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By M. Souley, G. Armand, M. Ghoreychi, D. Seyedi, J. -B. Kazmierczak

"An anisotropic constitutive model is proposed in this paper accounting for both structural anisotropy and induced anisotropic plasticity. It is assumed that the rock is composed of a matrix and of potential planes of weakness. The matrix is assumed to be linear, transversely isotropic and the plasticity is described by a non linear yield function where the parameters are deduced from the nonlinear Hoek-Brown envelopes in pre- and post-peak, and derived from the laboratory characterization. A non-associated flow rule is used with a distinction between compression and extensional stress paths, as well as the absence of volumetric strain beyond large plastic distortion.The planes of weakness are considered as known a priori or assumed to be oriented perpendicular to the direction of the current minor principal stress. An elastic-perfectly plastic behavior according to the Mohr-Coulomb criterion is assumed in the planes of weakness; while the elastic part is considered as linear and transversely isotropic.Finally, the proposed model was implemented in FLAC3D and used to simulate triaxial compressions to provide a verification of the implementation. The applicability of the implemented model to reproduce damage (pre-peak) and/or failure developments around a circular opening is checked. The GCS drift, one of the mine-by experiments set up at the main level of the Meuse/Haute-Marne Underground Research Laboratory, is selected for this first application."

Jan 1, 2015

By I. Nolet

Matte tapping is one of the most critical operations of a matte smelter. Every smelter has developed particular matte tapping practices and tap-hole repair strategies that are often based on the experience of the smelter personnel. As a result, very little information has been published on these critical topics. This paper aims at providing the information gathered from a survey conducted in the PGM and Ni industries on matte tapping practices and tap-hole maintenance strategies for different PGM-Ni matte smelting operations.

Jan 1, 2014

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Jan 1, 2011

I am interested in experiences people have had with blasting near existing railroad tracks: What kind of vibration criteria are used, and how successful you have been in gaining approval for blasting in the vicinity of railroads. This would both be for passenger and freight railroads

Jan 1, 2008

By Jeff Poe

When a takeup weight is offset to the side of a conveyor, a wire rope is commonly routed from the takeup weight through a series of sheaves to the takeup carriage. The friction developed from a series of wire rope sheaves can effect the actual belt tension at the make up carriage. Through real-mime testing the actual tension on a makeup carriage was recorded with data acquisition using an internally strain-gauged turnbuckle. The wire rope used to transfer the makeup weight to the makeup carriage was routed through a series of lubricated bronzed bushed sheaves. The tension on the makeup carriage was found to be as much as 26% higher than the weight of the makeup due to the friction in the sheaves. A detailed discussion of the testing and results will be made.

Jan 1, 2000

By William P. Kimball, Morris Cohen

This paper deals with the calculation of the composition of the infinitesimal trace of alloy that transfers from the liquid to the solid state at each temperature during the equilibrium solidification of a binary solid solution, and which thereby enables the compositions of the coexisting liquid and solid phases to follow along the liquidus and solidus lines respectively during the cooling. A general expressioll for the composition (6) of the liquid-to-solid transfer at each temperature has been derived: , = Zlxl{xl - P) + ZsXsiP - xs) [I] Zl(xl -P)+ Zs{P - xs) where ZL and Zs are the slopes of the liquidus and solidus lines at the given temperature, xl and xs are the compositions of the stable liquid and solid phases at the same temperature, and P is the composition of the alloy under consideration. The d equation is valid for the process of equilibrium heating as well as cooling, and may be applied to vapor-to-liquid, vapor-to-solid, liquid-to-liquid, and solid-to-solid transfers if the corresponding equilibrium diagrams are known. The d-versus-temperature curves for two alloys P1 and P2 are shown in Fig. I. In each case, the liquid-to-solid transfer has the same composition as the stable solid phase only at the temperature of the beginning of solidification, and has the same composition as the stable liquid phase only at the temperature of the completion of solidification. At each intermediate temperature, the composition of the liquid-to-solid transfer differs by a finite amount from those of the two coexisting phases. Therefore, during equilibrium cooling. the transfers must merge into the solid phase from the liquid without forming a transition phase; otherwise the condition of equilibrium would be violated. At any given temperature during solidification, the composition of the liquid-to-solid

