Search Documents

By W. C. Thurber, C. J. McHargue

THE deformation textures of metals have been extensively studied because of both the practical implications in metal fabrication and the fundamental insight into the behavior of the metal during deformation. Most studies have been concerned with pure metals or solid-solution alloys and relatively little attention has been directed toward more complex alloy systems. Barrett' and Brick2 have summarized work on multiphase systems in terms of the relative ease of deformation of the phases. Thus, if the phases are deformed about equally, each develops its usual texture for the particular fabrication process. This behavior has been reported for Ag-Cu and Cd-Zn alloys3 and 62-38 brass.4 A dispersion of hard particles in a softer matrix interferes with the normal reorientation of the matrix lattice during deformation, thus distorting the texture or in some cases preventing its development. Increasing the amount of carbon in steel has been observed to increase the randomness of the ferrite5 and almost random textures have been reported for A1-12 pct Si eutectic alloy wires.6 It has also been noted that if the second phase is platelike, deformation causes the platelets to line up in a common direction. The present work concerns a quantitative study of the deformation textures in aluminum-uranium alloys containing 5 or 13 wt pct U. Alloys in this composition range are characterized by a pure aluminum matrix in which the intermetallic compound UA14, is dispersed. From a fundamental standpoint, this system affords the opportunity for studying the effects of a dispersed phase on texture development in an essentially pure metal matrix. Further, these alloys are commonly used as fuels for nuclear research and testing reactors, and a knowledge of the preferred orientation may be useful in evaluating physical and mechanical properties as well as response to fabrication. The deformation textures in sheets of aluminum (99.996 pct), Al-5 pct U alloy, and Al-13 pct U alloy were determined for reductions in thickness of 90 pct by cold rolling. The texture in an alloy rod of the latter composition which had been extruded at 455°C was also determined. It can be readily ascertained from the aluminum-uranium phase diagram that the 5 pct U alloy is hypoeutectic and contains 7.3 wt pct (3.4 vol. pct) UAl4. The 13 pct U alloy has the eutectic composition and contains 18.9 wt pct (9.4 vol. pct) UAl,; however, because of nonequilibrium solidification conditions, the typical eutectic microstructure was not developed. The alloys were prepared by open-air melting in graphite crucibles. The slab castings were cropped and cold reduced 90 pct in thickness on a two-high mill. The slabs were reversed after each pass. Although the uranium-bearing alloys exhibited edge cracking, they could be cold rolled to the desired thickness. Small coupons 1.5 by 0.75 by 0.1 in. thick were sheared from the center of the rolled plate with a major axis parallel to the rolling direction. To obtain a specimen of sufficient thickness for machining of the X-ray diffraction specimen, three coupons were laminated into a single three-ply sandwich with a suitable adhesive. Adhesives which could be cured at room temperatures were selected in order to prevent recrystallization during bonding. Bondmaster M-648 (Rubber and Asbestos Corp.) was used for the pure aluminum laminate and Armstrong A-6 (Armstrong Products Co.) was used for the alloys. The alloy for extrusion was cast in a cylindrical graphite mold! cropped and machined into a 3-in. diam by 4-in. long billet. This billet was preheated to 455°C, upset to 3.125 in. diam, and extruded on a 700-ton hydraulic press at a speed of 6 fpm through a flat-face, 1-in. diam die. The rod was quenched with a water spray as it emerged from the die. Pole-distribution data were obtained using the diffractometric method described by Jetter and Borie.7 A spherical X-ray diffraction specimen of

