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By D. W. Tully

Sedimentary and volcanic rocks of the Timiskaming series trend west-ward across the property of the Upper Canada mine, forming part of the north limb of a regional syncline which is truncated on the south by the Larder Lake fault zone. A wide zone of schistosity strikes east-west through the sediments and volcanics. The central portion of this zone is occupied by a system of narrow gouge-filled fractures in highly altered rocks, mainly tuff, trachyte, trachyte porphyry and syenite porphyry. The alteration zones have developed through repeated stages of deformation and accompanying metamorphism. The initial stages were marked by carbonatization and sericitization, followed by silicification and the introduction of minor amounts of sulphide minerals. Gold was finally de-posited in relatively late fractures within silicified zones and pyrite. Repeated movements along the Up-per Canada break have offset post-ore dykes of minette and diabase, and have resulted in minor displacements within the orebodies.

Jan 1, 1963

By L. Tafaghodi, S. Sarbishei

"Following the depletion of nickel sulfide ores, laterites are considered as one of the main sources for producing nickel and cobalt. One of the primary steps in pyrometallurgical extraction of nickel from laterites involves partial reduction of the ore at temperatures below 1000°C. Studying the kinetics of laterite reduction provides valuable information on reducibility of nickel laterites using coal as a reductant. The present study investigates the reduction behavior of a nickel laterite by thermal gravimetric analysis and differential scanning calorimetry. Non-isothermal reduction tests and X-ray diffraction analysis were performed to find the main reactions that occur during the process. The kinetic parameters were calculated by Coats-Redfern method. The general model of simple reaction f(a)=(1-a)n was used for calculating kinetic parameters.INTRODUCTION Nickel (Ni) naturally occurs in the form of sulfide and laterite ores. Laterite ores are mainly available in tropical regions. Weathering phenomena during millions of years has led to the formation of several layers with specific mineralogy (Butt & Cluzel, 2013). The three main layers downward are limonite, smectite, and saprolite (Crundwell et al., 2011). Although laterites are the most available nickel ore resource (70%), their share in the overall Ni production is limited to 40% (Crundwell et al., 2011). The main reasons for not materializing the full potential of laterites are as follows (Ray et al., 1985): •,Laterite ores contain high water content (free and combined) which in turn results in high energy consumption of the extraction process. •,As Ni is present in the form of solid solution in laterites, physical upgrading methods such as flotation are not effective in increasing the Ni content before the extraction process. •,Extracting Ni from sulfide ores involve exothermic reactions which are beneficial due to decreasing the overall energy requirement of the process, however there is not efficient release of energy in the process of extracting Ni from laterite ores."

Jan 1, 2017

By Hans E. Madeisky, Bruce W. Downing

The acid rock drainage (ARD) buffering capacity of mine waste is governed by its bulk chemical composition and its mineral assemblage. Conventional acid base accounting (ABA) tests can be used to predict short-term buffering, but should be used with caution to model or predict the long-term ARD buffering capacity of mine waste. ABA data for mine waste samples from four different deposit types: an orthosite-hosted magmatic sul?de nickel, dunite-hosted magmatic sul?denickel, porphyry copper-gold, and volcanic-hosted high sul?dation epithermal silver, were compared with whole-rock geochemical and petrographic data obtained from the same samples. Our objective was to devise and test a method of predicting ARD buffering capacity based on bulk chemistry and modal mineralogy. In other words, we developed a quantitative theoretical method for calculating neutralizing potential, relying on easily obtainable and relatively inexpensive whole-rock and sulfur XRF analyses, which could substitute for conventional ABA procedures in situations where these procedures are either too costly or too time consuming. For each deposit type, models were devised that reasonably approximate the net neutralization potential estimated by the ABA data obtained from laboratory tests. Once an appropriate model has been developed and tested, portable X-ray ?uorescence equipment, already used for grade control at many mine sites, could provide the whole-rock and sulfur analyses required to estimate the ARD potential of the waste material. Although this approach is not intended as a substitute for the legally mandated ABA test procedure, it could facilitate the timely and cost-effective management and mitigation of ARD at a mine site.

