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By Erol Yilmaz

This poster presents results obtained from an experimental test set-up called CUAPS (curing under applied pressure system) allowing to assess the performance and quality of cemented paste backfill (CPB) materials. CUAPS can be used for carrying out a series of tests: one-dimensional consolidation, saturated hydraulic conductivity, pore-water pressure dissipation, and curing under uniaxial stress. The poster presents CUAPS testing preliminary results including uniaxial compression tests (UCS) and mercury intrusion porosimetry (MIP). It was demonstrated that CUAPS-consolidated CPB materials cured under time-dependent stresses provide higher UCS values than standard mould (undrained and unconsolidated) CPB samples. This may be explained by the release of excess pore water within backfill that gives rise to a significant reduction in the total porosity of fresh CPB materials prior to hardening and thus, to higher mechanical strength (UCS) gain. The results obtained from MIP highlighted a good correlation between CUAPSconsolidated and mould-unconsolidated CPB samples. MIP results were also related to UCS values in order to predict the final CPB mechanical strength.

Nov 1, 2008

By H. Nejati

Experimentally, the Semi-Circular Bending Moment (SCB) fracture toughness test results of Basalt Specimens Under Illumination (BSUI) have shown reduction in fracture toughness and failure force as a result of micro-cracking. The fractographic investigation of BSUI confirms initiation and extension of micro-cracks. Determination of a relation between strength reduction (herein fracture toughness reduction) and density, location and orientation of thermo-mechanical induced cracks due to microwave radiation is critically important. Numerical modeling can bridge strength reduction and the density of thermal microcracks generated due to microwave radiation. Discrete Element Modeling (DEM) is a power tool for simulation of micro-crack initiation, coalescence into macroscopic fractures and failure mechanisms in brittle rock specimens under different stress regimes (herein combination of pure tensile stress due to SCB testing, and thermal stress due to microwave radiation). For this purpose, a DEM modeling of SCB testing of basalt using EDEM (i.e., a DEM based software). DEM models are calibrated according to results of experimental SCB testing as reported in previous papers. This paper first reviews some background data on SCB fracture toughness, and experimental results of fracture toughness of microwave treated SCBs. Then it proposes a new methodology for simulation and validation of rock samples exposed to microwave illumination with regard to experimental results, and discusses the correlation between fracture toughness of BSUIs and density of micro-cracks, which is attributed to applied microwave power and duration of illumination.

Aug 1, 2013

By M. Sider, D. L. Piron

"The cost of energy is a growing factor in the cost of zinc production. Electrowinning from sulphate solutions with periodical reverse current or from acidic chloride or alkaline solutions could be the way to lower costs. A comparison of the energy consumption for electrolysis shows that the alkaline solution may be considered the most promising alternative. IntroductionZinc electrowinning from sulphate solutions is a very energy intensive process (typical specific energy usage : around 3.1 kWh / kg zinc). This energy consumption represents 35% of the total cost of electrolysis and is the most important cost component.In order to lower costs, the energy efficiency of the sulphate process should be improved or alternative processes should be developed. In this laboratory, that objective has resulted in research on electro winning in sulphate solutions with the use of periodical reverse current (PRC) and in acidic chloride and alkaline solutions.Some of the results obtained in this laboratory and already published and new experimental results reported in this paper make it possible to present a more critical comparison of the energy consumption of the three processes mentioned.ExperimentalThe experiments at constant current were carried out as described previously. Specific conditions are indicated in Table I. Current efficiencies were obtained by weighing deposits.ResultsThe best energy consumptions for electrolysis obtained in this laboratory for the three media are presented in Table I. More detailed results can be found in the papers already mentioned."

