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By E. Miller, J. M. Eldridge, K. L. Komarek

The thermodynamic properties of liquid Mg-Ge alloys have been determined between 1000°and 1500°K by an isopiestic method. Germanium specimens, heated in a temperature gradient and contained in covered graphite crucibles of special geometry, were equilibrated with magrtesium vapor in closed titanium tubes. The crucible design allowed free access of magnesium vapor to the samples during the equilibration to form alloys of magnesium and germanium, but prevented magnesium losses from the crucibles on quenching the titaniuin tubes to terminate the experimental runs, thus preserving the equilibrium alloy compositions. The activities and partial molar enthalpies of magnesium and the integral thermodynamic properties of the system were calculated from the experimental data. THE Mg-Ge phase diagram' shows one congruent melting compound, Mg2Ge, of essentially stoichio-metric composition, two eutectics, and very limited terminal solid solubilities. Very little information is available on the thermodynamic properties of the Mg-Ge system. The free energy of formation of Mg,Ge was recently deter-mined2 by a Knudsen cell technique in the temperature range 610° to 760°C. The standard enthalpy of formation of Mg,Ge was measured calorimetrically by Bever and coworkers.3 The present study was undertaken as part of a general investigation of the thermodynamic properties of the homologous series of Mg-Group IVB systems, i.e., Mg-Pb,4 Mg-Sn,5 Mg-Ge, and Mg-Si. An isopiestic technique was used which was developed by the authors5 for investigating the thermodynamic properties of liquid Mg-Sn alloys. Specimens of the nonvolatile component, contained in covered graphite crucibles, are heated in a temperature gradient in an evacuated and sealed titanium reaction tube, and equilibrated with magnesium vapor of known pressure. The method employs crucibles of special geometry which preserve the high-temperature equilibrium composition of liquid alloys having a highly volatile component such as magnesium on termination of the experimental runs by quenching the crucibles to room temperature. EXPERIMENTAL PROCEDURE First reduction germanium of 99.999+ pct purity (Eagle-Pitcher Co., Cincinnati, Ohio) and 99.99+ pct magnesium metal (Dominion Magnesium Ltd., Toronto, Canada) were used. The graphite crucibles were machined from high-density (1.92 g per cu cm) graphite rods (Basic Carbon Corp., Sanborn, N.Y.) which had a maximum ash content of less than 0.04 pct. The non-reactivity of graphite with germanium at the temperatures used in this study had been previously established by Scace and Sleck.6 The experimental procedure has been previously described in detail.5 The selection of a particular crucible geometry for a run was determined by a combination of imposed experimental conditions, the principle being that more tightly covered crucibles were required to preserve alloy compositions during quenching when higher magnesium pressures and higher specimen temperatures were used. Depending upon the composition range of the equilibrated alloys the source of the magnesium vapor was either pure magnesium or a two-phase mixture of Mg2Ge + Ge-rich liquid of known magnesium pressure. The experimental runs can be divided into the following three groups on the basis of crucible geometry and magnesium source material. Crucibles with Small Holes and Pure Magnesium Reservoirs. The crucible dimensions were identical to those of the Mg-Sn investigation5 except that the hole diameters were reduced to 0.010 in. because of the higher temperatures and higher magnesium pressures involved in the Mg-Ge system. During an equilibration run, magnesium vapor diffused from the reservoir to each specimen through the small holes, one drilled through the crucible lid and two others drilled through graphite baffles positioned vertically inside the crucible between the lid hole and the specimen. Since the magnesium pressure was high, i.e., in the range 117 to 277 Torr, during the equilibration time of approximately 24 hr, equilibration was not impeded by these holes. A specimen composition at equilibrium was fixed by the relative temperatures of the specimen and the reservoir, and by the thermodynamic properties of the system. Upon brine quenching the titanium reaction tube to end a run the vapor pressure of magnesium above the liquid alloys decreased exponentially with decreasing temperature, and the small cross-sectional areas of the holes (4.9 x 10"* sq cm) drastically reduced magnesium losses from the crucibles. Because of its low vapor pressure, germanium losses from crucibles during a run were at most 0.2 mg for pure germanium and correspondingly less for the alloys. This crucible geometry satisfactorily retained the equilibrium alloy compositions on quenching for magnesium-rich (from 3 to 33 at. pct Ge) alloys provided their temperatures were below the melting

