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By Yongsheng Song, Yong Chen, Wenjuan Li

"Detailed characterization and improved recovery of a micro-fine, low-grade refractory gold ore from a beneficiation plant in Yunnan Province, China, was investigated. Different analytical techniques, such as videlicet size analysis, scanning electron microscopy, and mineral analysis by a Mineral Liberation Analyser, were used for mineralogical characterization of the ore. This paper describes an experimental study on the effect of the grinding fineness and the use of various flotation reagent regimes involving collectors, depressors, and activators on the flotation behaviour of micro-fine auriferous pyrite. Two-stage flotation resulted in high gold grade (32.71 g/t) and gold recovery (80.02%) of the bulk gold concentrate. This process opens the possibilities of maximizing Au recoveries in a micro-fine low-grade refractory gold ore where low recoveries are common.INTRODUCTIONWith the rapid development of the gold industry, refractory gold ores are becoming increasingly important resources for gold production (O’Connor & Dunne, 1994; Agorhom, Skinner, & Zanin, 2012). Among these refractory gold ores, sulfur and micro-grained native gold ores represent a significant challenge, since they are not amenable to gold recovery by direct cyanidation. The gold contained in this type of ore is very finely disseminated and encapsulated in sulfides, which makes it inaccessible to the cyanide solution during direct cyanidation (Bulatovic, 1997; Allan & Woodcock, 2001). Autoclave pressure oxidation (AC POX) of the sulfide ore and subsequent cyanidation is a common technology used to recover the gold in such types of ore. Flotation is often used as a pre-treatment method to refractory gold ore (Monte Lins, & Oliveira, 1997).As the gold-carrier minerals are usually finely disseminated (sometimes under 10 µm) and intimately associated with gangue minerals, a fine grind is necessary to get a high degree of liberation. At the same time, the fine grind usually causes the production of significant quantities of slimes, which comprise easy-to-float gangue minerals such as mica, dolomite, and calcite. These mineral phases are naturally hydrophobic and float readily, competing with the value minerals in the froth phase and resulting in lower concentrate grades."

Jan 1, 2016

By K. Östcrdahl

We are developing a cost-saving and environmentally friendly process with low energy consumption to recover the metal content and recycle the acid from spent pickle baths. The metals are precipitated as metal fluorides, which can be turned into metal oxides and hydrofluoric acid by thermal hydrolysis. In the proposed process the solution is concentrated by nanofiltration and the metals are precipitated in a subsequent crystallization step. Different nanofiltration membranes have been evaluated in experimental batch tests. The membrane MPF-34 from Koch Membrane Systems offered both good flux and acceptable metal retention even after long periods of acid storage. The solubility of B-FeF3.3H20 has been measured at varying acid concentrations and at different temperatures. The solubility decreases with increasing hydrofluoric acid and nitric acid concentrations. This suggests that the crystallization process should be operated at high acid concentrations. The crystal growth rate of B-FeF3.3H2O is found to be quite low but increases significantly with temperature. This suggests that the crystallization process should be operated at higher temperatures (40 - 60°C). Within the project a pilot plant has been constructed where the spent pickle bath solution is concentrated by nanofiltration to the desired metal fluoride supersaturation, and B-FeF3.3H2O is crystallized in a 10 L continuous crystallizer.

Jan 1, 2006

By G. M. Dunn

Pressure leach autoclave circuits are employed in the leaching of ores, concentrates, mattes, alloys and intermediates for the recovery of metals into solution. Once the metals are extracted into solution the value metals can be recovered by hydrometallurgical means such as by purification followed by electrowinning, hydrogen reduction, pyrohydrolysis, crystallisation and other unit operations. In many of these integrated flowsheets the pressure leach step is pivotal to the recovery of the value metals from the host material. In an expansion or a new project, a large component of the capital cost can be associated with the pressure leach step and process engineers frequently look for opportunities to mitigate costs. This paper examines a new route for increasing the capacity of an exothermic pressure oxidation leach for sulphide concentrates, sulphide and phosphide mattes and alloys.