Jan 1, 1961

By H. J. Gerritsen

From presently available components a portable, rugged, reliable apparatus can be built which will be able to detect methane concentrations of 0.1% and lower in air. Sensitivity and design considerations of the apparatus are given. Ways to build even simpler and cheaper apparatus in the future are indicated. Traditionally the problem of detecting and monitoring the presence of dangerous gases present in small concentrations in air has been a difficult one. Whether canaries,' or sophisticated chemical or thermal conductivity methods are used, they seem to be either slow, cumbersome, or otherwise not completely satis factory. A seemingly natural way of gas detection is to measure the strength of absorption undergone by light of a wavelength that the gas to be detected absorbs. This method has been difficult to apply in view of the low power output over a narrow wavelength range in the infrared from traditional glow bar light sources. Besides, the apparatus to select such a narrow band of wavelengths is quite costly and complex. The use of a laser solves these problems simultaneously. Although a variety of lasers are available to detect a variety of gases (such as carbon monoxide or methane) whose early detection would be desirable, this paper deals exclusively with the use of lasers for methane detection. ABSORPTION SPECTRUM OF METHANE IN AIR The absorption spectrum of pure methane is very well known.2-4 There are four fundamental vibrations, all occurring in the infrared of which only two absorb strongly. They occur at 3.3 µ and 7.7 µ. The absorption spectrum is quite complicated due to the fact that the methane molecule can rotate in a variety of rotational quantum states. These rotational states are furthermore split by anharmonic forces such as the Coriolis force.4 When methane is present in air at atmospheric pressure, much of the very fine structure due to the anharmonic force is washed out since the pressure broadening is of the same magnitude as the splitting due to anharmonicity. For most of the spectrum at 3.3 µ, the rotational lines are typically separated by 0.01µ while the individual linewidth is 0.0001 µ. It is thus very important that the laser emit at a wavelength that coincides with one of these rotational lines. Before elaborating on the point that such a laser exists and how it performs, it is necessary for later discussion to point out that in the

Jan 1, 1967

By K. A. Grace

K.A. Grace's article advocates usage of the terms describing resources as defined by the US Bureau of Mines (USBM) and the US Geological Survey (USGS). Originally recommended in 1943, the terms were revised in 1976 and again in 1980. Mr. Grace further recommends that the term reserves should be severely restricted and that it should encompass mining and processing factors. The list of references quoted in the article does not, however, mention the study concerning reserves that was made over a period of almost four years by a joint Committee on Definitions. This committee was sponsored by three societies - the American Institute of Mining, Metallurgical, and Petroleum Engineers (AIME), the Society of Economic Geologists (SEG), and the American Institute of Professional Geologists (AIPG). Most of the detailed work was done by a working party appointed by the Society of Mining Engineers (SME) of AIME. After studying previous definitions in the US as well as in the United Kingdom, Canada, Australia, France, and other European countries, the following definitions were proposed and accepted in 1974 by the SME board of directors. (Ore reserve definitions in Rhodesia, South Africa, or New Zealand were not mentioned as having been reviewed.) Table 1 lists definitions from MINING ENGINEERING, July 1975. Notice in Table 1 it was mentioned, under Resources, that the definitions of reserves were recommended for the purposes of commercial evaluation. It was further stated that the terminology adopted by the USBM and the USGS, as released to the public, covering resources, was recognized as such. Mr. Grace also recommends that reserve calculations should allow for future losses in ore tonnage, ore grade, milling, and smelting. Unfortunately, it might take years to develop a prospect after it has been discovered. Mining methods have to be tested before being finalized in order to know what tonnage of ore is going to be lost, or how much dilution is likely to occur. The same problems are present in concentrating or in smelting. Operating mines usually know how mined grade compares with ore reserve grade. They can usually predict fairly closely what their final metal recoveries are probably going to be. It is, however, expecting too much to have to delay quoting reserves from say a recently drilled prospect. Any responsible mining geologist or mining engineer (or company for that matter) should state several things about an ore reserve calculation: that it is or is not a preliminary estimate only; that estimated grades have not been diluted to allow for a future, as yet unknown, mining method; that metal recoveries are or are not only tentative, based perhaps on drill hole samples only or on no metallurgical testing at all. Reserves of a partially explored prospect are always quoted when it becomes necessary to raise capital. If they are described as preliminary only (really just tonnage and grade estimates and not true "reserves") in addition to being accompanied by some of the items just mentioned, nobody is going to be misled. If, on the other hand, a report by a member of the SME or SEG is intentionally misleading, the individual must be reported and eventually disbarred from his or her society. To summarize, let us stay with the definitions of ore, ore reserves, and related terms as recommended and accepted by SME. I might add that on occasion I have used a fourth category in my ore reserve calculations, namely Geologically Possible Ore. This category is more or less the same as Inferred ore defined by the USBM and USGS. I have never given a grade to this category, which may be based only on geological mapping, geochemical sampling, or geophysical traversing.