Jan 1, 1997

By Hans E. Madeisky, Bruce W. Downing

Abstract-The acid rock drainage (ARD) buffering capacity of mine waste is governed by its bulk chemical composition and its mineral assemblage. Conventional acid base accounting (ABA) tests can be used to predict short-term buffering, but should be used with caution to model or predict the long-term ARD buffering capacity of mine waste. ABA data for mine waste samples from four different deposit types: anorthosite-hosted magmatic sulfide nickel, dunite-hosted magmatic sulfide nickel, porphyry copper-gold, and volcanic-hosted high sulfidation epithermal silver, were compared with whole-rock geochemical and petrographic data obtained from the same samples. Our objective was to devise and test a method of predicting ARD buffering capacity based on bulk chemistry and modal mineralogy. In other words, we developed a quantitative theoretical method for calculating neutralizing potential, relying on easily obtainable and relatively inexpensive whole-rock and sulfur XRF analyses, which could substitute for conventional ABA procedures in situations where these procedures are either too costly or too time consuming. For each deposit type, models were devised that reasonably approximate the net neutralization potential estimated by the ABA data obtained from laboratory tests. Once an appropriate model has been developed and tested, portable X-ray fluorescence equipment, already used for grade control at many mine sites, could provide the wholerock and sulfur analyses required to estimate the ARD potential of the waste material. Although this approach is not intended as a substitute for the legally mandated ABA test procedure, it could facilitate the timely and cost-effective management and mitigation of ARD at a mine site.

Jan 1, 1997

By W. Griffith

"While oxalic acid precipitation practices are commonly applied on purified rare earth elements (REEs) pregnant leaching solutions (PLS), few efforts have been reported using oxalic acid (OXA) to recover REEs directly from untreated PLS. High oxalic acid demand is the most significant issue in all these direct oxalic precipitation (DOP) practices. Preliminary research on DOP of REE from a dilute PLS originating from an acid bake followed by water leaching was conducted. At 75°C, the oxalic acid demand was between 12 and 18 times the stoichiometric requirement for complete REE precipitation. Conducting DOP at 25°C had superior REE recovery but the product purity and filterability declined. Between 75°C and 100°C, REE recovery increased with increasing temperature. When sodium hydroxide was used to adjust the pH during DOP, in the range of 0.5 to 1.5, higher pH benefited REE recovery. Using ammonium hydroxide in place of sodium hydroxide, the REE recovery was reduced by close to 5%. Sodium chloride addition to the PLS was also investigated and the results showed that at both 75°C and 25°C, the REE recovery was enhanced substantially. At 25°C, when the oxalic acid addition was controlled at 4.7 times the stoichiometric requirement, the overall REE recovery reached 95% with the addition of 240 g/L NaCl. Other reagents such as KCl, NH4Cl also showed significant enhanced REE recovery at the same oxalic acid addition. The addition of sodium sulphate appeared to largely enhance the precipitation of LREE but inhibited the formation of HREE oxalate. All results indicate the formation of rare earth double salts may play an important role in REE DOP sulphate media.INTRODUCTIONThe conventional routes for recovering rare earth elements (REEs) from a sulphuric acid pregnant leaching solution (PLS) involves an initial purification step for removal of iron, aluminium for example by precipitation through neutralization, followed by precipitation of REEs as hydroxide, oxalate or carbonate salts, or by bulk solvent extraction for REE recovery (Xia, 2013). Another route was used by Bayan Obo in their historical flowsheet, where the PLS was first subjected to double sulphate precipitation, then the remaining REEs, mostly heavy rare earth elements (HREE), were recovered by oxalate precipitation (Li, 2011)."

