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By Axel Hultgren

In ingots of silicon-killed carbon steel-made without addition of aluminum, transparent spherical or nearly spherical inclusions, up to about 0.15-mm diameter, are generally present. They may be glassy or crystallized, wholly or partially In steels of high carbon contents they consist of silica,4,12 in steels of lower carbon contents and, in consequence, higher oxygen contents, of iron-manganese-silicate with higher iron and manganese contents the lower the carbon content. Small silicate inclusions, about 0.01 mm or smaller, may occur, just as sulphide inclusions do, in portions that have been the last or nearly the last to freeze,l0,12,17 this location being verified by primary etching.10 Hence, it may be concluded that they have separated from liquid steel that has become enriched in oxygen through segregation during freezing. Silicate inclusions with a diameter larger than about 0.02 mm appear, at least in ingots of moderate size, to be distributed at random in relation to the primary structure and therefore may be assumed to have existed as droplets in the liquid steel before freezing started at the point in question. The average size of such large inclusions has been found to increase toward the axis1' and their amount to increase from the top downward in the axial region.2,10 It has been assumed that this accumulation is a result of the droplets being brought down by the slowly settling metal crystals.2 It has been suggested that large silicate inclusions generally form,6 or may form,'' by coagulation of smaller ones, or by small drops being caught by larger ones during the rising of the latter.9 The occasional occurrence of a large inclusion surrounded by a swarm of small ones has been taken as a proof of this mechanism but, as suggested later, may be explained in another way. It has been pointed out that an inclusion droplet in the liquid steel may grow at the expense of an adjacent one, which implies diffusion of soluble inclusion-forming elements in the liquid steel.'= Further, it has been pointed out that, on cooling in ladle or mold, the metal-silicate equilibrium is displaced in the direction of separation of silicate. Thus inclusions may grow in size and amount.2,3,8,12,13,17,19,20 The exposure of the liquid steel to the oxygen of the air during tapping and pouring, and the consequent reaction, has been mentioned as a source of the formation of inclusions19.20 but this possibility has not received particular attention. Further, the admixture of furnace slag during tapping1 or of refractory material from various sourcesl,20 has been considered. The inclusions due to the latter may be recognized, however, through their resistamce to hydrofluoric acid, imparted by the alumina content. Silicate inclusions are often referred to as the natural and inevitable product of reaction on adding silicon or silicon

Jan 1, 1949

By R. J. Blackwell, J. R. Rayne, J. R. Henderson, W. M. Terry, D. C. Lindley

This paper presents the results of a laboratory investigation of the efficiency of water-solvent mixtures in recovery of oil. These mixtures may have the high displacement efficiencies characteristic of solvent floods and the high sweep efficiencies characteristic of water floods. Thus, the water-solvent process may increase the number of reservoirs in which a miscible-type displacement can be used profitably. The experiments on the use of water-solvent mixtures for recovery of oil were conducted to find the general applicability of the process. These studies demonstrated that, in flowing through sands, water and solvent segregated into a solvent layer on the top and a water layer on the bottom rather than flowing through the sands as a uniform mixture. Calculations based on the simultaneous flow of the water and solvent in layers were used to predict the effective mobility of the mixtures and the optimum operation of the process in steeply dipping, homogeneous reservoirs. As most reservoirs are not suited for the operation of the process under ideal conditions, experimental studies were conducted with sand-packed models scaled to represent more realistic reservoirs. These studies included the effects on recovery of oil of rate of injection, viscosity of oil, variations of permeablity within a formation and variations in water-solvent ratio. For the range of con-ditions studied, higher recoveries of oil were obtained with water-solvent mixtures than with water or practical volumes of solvent alone. INTRODUCTION A group of intriguing—because of their great possibilities—new oil recovery methods at the disposal of the petroleum engineer are the miscible displacement processes. These processes (high-pressure gas drive, en-riched-gas drive and LPG bank driven by methane) displace all of the oil from the portions of the reservoir swept by the injection fluid. The key question confronting the engineer applying one of these techniques to a particular reservoir is, "What fraction of the reservoir can be swept by injection of a practical volume of solvent?"." Intensive laboratory studies have been made during the past several years in seeking answers to the question of the sweep efficiencies which can be expected in solvent floods.' - These studies provided the answer that low-viscosity, low-density solvents channel and by-pass oil in sands with no dip. In horizontal sands, solvent flooding becomes less efficient as the viscosity of the oil increases, the recovery of oil at solvent breakthrough decreases and larger volumes of solvent are required to achieve a given recovery. More efficient displacement of oil by solvent is observed under certain conditions in sands with dip.".' If the permeability and dip of the sand are sufficiently high, gravity segregation of low-density solvent injected updip can prevent channeling. At rates of depletion below a critical rate,"." no channeling occurs. Unfortunately, in many reservoirs, the critical rate is so low that production of oil at rates below this rate is not economically attractive. And at rates over four times the critical rate, channeling is almost as severe in sands with dip as in horizontal sands. These findings pointed out that, for solvent floods to be generally applicable in recovering oil from all types of reservoirs, new methods of improving their sweep efficiencies are needed. Simultaneous injection of water with the miscible fluid was suggested by Caudle and Dyes7 as a method for improving sweep efficiencies. They theorize that water flowing with the solvent would decrease the effective mobility of the solvent and cause it to contact more oil sand. If, indeed, the water and solvent flowed as a uniform mixture, the process should have the advantages of the high displacement efficiencies characteristic of miscible floods and the high sweep efficiencies of water floods. Thus, the method (at least in theory) would greatly increase the number of reservoirs in which miscible-type displacements would be feasible. A research program was conducted to probe the technical feasibility of the water-solvent process. It