Jan 8, 1985

[Module 10. Internally Cued Operating Contingencies Prerequisite: Modules 1-7 - Estimated Instructional Time: 2.42 hrs. Module Abstract Trainees, using programmed instruction, describe the cues, consequences, operational sequences, and possible causes resulting from loss of air pressure from power fluctuations, and from a failure of the automatic lubrication system. Although some of the contingencies may be initially signaled by an alarm or from the oiler, each trainee will demonstrate proficiency in the appropriate operating techniques and procedures. This module also includes a review component for those trainees identified in the assessment exam as requiring less detailed instruction.]

Jan 1, 1976

By A. Vifian, I. lwasaki

Several samples of copper-nickel ore from the Duluth Gabbro were studied to relate their mineral-ogical characteristics with their amenability to concentration by flotation. The most common rocks in the samples were norites, gabbros, and troctolites. The texture of the rocks varied from coarse pegma-titic to very fine. The copper contents ranged from 0.2-0.7% and the nickel contents from 0.07-0.15%. The ore minerals were chalcopyrite, cubanite, pyrrhotite, and pentlandite. Most of the ore minerals occurred in interstices between laths of plagioclase, but some were included in the gangue grains. The samples responded readily to xanthate flotation. After a 65 to 100-mesh grind about 95% of the copper and 85% of the nickel were recovered in a bulk sulfide concentrate. Such recoveries remained constant even with grinds as fine as 270 mesh due to the extremely fine inclusions of sulfides in the silicate gangue. A finer grind than 270 mesh was necessary to free the copper-iron sulfides from the nickel-iron sulfides. A combination of time and British gum* was quite effective in producing concentrates analyzing 25-30% copper. The nickel concentrate averaged 1.5% nickel. Most of the nickel lost in the tailing was present as un-liberated sulfides and not as nickel in the olivine structure. The nickel content of the olivine was only about 0.05 or 0.08%. Western world nickel usage in 1967 is expected to be the second highest on record, amounting to 825 million lb. Despite a rise in mine production, however, demand is expected to exceed supply in 1968. ' Hence, exploration, development, and production expansion projects are being accelerated on a world scale. In Minnesota the copper-nickel bearing Duluth Gabbro is receiving much attention. During the past year 11 companies have been granted leases in the region, and the 1967 Minnesota legislature enacted a tax law designed to encourage the development of a copper-nickel industry. Although it is only during the past few years that mining companies have shown an active interest in the Minnesota copper-nickel deposit, copper-nickel minerals have been known to exist in northeastern Minnesota since 1899.3 At the 28th Mining Symposium, Dr. P. K. Sims, from the Geological Survey of Minnesota, gave a review of copper-nickel exploration in northeastern Minnesota,4 in which he emphasized that the potential of the Duluth Gabbro complex is large. Tests on samples from the Kawishiwi River area were reported by the United States Bureau of Mines in 1955.' In 1964, the Mines Experiment Station described tests on samples from the Gunflint Trail: and in 1966 the Minnesota Geological Survey published a geological map of the Gabbro Lake Quadrangle. The Mines Experiment Station published a study of the concentration of a Minnesota copper-nickel ore in 1967.7 The present paper is divided into three parts. The first describes the mineralogy of some rock samples from the Kawishiwi River area of northeastern Minnesota; the second summarizes the results of numerous flotation tests that were made on the samples, and the third explains the loss of nickel in the flotation tailing. MINERALOGY OF KAWlSHlWl ROCKS Recently five samples of mineralized gabbro from the Kawishiwi River area, a few miles east of Ely, were received at the Mines Experiment Station. One of the samples was a 175-lb composite, which was designated Ore 1556. The other four samples were 7-lb grab samples that had been obtained from four different exploration pits. They were designated Ore 1626, Lots 1 through 4. Basic Mineralogy: Many authors have reported that the rocks of the so-called gabbro of the Duluth Intrusion vary widely in composition.8-9 The samples on which our studies were based also contained a wide range of rocks. A gabbro, olivine-gabbro, anorthosite-gabbro, norite, and a troctolite were represented. The essential minerals in the samples were plagio-clase, pyroxene, and olivine, in variable proportions.

