterites and bauxites but the prime application is control of pyrite and pyrrhotite (iron sulphides) in the treatment of sulphide ores by flotation. In this paper some new concepts and innovations introduced in sulphide flotation plants over the past ten years or so (i.e., since a previous review [1]) will be outlined with examples of how they have contributed to improved “iron control”.To increase rejection of iron minerals, first the recovery mechanism should be understood. There are four principal ones, two physical in origin: 1, via locked particles (i.e., flotation of composite particles due to insufficient size reduction to liberate the minerals) and 2, through entrainment in the water reporting to the float product (an unselective process); and two originating from a response to the system chemistry: 1, true flotation (i.e., iron sulphides becoming hydrophobic) and 2, as a result of iron sulphides forming aggregates with hydrophobic particles (e.g., ”slime” coatings). There has been considerable effort to diagnose which of these dominates iron sulphide recovery; in many cases it is due to “chemistry”, in particular true flotation. The flotation effect often appears related to contamination by metal ions [1]. This can reduce selectivity in two ways: 1, metal ion species may activate iron sulphides, or 2, cause general depression which demands more vigorous flotation conditions (e.g., additional collector) with consequent loss of selectivity. This is where the review will start, by considering ways to control contaminant ions."

Jan 1, 2007

By G. R. Rawsthorne

"The Con-Rycon mine is located on the north shore of Great Slave Lake in the Northwest Territories within the present city limits of Yellow-knife. Development work on the Con property started in 1936, resulting in a mine-mill complex which produced its first gold brick on September 5, 1938. Since that date 5,000,000 tons of ore have been mined yielding 3,000,000 ounces of gold. At present there are 220 employees at the mine.MineThe mine has 1.22 miles of vertical shaft, 41 miles of level development and 10 miles of inclined raises. There has been 1,000,000 feet of diamond drilling in search of ore. The temperature at the 4900 level (4400 feet below sea level) is 68 deg. F.The present mining method is cut and fill. The majority of the ore mined today passed through an internal winze before reaching the main shaft. The long haulage and double handling will be eliminated with the completion of a new shaft presently under construction to the 5800 level.MineralizationThe ore treated today is from the Campbell shear 1vhich is characterized by lenses and veinlets of coarse textured quartz containing some carbonate, sulphides and finely disseminated as well as visible gold. At the lower levels of the mine the arsenopyrite content in the ore is de-creasing with a~ increase in the levels of chalcopyrite and pyrrhotite."

Jan 1, 1974

By Ben Murphy, Tomi Maksimainen, Antti Rinne, Ishmael Muzinda, Toni Mattsson

"A 500m3 flotation cell was commissioned during 2014 at the Kevitsa Mine in northern Finland working as the first rougher in the copper flotation circuit. Treating around 850tph with an average grade of 0.3% Cu, the TankCell® e500 has run with an average copper recovery of 71% and an average enrichment ratio of 50 during early optimization test work. In addition the cell has proved to be very energy efficient, being operating with a specific energy of around 0.4kW/m3 under production conditions.A review of the installation and operation of the TankCell® e500 at Kevitsa will be presented. Metallurgical performance of the cell during the operating period will be discussed on the basis of on-line analysis and the metallurgical samples collected during early optimization campaigns.IntroductionThe past three decades have seen a rapid increase in the size of mechanical flotation cells. This has been largely driven by demand from the copper industry to process larger volumes of material due to combination of declining head grades and the desire to benefit from the economies of scale (Lynch, Harbort and Nelson 2010). During this period Outotec has been at the forefront of development of large flotation technology with an enviable scale-up track record; 100m3 (1995), 160m3 (1997), 200m3 (2002), 300m3 (2007), 500m3 (2012) to the largest commercially available cell at 630m3 in 2013. This scale-up methodology has been developed, tested and validated (Elgueta et al., 2009; Coleman & Dixon, 2010) by the large installation base of cells in the 300m3 size class operating (over 400). The learning from large cell operation are then reviewed and incorporated back into new designs and scale-up practices (Murphy, Miettinen, and Yanez, 2014)."