Jan 1, 1967

By W. L. Saunders

This paper deals specifically with heading-driving as distinguished from the broader term tunnel-driving. A heading is a pilot or path-finder for the main tunnel. Some headings are complete tunnels in themselves; that is, conditions at times warrant driving a heading the full diameter of the proposed tunnel. A tunnel 10 ft. in diameter might be driven to advantage through a single heading, provided the nature of the material admits, but it usually pays to drive rock tunnels of large diameter through the heading-and-bench system. Much depends upon the material, as in soft ground tunnels of much larger diameter are driven in a single heading. In mines, the tunnels, drifts, cross-cuts, adits, etc., are usually of small diameter, hence the driving of headings applies to completed tunnels in mining-work, except, of course, in cases like drainage-tunnels and main entries leading into coal-mines, where conditions call for tunnels of large diameter, approximating those driven for railway-service. In building aqueducts the tunnels are driven from 6 to 8 ft. in diameter up to 15 or 20 ft., while in railway-service the dimensions vary from 18 to 30 ft. in diameter. There is no department of rock-excavation so difficult as that of heading-driving. This work also belongs to the most expensive class in rock-excavation. A heading is driven directly in the solid. There are no lines of lesser resistance towards which to direct the energy of the blast, but the material must be blown out by main force. The completion of the heading simplifies all the rest of the work of excavation. It is easy, for instance, to enlarge a heading either from above, below or on the sides, by breaking towards the open face through holes approximately at right angles to the axis of the heading; or the enlargement sometimes takes place by holes driven outside of the heading running approximately parallel with it, the rock

Jan 1, 1910

By F. L. Kadey

A discussion of fillers would not be complete without allotting adequate time to that unique and different mineral material, diatomaceous silica. Historical references to its use challenge the aesthetic aspect of man's curiosity; whereas modern scientific tools have provided a dimension of measurement that has motivated many of its more recent applications. A brief review of what is already known is in order. The diatom is a unicellular, plant-like, microscopic organism, related to the algae1. Its natural function appears to be that of a food for other marine or lacustrine organisms, The diatom's "claim to fame" from an industrial viewpoint is dependent on the fact that it secretes an opaline siliceous skeleton that must survive, chemically and structurally, the geologic processes that caused its formation in the first place; its subsequent demise, burial, consolidation, and alteration; and finally, its ultimate exploitation by man--in somewhat that order. In addition to its siliceous nature, the skeletal structure of the diatom possesses a variety of shapes, patterns, and sizes that have not only been the intrigue of microscopists for decades, but have served to set diatomite apart as an industrial mineral; and have reserved for it applications that cannot be satisfied by any other material. Its chemical inertness, its intricate structure, and its intrinsic hardness-coupled with fineness and friability-- are generally known and have accounted for its application in a wide range of filtration and filler uses. The increasing use of the electron microscope has considerably expanded our knowledge of heretofore unseen secondary and tertiary structure of submicron dimension in the diatom valve. The development of electro-sensing methods of particle size measurement which seem to supersede sedimentation and microscopic techniques has portrayed size distribution in a more realistic light. Interest in the surface chemistry of these silica structures has further strengthened our understanding for certain uses. Thus, new tools and ideas have made possible the introduction of diatomite into such sophisticated technologies as that of chromatographic supports, its improved utility in paint applications for the control of gloss and sheen, and its usefulness in the entirely new concept of "Dry Acid Concentrates." Each of these applications has been discussed elsewhere before people particularly interested in analytical chemistry, in paint, and in liquid handling, respectively, where the diatom has been a means to an end. The emphasis here, before the mining fraternity, however is on diatomite as a raw material with reference to the uses mentioned above. It is therefore toward an awareness and understanding of diatomite properties pertinent to these newer filler applications that this paper is directed.

Jan 1, 1967

By Gabriel E. Zárate, Juan C. Guzmán

In Chile there are significant amounts of gold in flotation tailings that could be recovered, being heap leaching the most promising alternative. In order to determine the feasibility of treating these tailings by this method, a detailed study is carried out using 4 representative samples. After defining the best conditions an economic evaluation is done for a plant treating 200 tpd, sensibilizing with respect to parameters such us gold price, tailing grade and capital costs. The results show that it is technically feasible treat these gold tailings by heap leaching with gold extractions over 70 %. It is also economically feasible for gold grades higher than 0.5 g/t and gold prices over 350 US/oz.