Jan 1, 1969

By Helen Francis

Over the last decade, lower ore grades and deeper mine environments have demanded that mining companies develop new technologies (tools and processes) to ensure a mine?s economic sustainability. At Inco Ltd., this demand has lead to the focus of technological development within the area of automation. Delineation (Drilling and Interpretation) is a key step in the determination of ore grade and location and therefore was chosen as a priority area for enhancement through automation. Focusing on the tools (the diamond drill) automation, significant benefits have been consistently achieved in terms of productivity, process efficiency, operational safety, cost reduction, component life cycles and machine reliability. It is suggested that optimum benefits have yet to be achieved, and will not be achieved until the human aspects of automation have been properly addressed. Since project inception, the human aspects of automation were recognized as critical to the success of the diamond drill automation project. However, it was only upon the installation of the Automated Diamond Drill in the field and its continued use on a daily basis that this issue received due attention. The human aspects of automation can be broadly addressed as economic, physical and psychological. This paper, with examples from the field, will discuss the human aspects of automation in each of these areas. The paper will draw from observations and experiences acquired throughout this project and recommend implementation strategies that should be utilized in future automation projects to enhance both the acceptance of automation and the success of the automated process itself.

May 1, 2002

By Steve Meador

West Virginia's basic coal mining regulatory programs are not unique when compared to Federally approved programs implemented by other coal producing states. In fact, the language in The West Virginia Energy Act governing subsidence mirrors that written in the Federal Surface Mining Control and Reclamation Act as well as other states' statutes. Certain legal and administrative issues, how¬ever, combine to make West Virginia's regulation of surface subsidence a singular approach.

Jan 1, 1986

By H. Kim, R. W. Plummer, T. J. Stobbe, W. G. Jones

"Exposure to airborne dust in coal mines has caused respiratory disease in coal miners. The causal agent in these diseases remains unknown, although the minerals found in and around the coal seam have been implicated. Toxicological studies designed to establish the effect of different mineral combinations and concentrations are lacking because the mineral content of coal mine dust has not been defined. This paper reports the initial results of a study done to characterize the mineral content of the coal mine dust found in a mine in the Upper Freepott seam in Northern West Virginia. Coal mine dust samples were collected using 4- stage cassette impactors at nine locations in the mine. The samples were analyzed by an x-ray microdiffraction film technique supplemented by energy dispersive scanning electron microscopy. Primary size modes of about 17 to 20 micrometers from the samples collected at the haulageway and the returnside of the continuous miner were found. From the returnside of the roof bolter samples, a primary size mode of 12to15 micrometers was measured. The minerals found in high percentages were illite, kaolinite, quartz, and calcite. The distribution of minerals in inter· and intra- location samples showed wide variations. Trends in these variations are presented. Stobbe, T. J.; Plummer, R. W.; Kim, H.; Jones. W.: Characterization of coal mine dust. Appl. Ind. Hyg. 2:95· 100; 1986. IntroductionDisease-associated dusts in coal mine environments continue to afflict miners in today's modern mines. The relationship between exposure to coal mine dust and the development of various respiratory diseases has been well recognized; however, the reasons for variation in occurrence of these diseases among miners in different geographic areas,(1) rank of coal seams,(2-3) and jobs (4) remain uncertain. Several papers have reported on the causal agent in the development of these diseases. Davis (5) found that the percentage of the non-coal minerals present in the lungs of deceased miners increased as the level oi pneumoconiosis became more severe. Reisner (6) reported that cell damage caused by coal mine dust increases with the mineral and quartz content of the dust and with the geo· logical age and rank of the coals. Two studies, investigating the in vitro effects of mineral dusts. reported by Seemayer and Manojlovic (7) also obtained positive correlation between cytotoxicity and the mineral content of coal mine dust. Gormly (8) postulated that destruction of macrophages by certain mineral dusts is the first step in the development of pneumoconiosis. In 1980, Adamis and Timar (9) reported that quartz. sandstone, slate. respirable colliery dust. bentonite, illite, and some kaolins had cytotoxic effects on rat peritoneal macrophages""in vitro."" Walton,(10) however, reported that the presence of high clay, illite, and kaolin mineral exposures reversed the hazards associated with increasing quartz exposures. This finding was supported by Le Bouffant and coworkers.(11) They reported that illite present in coal mine dust strongly inhibited the pathogenic effects of quartz while kaolin had a weaker inhibitory effect."