Jan 1, 2016

By A. E. Flynn

Introduction The purpose of this paper is to record the results of researches on anhydrite that have been undertaken at the Nova Scotia Technical College during the last two or three years. The work was carried out in the Mining Department of the College, and several investigators were engaged on the work from time to time. Messrs. R. E. Hanson and K. M. Ed devoted their mining theses to the investigation, and Messrs. J. C. Hall and G. W. Baker were engaged on the work during the summers of 1928 and 1929, respectively. The writer was responsible for the initiation and direction of these researches, which were made possible by the financial assistance of the Department of Mines of the Province of Nova Scotia. Nova Scotia is the leading Province in Canada for gypsum production, its output being about a million tons per year, or over 80 percent of the production for the whole of Canada. The commanding position of Nova Scotia in this regard is due to the abundant supply of gypsum and its proximity to areas of intense utilization in the eastern United States. The resources of gypsum are contained in a total area of 627 square miles of gypsiferous rocks, so that an enormous reserve of this mineral is available. Present operations are confined to the highest grade and most easily worked deposits close to water transportation.

Jan 1, 1930

By J. Peterson, D. D. Tannant, J. Todd

"A review of blast designs and improved blasting practices at the EkatiTM diamond mine is presented along with the results of three blast-monitoring experiments. Blast monitoring was undertaken to investigate blast damage mechanisms in the mine’s well jointed rock mass. Rock mass damage caused by production, pre-shear and wall control blasts was measured. EkatiTM uses 270 mm holes for production blasting, and 165 mm holes loaded with a decoupled charge for pre-shearing. The production blasts are loaded with bulk emulsion/ANFO blends. The mine plan involves using 30 m double benches. A 30 m high pre-shear is drilled and blasted prior to the production blasting that is done with sequential 15 m benches. This paper summarizes the monitoring equipment and data gathered to date. IntroductionThe EkatiTM diamond mine is located about 300 km northeast of Yellowknife, Northwest Territories, Canada. The mine is a joint venture of BHP Diamonds Inc. (80%) and geologists Charles Fipke and Stewart Blusson (10% each). The EkatiTM mine is Canada’s first diamond mine and BHP is the operator. The mine is accessed by air and by a winter ice road.The first kimberlite pipe mined at EkatiTM is known as the Panda pipe. The Panda open pit mine will have a total depth of 315 m (Fig. 1), a width of about 600 m, and a 50° overall slope. Initial ore production is 9000 tonnes per day.This paper summarizes the pre-strip blasting and the evolution of the production and wall control blast designs from 1997 to 2000. The initial blast designs and the rationale behind subsequent changes that took place are presented along with the techniques currently in use.The paper also briefly presents an experiment done to measure the rock mass response to blasting. Three blasts were instrumented with geophones, gas pressure sensors and time domain reflectometry cables. A brief summary of the instrumentation and data gathered is included."

Jan 1, 2001

By A. R. Sorta, D. C. Sego

ABSTRACT Temperature, solute, texture, and thixotropic effects on time domain reflectometry (TDR) measurements of water content in oil sands tailings are investigated. Results indicate that TDR water content measurements are influenced by temperature, residual bitumen, and percent clay in the tailings. A correction of 0.03–0.10/ °C must be applied to the apparent dielectric constant (Ka) to account for the effect of temperature. The Ka-volumetric water content relationships of slurries differ significantly from the commonly used Topp calibration equation. The age of the tailings and the addition of phosphogypsum did not influence TDR water content measurements.