Jan 1, 1988

By B. Schulz, T. Leißner, R. G. Merker

"Ores enriched in critical rare earth elements (CREE) keep on being of particular interest. The CREE are Nd, Eu, Tb, Dy, including Y. In conjunction with beneficiation tests on Y-mineral bearing flotation tailings of complicated mineralogy, the Mineral Liberation Analyses (MLA) was used to work out processing characteristics fraction-wise in terms of modal mineralogy, mineral recovery, locking and liberation, and particle size.Size fractions from mechanical sieving (“real fractions”) as well as fractions created virtually by using filters on the MLA-data of the unsieved head material (“virtual fractions”) were investigated comparatively thereby. The ""virtual-sizing-approach"" features a much simpler sample preparation procedure and considerably lower cost if a fraction-wise investigation of a head sample is needed. It can be conveniently applied to intimately locked ores needing a fine grinding to a particle size distribution (PSD) range below 25 microns for instance. However, its practical use is sometimes disputed in terms of the reliability of the results. Results are shown for the main Y-bearing mineral xenotime and selected main gangue minerals of a low grade REE-ore.The influence of sample representativeness, sample preparation and particle counting statistics turned out to be of crucial importance. Given that all preconditions for reliable results are met, locking characteristics, modal mineralogy, and even PSD trends can be read from virtual fractions.The authors encourage processing engineers to use the virtual approach. They propose practical measures to improve the reliability of the MLA results, identify and tackle conflicting results, and mitigate evaluation problems that may arise."

Jan 1, 2016

By Alok Porwal, E. J. M. Carranza, M. Hale

Abstract- Approaches to mineral potential mapping based on weights of evidence generally use binary maps, whereas, real-world geospatial data are mostly multi-class in nature. The consequent reclassification of multi-class maps into binary maps is a simplification that might result in a loss of information. This paper describes results of using multi-class evidential maps in an extended weights-of-evidence model vis-à-vis results of using binary evidential maps in a simple-weights-ofevidence model. The study area in the south-central part of Aravalli province (western India) hosts a number of SEDEX-type base metal deposits in Proterozoic supracrustal rocks. Recognition criteria for base metal deposits were represented as both multi-class and binary evidential maps. The known mineral deposits were divided into two subsets, viz., the training and the validation subsets. The training subset was used to calculate, for the evidential maps, the weights, contrasts, and posterior probabilities and their variances. The distributions of expected frequencies of base metal deposits estimated from the posterior probabilities and the observed frequencies were compared using standard goodness-of-fit tests to verify conditional independence of the input evidential maps. The posterior probabilities from both the models were mapped and interpreted to classify the study area into zones favorable, permissive, and non-permissive for base metal deposit occurrence. As compared to the simple weights-of-evidence model, the extended weights-of-evidence model results in more robust and finely differentiated posterior probabilities in favorable and permissive zones and has a better prediction rate. The results also reveal that the statistical properties of the weights of evidence, the contrasts, and the posterior probabilities are not significantly degenerated by using multi-class evidential maps in weights-of-evidence modelling.

Jan 1, 2001

By Adam Jordens

Along with gravity separation and froth flotation, magnetic separation is one of the most common techniques for separating minerals. Minerals are separated by a magnetic force that is dependent on a combination of the applied magnetic field, the magnetic field gradient and the magnetic properties of the minerals themselves. Minerals may be classified as diamagnetic, paramagnetic or ferromagnetic based on the way they interact in a magnetic field. The magnetic susceptibility of a material may be determined by using a vibrating sample magnetometer to determine the magnetic moment of the material in an applied magnetic field and then converting this measured moment into the material?s magnetic susceptibility. In the case of rare earth minerals, magnetic separation is commonly used to remove ferromagnetic iron-bearing gangue minerals and to concentrate paramagnetic valuable minerals such as monazite and xenotime. These magnetic separation steps are frequently preceded by a gravity separation step as many iron oxides and rare earth minerals have specific gravities that are much larger than those of typical gangue minerals. By varying the magnetic field intensity it may be possible to use magnetic separation to selectively separate the paramagnetic and ferromagnetic minerals that report to the concentrate after gravity separation. Test work on a Canadian rare earth ore has shown that a gravity concentration step using a Knelson concentrator was able to selectively concentrate heavy magnetic minerals as evidenced by increased recoveries, relative to the Knelson tailings and fresh feed, after processing through a wet variable intensity magnetic separator. Mass pull to the Knelson concentrate was 4.5 % and the magnetic recovery of this concentrate, at 0.94 T, was approximately 40 % higher than the magnetic recovery from the Knelson tailings (70 % versus 30 %). The products of the wet high intensity magnetic separation were also analysed using a vibrating sample magnetometer to determine the magnetic susceptibilities of the various concentrate and tailings products.