Jan 1, 2004

By R. Drozd

"The Bui/moose Mine is in the northeast coal area of British Columbia, and less than $300 million was spent to bring this 2.3 million-tonnes-per-year metallurgical and thermal coal mine into production by November 1, 1983, one month ahead of schedule. The project is an unincorporated joint venture undertaking, with Teck-Bullmoose Coal Inc., a wholly-owned subsidiary of Teck Corporation, holding 51%, Lornex Mining Corporation holding 39%, and Nissho-lwai (Canada) Ltd. holding 10%. Bullmoose Operating Corporation will be the manager and operator of the project. This paper will broadly cover the history and exploration work on the property, the geology, open-pit mining plans and equipment, the preparation plant flowsheet from breaker station to clean coal silo, the 35 km highway truck haul, the railroad loadout system, manpower requirements, approvals and permits, and general financial arrangements. Introduction The Bullmoose Mine Project is located about 87 km by road south of Chetwynd, B.C. and consists of an open-pit coal mine, a plant facility in the Bullmoose Creek valley below the mine, and a separate rail loadout facility on the British Columbia Railway Branchline near the new town of Tumbler Ridge some 35 km by road to the east. The plant will be capable of producing a total of 2.3 million tonnes per year (mty) of washed metallurgical and thermal coal. Long-term metallurgical coal contracts for high-quality, medium-volatile, bituminous coal have been signed with nine Japanese buyers for 1.7 mty, and the balance of0.6 mty will be for thermal coal production. Plant construction began in the spring of 1982, and in May 1983, preproduction waste and overburden removal in the mine was started. The plant began tuning up in October and was fully operational by December 1983. Metallurgical coal began moving to the coast in unit trains during November, and first vessel loading at the Ridley Island port began in late December 1983."

Jan 1, 1985

By Gui Hua Wang, Xin Chang, Zhong Ying Han, Yuan Fang Cheng

"According to the geometrical characteristics of triangular meshes, Triangular Mesh Split Model (TMSM) based on a finite element method is proposed to simulate cracks propagation of fractured rock. With TMSM, stress intensity factors at the tip of cracks are obtained by far-field contour integral. Meanwhile, vector cross multiplication is adopted to fix the cracking unit, and maximum circumferential stress criterion is used to determine the crack propagation angle. Then crack dynamic propagation simulation is realized through local mesh split or node movement at crack tip. With this method, the crack can split apart a unit or extend along the mesh boundary, so the crack could extend along an arbitrary path and is not restricted by the initial mesh. The applicability of the mesh split algorithm is verified by the crack propagation examples of wing crack and multi-fractured rock mass with holes. The results show that this method not only has higher accuracy of fracture parameters calculation, but also can accurately simulate the crack initiation and propagation of rock under complex stress state such as tension, compression and shear. In addition, the far contour integral is introduced to determine the crack tip stress intensity factor and there is no need to reconstruct a singular element at the crack tip or to intensify the meshes. Compared with the existing mesh reconstruction algorithm, this method significantly improves the computational efficiency.INTRODUCTIONGenerally, rock is neither the full continuum, nor entirely discrete body, but rich in geological defects, such as joints and cracks. A large amount of geotechnical engineering practice shows that most fracture of rock is closely related to the internal joints, cracks and crack expansion (Horii et al., 1985; Wong et al., 1998). Although the classical theory of fracture mechanics established the theoretical foundation of crack initiation, propagation, so far it is still extremely difficult to characterize the propagation of airfoil-typed crack and the interference problem of multiple cracks. Up to now, the micromechanics mechanism research of crack propagation of fractured rocks has still been a hot research topic of rock fracture mechanics."