Jan 1, 2013

By A. K. Ratte, C. M. Brown

"IntroductionIn autogenous mills, grinding of the ore feed is accomplished by the direct action of the variously sized rocks making up the feed on each other and the mill interior. Size reduction is achieved without the addition of grinding media foreign to the mill feed, differing in this respect from rod, ball, and pebble milling. The term ""semi-autogenous"" has entered common usage in the industry to describe autogenous mills to which supplementary grinding ball additions are made to increase the rate at which size reduction of the mill feed occurs, beyond that achieved solely by the action of rock on rock. In a semi-autogenous grinding (SAG) mill, 50Jo to 10% of the mill volume may be occupied by grinding balls, whereas a range of 25% to 45% is more usual in ball milling.The comminution of mineral ores using ""conventional"" techniques has been described in previous papers. In conventional operations, size reduction of mined ore, and liberation of the valuable mineral content, is typically accomplished using flowsheets employing two to four stages of crushing. Crushing is followed by one to three stages of grinding and classification using various combinations of rod, ball, and pebble mills. This paper describes operations employing autogenous or semiautogenous grinding as the principal stage in the comminution process. Primary crusher discharge, or in some instances run of mine ore, is fed directly to an autogenous or semi-autogenous grinding circuit. Final product sizing may be accomplished in a single grinding stage, the mills operating in closed circuit with classification equipment, or in several stages using autogenous and semi-autogenous mills for primary grinding, and ball or pebble mills for secondary grinding. In several operations a portion of the primary mill discharge is crushed in an intermediate comminution step, these circuits being called autogenousball- crusher circuits (ABC). Many different flowsheets exist, but common and important to most are crushed ore stockpiles, reclaim feeder and mill feed conveyor systems, pumps, screens, and classifiers. Autogenous grinding is employed in both wet and dry processes, although today wet predominate."

Jan 1, 1989

By Forrest A. Kerr

Northern British Columbia is a term which probably conveys to various individuals quite different meanings. To the average person it signifies an area of indefinite size and indefinite position which may start at Prince Rupert, Stewart, or somewhere else, and end at Alaska or .the Yukon. Many people, in their conception, fail to encompass in their mind's eye that vast tract of land lying ~o the east of the panhandle of Alaska. Few realize that the area north of an east-west line through Stewart represents more than one-third of the Province and is, of course, the part most appropriately designated as northern British Columbia. This section, more especially the northernmost 100,000 square miles, from political, geographic, and economic viewpoints stands isolated perhaps more than any other part of British Columbia. Unlike the rest of the Province, it has no coast line, being cut off from the Pacific ocean by a narrow strip of Alaska. It is bounded on the north by the Yukon, and on the east by Alberta. In the south this northernmost 100,000 square miles is limited by a divide of rugged mountainous country separating the drainage basins of the Stikine and Liard rivers from those to the south.

Jan 1, 1929

By Philippe Poirier

Les parcs à résidus miniers représentent un risque environnemental qui nécessite une gestion adéquate. Suite à la publication du « Guide sur la gestion des parcs à résidus miniers » en septembre 1998, les membres de l?industrie minière ont appliqué les éléments de gestion du Guide, permettant ainsi une amélioration continue des opérations. En février 2000, l?industrie minière a indiqué le besoin d?établir des lignes directives plus spécifiques pour l?opération sécuritaire des parcs à résidus. Dans cette optique, l?Association Minière Canadienne a publié, en septembre 2002, un Guide pour l?exploitation, l?entretien et la surveillance (EES) des parcs à résidus miniers. Suite au développement du Guide et à son application sur plusieurs sites miniers, il apparaît que la mise en oeuvre d?un Manuel ESS des parcs à résidus miniers permet : ? d?établir les responsabilités des intervenants; ? de s?assurer que les critères d?opération sont appliqués; ? de mettre en oeuvre le plan de déposition des résidus et de gestion de l?eau; ? de rencontrer les normes et exigences environnementales; ? de définir un programme d?entretien et de surveillance des ouvrages; ? d?identifier les risques et de mettre en oeuvre un plan d?intervention d?urgence; de démontrer une opération diligente.

Nov 1, 2002

By R. D. Westervelt

IN JULY 1954, the first iron ore from the deposits in the Knob Lake vicinity was mined and shipped to Seven Islands. The beginning of the production may well be the greatest advance in Canadian economy in the past fifty years. In 1955, the Canadian iron ore output was more than twice that of 1954, an in-crease due principally to the iron ore shipments from the Knob Lake area. Previous to this, Canada's iron production was 1.85 per cent of the world's output. Now, this percentage has risen to approximately 5.5 per cent and shows every indication of increasing further as the production rate from the area is increased. In 1955 the Iron Ore Company of Canada shipped over 7 million tons of iron ore and expects to ship over 10 million tons in 1956. Present facilities may be enlarged to handle up to 20 million tons per year. If the expected I 0 million tons per year are exported, Canada's iron exports will be second only to her wheat and newsprint exports. Before production began from the Knob Lake deposits, years of development were necessary in the area. Geologists spent several summers mapping the general area in order to unravel the complicated stratigraphy and petrology. Particular attention was paid to the mapping of the iron formation and all stages of alteration from fresh iron formation to ore were observed.