Mar 1, 1989

By R. D. Springer, R. L. Wilcox, P. A. PALMARTAN, R. K. Waring, E. A. Anderson

A knowledge of the mutual liquid solubility of zinc and lead is of importance in various phases of zinc metallurgy. The determination of this solubility has been the subject of numerous investigations, but unfortunately the results vary widely, particularly at higher temperatures, and it has been impossible to decide what data are most reliable. It was decided to investigate the matter and in view of the disagreement in the published results, it was felt necessary to take rather unusual precautions to insure the reliability of the data. For this reason determinations were made by four different methods, and the procedures of several of the previous investigators were duplicated. The previously used methods can be divided into two classes: methods involving freezing the entire melt and taking samples of the lead-rich and zine-rich phases in the solid state,l,2 and methods in which the samples are taken while the bath is liquid and at the desired temperature.3,4 Data obtained by methods involving freezing the entire melt show much lower solubility of lead in zinc than methods in which the liquid is sampled at temperature. The latter method was used by Spring and Romanoff5 and by Hass and Jellinek.6 The results of these two investigations are in agreement at temperatures near the melting point but at higher temperatures Hass and Jellinek find a higher mutual solubility. The low-temperature section of the zinc-lead equilibrium diagram is given by Peirce and Anderson.7 The diagram is based on the data of Hodge and Heyer,8 Arnemann,9 and Peirce.10

Jan 1, 1934

By Dimitri Frantzos, Stewart Thompson, Ed Skiba, Hans Tammemagi

On behalf of Acres International we would like to thank Frank Chiappetta of Blasting Analysis International for inviting our organization to participate in this state-of-the-art course on blasting analysis techniques.

Jan 1, 1989

By Motohisa Kanda

The time and amplitude statistics necessary to adequately describe electromagnetic (FM) noise in mines are illustrated through computer software techniques. They are 1) Allan Variance Analyses (AVA), 2) Interpulse Spacing Distributions (ISD), 3) Pulse Duration Distributions (PDD), 4) Average Crossing Rates (ACR), and S) Amplitude Probability Distributions (APD). These statistics are illustrated using data taken from a rather large store of raw analog data recorded in operational mines. The curves generated for the illustrations characterize the noise environment in the mines from which the corresponding data were taken, and should aid in the design of reliable communication systems for such mines.

Jan 1, 1974

By J. Kretschmann, N. Nguyen

"The Vietnamese economic growth of more than 5% per year leads this country to an enormous demand for raw materials. In the duration of ongoing expansion, the big challenge in the coming decades will be to make this process as sustainable as possible when it comes to economic, environmental and social objectives. Risk management especially in occupational safety and health (OSH) in enterprises is one of the most important tools to achieve this aim. But until now, this has not been taken into priority by Vietnamese coal mining enterprises. Main reasons for this situation can be found in deficient legislative conditions and in non-synchronized risk management systems. In contrast, Germany has implemented an adequate legal system and the German mining industry has gained lots of experiences in risk management, especially in occupational safety and health. This paper seeks to point out specific challenges how to improve occupational safety and health in Vietnam by transferring and applying German methods and standards."

Jan 1, 2011

Geotechnical conditions in the Lochiel Deposit are somewhat unusual, and thus geotechnical and hydrological investigations have assumed a high priority. The cost of mining the deposit has been shown to be very sensitive to geotechnical design parameters, and it is now seen to be of paramount importance to establish a high level of confidence in the design parameters used for mining cost estimation. Geotechnical investigations in the Lochiel Deposit reached a crossroads in 1985, when there appeared to be several quite different and equally attractive routes to follow to establish a satisfactory level of confidence in geotechnical mine design parameters. Three main options were identified, referred to as the No Further Work Route, the Drill and Trench Route, and the Trial Pit Route. The No Further Work Route was ruled out as being technically unsound and as satisfying none of the Electricity Trust of South Australia's (ETSA's) or the State Government's requirements. The planning process for further geotechnical investigation involved: clearly identifying the current status of work and significant geotechnical and geological factors planning and costing the alternative schemes for improving knowledge assessing the increase in knowledge to be gained from the alternative schemes . choosing the preferred scheme This process resulted in ETSA's