Jan 1, 1941

By H. W. Schreiber, H. R. Babitzke, R. E. Rodriquez, Patterson J. A., C. D. Masters, J. J. Schanz, R. B. McCammon, E. A. Noble, P. M. Krishna

A five-year effort by the ASTM subcommittee E10.14 on uranium resource evaluation has culminated in the approval of E901-82, Classification System for Uranium Resources (ASTM Standard E901-82, April 1983). The standard classification was prepared with participation of representatives from the US Department of Energy, the US Department of Interior's Geological Survey and Bureau Mines, utilities, and consultants. The classification a resource is based on level of knowledge and the economic producibility of the resource at a given time. Alphanumeric identifiers are used rather than words to avoid confusion with existing terms, definitions, and classifications. An example demonstrates the dynamic nature of the process of classification and attests to the necessity of addressing both the aggregate needs of broad-scale resource planning and the specific needs of individual property evaluation. Problems that remain infixing the classification a given uranium resource include uncertainty in estimating the quantity of undiscovered resources and resolving differences that may exist in deciding when the drill-hole spacing is adequate to determine the tonnage and grade of discovered resources.

Jan 1, 1985

By G. W. Luxbacher

Coal mine infrastructure is rarely designed solely on the basis of ventilating the underground workings. Instead the optimization of a mining sequence, the minimization of coal quality variation, reduction of coal losses by adapting to property boundaries and geologic structure (seam inclination, roof conditions, coal thickness, etc.) or other similar considerations have formed the objective for mine plant design and layout. Infrastructure development may also reflect a desire to maximize production early in the life of the mine or to permit flexibility in mining according to market conditions. Requirements for coal mine ventilation, established by the 1969 Coal Mine Health and Safety Act, have mandated changes in the traditional concepts of mine plant design in the United States. Mines, in operation prior to the passage of the Act have experienced severe limitations when attempting to adapt the existing mine infrastructure to meet the requirements of the law. In addition the increased quantities of air required, velocity limitations on trolley haulage and belt lines, the establishment of separate intake escapeways and gob area ventilation have com- pounded the inherent complexity in planning mine ventilation systems. Past practices in mine ventilation planning and design are no longer adequate. Given the complexity of today's ventilation systems, effective planning and design approaches are required. As such, the interrelationship between the overall coal mine plant design and ventilation system design is examined. An approach to ventilation system design constrained by global considerations of mine plant design, is presented. The proposed approach is, however, applicable to the design of new mine ventilation systems as well as to the modification of existing systems.

Jan 1, 1980

By J. D. Bittner

The ST separator utilizes electrical charge differences between materials produced by surface contact or triboelectric charging. When two materials are in contact, material with a higher affinity for electrons gains electrons and thus charges negative, while material with lower electron affinity charges positive. This contact exchange of charge is universally observed for all materials, at times causing electrostatic nuisances that are a problem in some industries. Electron affinity is dependent on the chemical composition of the particle surface and will result in substantial differential charging of materials in a mixture of discrete particles of different composition. In the ST separator (Figures 1 and 2), material is fed into the thin gap 0.9 ? 1.5 cm (0.35 -0.6 in.) between two parallel planar electrodes. The particles are triboelectrically charged by inter-particle contact. For example, in the case of coal combustion fly ash, a mixture of carbon particles and mineral particles, the positively charged carbon and the negatively charged mineral are attracted to opposite electrodes. The particles are then swept up by a continuous moving open-mesh belt and conveyed in opposite directions. The belt moves the particles adjacent to each electrode toward opposite ends of the separator. The electric field need only move the particles a tiny fraction of a centimeter to move a particle from a left-moving to a right-moving stream. The counter current flow of the separating particles and continual triboelectric charging by carbon-mineral collisions provides for a multi-stage separation and results in excellent purity and recovery in a single-pass unit. The high belt speed also enables very high throughputs, up to 40 tonnes per hour on a single separator. By controlling various process parameters, such as belt speed, feed point, electrode gap and feed rate, the ST process produces low carbon fly ash at carbon contents of 2 % ± 0.5% from feed fly ashes ranging in carbon from 4% to over 30%.