Aug 1, 2013

By Jocelyn Koshinsky

The uranium milling process at the Key Lake Operation of Cameco in Northern Saskatchewan, Canada results in the leaching of untargeted metal(loid)s such as arsenic, iron and nickel that are found at high concentrations in the ores. To mitigate the environmental impact of the leached As, Fe and Ni in the mill waste solution (raffinate), the pre-discharge mine tailings (rafffinate and residue) are neutralized with slaked lime, resulting in the formation of secondary As-, Fe- and Ni-bearing mineral phases which serve to control their concentrations to very low levels in the tailings porewater. As part of an ongoing investigation into the long-term evolution (10,000 year) of arsenic mineralogy in the Deilmann Tailings Management Facility (DTMF) at Cameco?s Key Lake Operation, it has become necessary to characterize secondary As-bearing mineral phases during the neutralization process prior to the discharge of the tailings and compare these to in situ As-bearing mineral phases in the tailings facility. This is primarily to determine if changes have occurred over time to As-bearing mineral phases in the tailings facility as well as to investigate whether alternate neutralization techniques may have a beneficial effect on the stability of mineral phases formed. Thus a bench-scale model was developed and preliminary testing conducted to simulate the mill neutralization process (i.e. neutralization of synthetic raffinate solutions with incremental amount of slaked lime to terminal pH values). The resulting ?tailings? (precipitates and supernatant) were then separated and analyzed. XRD analysis of the precipitates shows that the Fe occurred mostly as ferrihydrite. This was complimented with results from Raman spectroscopy which further indicated that the As appeared adsorbed onto the ferrihydrite. Results of geochemical modeling of the bench-scale test compared well with supernatant profiles from ICP-MS analysis.

Aug 1, 2010

By John E. Udd

The stability of a rock mass is influenced by the presence of discontinuities. In assessing stability, it is required that one should know the frictional and cohesive properties of these contained weaknesses. In the study performed by the authors, a number of samples of diamond drill core containing open discontinuities were tested in shear. The results obtained were compared with values which resulted from back analyses of instabilities in the field.

Jan 1, 1979

By G. E. Rodger, R. Jugoon

"SUMMARYSamples of Aunor gold precipitates which varied widely in gold content were fluxed to produce a bisilicate slag with various oxidizers and combinations of oxidizers. On the basis of some preliminary tests a flux composition was chosen which appeared to show the most promise.The different components of this flux were then incorporated into a matrix for a 23 factorial design experiment to determine what were the most important variables and what changes should be made to the flux composition to optimize the resultant total bullion fineness.The results of the 23 factorial design experiment showed that the flux composition as calculated was very close to an optimum as only very minor changes were required to give bullion in excess of 995 fine.INTRODUCTIONEarly in 1969 the results from the Aunor Gold Mines refinery showed that wide variations existed both in precipitate grade and in total bullion fineness. An analysis of this data showed that as the precipitate grade increased, the gold bullion fineness also increased (see appendix 1). Therefore, to improve the refinery results test work was initiated to: 1) improve the precipitate grade and 2) to determine, with the help of statistics, a flux composition which would be able to handle wide variations in precipitate grade and give a total bullion fineness in excess of 950 fine."

Jan 1, 1970

By Ahmad Hemami

A working code is developed to simulate the forces that are involved in the process of penetration of a bucket into bulk media during loading and excavation. The result of this work can be integrated with the simulation software that many manufacturers of heavy earthmoving machines introduce for training purposes. The simulator programs for machines such as backhoes, loaders, LHDs and the like, used for operator training, are usually only based on the kinematics of the machine. As a result, there is no indication of the effect of the medium when the cutting tool of such a machine touches the excavation material. This is like moving the machine in the air. Feeling the resistance force of medium to cutting by a bucket or blade during loading and excavation is more realistic, both mentally (by vision) and physically (feeling resistance in pedals and levers). The results from force simulation can be used by feedback in creating simulated resistance in these training gears, similar to what is done in flight simulators for aircrafts. In addition to serving for the above purpose, the results of this work can be used as a tool for examination and verification of the effect of a control policy for automation of loading and excavation of bulk media. So far no commercial product has been introduced that has such a capability, despite some extensive research on the subject.