By Herbert H. Fields

The Bureau of Mines has been promoting health and safety in mining since its establishment in 1910. One area of research covers methane control in coalbeds; this report describes a multipurpose borehole used for both degasification and the production of commercial-quality gas. The multipurpose borehole was drilled into a barrier pillar, and seven horizontal degasification holes and one pressure point hole were drilled radially out from its perimeter. On January 28, 1974, after 505 days of degasification, a compressor was installed to the exhaust stack of the multipurpose borehole (1), and pipeline--quality gas was introduced into a commercial pipe-line. As of February 4, 1975, over 171 million cubic feet of gas had been purchased by the gas company. The in situ pressure at a depth of 200 feet into the coalbed and the average gas flow and water flow from the degasification holes as they were completed and closed off were 203 psig, 1,200,000 cfd, and 6.8 gpm, respectively. Twenty-four hours after the seven holes were connected through individual pipes to the atmosphere, the total gas and water flows and the in situ pressure dropped, respectively, to 971,000 cfd, 1.3 gpm, and 18 psig. Five hundred days later, the corresponding values were 620,000 cfd, 0.15 gpm, and 10.5 psig. Daily flows have been as high as 800,000 cfd, and as low as 360,000 cfd, during the last 500 days. On February 4, 1975, after 882 days of degasification, 576 million cubic feet of gas had been removed from the Pittsburgh coalbed; apparently a much larger area of virgin coal is being degasified than is defined by the holes. This study indicates that horizontal boreholes drilled into a coalbed from the bottom of any shaft will effectively remove gas from the Pittsburgh coalbed,

Jan 1, 1975

By Richard L. Bullock

INTRODUCTION TO COAL MINE PRACTICES To illustrate the trends and the proportions of the various methods used to mine coal underground, Fig. 1 is useful. However, of all the underground coal mining practices in the US, room-and-pillar mining still accounts for 96.5%. This percentage is divided as follows: conventional, machine cut-33.0%; conventional, hand cut -1.7% : and continuous mining-61 .8%. As was stated in Sec. 1, Chap. 7, the term "development" means different things to a coal miner and a metal miner. Though some of the ideas and concepts presented in the first two chapters of this section are valid for all types of mines, they apply particularly to mines using noncoal methods. In coal mines (and many other mines using coal mining systems) development refers to that phase of the mining cycle which advances several entries through the virgin seam, forming pillars by cross- cutting between the parallel entries. Because in the past, coal seams (as well as seams of potash, trona, oil shales, gilsonite, salt, and some uranium) have been much more uniform in their thickness, continuity, and quality, and because they often cover large areas, it has been possible to develop the room-and-pillar mine completely from a preplanned, systematic approach with a minimum of surface drilling information. As the better coal seams of the midwest and east are being depleted (or eliminated because of high sulfur), seams of poor quality and continuity will have to be mined. Also, seams that are not nearly so well known at greater depth will have to be taken as well as some of the steeply pitching and folded seams of the west. The American coal miner is only now feeling the need to do a much more complete job of prospecting the geology of the coal seams than he has ever done before. This is the only way that the risk of unknown rock conditions can be minimized. In many ways, surprises in the rock or coal seams are much more of a disaster to the coal miner developing a room-and- pillar mine than to his metal mining counterpart. Be- cause of the multiple entries, the complex set of the Federal Coal Mine Health and Safety regulations, and the totally dedicated design of coal mining equipment, coal mining systems do not have the flexibility to easily overcome changes in the seams mined or the rocks around them. Thus, if a wide parting of rock shows up unexpectedly in the middle of a coal seam entry being milled away by a continuous miner, it will either have to slowly "chew" through it, convert to a conventional mining system, or stop that particular entry. Likewise, if suddenly the coal disappears completely and in its place are the remains of an old creek bed, the "want" or "wash out" as it's called will usually have to be mined right along with the coal, by the same methods and with the same soft rock equipment, or the entry will have to be abandoned. For reasons such as these, coal seams today should be drilled from the surface just as thoroughly as the ore bodies in metal mines. Having once determined a more precise location of where the coal seams are located and which are suitable for mining, the engineer must still take a systematic approach to planning the mine as completely as possible. The initial series of entries must be planned to be driven either from the side hill portal or from the bottom of the shaft or slope. In either case, they should usually intersect the area close to the midpoint of the mass to be mined.