Jan 1, 1986

By W. C. McClain

The amount of vertical subsidence occurring over a longwall operation is a function of the thickness of material removed, the quantity and quality of any fill material, the width of the extraction, and the depth. The relationship between these variables is empirically determined, and subsidence can be predicted using only a few easily established constants which are subject to local conditions, such as the angle of draw. Associated with the vertical subsidence are both tensional and compressional horizontal strains. The component of strain in the direction of face advance is temporary, while the component perpendicular to the direction of working is permanently retained by the ground. It is these horizontal strains and the vertical strain (tilt) which result in damage to surface structures. It is possible to ameliorate the consequences of mining subsidence by utilizing the well-proved techniques described in this paper in the underground operations and on the surface. INTRODUCTION In the past, American mining practice in stratified deposits has been mostly restricted to production by the room-and-pillar system where a sizable percentage of the salable product is left in place as pillars. These pillars are usually regularly spaced and provide a permanent support for the roof of the mine and the surface above. The longwall system employs an advancing face which may be as much as several hundred yards long and some sort of breaking and loading machine which moves back and forth along the length of the face. The roof is supported behind the face only to a distance required to provide working space for the production machine. The supports are then withdrawn and the roof is allowed to cave. The longwall system therefore possesses the various advantages of complete extraction, thus effectively increasing reserves, while at the same time eliminating the necessity and cost of installing artificial supports such as roof bolts and heavy steel girders over large areas. Longwall mining has long been practiced in Europe, where it has undergone an involved development, culminating in the recent perfection of powered-loading machines and automatic self-propelled supports, offering high productivity. With ever-in creasing

Jan 1, 1967

By Bruce E. Stauffer

The report recommends criteria for the design of safe surface mine haulage roads. Information relevant to road design at surface mine sites delineates vertical and horizontal alignment; subbase and surface materials; drainage; cross-slope, superelevation; runaway vehicle protection; and road and vehicle maintenance criteria. An appendix Is included which details the study methods used in compiling the final report information. A bibliography following the main text lists literature referenced for the report data and which is available as a valuable tool for use by individuals involved in haul road design.

Jan 1, 1976

Strategic planning in mining is an important value accretive process. One of the most important aspects during the planning phase is determining the correct mine and plant design. The traditional mine site design (mine, plant, stockpiles, dumps) develops a fixed system for one set of conditions or expected values. A robust design is a fixed system that is designed to deal with a large range of conditions without changing the system design. A flexible design changes the system dynamically in response to changes in uncertainties. It is hypothesised that a robust design generates less value than a flexible design as it fails to realise the value of actively managing the operation. One of the most critical aspects in creating strategic value from a project is embedded in the design of the system. A flexible design has components that allow the design of the system to change over time. An operational plan is proposed as a hybrid of a robust and flexible design. This paper seeks to compare these different design methodologies. An analysis of the same case study is undertaken to compare these approaches, showing an a traditional net present value (NPV) of $1678 M, robust design NPV of $1787 M and an operational plan NPV of $1867 M - an increase of $170 M.