Jan 1, 1987

By Manqiu Xu, Peter Wells, Gord Bacon

"A methodology is proposed for milling flowsheet design using flotation rate, selectivity and residence time distribution data. The flotation rate data and the difference in flotation rates between the main minerals to be separated are determined from laboratory batch tests. The residence time distribution (RTD) data is calculated using the tanks-in-series model assuming each flotation cell is a perfect reactor. The rate, selectivity and RTD data can be used to calculate the optimum flotation cell number in a bank for rougher/scavenger in a commercial plant.This methodology is demonstrated in a case study of a nickel sulfide ore. It is shown that miniplant rougher/scavenger data can be closely predicted from a batch test using the proposed methodology.INTRODUCTIONIt has been the trend in the past several decades that large flotation cells are increasingly used. This is largely driven by the economics as ore grades are getting lower and the rate of ore processing must be increased in order to keep mining operations economically viable. The development and evaluation of 250 to 500 m3 WEMCO SmartCell and Outokumpu TankCell flotation machines are well in progress. There are many advantages in using large cells: fewer cells, less floor space, lower power cost (although this is now being questioned for fine particle flotation), simpler circuit, simpler control, and lower auxiliary equipment cost (Arbiter, 1999). As flotation cells are getting larger, the importance of cell hydrodynamics, namely, air/slurry dispersion, power intensity, froth transfer, residence time distribution (RTD) and short-circuiting are being fully recognized (Jonaitis, 1999; Weber, et al., 1999; Gomez and Finch, 2002). It is not the intention of this paper to discuss flotation cell scale-up. The intention is to answer what is the optimum number and size of cells in a flotation bank for a given throughput. Figure 1 illustrates this problem facing design engineers. The same mean residence time required for a given throughput can be achieved in a single large cell, or many smaller cells in a row. A few cells of large size may be more cost-effective but they may not be as effective as more cells of smaller size in achieving the optimum metallurgy."

Jan 1, 2005

This paper gives a first-hand account of what a graduate mining engineer can learn by visiting Atlas Copco in Sweden and gives reasons why all mining engineering students should apply for this excellent scholarship. It will also (attempt to) refute the authorÆs mine managerÆs claim that the trip was a huge junket.

Jan 1, 2006

By J. L. Kohler

Component failures in the electrical control circuits of mining machines account for a large percentage of the total downtime of the machine, Once a failure has occurred it is always a tedious and usually a time-consuming task to locate the failed component. Moreover, the pressure to quickly locate the cause of a delay can lead to compromises in safety. Thus an onboard diagnostic system, essential for an automated machine, would be a very useful addition to existing machines. This report details U.S. Bureau of Mines development of a generic procedure for synthesizing diagnostic systems for electrical-control-circuit failures in mining machinery. A continuous mining machine is used as the testbed to illustrate the application of the developed methodology. Substantive differences among mining machine control circuits made it impossible to achieve a generic diagnostic system, but a generic approach for the synthesis of the diagnostic system was possible. As the research progressed, it became apparent that an algorithmic approach was better than an expert-system-based implementation. A prototype system was constructed and used to evaluate the diagnostic system. Prototype implementation issues are also examined.

Jan 1, 2010

By Phipps P. J

The rate of precipitation of ferric iron as ammonium jarosite, NH4.Fe3(SO4)2(0H)6 from acidic solutions near 100¦C, was found to increase with weight of jarosite seed crystals, and with concentration of ammonium ion, but to decrease with increasing acidity. The particle size of the product crystals was larger at low sulphate concentrations and at high ammonium concentrations. Particles appeared to be agglomerates, with diameters ranging from 0.5 up to 25 um, and composed of individual crystals of diameters ranging from 2 um apparently down to 0.1 um or Lass

Jan 1, 1977

Compared with modern longwall coal mining systems, automation has not been developed very far in the underground metal minin industry. One of the main reasons for this is that present har rock mining systems do not produce ore with a controlled degre of fragmentation. The often poor fragmentation from the stopescauses large extra costs and precludes material which 1-ends itself to automationThe paper investigates the possibility c technology for automatic loading and trapconveyor belt to a shaft dump pocket. With this concept, it will be necessary to control the maximum fragment size by careful blasting design. The degree of fragmentation will have to be finer than is common to-day. The study shows that the belt conveyor mining con-cept with controlled fragmentation may be economically competitive with conventional cyclic loading and transport, even for an annual production as low as 0.5 mill. tonnes.. Fragmentation design and cost calculations have been based on a combination of the Kuznetsov relationship between blasting parameters and mean frag-ment size, and the Rosin-Rammler formula for fragment size distribution.