Jan 1, 1957

By John D. Campbell

Alberta Plains coals exhibit an almost rectilinear system of vertical fracturing called "cleat", which is best traced by mapping true bearings of underground coal mine rooms. The cleat pattern is approximately conformable with a pattern of surficial features and a pattern of bedrock jointing in the Alberta Plains, and with the trend of the Rocky Mountain Trench, but not with the distribution pattern of coal rank nor with the trend of the eastern boundary of the Rocky Mountains and Foothills physiographic province. The processes resulting in cleat remain in doubt, but cleat is always of practical importance in extracting coal from deeply buried deposits.

Jan 1, 1979

By Y. Özçelik

Rocks show a variety of engineering and physical properties that affect typical quarry operation such as diamond wire cutting in the marble quarry. The physical and mechanical properties of rocks depend on their mineralogical and petrographical characteristics. The purpose of this study is to investigate the effects of mineralogical and petrographical properties of a rock, such as grain size and mineral content, and how these affect the wearing of diamond bead and cutting rate in a diamond wire cutting operation in a marble quarry. Statistical analysis is used to carry out the study. Three different rock types were examined, namely, hard rock (andesite) and softer rocks like limestone and marble. The study revealed that the mineralogical and petrographical properties of the rocks, such as percentage of quartz, biotite, plagioclase, matrix minerals, and opaque minerals, are important. In addition, the grain size of quartz and plagioclase in andesitic rocks and calcite grain size for carbonate rocks significantly influenced the wear rate and cutting rate in diamond wire cutting operations.

Jan 1, 2005

By Fidelis T. Suorineni

The Ashanti mine, located in Obusai, Ghana, West Africa, has been in production for over 100 years, and is currently the second largest single gold producer in Africa. The mine covers a concession area of 434 square kilometers with four main ore types: the quartz vein type, disseminated sulphide type, supergene sulphides and transition ore, and granitoids stock works (visible gold and quartz veinlets). Various mining strategies and methods have been employed for the mining of the narrow-vein ore at Ashanti. In recent years, the mine owner: Ashanti Goldfields Company Ltd. has embarked on a mine-wide expansion program to boost production to a target of one million ounces per annum. With the rapid depletion of its high-grade ore, a suitable mining method was required to produce a high tonnage at low cost to meet production objectives. This paper discusses how the Ashanti mine applied Canada experiences in mining technology to boost production, and the impact that the differing geological settings between the two countries had on the stope performance at the Ashanti operations. Experience developed at the Ashanti mine led to innovative modifications for stope layout and design methods in relatively poor ground conditions.

Oct 1, 2001

By J. Goyman, K. R. Rawling

"1.1 Location and Transportation ServicesNanisivik Mines Ltd. is located at latitude 73°, 3,200 km north of Montreal and 700 km north of the Arctic circle, on the south side of Strathcona Sound, which is an arm of Admiralty Inlet, in Northern Baffin Island. The orebody and mill site are 4 km distant from Strathcona Sound which provides deep water anchorage for ocean transportation of concentrate to Europe during the short northern shipping season from mid-July to mid-October.The operations and the resident community have direct jet aircraft service from Montreal two days per week, and local nights connect with other Arctic communities in the northern Baffin region. Bulk supplies and general cargo are delivered in the summer shipping season.1.2 HistoryThe Texas Gulf Sulphur Company, now Kidd Creek Mines Limited, staked claims in the Nanisivik area in 1957 and carried out exploration through to 1968. After the Kidd Creek orebody was found, Texas Gulf optioned the property in 1972 to Mineral Resources International Inc. (MRI), which undertook further exploration and completed a feasibility study.Nanisivik Mines Ltd. was formed in 1975, with MRI as the majority shareholder and minority shareholdings held by Metallgesellschaft, Billiton and Texas Gulf. Strathcona Mineral Services Limited was retained to manage the project during both the construction and operational phases."