Jan 1, 2015

By Finley A. Campbell

THE Quemont and Horne mines are located at Noranda in north-western Quebec. As shown in Figure 1, the Noranda area lies between the Cadillac-Bouzan fault zone on the south and the Porcupine-Destor fault zone to the north. Both of these major structural lineaments appear to be geometrically, and possibly genetically, related to the Grenville Front. The regional geology has been discussed by Wilson (12), Cooke et al. ( 2), and others. The geology in the Noranda vicinity, as compiled by Dugas and Gilbert ( 3), is shown in Figure 2. The Cadillac-Bouzan fault zone separates the dominantly Temiskaming-type sediments in the south from the Keewatin-type canic rocks to the north. These volcanics, which vary in composition from rhyolite to andesite, form a broad anticlinorium with a general eastward plunge. A succession of dioritic stocks, sills and dikes have been mapped in the area. The principal granitic rock masses include the Powell, Flavrian and Dufault stocks.

Jan 1, 1962

By Robert B. Parsons

"On August 8, 1989, the Minister of Finance released a technical paper which sets out many of the details of a proposed 9% federal value-added tax to be called the ""Goods and Services Tax"". The GST, which is scheduled to be implemented with effect from January 1, 1991, will affect all Canadians in different ways and to varying degrees. The mining industry itself should, by and large, benefit from the new tax.IntroductionOn January 1, 1991, it is proposed that the existing federal sales tax will be replaced by a new sales tax - the Goods and Services Tax (GST). GST will be leviedat a rate of 9% on most goods and services consumed in Canada.How it will WorkThe GST will be a tax on the final consumption of goods and services. It differs from a retail sales tax in that it will be collected throughout the production and distribution chain. However, the net effect of the tax paid on a particular good or service will be equivalent. The example above (next column) illustrates the application of the GST to the production and sale of a chair.Illustration of GST for a ChairEssentially, the total tax will equal 9% of the sale price to the final consumer. However, GST will be collected on the value added at each stage of the production and distribution chain . What will be TaxedVirtually all goods and services will be taxed. Items on which GST will not be charged fall into two categories: tax free and tax exempt."

Jan 1, 1989

By K. Eklund, T. Ahokainen

The Flash Smelting process is the most used production method for primary copper in the world. The basic idea of the Flash Smelting is to utilize the thermal value of sulfidic concentrates and melt the reaction product, in controlled conditions, without a need for external energy. This basic principle is utilized both in matte smelting and in Direct-to-Blister Flash Smelting Processes. In this paper, the behavior of normal (CuFeS2) concentrate of FSF-process is compared to the behavior of high grade feed material produced by Outokumpu Concentrate Upgrade Process with the aid of computational fluid dynamics (CPD) simulation using add-in subroutines developed by Outokumpu Technology. Results consist of the information concerning the behavior of the suspension, reaction efficiency and other parameters concerning the reactions of the feed material. It will be shown that, using the Upgrade Process, the throughput of copper from the Flash Furnace can be almost tripled.

Jan 1, 2007

By J. H. Baek, D. W. Lee, B. K. Kim

Nanostructured Cu-Al203 composite powders were synthesized by thermochemical process for an electrode application. The synthesis procedures are 1) preparation of precursor powder by spray drying of solution made from copper and aluminum nitrates, 2) heat treatments to evaporate volatile components in the precursor powder and synthesis of nanostructured CuO+Al2O3, and 3) CuO reduction by hydrogen into Cu. The suggested procedures stimulated the formation of the gamma-Al2O3. The mean particle size of the final composite powders was 20nm, and the electrical conductivity and hardness in the hot-extruded bulk were competitive with Cu/Al2O3 composite by the conventional internal oxidation process.

Jan 1, 2004

By S. Mathieu, R. Mennillod-Blondin, P. Cousin, N. Valle, H. Bouzahzah, P. Marion

The gold deposit of the Lapa mine has a complex mineralogical gold state with coarse and fine native gold, aurostibite and submicroscopic gold within arsenopyrite. A specific mineralogical characterization of the refractory fraction was conducted combining various techniques such as optical microscopy, electron microscopy, electron microprobe and secondary ion mass spectrometry. In parallel, the mineral processing behaviour and the recovery of the refractory gold were studied by pressure oxidation and cyanide leaching. This paper presents the mineralogical methodology developed to characterize the refractory gold distribution within arsenopyrite. These results are presented and interpreted in order to plot gold recovery versus sulphide oxidation. The comparison of the metallurgical behaviour yields predicted from the mineralogical data and the one calculated from leaching tests shows a good agreement and allows mineralogical explanations of the gold metallurgical behaviour.