Jan 1, 1985

Since 2001, the Hybrid Stress Blast Model (HSBM) project members have developed a software suite to model the complete blasting process from non-ideal detonation to muck pile formation. To preserve the physics and improve solution time, the breakage engine uses a combination of analytical models and 2D axisym-metric finite differences to model near-field crushing, coupled to 3D discrete lattice fracturing and distinct element numerical methods to model throw and muck pile development. The model has been validated by comparison with laboratory and field tests in kimberlite. Multiple blasthole simulations are used to demonstrate how changes to blasting parameters can influence downstream efficiencies. Case studies of wall control blasting show that the presplit design must balance the two opposing effects of increased damage with increased charge and decreasing attenuation of the seismic waves with decreasing charge. Modelling of decoupled explosives needs further development. Reducing the charge towards the back of the trim blast results in a much more significant decrease in back damage than altering the timing. The model demonstrates how separation along the weaker planes in a jointed rock can coarsen the fragmentation, leading to inefficient beneficiation, and extends the damage to a distance of at least twice the burden behind from the blasthole, severely compromising the wall stability.

Jan 1, 2012

By K. Baris

Methane (CH4) present in coal seams, generally known as Coalbed Methane (CBM), Coal Mine Methane (CMM) or Coal Seam Gas (CSG), is a major energy source, as it is the primary constituent of natural gas. At the same time, released methane is a potent greenhouse gas (GHG) and a major safety hazard in underground mining operations. Therefore, a considerable effort has been undertaken to develop coal bed methane extraction and utilization technologies that can achieve both environmental and safety benefits. Countries with significant coal reserves and gassy seams, i.e., USA, Australia, Canada, China, India, Poland and Germany have developed mature methodologies and practices to extract, before and during mining, and utilize significant amounts of methane from coal seams. Nevertheless, on a global scale, methane recovery remains untested and many technical and economic challenges require urgent attention. The research presented in this paper, sponsored by the United States Environmental Protection Agency (EPA), is focused on the Zonguldak Basin, the only hard coal region in Turkey. The basin is known to contain very gassy coal seams and consists of three major regions: Armutcuk, Zonguldak and Amasra. Among the three, the Amasra coalfield is considered as the most favorable for CBM recovery, as it comprises of deeper coal seams and where a private company, Hema Enerji A.S. (Hema) is already developing coalbed methane production facilities. This study aims at developing a reservoir model to predict the potential CBM recovery from the CBM-2 well, the only producing CBM well drilled by HEMA Enerji A.S in the Amasra coalfield. History matching was utilized to model the 15-day production data and uncertainty analysis was performed to estimate the optimum values for the reservoir parameters. The produced model yielded an average production rate of 143 m3/day for 10 years. The total amount of CH4 to be produced during this period was 0.53 million m3. Furthermore, the total amount of GHG to be avoided was estimated at 6585 tCO2e.

Aug 1, 2013

By Ernest Y. Baafi

This paper describes the results of a series of comparative studies in which polygonal, IDS, and kriging were compared for their local estimation accuracy. The specific objective of this study was to determine the impact that the particular nature of the mineralization has on the results of such comparative studies. The study was performed on simulated coal deposits and the evaluation parameter chosen was the coal thickness. For this study, many conditionally simulated deposits having different degrees of seam thickness continuity were first developed. Each simulated deposit covered an area of 30,000 ft x 30,000 ft which was, in turn, subdivided into 100 mine planning blocks of 3,000 ft x 3,000 ft in size. Thickness estimation was made for these mine planning blocks using surrounding DDH assays. In all cases tested, e.g., from a perfect continuity to a complete lack of continuity in coal seam thickness, kriging gave the best estimates among the three, followed by IDS and polygonal methods. As expected, the relative efficiency (i.e., superiority) of kriging over other estimators was highly dependent on the nature of mineralization.