Aug 1, 2013

By Desmond M. Levins

"Australia has almost twenty per cent of the Western World's low-cost uranium reserves, located mostly in the Alligator Rivers region of the Northern Territory. At present, only one uranium mill is operating in Australia, but a number of new mills are planned for the early 1980s.Details are given of Australian uranium mining and milling proposals and the measures taken to minimize their environmental impact. Major factors affecting environmental impact are discussed, including treatment of liquid wastes, water management, control of radon and other airborne releases, and disposal of tailings. IntroductionAustralia is entering a new phase in uranium mining and milling activity . In the first phase of operations, between 1954 and 1971, about 10,000 tonnes of yellowcake was produced at five locations. Most of the production came from the mines at Rum Jungle in the Northern Territory and Mary Kathleen in Queensland. In the 1960s, the demand for uranium slumped, and much of the yellowcake produced at Rum Jungle was stockpiled by the Australian Atomic Energy Commission. The early uranium processing operations in Australia were typical of overseas practice at that time. Environmental problems were generally given low priority and pollution control measures were unacceptable by today's standards. There was little or no recycling of process water, acid waste streams were not neutralized and tailings dams were poorly designed. At Rum Jungle, oxidation of sulphides in waste heaps, tailings dumps and ore stockpiles subsequently led to destruction of vegetation and heavy metal pollution in a nearby river.In the early 1970s, major uranium deposits were discovered in the Alligator Rivers region of the Northern Territory. Significant orebodies were also located in the Great Artesian Basin in South Australia, in calcrete formations in Western Australia and at several locations in Northern Queensland. According to latest available estimates. Australia has about 160/0 of the Western World's uranium recoverable below $US 80 per kg U. Mining these orebodies will require pollution control measures more stringent than practised previously in Australia."

Jan 1, 1980

By A. P. Annan, C. Vaughan, J. L. Davis

"The increase in the price of gold in the last decade has led to a significant increase in placer gold exploration around the world. Placer deposits occur in depressions along ancient river and stream channels. In many instances these channels have been covered with materials from subsequent erosion or glacial activity. Use of high resolution geophysical techniques to map the bedrock and soil strata can improve the probability of finding the placer deposits. The buried bedrock surface is often the factor controlling deposition. Ground probing radar (GPR) and seismic sounding are complementary high resolution techniques which can be used to map bedrock topography and soil stratigraphy. Current radar instrumentation operates best in coarse grained soils. Penetration depths of up to 30 m have been attained, and stratigraphic horizons have been mapped with spatial resolutions of better than 0.5 m. Although its spatial resolution is considerably lower than radar, seismic refraction achieves superior penetration in electrically conductive environments such as clay and silt soils. The seismic refraction method thus proves to be complementary to the radar method. The seismic reflection technique is analogous to the radar method. Current seismic instrumentation and energy sources however, provide limited spatial resolution and do not permit the use of seismic reflection for reflecting horizons lying at depths of less than 20 m to 30 m.This paper discusses the use of both radar and seismic methods for placer exploration. The radar and seismic techniques, the field methods, equipment, and data displays are presented. Field examples illustrate the high resolution mapping capabilities of the techniques. The radar data agree very well with the borehole data, and in fact they sometimes detect subtle soil strata variations associated with changes of permeability that were not initially observed in the borehole logging. The radar method can give a continuous profile of water table and bedrock in sandy soils. Seismic soundings are necessary to differentiate between the bedrock and water table reflections observed by the radar."

Jan 1, 1987

By J. M. Skeaff

"The main solution, ore and residue pulp streams in the Rio Algom Quirke I uranium mill at Elliot Lake have been sampled and analyzed for radium-226. Analysis of the ore and leached residues indicates that Ra-226 dissolves and precipitates within the first pachuca tank. A maximum of approximately 0.2% of the Ra-226 in the ore remains in solution during leaching.IntroductionConcern over the acid drainage and radionuclide dissolution problems associated with uranium tailings disposal sites in Canada has prompted investigation into methods for improving current uranium processing practice.Present commercial practice usually involves leaching of the uranium ore with sulphuric acid, ion-exchange separation of the uranyl sulphate complex from the leach liquor and"" precipitation of the uranium as yellowcake. Daughter products of U-238, such as Th-230, Pb-21O and Ra-226 remain with the tailings and undergo long-term dissolution.Table I summarizes the probable behaviour of the radionuclides during milling. The concern here is with Ra-226, because it will deposit in the bones if ingested and is ""thus considered a serious health hazard(1), Similarly, Po-218 and Bi-214, which are daughter products of Rn-222, can be retained in the lungs and are considered to be the principal sources of radiation dosage(2) arising from the inhalation of Rn-222. Previous investigations(3-9l indicate or imply that Ra-226 undergoes redistribution during milling. However, detailed published information on the behaviour of Ra-226 in the uranium milling process appears to be lacking. A knowledge of the behaviour of Ra-226 in uranium mills, particularly during leaching, could lead to improved methods for its control.This paper reports the occurrence of Ra-226 in the Rio Algom mill. The main objectives of the study were to determine (1) the distribution of Ra-226 within the solids, (2) at what point, and how rapidly, redistribution occurs and (3) whether there are high concentrations of Ra-226 in solution."