Jan 1, 2016

By J. R. Goode

ABSTRACT Aurora Energy Ltd. is considering a 10,000 t/d process plant to produce 97 million lb (43,998,460 kg) of U3O8 over a 17-year project from uranium deposits in coastal Labrador. This paper summarizes the test work, mostly done by SGS Mineral Services in Lakefield, Ontario, and the economic studies supporting flowsheet selection. The proposed flowsheet includes semi-autogenous grinding (SAG) and ball milling, acid leaching using air/SO2 as an oxidant, and resin-in-pulp (RIP) extraction of uranium from the leached slurry. Other unit operations examined include ore sorting, heap leaching, liquid- solid separation, solvent extraction, and nanofiltration for eluate upgrading. We also review the extensive environmental test work and studies that were completed.

Jan 1, 2012

By G. Ellie, D. F. Coates

"The mining of coal in thick and inclined seams is difficult, and there are only a few countries in the world that are really interested in the problems associated with their exploitation. In France a considerable amount of investment has been made in the mechanization of systems appropriate for such conditions.For seams inclined up to 50 degrees, of normal thickness, a new system was introduced in 1967. It consists of a longwall advancing to the rise with walking props, a shearer and hydraulic backfill. During the first two years the panel productivity was an average of 25 TN/MS, and they foresee the possibility of reaching an average of 30 TN/MS.A system for thick beds, both horizontal and inclined, is in p•rocess of being mechanized. In 1969 they started with a method consisting of a longwall in a footwall slice, or sub-level drifts with walking props, and caving and drawing up to a height of 20 m of coal. Recent experiments with a shearer have produced panel productivities between 9 and 11 TN /MS, and there is hope that with experience an average of 16 TN/MS will be obtained.In 1964 a method was initiated for beds inclined from 60 to 90 degrees, based on a mechanical cutter designed by the collieries in Lorraine. An overhand slice of 4.2 m in height is taken with a drum cutter mounted on the end of a boom. Hydraulic fill is placed at the completion of a slice. The face productivity, which has been from 13-20 TN /MS, can be perhaps increased to an average of 22 TN/MS. For those companies with important reserves in inclined and thick seams, these French systems may be worth the trouble of examining in some detail."

Jan 1, 1972

By Jason Mewis

Do you have it figured out? ? If so, I want to network with you! ? There?s a lot to learn as an industry The more we share the better we?ll all become at dealing with this subject If you don?t have it figured out ? You?re among the majority in the in0dustry ? Project Risk is a broad and complex issue that is dealt with through many processes ? And it?s not intuitive either ? And you don?t want to confuse one process with Another

Nov 1, 2011

By A. R. Miller, R. A. Stanton, M. J. Male, G. R. Cluff

"The Dubawnt Group consists of continental sedimentary and volcanic rocks which fill structural depressions - Baker Lake Basin, Yathkyed Lake and Angikuni Lake subbasins- whose evolution was controlled by late Aphebian extension, between 1.9-1.8 Ga. Structurally controlled uranium mineralization having the signature U + Mo + Ag + Pb + Cu ± Zn is spatially related to (i) intersecting fault systems that record multiple movements, (ii) chemically reactive lithologies containing sulphide, chlorite or graphite/ carbonaceous matter in the Archean and Aphebian basement complex and as alkaline volcanics of the overlying Dubawnt Group and (iii) areas containing thick successions of basal Dubawnt Group red clastic sediments.The most significant uranium occurrence found to date is the Lac Cinquante deposit, which contains 14 million pounds U3O8 of drill-indicated reserves. Mineralization is hosted in Archean meta-volcanic and metasedimentary rocks of the Henik Group adjacent to the basal Dubawnt unconformity in the northeastern A ngikuni Lake subbasin and occurs as plunging pods in and peripheral to a fractured and brecciated tuffaceous meta-sediment- sulphide+ carbonaceous exhalite. This deposit exhibits many features typified by the classical vein deposit at Beaverlodge, northern Saskatchewan.Epigenetic stratabound U + Cu + Ag ± Mo mineralization, typified by the 74-1E prospect in the eastern Baker Lake Basin, is hosted in thermally altered Kazan redbeds adjacent to alkaline dykes.At the eastern end of the Baker Lake Basin, small prospects having an elemental signature U + Se + Cu + Ag +Au are interpreted as unconformity-related mineralization associated with the Helikian The/on Basin.The Dubawnt Group rocks of the Baker Lake Basin, subbasins and adjacent basement complex contain uranium mineralization of two eras- Late Aphebian and Helikian."