Jan 1, 1982

By E. C. Mahan

THE text of my talk was suggested by the invita-tion of your secretary, who said that the excess productive capacity of the bituminous industry was a matter of common concern to engineers and coal operators. Cognizant of the depressed condition of the industry, he tactfully added that in common council we might find wisdom. Certainly it would seem that the application of the best engineering knowledge available would be helpful in solving mining problems. There is no doubt that clear thinking, characteristic of the engineering type of mind, is essential to untangle the shapeless pattern which increased complication, so often mistaken for progress, has developed. There is much thinking today which neither looks forward to a logical combination of ideas nor backward to simple fundamentals. The result of that sort of thinking is hopeless confusion. Lately we have heard much about excess capacity and overproduction, along with their attendant evils. Respecting the bituminous coal industry, these terms have been frequently misused. How many of us have seen them applied to the coal industry but how few of us have understood them. I am reminded of a state-ment which William Green, president of the American Federation of Labor and a former miner from the Ohio field, gave to the Associated Press last month, in which he was quoted as saying: "The coal men must realize that conditions have altered since the World War. To-day they are producing a billion tons a year, with a market for only half that tonnage." I cite this instance as a glaring example of the flood of. loose statements about the bituminous industry with which the public is being deluged. If the general public, or some of the newspapers, had any passion for the facts, this sort of misleading statement would not get very far. Unfor-tunately, in these days, opinions are more sought after than facts, and statements by such a person as Mr. Green and are promptly accepted by a credulous public.

Jan 4, 1928

By Mike Perkins

The Summitville Au-Ag-Cu district in the southeastern part of the San Juan volcanic field has been a major gold producer in Colorado. Mineralization is hosted by the composite South Mountain quartz latite porphyry volcanic dome, which is believed to have been localized by the intersection of the northwest margin of the Platoro-Summitville caldera complex and the northwest-trending Platoro fracture zone. K-Ar age dates indicate that alteration and mineralization were nearly contemporaneous with emplacement of the dome at 23-20 m.y. ago, 6-7 m.y. after collapse of the Platoroo-Summitville caldera complex. The dome consists of at least four major phases including strongly altered quartz latite porphyry; a late-mineral, coarse-grained sanidine quartz latite porphyry; and postmineral quartz latite dikes. The dome was emplaced into intracaldera rhyodacite and andesite. Gold-silver-copper mineralization is concentrated near the top of the dome and is associated with pods, pipes and steeply dipping tabular zones of silicification and advanced argillic alteration. Alteration is zoned outwards from the center of mineralization through argillically and propylitically altered rocks, and is largely confined to the area of the dome. Trace element geochemical patterns for the deposit indicate a central zone of anomalous As-Au-Pb-Ag and low Mn within an exterior zone of anomalous Ag-Au-Pb-Hg and high Mn. The chief ore mineral is native gold which is associated with goethitic iron oxide in the oxide zone. In the sulfide zone, native gold is associated with copper sulfide-sulfosalt mineralization. A genetic model for the Summitville deposit is based on recent models of modern geothermal systems, with which it shows similarities in alteration, ore mineralogy, zoning, and structural patterns.

Jan 1, 2013

By M. A. Whiting, H. Kenyon Burch

One of the advantages presented by electric drive in many classes of work is the ease with which the electric motor can be controlled automatically. In a large number of cases certain features of the control are automatic—for example, the rate of acceleration may be limited automatically or the equipment may be stopped automatically at the limit of travel—but the equipment is started and ordinarily is stopped by hand. In other cases the motion of the machinery is utilized to start, control the speed, and stop the motor automatically, independently of any operator. A considerable proportion of the large mine hoists now in use have certain automatic features, particularly protective devices against overwinding, and, in some classes of electric hoists, devices for preventing excessive acceleration or retardation. The large automatic hoists discussed in this paper, however, are completely automatic, i.e., capable of making their trips without the presence of an operator at the control levers. According to circumstances, various advantages may be obtained . by automatic control, chief of which are decreased power consumption, increased precision and safety of operation, and decreased cost of attendance. The first step in the analysis of a prospective automatic mine hoist is to determine whether automatic operation is feasible at all. If men are to be hoisted, or levels changed, the attention of an operator is required for these purposes; but under some conditions it may be entirely practicable and advantageous to build the equipment so that, while provision is made for hoisting men or changing levels, ore can be hoisted automatically from any one level. If, however, an operator's attention is required every few minutes for changing levels, handling men or drills, or other work requiring hand control, it is obvious that automatic operation between times will not be of any practical benefit.