Dec 6, 2010

By W. G. Pariseau

Conventional barrier pillars in underground coal mines are intended to protect mains and sub-main entries for relatively long periods of time. Inter-panel barrier pillars that may be used in especially deep coal mines, say, below 800 m (2,400 ft), are intended to protect longwall panel faces and adjacent panel entries for the life of nearby panels. Barrier pillar failures have serious consequences, especially when failure is sudden and catastrophic. Design of barrier pillars is thus of considerable importance to safety and efficiency of operations. Locally, failure occurs at the limit to elasticity when stress tends to exceed strength. Strength of rock is stress dependent, so an analysis of stress is central to rational design of barrier pillars. There are several approaches to stress analysis for coal mine pillar design that range from the simplified but popular tributary area method including variations that use empirical definitions of the tributary area (volume) to more sophisticated numerical methods such as the boundary element method and the finite element method. Despite ease of use, empirical methods are limited and always suspect when applied to conditions beyond their scope of definition, while boundary element methods do not generally include pillars in the solution domain. The finite element method is not so restricted and thus is a natural choice for conventional and inter-panel barrier pillar safety analyses. Caving of roof strata certainly affects the distribution of stress about longwall faces and nearby barrier pillars. Conventional wisdom that considers roof strata to respond as cantilever beams indicates the longer the span the greater are loads on longwall faces and face supports. Failure of roof beams, caving, shortens spans and relieves face and support loads. The same view is applicable to barrier pillars in that long roof spans outside of barrier pillars leads to higher pillar loads than short spans. Hence, caving should be beneficial to barrier pillar safety. However, the benefit may be small depending on the extent of caving and only in the form of a slight reduction in overburden weight transmitted to the barrier pillar, especially in case of inter-panel barrier pillars. FINITE ELEMENT MODEL Finite element models for stress analysis in rock mechanics are formulated in consideration of physical laws, kinematics, and material laws. These models meet the essential requirements for equilibrium of stress, satisfy strain displacements relations, and in a general sense, relate stress to strain through material laws, often referred to as constitutive relations. The last may be complex and replete with nonlinearities. A de facto engineering material law is the elastic model in the form of Hooke?s law for linear, homogeneous materials that may be anisotropic. A limit to elasticity is imposed by strength in the form of a failure criterion or yield function that accommodates triaxial stress states. Examples used in rock mechanics are the well-known Mohr-Coulomb (MC)criterion, the popular Hoek-Brown (HB) criterion, and the Drucker-Prager (DP) criterion. The first is linear in stress, while the second is nonlinear. Both do not include any effect of the intermediate principal stress. The last is also linear in stress but includes the intermediate principal stress. All show an increase in the major principal stress at failure with confining pressure. In the UT2/3 finite element programs, the failure criterion Y is a nonlinear, anisotropic form and is also the plastic potential [1]: [YJI=+ 11NYJI=+- (1) ] where N is an exponent, usually 2, and J2 and I1 are nonlinear anisotropic forms of the second invariant of deviatoric stress and the first invariant of principal stress. The form (1) reduces to DP when N=1and isotropy prevails. When Y = 1, the limit to elasticity is reached and plastic deformation is possible. The material constants that enter (1) are all computed from unconfined compressive, tensile, and, in case of anisotropy, shear strengths. These latter strengths may be estimated from conventional laboratory test data. Of course, joints in field-scale rock masses induce different values. How different is often a matter of conjecture. For comparative studies, laboratory rock properties seem adequate. Inelastic, that is, plastic behavior occurs with strain beyond the elastic limit. Strain hardening, ideal plasticity, and strain softening are possibilities. Ideally brittle behavior with complete loss of strength is an extreme form of strain softening [2]. Caving may also be considered as strain softening with a less drastic reduction of material properties. As intact roof strata fail, fall, and form a rubble pile below, a new material is formed with greatly reduced elastic moduli and strengths, say, 1/100th of the original intact moduli and strengths. From another viewpoint, caving is simply a natural cut and fill process. Input for stress analysis by the finite element method includes: 1. Preexcavation stress state 2. Geology (stratigraphic column) 3. Structure (major faults, joint sets) 4. Elastic moduli 5. Strengths, and specific gravities of each rock type 6. Excavation sequence Output from an analysis includes distributions of: 1. Stress 2. Strain 3. Displacement 4. Local factor of safety (element fs) The last is defined as the ratio of strength to stress and in view of the failure criterion (1), is the ratio [fsJstrengthJstress=] where the numerator is the maximum possible value and the denominator is the actual value of the square root of the second invariant of deviatoric stress. This invariant is a measure of shear stress intensity but also reduces to uniaxial tensile and compressive strengths as the case may be. An fs greater than 1 indicates an elastic state of element stress; a value of 1 indicates the limit to