Feb 27, 2013

Proper technical appraisal of a coal deposit and the determination of optimum open- pit mine design can only be made after detailed examination of a number of interrelated geographical and geological factors. Initial evaluation of the geography and the overall attitude of the coal measures can often determine the most likely coal mining technique. Conclusions reached in the early stage of the investigations frequently dictate the quality and quantity of particular types of data required from further drilling of the deposit. In many cases, insufficient information relating to the overall engineering aspects of a coal deposit is derived from drilling. Little or no attention is paid in the exploratory phase to such things as the specific gravity of selected overburden samples, floor conditions, clay content of the overburden and selected strength tests on cores. This is a direct result of geologists treating coal deposits purely in geological terms without sufficient consideration being given to the engineering parameters required for a proper evaluation of the deposit.

Jan 1, 1977

By B C. McLellan

The decline of ore grades on land, increasing demand and high prices for various minerals have led to an increased interest in mining the deep ocean sea floor. The potential deep ocean reserves are vast and in some circumstances the grades are high enough to consider mining. However, the consequences of such mining are as yet uncertain. This paper will present the results of a comparative assessment of life cycle impacts of deep ocean minerals. Two alternative extraction technologies from different deep ocean resource types were examined and key commodities are considered. Initial results indicate that the energy requirements will be higher than the average land-based resources, but may fall within the range of the deepest terrestrial mines.CITATION:McLellan, B C, 2015. Comparative life cycle impacts of deep ocean minerals and land-based counterparts, in Proceedings Third International Future Mining Conference, pp 253–260 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 4, 2015

By T. Koulomzine

The theory of the Askania mag-netometer, as well as a complete discussion of all factors influencing magnetometer readings, is very ably described by J. Wallace Joyce.1 We will assume that the reader is thoroughly familiar with this theory and we will use whenever possible the same notations for the physical values involved in the theory. On the other hand, we believe that the principal formulas of the magnetometer can be deduced much more easily by the use of differential calculus than by the methods employed by Joyce. Consequently, we are presenting herewith a resumé of the theory of the vertical magnetometer in order to determine the correct formula of the temperature coefficient which we propose to compensate. Short Theory of the Vertical Magnetometers We shall accept the following notations: m = total mass of the magnet system, in grams. g = acceleration of gravity, in dynes. a = projection on the axis of the magnet system of the distance between the axis of support and the center of gravity of the magnet system, in centimeters. h = projection on the perpendicular to the axis of the magnet system of the distance between the axis of support and the center of gravity of the magnet system, in centimeters. M = 2pL = M, magnetic moment of the magnet system. Z = vertical component of the earth's magnetic field. ? = deflection angle of the magnet system. = focal length of objective lens in the optical system S = scale reading. In a vertical magnetometer the axis of the magnet system is kept horizontal within 2". Therefore, referring to Fig 1, we see that the magnetic torque of the magnet system influenced by Z must be counterbalanced by the gravity torque. MZ cos ? = mga cos ? + mgh sin ? or dividing by cos ? MZ = mga + mgh tan ? If we now move to another magnetic station and the temperature conditions change, the values of all the variables will change also: M, a, and h under the influence of temperature. 2, both with time and space. ? and S, as a result of all the above changes. Taking increments in Eq 2 we obtain: (Mo + ?M)(Z + ?Z) = mgao + mg?a + mg(h + ?h) tan (? +??) but tan (? +??), because ? is small, can be replaced by tan ? + tan ??; in fact tan (?+ ??) = sin ?.cos ?? + cos ?.sin ??/ cos ?.cos ?? — sin ?sin ?? but sin ? sin ???0 tan (8 + ??) ? tan ? + tan ??. Fur-thermore, neglecting all terms of the equation containing two differentials becomes: or a differential and a tan ??, Eq 3