Jan 1, 1986

By Sylvie Lévesque, Frederic Couët, Laura Pérez-Barnuevo, Cory Hennessey

Determining mineralogical characteristics of samples throughout the iron ore concentration process – from drill core samples up to the concentrate – is critical to process improvement and remediation. However, assessment of mineral proportions, liberation, and texture, although essential, can be time-consuming and costly. Thus, a mineralogical study is often not performed or only restricted to a limited number of samples. New iron ore deposits or new ore zones in active mines will bring new challenges as they are known to be problematic due to their lower grades, complex textures and presence of mineral impurities. In those cases, applied process mineralogy will be necessary to understand and to diagnose potential processing issues. To take full advantage of mineralogical information along the iron ore concentration process, new tools are being developed to provide mineralogical data based on texture, composition and liberation. In an early stage at the mine, the automated identification of drill core textures could provide essential information for the prediction of ore concentrability and grindability. In a later stage, the automated modal tool, combining the knowledge of the iron ore mineralogy of a specific deposit and the results of chemical analyses, can quantify the mineral proportions of the ore; results show good agreement between mineral estimation via chemical analysis and modal analysis using a Mineral Liberation Analyser. A bulk mineralogical analyser allowing semi-quantitative evaluation of the liberation of iron oxides is being validated. The results from the bulk mineralogical analyser will eventually be linked with the mineral proportions issued by the automated modal tool to provide complete mineralogical characteristics of iron ore samples. Ultimately, these mineralogical tools could provide valuable and timely insight for predicting concentrators’ performances or troubleshooting gravity separation processes.

Jan 1, 2016

By L. E. Djingheuzian

Introduction Cylindrical grinding mills have been used for a long time, but during many years of testing no methods were established that gave consistent results. The work by both scientific investigators and operators led very often to contradictory conclusions, and each millman was very much 'on his own' when trying to find the limits for working of his ball-mills. Fred C. Bond, in the introduction to his paper on Wear and Size Distribution of Grinding Balls, freely states that the mathematical immaturity of the grinding process "is the cause of many inefficient designs and installations as well as of much poorly oriented experimental work", and evidently he sees the solution of all our grinding problems in the discovery of the fundamental laws "which can change the empirical art into a mathematically exact science". While the writer is in complete agreement with Mr. Bond that grinding is governed by fundamental laws which can be expressed by mathematical formulae, there are so many variables in grinding, all interwoven with each other and mutually influencing each other, that the coalescing and merging of all these variables into a harmonious mathematical whole appears to be a homeric job. However, a light appeared on the horizon when it was proposed that power be used as an index of the grinding operation. It is pertinent to quote Gow, Guggenheim, Campbell, and Coghill from their paper on Ball Milling, in which they say: "Much of the past literature devoted to the subject has been concerned with the Kick-Rittinger dispute over evaluating grinding, or with dissertations on ball paths and the closely allied impact versus attrition hypotheses or with the exclusive consideration of certain specific variables, such as mill speed or circulating load, irrespectve of the other factors in operation. Such subjects may be dismissed with but little discussion in what may be called, by comparison a new philosophy in the stud of ball milling. Briefly this 'new philosophy' consists ' of using power as an index of the mill operation, and investigating grinding from the standpoint of set and induced variables with emphasis on the latter". But it is safe to assume that it was the work by E.W. Davis, H. E. T. Haultain, F. C. Dyer, A: M. Gaudin, J. Gross, S. R. Zimmerley, and others, that cleared the path for the authors of Ball Milling and induced them and A. W. Fahrenwald to look farther afield and thus to establish a far-reaching relationship between the power and capacity and the diameter of the ball-mills.