Jan 1, 2011

By A. J. Williams

"IntroductionAN EXCELLENT paper entitled Potash Discoveries in Western Canada (l) was presented to the 1948 Annual Meeting of the C.l.M. by L. Heber Cole, of the Department of Mines and Resources, Ottawa. In the intervening three and a half years, the search for oil and gas in Saskatchewan has added considerably to the information available on the occurrence of potash. It is the purpose of this paper to summarize the additional data. For greater clarity, it may be useful, first, to recapitulate some of the highlights of Mr. Cole's paper.Underlying the western plains of Canada there is a large deposit of sodium chloride, believed to be of early Devonian age. Associated with the sodium chloride are certain potash salts, primarily 'sylvinite,' a mechanical mixture of sodium chloride (halite) and potassium chloride (sylvite) crystals. The potash salts, when present, are usually near the top of the salt formation, though in at least one well they were found near the central part of the deposit. Cole reported on some seven wells in Saskatchewan which had encountered appreciable amounts of potash; these were the Ogema, Radville, Davidson Nos. 1 and 2, Prairie Salt Nos. 1 and 2, and Verbata wells. Depth of potash from surface ranged from 7,653 feet in the Radville well to a minimum of 3,466 feet in the Verbata well. Such depths do not compare favourably with those at which the potash is found in mines in the United States, France, and Germany, where about 1,000 feet is the average. The greatest concentration of sylvite encountered was in the Verbata well, where a bed 11 feet thick carried the equivalent of 21.64 per cent K20."

Jan 1, 1952

By H. P. Pew

THE Pembina oil field is in west-central Alberta, about 88 miles west-southwest of Edmonton. The field, about 35 miles long by T HE Pembina oil field is in west-central Alberta, about 88 miles west-southwest of Edmonton. The field, about 35 miles long by 85 miles wide, is the largest in Canada in areal extent, in the amount of oil in place and in the current producing rate. The discovery well was completed in the middle of 1953. The field has grown until its present area is 369,820 acres. There are 74 working interest owners in the field. Early in the producing history of the Pembina field a committee of operators was formed on a permanent basis. The Pembina Operators' Committee has established a permanent staff, which compiles and distributes pertinent field information on a regular basis to all participants. The Alberta Oil and Gas Conservation Board has dealt with field-wide problems primarily through this committee. Expansion of the field was rapid as shown by the number of wells completed -from 135

Jan 1, 1961

By James D. Hall

THIS essay deals with the materials and methods used in the waste filling of scopes in Canadian mines, with especial reference co the methods used at the Sullivan mine, Kimberley, B.C. The author would like to thank members of the staff of the Sullivan mine for information obtained, and acknowledges material from the following publications: Waste Filling of Stopes: Information Circular 6816, U.S. Bureau of Mines. Mining Methods and Practices at Lake Shore, by Leslie S. Weldon; Trans., C.I.M. & M., Vol. XXXIX, 1936, pp. 142-202. Development and Mining (at the Mcintyre Mine), by D. E. Keeley; Eng. & Min. Journal, November, 1933. INTRODUCTION Although mining in Canada is a relatively youthful industry, many of our mines have reached a stage of extraction which requires waste filling, both as a means of support and co make possible the winning of ore that would otherwise have co be left in the mine. Fill may be used at regular intervals as one of the cycle of operations, as in cur-and-fill and square-set scoping; or it may be introduced into a completed stope co provide permanent support for overlying rock formations, and co permit mining of pillars of valuable ore. Waste fill is sorted in resuing, which is a form of cut-and-fill scoping employed in narrow veins.