Jan 1, 1917

Jan 1, 1932

By D. E. Light, J. Y. Somnay

The use of alkyl acid phosphates and their alkali salts as collectors for the uranium minerals brannerite and uranothorite was investigated. In particular a detailed flotation study was carried out using isooctyl acid phosphate as a promoter. Tests were conducted on an Elliot Lake ore analyzing 0.11% U 3 O 8, chiefly as brannerite, and approximately 9% pyrite. Pyrite was floated with conventional reagents; the flotation of brannerite from the sulfide tails was studied to determine the optimum pH and pulp density to employ with isooctyl phosphate. 4 uranium recovery of over 90% was possible with a selectivity index of 8.95 when these tails were floated at 17% solids and a pH of 1.7. The effect of various modifying agents was also investigated. Aluminum sulfate, lactic acid and sodium silicate improved the selectivity of collection yielding selectivity indices of 9.5 to 11.5. Ferric chloride was found to be a depressant for brannerite. A Bancroft ore which analyzed 0.13% U308, and contained uranothorite was also amenable to flotation using isooctyl acid phosphate. The composite uranium concentrate obtained assayed 0.41% U308 at a recovery of 95.3% and a ratio of concentration of 3.24. Previous investigations1,2 on the flotation of brannerite, pitchblende, uraninite and/or uranothorite from various Canadian ores showed that fatty acids, including oleic acid and tall oils, petroleum sulfo-nates and alkyl acid phosphates, were suitable collectors for these uranium minerals. Additional work carried out by the Mines Branch in Ottawa on brannerite ores from the Elliot Lake area3 used the tall oil Acintol FA-1, or FA-2, as a collector. A recovery of 92% of the uranium in a concentrate which contained about 55% of the weight and assayed 0.2% U3O8 was obtained from a feed of 0.1% U308 in this latter instance. Eigeles et a1.4 carried out flotation on pitchblende ores and found that the most promising results were obtained with alkyl acid phosphates and their alkali salts as collectors. Within the homologous series of alkyl phosphates, the best flotation characteristics were displayed by isooctyl phosphates. Further, isooctyl phosphate derivatives were much less sensitive to hardness forming salts than oleic acid, enabling all flotation tests to be made with tap water. This paper describes tests employing alkyl acid phosphates for the flotation of ores containing brannerite and uranothorite. The flotation work described herein was conducted in conjunction with the development of a leaching process for Elliot Lake, Ontario, brannerite ores. In this process pyrite in the ore was concentrated by flotation and roasted, providing sulfur dioxide which was used to produce sulfuric acid in situ in a liquid-solid extraction process. A 40% conversion to sulfuric acid was indicated at a leaching temperature of 80°C; this resulted in the extraction of 93% of the total uranium after a contact period of 6 hr. A rotary kiln was used as a leaching vessel. Preconcentration by flotation was studied in conjunction with the above process to reduce the bulk of solids for leaching and, consequently, capital expenditure. The materials and methods used in this study will be described initially. Then the results will be presented in two sections, the first dealing with the brannerite ore, and the second with the uranothorite. Finally an economic analysis of the results on the brannerite material will be made. MATERIALS AND METHODS The two ore samples examined were from uranium-producing mines. One was a brannerite ore, a uraniferous quartz pebble conglomerate from the Elliot Lake area, which contained 0.11% U3O8, 4.3% S and 4.5% total Fe. Mineralogical studies of similar samples indicated the presence of quartz, microcline, sericite, pyrite and minor quantities of rutile, pyrrho-tite and zircon. The other ore was from a biotite-rich pegmatite in the Bancroft area; feldspar, quartz and biotite formed the major gangue components. Uranothorite was the principal uranium mineral; the head assay was 0.13% U 3 O 8. Preparation of Feed: The ores were crushed to —10