Jan 1, 1949

By D. B. Dreisinger, M. R. Tavakoli

"Vanadium has been produced from a variety of sources. This metal is known to make aqueous solutions of complex speciation as a function of vanadium concentration and pH. Considering the method of leaching, the impurities in the pregnant leach solution, and the planned vanadium product, different extractants and flowsheets can be applied for solution purification. This paper reviews the purification of vanadium leaching solution using organophosphorus, solvating, and amine extractants. This review focuses on vanadium separation from iron, molybdenum, and uranium, which are often present in vanadium resources such as primary ores, fly ashes, and spent catalysts.RÉSUMÉ Le vanadium est produit à partir de diverses sources. Ce métal est connu pour permettre des solutions aqueuses de spéciations complexes en fonction de la concentration du vanadium et du pH. En tenant compte de la méthode de lixiviation, des impuretés dans la solution mère de lixiviation et du produit de vanadium planifié, divers agents d’extraction et schémas de traitement peuvent être mis en oeuvre pour purifier la solution. Le présent article étudie la purification de la solution de lixiviation du vanadium au moyen d’extractants organophosphorés, solvants et aminés. L’étude cible la séparation du vanadium du fer, du molybdène et de l’uranium, lesquels sont souvent présents dans les ressources de vanadium telles que les minerais primaires, les cendres volantes et les catalyseurs usés."

Jan 1, 2016

By M. Tanaka, T. Oishi, K. Koyama, S. Alam

"A novel hydrometallurgical process for copper recovery from waste printed circuit boards has been proposed, which consists of selective leaching of copper metal as Cu(I) by Cu(II) in ammoniacal alkaline solution, purification, and electrowinning. The distinct feature of this process is much lower power consumption during the electrowinning than that during the conventional one from acidic copper sulfate media, because the deposition of copper metal from Cu(I) at the cathode and the oxidation of Cu(I) to Cu(II) at the anode take place in the present method. In the laboratory-scale tests using waste printed circuit boards, high purity copper which contains 1.1ppm of impurities was obtained in a chloride system. INTRODUCTIONThe recovery of various metals from waste materials is an important issue in order to conserve the environment and metal resources. Copper is used for cables, electric appliances, etc., and is currently a very important material for industries and will continue to be in the future. Among the wastes that contain copper, one of the most complex recycling materials is electronic scrap like printed circuit boards (PCB), which generally contain 10-30% copper as well as plastics, fibreglass and other metals. There are two ways to recover copper, either pyrometallurgical or hydrometallurgical. Hydrometallurgical treatments generally have more flexibilities in the process scale and control. The separation processes tend to become complicated because of the poor selectivity during the leaching in the sulphuric and hydrochloric acid media. On the other hand, the use of ammonia as a complexing agent, gave a higher selectivity for copper leaching (Nigo, Majima, Hirato, Awakura, & Iwai, 1993; Majima, Nigo, Hirato, Awakura, & Iwai, 1993; Zhou, Shinme, & Anezaki, 1995; Koyama, Tanaka, & Lee, 2006) .In this study, recycling process containing leaching, copper deposition at a cathode and oxidation of copper(I) at an anode from ammoniacal alkaline solutions containing Cu(I) or Cu(II) is introduced. Furthermore, copper recovery from a waste printed circuit boards and impurity behaviour was examined in ammonia-ammonium sulfate (sulfate system) and ammonia-ammonium chloride (chloride system) systems."

Jan 1, 2012

By L. E. Djingheuzian

"IntroductionDURING the past two decades an enormous amount of empirical data has been accumulating on grinding. These data have been obtained by mill operators on grinding ores with a diversity of characteristics, and generally they have been published. However, whenever an operator, who has started milling on a new ore, has tried to find a case in literature which would exactly fit his particular grinding problem, he has failed. Evidently, something fundamental has been lacking in all these empirical data, and, as a consequence, each new operator has had to undertake his own experimenting, get the best results he can, maybe publish these results, and let it go at that. This, in the writer's opinion, has been• due to the fact that even the ablest and most inquisitive millman has been forced to look on grinding as an art, and so, except very rarely, has not been able to find the time to stop and think; if he did so, it was rather in a philosophic way, that grinding is a fundamental process, a process that takes place in river beds, on the sea shore, etc.Nevertheless, when all the data so far published are correlated and examined as a whole, one is struck with the advances that have been made in grinding. The writer (7) had made such a study in 1949, and, when the correlation was completed, it showed that there was nothing heterogeneous about the accumulated data and that when brought together they were in harmony with each other. The writer's conclusion was that, if grinding criteria were defined and were expressed in such a way as to make integration of any data into basic formulae possible, grinding would be established as a mathematically exact science, and grinding worries of the harassed mill operator would be at an end!The first attempts to rationalize comminution were made by Rittinger and Kick. However, their theories led to endless disputes which did not accomplish much, except possibly clearing the path for subsequent investigators."