Jan 1, 1937

By Y. M. Zhao, C. L. Duan, Z. F. Luo, X. L. Yang

"Fine coal beneficiation in a dry way is considerably imperative for the efficient utilization and conservation of coal resource in drought regions, especially for China. In this study, a continuous vibrated gas-fluidized bed separator was established and developed based on the results of laboratory batch separation results. The fluidization behavior and the effects of several main parameters including superficial air velocity, vibrational parameters, bed height and separating time were systematically investigated. The results showed that the bed expansion increased linearly with the superficial air velocity and theoretically, higher bed expansion benefited the hindered settling process of fine coal particles. However, larger superficial air velocity would induce intense fluctuation in bed surface and a poor performance of density segregation. The conveying velocity of particles was mainly determined by the vibration angle and a predicted model was established by the nonlinear regression analysis. A theoretical model of bed height distribution was established and this model fitted the measured value well. The beneficiation performance of this separator was evaluated by separating -3+1 mm Wuhai fine coal. The results showed that the probable error, E, is 0.225, indicating a continuous vibrated gas-fluidized bed separator could achieve a satisfactory performance.INTRODUCTIONFine coal beneficiation in a dry way is considerably imperative for the efficient utilization and conservation of coal resource in drought regions, especially for China. Meanwhile, it is also conducive to environment protect and low-carbon, green development of coal sector. In recent years, the increasing academic interest in the field of fine coal dry beneficiation has been aroused in most of big coal countries. In summary, recent processes of fine coal dry beneficiation are, according to their processing principles, mainly classified into two categories: dense medium beneficiation (Dong et al., 2015; Fan, Chen, Zhao, & Luo, 2001; Luo et al., 2008; Macpherson, Iveson, & Galvin, 2011) and improved pneumatic beneficiation (Patil & Parekh, 2011; Yang, Zhao, Luo, Song, & Chen, 2013; Yang, Zhao, Luo, Song, Duan, et al., 2013; Zhao, Zhao, Chen, & Luo, 2015). The former deploys artificial dense medium, such as magnetite powder and fine sand, to form a stable gas-solid suspension with certain density under the function of the upward fluidizing air flow and in some case, the external force field like vibration and magnetic field. Generally, it is believed that the process of air dense medium fluidized bed (ADMFB) is suitable for separation of coarse coal (-50+6 mm) and can't obtain a satisfactory result for fine coal separation (Luo, Chen, & Zhao, 2002; Luo et al., 2008). However, some recent laboratory efforts (Patil & Parekh, 2011; Zhang et al., 2014) attempt to find an improved performance for fine coal beneficiation in ADMFB with some modifications. The point of these works is to use a comparatively shallow bed and finer magnetite powder with appropriate size distribution. It is known that fluidization of finer magnetite powder is prone to arouse adverse fluidization properties such as channeling and slugging. Thus, the external force is introduced to an ADMFB in order to improve the fluidization quality of a magnetite powder bed. Some recent literature on air dense medium magnetically stabilized fluidized beds (Fan et al., 2001; Luo, Zhao, Chen, Fan, & Tao, 2002; Song, Zhao, Luo, & Tang, 2012), air dense medium vibrated fluidized beds (Luo et al., 2008) and Reflux Classifier (Macpherson, Iveson, & Galvin, 2010) reports that the separation efficiency, namely the probably error E value, of dense medium separation of fine coal ranges from 0.066 to 0.07, indicating a good separation performance. But, the main unresolved problems of dense medium beneficiation are located in the processes of product purification and medium recovery and are urgent to be solved befor"