Jan 1, 1963

By Cris Sideco

This paper describes the concentration processes of the Sullivan. Concentrator, including crushing , transportation, sink-float, grinding , flotation and dewatering. Handling of hot muck has had a deleterious effect on most of the processes, resulting in significant operating problems and poorer metallurgy. These problems are described -in detail. THE SULLIVAN CONCENTRATOR is situated in Kimberley, high up in the Purcell Range of the Selkirk Mountains, at an elevation of 3,600 feet, approximately 4 miles from the Sullivan Mine (see Fig. 1). The original plant went on stream on August 23, 1923 at a designed capacity of 2,500 tons per day. This has been increased in stages to its present rated capacity of 11,000 tons per day. The latest stage was in 1949, when several major projects, including the underground crushing plant, the sink-float plant and the installation of a rod mill, were completed. Some 112,800,000 tons of ore have been treated in this concentrator to November 1975. The plant started on a 48-hour/ week schedule, but was reduced to 44 hours/week in October 1946 and down to 40 hours/week by the end of 1949. Except for a 6-month period in 1952 when the mill operated continuously, the 40-hour/week, Monday to Friday schedule was in effect.

Jan 1, 1977

Chairman James F. Kemp :—I call upon Prof. Albert Sau-veur to open the general discussion on the subject of sound steel ingots. Albert Sauveur, Cambridge, Mass.:—I believe that I have the privilege, this instant, of opening one of the most important, significant and far-reaching discussions ever held before this Institute. I believe that our talk this morning will be heard far and wide, not only on this side of the water, but also on the other side. Everywhere, in short, where steel is manufactured. I welcome the opportunity of expressing my mind in regard to the production of sound steel ingots, a subject about which I entertain Rome very decided opinions. I shall not attempt to discuss the relative merits of the various processes that have been described. I shall, on the contrary, confine myself to a general consideration of the important questions before us. So intensely do I feel regarding the production of sound steel ingots that I have put my remarks on paper in my desire to present my views. I have always remembered my astonishment, not to say dismay, some twenty years ago, when I beheld for the first time a steel ingot, a rail steel ingot, that had been split open from top to bottom. There was the pipe! There was an enormous cavity extending from a short distance from the top to nearly one-third of the length of the ingot; and when I was told that manufacturers seldom discarded more than one-tenth of the total length of the ingot it became evident to me that a large proportion of piped, and therefore, defective, or at least weak and dangerous, rails were manufactured; in fact, probably one rail out of six, nearly 17 per cent., of dangerous rails.

Jan 1, 1914

By B. Chareyre, H. Pourpak, L. Scholtès, E. Papachristos

"Numerical methods used to study hydraulic fracturing in low permeability rock masses need to take into account the interaction between the progressive failure mechanism of the intact rock matrix and the pre-existing natural fractures. A fully coupled hydro-mechanical 3D discrete elements model dealing with such interaction is presented here. The ultimate objective is to grasp the role of the rock-matrix failure on the propagation of the main fluid driven fracture under different stress fields and pre-existing fracture orientations. The proposed method provides the spatiotemporal distribution of the induced crack events which can be tracked during the simulations, making possible the identification of the local mechanisms leading to macro-failure in the rock matrix or reactivation of the discontinuity planes. Fluid flow inside a fracture, pressurized crack and hydro-fracking experimental tests were simulated as part of the validation process of the model and, are presented in this paper.INTRODUCTIONWith the expansion of shale gas and geothermal reservoirs exploitation, the technique of Hydraulic Fracturing (HF) has been widely used over the past years (McClure, 2012; Britt, 2012). Despite recent advances in micro-seismicity analysis or numerical modeling (Nagel et al., 2013), the way progressive failure of rocks affects the fluid flow depending on the in situ stress conditions or the rock heterogeneities is not fully controlled yet (Zoback, 2010).In this work we focused on setting up a three-dimensional discrete elements model to study fluid driven fractures propagation. The intact rock is modeled by a discrete element method (DEM) able to reproduce fracture initiation and propagation (Scholtès and Donzé, 2013, Garcia et al., 2013). The pre-existing fractures are explicitly modeled and can be individually defined or be part of a discrete fracture network (DFN) (Harthong et al., 2012). Their mechanical response is fully coupled with the fluid within a HM-DEM framework. The fluid flow and it’s interaction with the intact matrix and the pre-existing discontinuities is modeled through a Pore-scale Finite Volume (PFV) scheme specially enhanced for fractured rock modeling (Chareyre et al., 2012, Catalano et al., 2014) ."

Jan 1, 2015