Jan 1, 1952

By A. Hashempour, D. Javani

"Selection of an optimal bit is a critical issue in drilling engineering. Specific energy (SE) and Cost per Foot (CPF) are the most common methods for evaluation of bit efficiency and optimal bit selection. The major shortfall of both methods is that they do not directly account for geological and geomechanical parameters of formations being drilling. In this paper, results of using the two foregoing methods are compared using Formation Drillability (FD) method, which directly accounts for geological and geomechanical parameters of a formation. This method combines various geological and geomechanical parameters and yields a dimensionless parameter indicating the formation drilling hardness (drillability). Results of this study show that the amount of energy, cost and time spent for drilling a unit volume / length of an identical formation may vary in different wells. This is due to the fact that geological and geomechanical characteristics of a formation may vary among wells and even within short intervals of a single well. Therefore, an optimal bit may not be appropriately selected, if drilling and economical parameters are the only factors considered and the formation geological and geomechanical specifications are ignored.INTRODUCTIONMany authors such as Bilgesu, Al-Rashidi, Aminian, and Ameri (2000) and Clegg (2006) conducted research in the evaluation of bit performance and optimum bit selection. Almost all of them tried to establish a good relationship based on geomechanical parameters. Primary examples are the research done by Ubaldi (1999), Macini, Magagni, Da Dalt, and Valente (2007)."

Jan 1, 2015

By Rob Kelso

What is a Lunabot? ?Robot Controlled Remotely or Autonomously ?Visual and Auditory Isolation from Operator ?Excavates Black Point 1 (BP-1) Simulant ?Weight Limit - 80 kg ?Dimension Limits - 1.5m width x .75m length x .75m height ?Designed, Built and Tested by University Student Teams

May 1, 2013

By Reis. E., P. Pino Véliz

"Along the process of defining the fundamentals of a ferronickel project, several questions require proper answers from the very beginning such as the size and characteristics of the laterite deposit (oxide or silicate), the mineralogical variability, and the selection of the most valuable process route. In general, these elements will be defined along the progress of the exploration and feasibility studies, and definitive engineering. Also, from the very preliminary estimates, three questions are very common: what should be the plant capacity (annual tonnes of contained nickel), what will be the probable capital required (CAPEX), and what will be the estimated timeline (schedule) between the initial stage, say the feasibility study and the plant start-up and ramp-up to designed capacity. The accuracy of the capital cost estimates also progresses along the different stages of development from ±50% at the Preliminary Economic Assessment (PEA) stage down to ±5% for the Detailed Engineering and Construction stage. These accuracies of estimates have been, in some ways, accepted by the mining community and investors. The industry accepted benchmark for ferronickel projects in terms of CAPEX are then in the order of US$50,000 per tonne of nickel contained in the FeNi product. In this paper, the authors analyze in some details three ferronickel projects for lateritic ores that have selected the Rotary Kiln - Electric Furnace (RKEF) as the suitable technological route and have started production in the period 2006–2016. Some key aspects of the ferronickel projects are examined: 1) size and type of deposit; 2) flowsheet selection; 3) evolution of capital cost as a function of capacity from the feasibility stage to construction; 4) comparison between planned production scale and actual ramp up; 5) estimated capital expenditure (CAPEX) compared with actual figures; 6) nickel price evolution and timing of the approval for the project to proceed to construction and production start-up. The objective of this paper is to identify key elements that should be addressed in future projects and that are proposed as benchmark figures for similar projects."