Jan 1, 2016

By K. Musgrave

"The ISO 9000 Quality Management System Standards have become internationally accepted as the way to manage quality assurance. The author has developed a Maintenance Management System Standard using ISO 9001as a model and is working with a major basemetal mining company to implement this. This paper describes the structure of the standard and the steps required for its implementation.Using a management standard ensures that a system is maintained, reviewed by internal audit, and when instances of failing to follow the procedures (nonconformances) occur, that corrective action is taken. It also ensures that regular reviews are made and that the sys-tem is continuously improved.The Standard is designed so that if an organization decides to register to an ISO 9000standard, the Maintenance Management System will be in a format that is compatible and can be incorporated with little effort into the Quality System.IntroductionIn reviewing maintenance operations at many mining companies, and often seeing those same organizations some years later, two recurring issues have been noticed..The first issue is how dependent effective maintenance is upon individual supervisors and crews. In one situation, a mine that had an excellent maintenance system in the 1980s. Now, with new supervisors and crews, it is starting to develop a system that will bring it back to where it once was. No written procedures remain from the earlier excellent system.The second issue is that most operations feel that maintenance is the sole responsibility of the maintenance supervisors and crews. Nothing could be further from the truth. Maintenance, like safety, is everyone’s responsibility. If the operator does not report problems and perform daily checks, and if the production supervisor runs equipment until it fails, then nothing the maintenance crew does will prevent emergency breakdowns and lost production. More importantly, equipment will be operated in unsafe conditions, putting employees at risk and, with today’s regulations, leaving supervisors open to prosecution.The Maintenance Management System outlined in this paper is designed to establish the present best practices as standard procedures, to provide a continual review so that these practices are improved in the future, and to identify instances where the system fails so that corrective action is taken. The mine referred to above is using this approach to make sure that, this time, the system it implements will survive."

Jan 1, 1997

By Kristopher John Shelswell

The work in this paper describes the theory, development and analysis of a computer model simulating a mine operation moving material using concurrent decline ramp truck haulage and skipping practices. This simulation was designed specifically to assess the performance of the truck haulage component and to expand on the scope and capabilities of traditional TKM fleet calculations. Haulage fleet simulation results were similar to TKM calculations for the lower to moderate production rate cases. However, comparisons of ramp-up and high production rate scenarios indicated that the simulation model was significantly more sensitive to operating practices and factors not accounted for by calculations derived when using a single TKM figure. Given the proper calibration of inputs and logic, the simulation model was additionally capable of providing information associated with general mine operations efficiency. Further benefits of the simulation model over a TKM approach included logistical assessments such as production schedule feasibility, productivity trade-off quantifications for proposed operating practices, and the validation of mine resource utilization scheduling. The simulation model was able to expand on traditional methods to approach planning and logistics by considering factors such as truck queuing, real-time traffic interactions, random events, and competition for resources on mine productivity.

May 1, 2013

By The Hollinger Mill Staff

MR. E. L. Longmore (l), in presenting this paper at the Annual General Meeting, suggested that it was neither desirable nor necessary to read any but the briefest portion of the paper. He wished to emphasize that the scale of the milling operations at the Hollinger not only provided several units in continuous operation to ensure that the test work was being carried on under normal conditions, but also provided a large and competent staff to supervise such an extensive experiment and to criticize intelligently the recorded results. He wished to acknowledge the enthusiastic co-operation of the group of assistants whose excellent work was responsible for the confidence with which the conclusions could be presented. The following data were given as a supplement to the published paper. Following conclusion of the test work described in the paper, the size of the balls fed to the mill was changed from 3-in. diameter to 2Yz-in. Within a short time, a gradual rise in the tonnage rate took place and, as no other condition had been changed, the increased tonnage was credited to the changed character of the ball load. At present, the average daily tonnage for the mill with low discharge is 1,507 tons, an increase of 77 tons over the 1,430 tons ground when feeding a 3-in. ball. A similar increase in tonnage should be credited to the high-discharge test, if the 2;Y2-in. ball had been used at that time. This would raise the 1,117 tons ground at that time to a new figure of 1,177 tons. From a study made to determine those factors influencing the rate of pulp flow through the low-discharge mill, it became evident that the ball screen used on the high-discharge mill must have been restricting the flow through the mill, although this was not recognized at the time. Accord-ingly, it was decided to repeat the high-discharge run. Additional length was added to the 21-in. ring grizzly originally used, and the length of the reverse spiral was increased. The end of the grizzly was left open so that, if necessary, excess pulp might flow out. A coarse mesh 3 ft. by 3 ft. flat screen was provided to catch stray balls.