Jan 1, 1950

By S. K. Paul

"This paper briefly describes the underground crushing and conveying system at Sherritt Gordon Mines' Ruttan Mine. A series of three steel-cord belted conveyors, with a total length of 2J55 metres and a total vertical lift of 407 metres, convey the muck from the underground crushing station to the mill at the surface. The maximum size of the crushed rock is 6 in. All the conveyors are provided with numerous safety and fire suppression devices. IntroductionThe Ruttan Mine of Sherritt Gordon Mines Limited is located 20 km east of the town of Leaf Rapids, 760 km north of Winnipeg. The orebody is multilensed, with lens width varying from 6 to 45 m. Ore production started in 1973 with the opening of a 10,000-tpd open-pit mine. Initial development of the underground mine was started in 1972 with the driving of parallel conveyor and access declines at minus-2507o gradient to the I 220-m level, and then the access ramp was further extended to the 1160-m level. Exploration drifting was done on the 1280-m and 1160-m levels from the access ramp. The final development plan for the underground project was approved in January 1977, and full-scale development and constructionwere resumed.At present, Ruttan Mine is in a period of transition from open-pit to underground mining. Ore production is now coming from both operations. By mid-1980, open-pit production will cease and all production will come from the undergroundmine.The underground mining method at the Ruttan Mine is sublevel open stoping with delayed backfilling. The orebody is divided along strike into a series of 27-m-long stopes and 15-m pillars, both 60 m high."

Jan 1, 1980

By Norbert Patzelt

The high-pressure grinding roll was introduced about 10 years ago and has established itself in many applications. Most of these machines are used in grinding circuits in the cement industry. A few operate in conventional mineral-processing plants. Polysius has sold more than 350 Polycoms worldwide. An exception is diamond-ore processing in South Africa and Australia. This type of grinding system increases diamond liberation and reduces the capital costs for the downstream processing plant. There are several reasons why high-pressure grinding rolls have not been used more extensively in the mining industry: - The mining industry adheres to conventional grinding technology and regional preference for certain comminution circuits. - Design, scale-up and flowsheet assessment procedures are viewed as being immature and reduce confidence in the potential application of the technology. - Downstream process benefits are unproven on a plant scale.

Jan 1, 1995

By G. E. Blight

The mining industry is making increasing use of pumped tailings fills. The properties that affect the performance of tailings fills are described by means of typical examples. They are the particle-size grading for maximum density, the water content for pumpability, settlement after placing, compression and drainage under stress, shear strength, and resistance to shock loading.

Jan 1, 1979

By John G. DiGenova

The Allwell Center Ice Arena at Plymouth State University, Plymouth, New Hampshire is located in the flood plain of the Pemigewasset River. To maintain the arena near the level of the surrounding area, foundation design required use of uplift elements to resist hydrostatic loads caused by frequent flood conditions. Soils beneath the arena consisted of very loose to medium dense alluvial deposits overlying thick glaciolacustrine deposits. Bedrock was encountered at depths of 115 feet (35 m) or more. A mat foundation with a sealed building system was developed for the arena to guard against flood events. Several alternatives were investigated to provide uplift capacity for the mat foundation, including the Rammed Aggregate Pier® (RAP) system, driven precast-prestressed concrete piles, driven H-piles, and deep drilled soil/rock anchors. Evaluations indicated that the most cost effective uplift solution was a tension RAP system that developed uplift resistance in the shallow alluvial deposits. A total of 344 55-kip (245-kN) capacity RAP elements were required for the project. The RAPs were designed to resist tensile loads using threaded steel bars. Dual corrosion resistance was provided to the uplift elements by including sacrificial steel and galvanizing the oversized bars and connection plates. A pre-construction test program was conducted on RAP elements drilled to a depth of approximately 25 feet (7.6 m) and installed using various shaft construction procedures and materials. Ten test piers were constructed and six were tested to verify that the piers could resist 200 percent of the design load with measured deflections of less than 1 inch (25 mm). The RAP system also provided improved bearing conditions for the perimeter wall and mat foundations by densifying and stiffening the upper alluvial soils.