Jan 1, 2009

By J. E. Sepulveda, R. Morrow

"The Linear Wear Theory provides a reliable framework for the assessment of grinding media performance at full scale. This theory can be applied for analyzing grinding media consumption rates in any given grinding operation as well as the evaluation of alternative grinding media products, based on actual operational data, correcting for conditions like ore hardness, feed and product sizes, mill power and circuit throughput; all of them unrelated to the grinding media intrinsic quality attributes. A series of Moly-Cop Tools 3.0 applications have been developed to facilitate the calculations involved. Moly-Cop Tools 3.0 is available, free of charge, to all interested parties through the corresponding author.INTRODUCTION Mineral grinding circuit operators recognize the significance of grinding media (balls) consumption on the cost structure of any full scale grinding facility. The effective cost associated with this essential consumable depends on two factors: delivered price and durability (quality) of the grinding media. When evaluating alternative grinding media varieties, there is a well accepted evaluation criterion by which a given media type is considered to be “cost-effective” when its unit application cost – normally expressed in $/ton of ore ground – is shown to be lower than that of a nominal reference condition: Grinding Media Cost = Ball Price x Ball Consumption (1) ($ /ton ground) ($/ton balls) (ton balls/ton ground) Under this criterion, an alternative, higher-price grinding media product could be “cost-effective” if its associated consumption rate is sufficiently lower than the reference media, enough to yield a lower grinding cost, as dictated by Equation 1. For the proper application of the above criterion, evaluators must rely on representative indicators of grinding media quality, ideally independent of most mill operating conditions. In each particular case, ball price is always a known component; however, it is not so evident how media performance (quality) differences amongst alternative product types could be assessed with a reasonable degree of accuracy and precision."

Jan 1, 2014

By W. R. Hardwick

This paper describes mining methods and practices of the Morenci Branch, Phelps Dodge Corp., Greenlee County, Ariz. It is one of a series being prepared by the Bureau of Mines on mining methods, practices, and costs in various mining districts in the United States. A similar program, conducted from 1928 to 1939, was generally conceded to have contributed substantially to improving the efficiency of mining operations. The morenci open-pit mine (see photograph, fig. 1) is in southeastern [zona] (fig. 2), on U.S. Highway 666 about 5 miles north of Clifton, county at of Greenlee County. The concentrator and smelter are near the mine at an altitude of about 4,400 feet. Figure 3 is a map of the mine and plant.

Jan 1, 1959

By H W. Wu

The objective of this paper is to examine some modern forms of auxiliary ventilation fabric ducting with streamline design fittings and the efficiency of construction practices. The most dynamic form of auxiliary ventilation is a moveable fan and ducting, which is easy to install and uses light weight fabric. The porous nature of some ducting construction results in leakage and loss of efficiency. Flexible fabric ducting is primarily associated with metalliferous mining and industrial material quarrying. Some applications are found in coal in stone/rock development and with use of dust scrubbers.Forcing auxiliary ventilation systems employing circular fabric ducting have been examined. Leakage of air from ducting is affected by ducting material and construction methodology (such as welded or sewn), quality of installation, number of joints, total length, pressure differences between the inside and outside and diameter. The best ducting system with welded construction can have up to four times less leakage as compared with some systems with sewn construction. It is also evident that the sewn type ducting will deliver less air to working faces and require more fan power, hence higher operating costs to run when connected to the same fan.A case study has also been undertaken to compare predicted operating costs of auxiliary ventilation systems connecting various types of ventilation duct fittings. The study also assesses the technical importance of streamline design with low aerodynamic resistance to reduce the shock losses of the duct fittings. It was found that the system connected with streamline designed fittings is able to provide some power cost savings and deliver more air compared with the system fitted with conventional fitting pieces.Using a ventilation ducting system with a lower leakage and frictional characteristic represents a sound solution to new challenges faced by the industry for increasing demand for better quantities and quality of air delivered to working faces; this is due to higher production schedules and stricter mine health and safety regulations within the Australian mining industry.CITATION:Wu, H W and Gillies, A D S, 2017. A comparison of auxiliary ventilation systems and their predicted operational performances, in Proceedings Australian Mine Vent Conference 2017, pp 207–214 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017