Jan 1, 1937

By M. Alvarez-Silva, A. Uribe-Salas, J. A. Finch, M. Mimezami

The point of zero charge (PZC) of electrokinetically anisotropic minerals (chlorite and serpentine) was determined using the Mular-Roberts (M-R) titration technique, and compared to isoelectric point (IEP) obtained from electrophoretic mobility measurements. To validate the M-R technique, the point of zero charge and isoelectric point of two isotropic minerals (alumina and silica) were shown to be comparable. For serpentine and chlorite, respectively the PZC was found at pH 4.3 and 4.6, while the IEP occurred at pH 3.2 and <3. Settling tests showed that the maximum settling rate, which corresponds to aggregated particles, for chlorite was around pH 3, i.e., close to the IEP; lizardite did not show a conclusive maximum.

Jan 1, 2009

By Duanyou Li, Qing Xiao, Zhiwang Wang, Xiaoqing Gan

"Seepage is a major cause of earth-rockfill dam failure. Numerical simulation of infiltration process, distribution of drift, flow velocity and direction after the impoundment can guide seepage control and evaluate its effects. A double field coupling model is established based on soil consolidation theory and the interaction between the seepage field and the temperature field. The simulation is implemented under the situation of core wall setted, core wall not setted, and core wall setted with concentrated leakage. The results indicated that the saturation line height was significantly reduced due to impervious core wall, and the temperature distribution is uniform above the saturation line. If concentrated leakage occurs, temperature distribution of dam was significantly changed, which is consistent with theoretical inference.INTRODUCTIONEngineering seepage field distribution must be calculated in seepage control design, optimization, and correction of new engineering construction of earth dams, as well as in seepage monitoring and evaluation of existing earth-rockfill dams. The aim of calculation is to identify the important parameters in seepage field, such as the coefficient of permeability, seepage flow, pressure, head, and allowable seepage slope, for effectively implementing seepage control design, correction, safety assessment, and anti-seepage reinforcement. These seepage parameters usually can be obtained by in-situ test and back-analysis based on reliable monitoring data.In order to obtain reliable monitoring data and reflect the earth-rockfill dam seepage state and change process, appropriate monitoring technology and monitoring method is necessary. The traditional seepage monitoring method is through site inspection and sensors (e.g., osmometer, piezometric tube and measuring weir) that buried in the structure and its foundation. However, there are still some disadvantages:"

Jan 1, 2015

By Rod Maclean, George Stewart

"This paper covers the planning, development, and mining of an underground pillar complex using a bulk mining method. To achieve the long-term goal of continually improving our operating processes, all cut-and-fill and undercut-and-fill workplaces at Thompson Mine have been evaluated for their conversion potential to the bulk method referred to as VerticalBlock Mining (V.B.M.).Since December 1979, conversion has occurred to the extent where 80% of the ore hoisted is mined using Vertical Block Mining (V.B.M.).Productivity and profitability have been substantially improved by converting to this bulk method, especially in pillar areas.IntroductionThompson Mine is a 6000 ton per day underground nickel mine located 3 miles outside the City of Thompson, 400 air miles north of Winnipeg . The mine has two shafts, T-l and T-3 (Fig. 1). The T-l shaft is 4427 feet deep with the deepest mining taking place on 4000 level. The T-3 shaft, which is the area of this presentation, extends to a depth of 2607 ft with the deepest operating level at 2400 ft. The two shafts, some 10 000 ft apart, are connected underground on numerous levels. The ore mined at T-3 is trammed on 2400 level to the T-I main ore pass and hoisted through the T-l shaft. Eight hundred and fifty thousand cubic feet of fresh air per minute ventilates the mine.The 1400 level 310 pillar complex is located in the T-3 mine area. There are six rib pillars, two crown pillars and one main pillar (boundary pillar between cut-and-fill stopes) within the complex (Fig. 2). The majority of the ore in this area has been mined using mechanized cut-and-fill techniques .Undercut-and-fill mining was the method originally planned to recover these pillars. It is slow, labour-intensive, and requires extensive use of timber during the mining cycle. The extraction drifts in the area are heavily timbered due to their proximity to the ore and the age of the facilities. Increasing drift maintenance costs and the successful mining of two main pillars with V.8.M. techniques provided the stimulus to review the mining plans for this area."

Jan 1, 1988