Jan 1, 1981

By P. W. Percival, J. J. Janssens

"Whitehorse Copper is a 2600-ton-per-day underground copper mine located within the city limits of Whitehorse, Yukon . In early 1978, due to increasingly difficult ground conditions. it was decided to investigate the conversion from longhole open stoping to vertical crater retreat mining. On-site crater testing by C-I-L demonstrated the feasibility of V.C.R. mining at Whitehorse Copper.After some investigation. a down-the-hole drilling rig was chosen and drilling began in November /978. Blasting began in June /979. and to September 1980 approximately one million tons of ore had been blasted. In-field changes to both drilling and blasting techniques have been implemented. Some of these changes were dictated by problems encountered; others were made to improve efficiencies.To date. V.C.R. mining has proven successful from a drilling and blasting aspect.IntroductionWhitehorse Copper Mines Ltd ., a division of Hudson Bay Mining and Smelting Co., Limited , is a 2600-Lp.d. underground copper mine. The plant site is located I mile west of the Alaska Highway, and approximately 7 miles south of Whitehorse city center. Production is from the Little Chief orebody, and total output is from vertical crater retreat stopes.The Little Chief and Middle Chief orebodies at Whitehorse Copper had original reserves in excess of 6 million tons, most of which graded 2.00/0 copper or higher . The orebodies are typical skarn deposits that have developed along the contact between an intrusive diorite and a limestone-quartzite flank pendant. The limestone has been metasomatized to form a calcium-magnesium skarn which, along the limestone / quartzite contact, has locally further developed into a mineralized magnetite-serpentine skarn. The principal copper minerals are bornite and chalcopyrite, with minor chalcocite and valleriite. Skarn gangue minerals are serpentine, diopside, garnet, epidote, calcite and wollastonite. Appreciable values in gold with some silver are present and recovered as well."

Jan 1, 1981

By Claudio Rocha Rosales

"Peñoles is a mining group with integrated operations for the smelting and refining of non-ferrous metals, and it is also a chemicals producer. Peñoles is the world’s foremost producer of refined silver, metallic bismuth and sodium sulphate. It is also the leading Latin American producer of refined gold, lead and zinc. The Met-Mex Peñoles Zinc Plant commenced operations in 1973 with an annual production capacity of 105,000 t/y zinc. Since then, production capacity has increased, and at present, it stands at 240,000 t/y zinc. In this paper, the leaching process and particularly the operation of the jarosite circuit is reviewed, specifying process improvements to treat high iron calcines.INTRODUCTIONTwenty years ago the Electrolytic Zinc Plant in Peñoles was treating high quality zinc concentrates with low concentrations of iron (~7.5 wt.%) and silica (1.8 wt.%). Nowadays the quality of zinc concentrates is poor with a typical iron content of around 11 wt.% and silica at 2.0 wt.%. This has created numerous processing problems in our roast-leach-purification- electrolysis sections. In order to meet customer requirements it is necessary to find processing solutions to assure the operational continuity of the plant. These processing solutions are the focal point of this paper.REVIEW OF THE LEACHING PROCESSThe purpose of the leaching process is to recover the metals, mainly zinc, through a series of extraction stages with sulphuric acid solutions, which progressively increases the zinc dissolution. After the series of acidification stages, the solution, which is rich in iron, is sent to the jarosite circuit to separate zinc from iron (Garcia & Valdez, 1996). The first step of the leaching process is the neutral leach circuit which produces a zinc sulphate solution that is sent to the purification section for the removal of impurities such as cadmium, cobalt and copper. Then this solution is sent to the tank house to feed the electrolytic cells."