Jan 1, 1984

By Earle J. Klohn

"The design, construction and operation of tailings dams is rapidly becoming a major consideration for most new mining developments, as well as for many existing operations. In North America, under pressure from public opinion, various government agencies have in recent years enacted legislation covering the design, construction and operation of tailings dams. In British Columbia, comprehensive legislation has been passed during the past year (1971). All of these new regulations may be placed into two broad categories: regulations concerning the safety of the dam and appurtenant structures, and regulations concerning the prevention of pollution. Both types of regulation affect the design of the tailings dam facilities. Consequently, they must be considered jointly by the designer. A good tailings dam design must satisfy the basic requirements of safety, pollution control, storage capacity and economy. To achieve this, 'the design must be based on a thorough understanding of both the geotechnical problems involved and the requirements of the mining development. Obviously, this requires close cooperation between the designing geotechnical engineers and the mining operators. This paper outlines the major engineering factors that should be considered in the investigation, design, construct ion and operation of tailings dams. Basic methods of tailings dam construction are reviewed and the advantages and disadvantages of each briefly discussed. Comparisons are made between the design and construction of conventional water supply dams and tailings dams. Examples are presented of some of the problems that may arise during construction and operation of a tailings dam. The paper also explains the need for basic instrumentation of tailings dams and for carrying out simple field tests and measurements during construction and operation. Of prime importance is the downstream monitoring of surface and subsurface seepage flows. These observations are invaluable where contamination of surface and/or subsurface water may become a contentious issue at some future date. Each tailings dam design must be based on a study of the particular requirements of that project. To make a proper evaluation, adequate data must be available concerning a large number of items•, including such things as: topography, hydrology, geology, subsoil properties, seismicity, government regulations, fineness of mill grind, mill water balance, availability of suitable construction materials, volume of water to be stored behind starter dam, downstream development or liability, and many others. The large number of variables involved in such an evaluation often result in the designer considering two or more alternative designs. Selection of the final design, which is usually based on detailed cost analyses of the alternatives, must be made in close consultation with the owner."

Jan 1, 1972

By Mika Muinonen

"Dewatering of nickel concentrate at the INCO Thompson integrated mine-metallurgical site takes place within the smelter plant. The eight original 60 foot diameter cable torque thickeners of which four were operational and eight Dorrco internal drum filters have been replaced with a single Outokumpu high rate thickener and a single Lasta pressure filter. The modernization has resulted in a filter cake which has a more stable and lower moisture content. Control of the entire dewatering process has been improved and costs have been reduced.Typical filter cake produced from the Dorrco filters during processing of high tonnages of Birchtree mine ore was ~20% moisture vs. ~13% moisture for concentrate produced from the Thompson mine ore. The high moistures were correlated with the amount of fine magnesium silicate gangue in the concentrate. As Birchtree ore processing rates have increased, this filter cake moisture became unacceptable because of material handling problems and the impact on furnace metallurgy. The Dorrco filters were replaced with the Lasta filter in 2003 that has consistently produced filter cake at ~11% moisture.This paper summarizes the piloting, design, installation and commissioning of the modernized dewatering circuit.INTRODUCTIONThe Thompson integrated nickel processing facility was built in 1958 and originally consisted of a mine-mill-smelter and, soon after, a refinery.Nickel concentrate is pumped from the mill as a dilute slurry directly to the smelter facility ~200 meters along an unheated pipe trestle. Originally, nickel concentrate was de-watered using up to 8 conventional 60-foot diameter thickeners and 8 internal vacuum drum filters."