Jan 1, 2011

Map

Jan 1, 1952

By Frank Otten, Jochen Greinacher

"In July 2017, Deilmann-Haniel signed a contract with IOO Slavkaliy to sink two 750 metre (m) deep freeze shafts for the Nezhinsky Potash mine in Belarus. This is the second shaft project where this technology is employed after two over 1,000 m deep shafts for a potash project in Saskatchewan, Canada, have been sunk by the application of the Herrenknecht Shaft Boring Roadheader (SBR) technology. Developed to sink shafts for green field projects in soft to medium hard rock, in frozen or unfrozen geology, the machine technology utilizes a roadheader-type cutterhead on a telescopic boom to mechanically excavate the rock, and a pneumatic mucking system to convey the muck from the face into the buckets. This leads to a manless face during the sinking cycle and thus a significant improvement in terms of workplace safety. For the Nezhinsky project, the SBR design was significantly modified in close cooperation between Deilmann-Haniel and Herrenknecht, based on the available experience and the special requirements of the project. Such modifications have been for example the adaptation of the Pneumatic Mucking System to cope with wet material as well as dry as originally planned or the shaft lining moving from shotcrete application to jump for and tubbing installation for which a step-change hands-off tubbing erector for the placement of segments was designed. Further modifications to optimize the machine technology for the requirements during shaft sinking were applied on the two machines for the Slavkaliy project as well as several revisions to match the special project requirements. The SBR is suitable for soft and medium hard rock up to 100 MPa UCS. The main design limits are maximum weight of the machine with 400 tonnes and length or height of 50 m due to the depth of the pre-sink. The Pneumatic Mucking System was changed in all components, sheaves and rope arrangement were redesigned and the concept of the hoist system was also re-engineered. The SBR in Belarus is configured for a daily sinking rate of three metres and all systems behind the cutting boom are not the bottle neck of the whole system. The works on site started in November 2017 including freeze hole drilling, construction works, site setup, installation of hoists winches, permanent headgears and other components required for sinking. The SBRs arrived on site in June/July 2018 and were preassembled on surface and then lowered into the 50 m pre-sink. SBR #1 started full face sinking in mid-December 2018, and SBR #2 in late January 2019. By the end of February 2019, the machines were at 150 m and 100 m depth respectively, and achieved up to 3.6 m/day concrete lined shaft. The SBR and the sinking system, the experiences, problems, technical solutions and performances of this challenging project will be presented in this paper."

Jan 1, 2019

By Y. David

In 1990, the United States became only the second country (after China) to produce more than 900 Mt (1 billion st) of coal in a single year. US production of bituminous and subbituminous coal, lignite and anthracite totaled 933.5 Mt (1.029 billion st) in 1990, 5% more than in 1989. The production gains were largely due to the rebuilding of depleted coal stockpiles by electric utilities and a surge in US coal exports. Domestic coal consumption essentially remained at its 1989 level. A slight increase in electric utility coal consumption was partly offset by a decline in metallurgical coal use. The use of coal in other industries and in the residential-commercial sector remained virtually flat.

Jan 1, 1991

By M. C. Mound

A recent development using XRD techniques, a unique fixed detector, specifically permitting full mineralogical information simultaneously for the first time, specially designed sample preparation and presentation system and Rietveld codes provide a continuous real- time analysis on the true contained mineral polymorphic composition of powder streams such as ore materials, cement, ground clinker or any other mineral stream conducive to XRD analysis. There are several COSMA systems currently operating in the US, Canada, Italy and Australia, proving service and reliability of this concept. The mineralogical information generated continuously in real time is used for automated efficiency, product quality and reduces costs. Quantitative Phase Analysis information is transmitted to the plant DCS via Modbus over RS422 or ethernet. Some limited information is stored in the on-board computers for optional future use at the operator?s option.

Jan 1, 2010