Jan 1, 2016

By Chri Langille, Dougal McCreath, Samantha Espley

"In 1993, trials of two innovative applications of liner support involving the replacement of bolts in the ground support system were undertaken at Inco mines. The first trial involved an evaluation of the effect of dynamic loading on mesh-reinforced boltless shotcrete in a sub-level cave mining area on 2400 level at Stobie Mine. The second trial involved an evaluation of the MIROC Mineguard spray-on liner as a replacement for bolts and screen support, or as a replacement for screen support alone, in a topsoil heading of a bulk, VRM (vertical retreat mining) stope on 3240 level at Lower Coleman Mine.In the Stobie Mine trial, an uppers retreat heading was supported using a two-pass system. First, shotcrete was applied 2 inches thick, followed by pinning of mesh screen to the back and walls. A second pass of another 2 inches of shotcrete was then performed. For recovery of the ore zone, a vertical slot was created using a Borepak blind boring machine. The uppers rings were blasted, one or two rings at a time, starting at the slot and retreating out of the heading. During blasting of the production rings, blast monitoring was used to record the ppvs (peak particle velocities). For a series of 5-ring blasts, the maximum ppvs were measured to be 0.4 m/s (16 in./s) with a maximum charge weight per delay of357 kg (800 Ib). Six 5-point convergence measurement stations were set up and after each blast the total rockmass convergence was measured using a tape extensometer. Although a maximum displacement of 100 mm (4 in.) was measured, the shotcrete still managed to provide some support junction. A support damage assessment was also done to try and relate blast vibration to effectiveness of the mesh reinforced shotcrete liner. This trial demonstrated the effect that blasting has, on a support system, and it also demonstrated the necessity of having the mesh reinforcement properly placed within the shotcrete."

Jan 1, 1995

By R. J. McLeod

"Grinding at Giant Yellowknife is done in two stages. The first stage consists of two ball mills in closed circuit will: mechanical classifiers, the classifier overflow goes to flotation. The tailings from this flotation circuit are separated into slime and sand portions, the slimes go to an additional flotation circuit while the sand portion goes to the secondary grinding mills.Milltronic automatic sound control units were installed on each of the two primary ball mills and put into operation September 3, 1966. After about two months adjustments to the control units the operators were taken off the grinding circuit on November 1, 1966. The operators duties were taken over by the automatic control units while miscellaneous work was taken over by the helpers in the mill. The labour saving by the installation is approximately $2250.00 per month.After a short period of operation it was apparent there was a slight change in grind. The screen analysis of the primary classifier overflows for the period May to August 1966 averaged 62.23% minus 200 mesh while for the period September to December 1966 averaged 62.0% minus 200 mesh. This is a slight coarsening of the grind. The change in the primary grind has manifested itself also in the screen analysis of the final grind. Before automation the final grind was 80% minus 200 mesh, while after automation it was 7 5% minus 200 mesh. A change of grind was also noted in the flotation concentrates. Before automation the screen analysis of the concentrates was 35% plus 200 mesh and 50% minus 3 25 mesh. A ft er automation it was 35% plus 200 mesh and 47 .5% minus 325 mesh. In other words the grind got coarser or perhaps we should say there was less overgrinding. Another significant change took place in the assays of the slime and sand portion of the primary flotation tailings. It was usual under manual control for the slime portion to assay higher in gold than the sand portion. The reverse is now true."