Jan 1, 2006

By Jan E. Nesset, James A. Finch, Frank Rosenblum, Rebecca Payant

"Under certain conditions of moisture and oxygen, sulphides can spontaneously heat, a phenomenon known as self-heating. In this paper, we test a hypothesis that galvanic interaction between some sulphides can promote self-heating. Galvanic interaction is controlled by restpotential difference between minerals and their surface area of contact (particle size). In order to decrease rest-potential, four sulphides were tested: pyrite, chalcopyrite, sphalerite and galena. Two series of samples consisting of mixtures of two sulphides, each representing a mass fraction of 50 %, were run in standard self-heating tests. The first series comprised five mixtures, two with low rest-potential difference (pyrite-chalcopyrite and chalcopyrite-sphalerite) and three with high rest-potential difference (pyrite-galena, chalcopyrite-galena and pyrite-sphalerite). The second series was performed on a pyrite-sphalerite mixture at four particle sizes (80 % passing 850 µm, 300 µm, 75 µm and 38 µm). The first series showed that the individual sulphides and the mixtures of low rest-potential difference did not self-heat but the mixtures of high rest-potential difference did self-heat. The second series showed that self-heating increased inversely with particle size (increased specific surface area) and revealed that it was the fineness of the pyrite (high rest-potential sulphide) that governed the self-heating effect, indicating the rate-limiting step is the reduction reaction on pyrite (cathodic mineral). The increase in self-heating with high rest-potential difference and increasing particle fineness supports the hypothesis that galvanic interaction contributes to sulphide self-heating. A possible mechanism based on the hydrogen sulphide hypothesis is proposed where the reduction reaction at the cathodic mineral surface is ferric to ferrous. The understanding gained will be of interest to those involved in storage, shipping and disposal of sulphide minerals.INTRODUCTIONSulphide mineral ores are the major source of most base metals such as copper, zinc, nickel and lead. Once the ore is extracted, the sulphide minerals are liberated by size reduction (crushing and grinding), concentrated by flotation, and the concentrates sent for further processing to extract the metal (smelting and refining). At the various stages from mining to concentrate shipment and tailings storage, the sulphides are exposed to a variety of environments, which under certain conditions of air (oxygen), humidity and temperature can cause them to heat. Since no external heat source is involved, this pyrophoric process is known as spontaneous heating or, more commonly, self-heating. The self-heating of sulphides is associated with oxidation reactions under moist conditions. If the heat generated is greater than the heat dissipated, the sulphides will self-heat and the material temperature will rise. This can make working conditions potentially dangerous as oxygen is consumed from the surrounding air and significant quantities of sulphur dioxide (SO2) can be released (Ninteman, 1978)."

Jan 1, 2011

By Y. Obara, A. Sato

"Limestone is one of the most important mineral resources, and in Japan it is only one mineral which can be fully supplied by the domestic product. Generally limestone mines in Japan located at mountainous regions, and the combination between bench-cut method and the carry-out system using vertical shaft is widely used in the domestic mines. During mining operations, the occurrence of the blockage in the vertical shafts is very serious problem. Once the blockage in the shafts is occurred, a lot of works and dangerous procedures are necessary to solve the problems. It is thought that blockage is influenced by the cohesion and friction coefficient due to clay minerals thrown into the vertical shafts, however, the quantitative and precise reason is not revealed yet.In this study, the process of blockage occurrence was simulated by 2-Dimensional Discrete Element Method (DEM). In the 2D-DEM analysis, discontinuous circular particles are generated, and the individual movement of each particle is calculated by forward difference method. In the analysis, several typical ore distributions are applied, and the DEM analysis was conducted under the various combinations of cohesion and friction coefficient. From analytical results, it was found that the friction coefficient strongly influences to the occurrence of blockage and there were a threshold values which blockage does not occur. Several kinds of shaft shape were also simulated towards different ore size distributions, and those influences to the blockage occurrence were also discussed."

Jan 1, 2015

By A. Becker, C. Gauvreau, L. S. Collett

"A scale model was designed to simulate the airborne Barringer INPUT time domain electromagnetic system. The amplitude of the response to a vertical conductive dyke varies considerably with the conductor's thickness conductivity product (TCP). The response increases from a minimum detectable level at a TCP of about 2 mhos, rises to its maximum value at a dyke TCP of 15 mhos and then falls again so that conductors with a TCP exceeding 500 mhos would probably remain undetected. Measurements made with both standard receiver coil configurations (maximum or minimum coupled) indicate comparable effectiveness for both systems, either from the viewpoint of vertical conductor detection or from the viewpoint of discrimination against overburden anomalies."

Jan 1, 1972