Jan 1, 2017

By P. T. Bloomer

Introduction Mining of the orebody of the Sullivan mine, which is owned and operated by the Consolidated Mining & Smelting Company of Canada, Limited, has reached a point where pillar extraction is rapidly becoming a major part of the operation. Very few of the larger pillars have been completely extracted; consequently, this paper contains much that is theory and expectation. Due to early mining methods and the physical characteristics of the Sullivan deposit, similar conditions may not exist in other older mines facing the problems of pillar extraction, but it is thought that the methods employed and those contemplated may be of general interest. History The Sullivan orebody, which is at Kimberley in the Fort Steele mining division of the East Kootenay district of British Columbia, was discovered in May, 1892, by Messrs. Sullivan, Smith, Cleaver, and Burchett. After passing through the hands of several owners the property was optioned in 1909, and purchased in 1910, by The Consolidated Mining and Smelting Company of Canada, Limited. Early mining was highly selective, the objective being a clean, high-grade silver-lead ore. By 1923 the difficulties of milling the very complex ore had been overcome and the first unit of a concentrator was completed with a capacity of 2,500 tons daily. Capacity was increased by stages to the present rate of 8,500 tons per day. Concentrates are shipped over the Canadian Pacific railway ? to the Company smelter at Tadanac, British Columbia, for final treatment. Approximately 45,000,000 tons of ore has been mined to date.

Jan 1, 1948

By Webber. G. R.

Variation in content of some of the major and minor elements in the Lake Dufault granodiorite near Noranda, Quebec has been studied. The element variation and specific gravity data show a pattern consistent with the concept of a concentrically zoned intrusive on the western side of Lake Dufault with a potassium-rich core and a calcium-rich inner zone. The part of the intrusive on the eastern side of Lake Dufault is of more problematical origin, but the suggestion is made that it has resulted from assimilation of volcanic rock by the intrusive. THE purpose of this investigation was to test the application of x-ray fluorescence analysis to the study of variation in chemical com-position of the Lake Dufault granodiorite. The Lake Dufault granodiorite is an intrusive body approximately six miles long by three miles wide located four miles north of Noranda, Quebec. This intrusive is of particular interest because of its -

Jan 1, 1962

By Zhiwei Wang, Chunsheng Qiao

"Aiming at the problems of the existing calculation methods for the support pressures of shallow tunnels in layered jointed rock strata, by using OAT (orthogonal array testing strategy) and DEM (distinct element method), 12 influencing factors including the thickness of the weathered overburden, and the states of the rock joints were considered. Then, the classification of the loosening zone was analysized. On this basis, by using the DEA (data envelopment analysis) theory, quantitative analysis on the influencing factors was conducted. By analyzing the influence of the key factors, a prediction formula of the boundary of the ringent trumpet-shaped loosening zone was established. According to principles of stress transformation, the formula of support pressures of the loosening zone which could considering the key factors were deduced. Finally, a practical case was introduced and studied to verify the proposed method.INTRODUCTIONIn recent years, various urban subway tunnels are increasingly carried out through jointed rock strata in many cities. Rock masses are discontinuous mediums with fissures, fractures, joints, bedding planes, and faults. The frequency of joints, their orientation with respect to the engineering structures, and the roughness of the joint have a significant importance from the stability point of view. However, there is a lack of understanding about the influence of the joints on the stability of tunnels. Therefore, how to measure the degree of the joints’ influence is a problem that needs to be solved.For the calculation of the support pressures, until now, many research results were focused on soil ground (Bierbäumer 1913; Terzaghi, 1943; Peck, 1969; Sabatini et al., 1999). In terms of site tests and model indoor tests, by measuring strut loads and calculating earth pressure based on strut spacing, the results proposed by Peck (1969) and Tschebotarioff (1973) have been widely used in practice for designing excavation support systems under sand (drained) and clay (undrained) soil conditions. Note that, it is unclear whether these findings could be extended to rock strata. For numerical simulations, how to simulate the states of the rock joints is a key problem that needs to be solved."

Jan 1, 2015