Jan 1, 1967

By K. Blower

"Since there was no representative from Wesfrob Mines Limited, the information on their infra-red concentrate drying operation, which had been prepared by Mr. K. Blower, was presented by Mr, Dorland. Mr. Blower, in submitting the information, emphasized that their objectives in this operation were (a) to guarantee less than 90%, of flow-moisture in their copper concentrates and thus satisfy marine shipping regulation• and (b) to minimize capital costs, The infra-red dryers have satisfied these objectives.To control final moisture in the copper concentrate, current practice is to preheat the filter cake with steam in the disc filter hood and then conveyor-0r7 the concentrate with infra-red calrod heaters. The application of steam aids the system as follows:1) Improves vacuum pump efficiency.2) Lowers viscosity of filtrate (increased temperature},3) Preheats the cake prior to radiant heating. (150°F)The hot cake discharging from the filter ranges from 10.6 to 10.9% moisture. After radiant heating the moisture level is reduced to 8.5% - 9.7% depending on production rate. Further moisture loss occurs during conveying and transferring to the stockpile for a final shipment moisture ranging from 8.1% to 9.5%. Production rate is frail J to 5 dry short tons per hour.1) #23 Conveyor 18"" flat belt covered from filter discharge to transfer point with 7 heaters end-to-end. Drive - Reeves Vari-Speed, 31:4 ratio, type 332-10-18, class ""D"", driving through a 24:1 ratio Dodge reducer. Speed of Conveyor - 3,9 feet per minute to 19.6 feet per minute. (Speed has been established and is not changed as a rule from the present level of a1:out 15.)2) #23 Conveyor 24"" wide, 20° troughed belt covered with 21 heaters, end-to-end from loading point, Speed - Constant at 54,2 feet per minute."

Jan 1, 1972

By Halet. R. A.

"The Quesabe Mine, owned by Brunston Mines Limited, is a gold deposit in the south central part of Duprat Township, northwestern Quebec, about 10 miles northwest of Noranda.It is one of the few gold mines in Canada developed since World War II. The occurrence of gold in the Quesabe fault structure has been known for many years, but serious exploration did not begin until 1945. By 1949, 15,000 feet of diamond drilling had been completed and a shaft sunk to a depth of 635 feet, with lateral development on the 450 and 600 foot levels. In 1949 and 1950 a new shaft was sunk to 1,050 feet, new levels were established at 800 and 1,000 feet, and ore was mined and shipped to the Noranda smelter. After November 1950 the ore mined was treated in the company's own mill. The grade of ore was marginal under existing economic conditions, and production was suspended in April 1952. The mine produced about 100,000 tons of ore with an average grade of approximately 0.25 ounces of gold per ton.GeologyThe bedrock formations are all Precambrian, and the general geology of the area has been described by W. G. Robinson (1). The rocks in the vicinity of the mine are acidic and basic volcanics of the Keewatin series (andesite, rhyolite, agglomerates, tuffs) invaded by the Flavrian Lake granite stock, and by dykes and irregular masses of intrusive felsite, quartz porphyry, syenite-porphyry, diorite, and lamprophyre.The mine workings are in andesite and felsite, about 2,500 feet west of the western margin of the Flavrian Lake stock.The principal structural feature of the area, from an economic point of view, is the Quesabe fault. This steeply-dipping fault cuts all the rock formations of the area, and has been traced for about 11 miles along a strike of north 65 degrees east from Lake Arnoux in the northeast corner of Dasserat Township to Duprat Lake near the east boundary of Duprat Township."

Jan 1, 1949

Unplanned open stope hanging wall dilution is a significant cost for many open stope mining operations. Significant advances in empirical and analytical approaches for estimating stope dilution have been made, however, many important factors are still either ignored or assessed in purely subjective terms. There are many hard to quantify factors that simultaneously influence opening stability and dilution. This creates a significant difficulty for determining the influence of a single variable, such as exposure time, on opening stability. Open stope exposure time has been found to be one of the important factors that influences overall stope stability and dilution. A study is currently being conducted to quantify many of the factors influencing open stope dilution. The assessment of a large comprehensive case history database of Hudson Bay Mining and Smelting Co. Ltd. (HBMS) operations forms the basis of this study. This paper presents the findings of a study on the influence of stope exposure time on open stope dilution. The study is based on applying existing empirical stope dilution prediction methods to the extensive HBMS database. Minimising stope mucking time has been recognised as a method to significantly reduce stope dilution. Factors which influence hanging wall dilution, such as drill and blasting methods as well as undercutting the hanging wall, have been applied to reduce the scatter in the database. Empirical design techniques based on rock classification attempt to relate opening stability to ?stand-up time?, however, these approaches have not frequently been applied to the mining industry. Data from a ?stand-up time? graph have been interpreted in terms of an equivalent average depth of hanging wall slough and compared to the HBMS database findings. To add to the existing HBMS database, several stope case histories have been taken from Geco mine. These cases include data from very large open stopes with long exposure times. Data from all three sources suggest that hanging wall slough increases between .14 and 0.3 metres per month for a rock mass Q? value between about 4 and 13.

May 1, 2003