Jan 1, 1982

ASHBEL WELCH, Lambertville, N. J.: Dr. Dudley has given the wear of steel rails under four different conditions. He arrives at the conclusion that the softer rails, or those that from their composition ought to be softer, wear better than the harder. But there is another condition which has an important bearing on the subject, and should not be overlooked,—the weight on a wheel. With the lighter weights of the past, the softer rails may have worn best; with the heavier weights of the future, the harder may wear best. Weights will probably be increased up to the capacity of steel to bear; then doubtless the harder steel will wear best. A leaden rail with ten pounds on a wheel might carry millions of tons, but with one hundred pounds on a wheel, it would be destfoyed by a few thousand tons. So in the days of iron rails, my experience was that the softer rails under light machinery stood better than some of the harder; but under heavy machinery the softer were muqh the most rapidly destroyed. It is doubtless the same with steel. The pounding motion of the wheels loosens or spreads the particles of a thin film of steel; the pull lengthwise on the rail de-

Jan 1, 1881

By Aug. J. Jr. Bowie

(Read at the Wilkes-Barre Meeting, May, 1877.) Brief Outline of the General Topography of the Gold Regions of California. THE topographical features of California, as demonstrated by the explorations of the State Geological Survey, are found to be exceedingly simple. Three equidistant parallel lines can be used in conveying a general idea of the physical geography of Central California. A straight or " main axial line," whose course would be north 31° west, passing through the culminating peaks of the Sierra for a distance of five hundred miles, can be assumed as the eastern boundary of the State. A second parallel drawn fifty-five miles west of the " main axial line" will skirt the western base of the Sierra Nevada, along the edge of the foot-hills. A third parallel run equidistant from the second will represent "as nearly as possible the western base of the Coast Ranges." These parallel lines divide the State into three belts, namely, the Sierra, the Great Valley of California, and the Coast Ranges. " This arrangement of the physical features holds good for a length of four hundred miles in the direction of the main axial line, comprising almost the whole of the agricultural and the greater part of the mining districts."* The section of the country which is of immediate interest to the miner is the western slope of the Sierras. These mountains, rising in a short distance from the Sacramento plains to elevations of over seven thousand feet, with occasional peaks ten and twelve thousand feet high, are cut by numerous deep and precipitous gorges or capons, through which drains the immense watershed of the Sierra, supplying the main rivers of the State and ultimately emptying into the Pacific Ocean. Between these capons, ridges or divides are formed, on top of * See vol. i, p. 5, Geological Survey of California. J. D. Whitney, State Geologist.

Jan 1, 1878

By Dennis R. Dolinar

A room and pillar coal operation in central Pennsylvania was experiencing roof cutters and long running roof falls caused by high horizontal stresses. The roof conditions created hazards for the miners, and caused several production panels to he abandoned prematurely. The mine requested assistance from the National Institute for Occupational Safety and Health (NIOSH) in applying the "advance and relieve" mining to reduce the affects of the horizontal stress during panel development. The "advance and relieve" plan that was developed called for removing a pillar on one side of the panel as it was being advanced, thus creating a cave. The caving then relieved a portion of the horizontal stress across the panel. Because the horizontal stress direction is key to the success of this method, stress mapping as well as mining experience was used to establish the direction of the horizontal stress. Subsequent field measurements provided an estimate of the stress magnitude. Three panels were mined using this technique, and good roof conditions were achieved in all these panels. Instrumentation was used to monitor the stress changes in the roof created by the cave. Stress relief of over 1.000 psi was measured 120 ft from the cave and to depths of 20 ft into the roof. The lateral extent of the stress relief zone across the panels appears to be about 400 to 500 ft. This case history has provided much insight into the practical application of the "advance and relieve" method discussed in this paper.

Jan 1, 2000

By Yu. V. Ponomarenko, A. I. Meka, P. A. Witherspoon, Yu. I. Volkov, J. E. Gale

"The mining of iron ore in the Kursk Magnetic Anomaly (KMA) region of the U.S.S.R. is done mostly by the open-cut method. An assembly of open pits capable of producing about 100 million tons of ore per year makes KMA the largest iron ore field in the world. Three gigantic open pits are currently in operation: Lebedinsky, Mikhailovsky and Stoilensky. These pits are as large as 13 km2 on the surface and as deep as 230 m. The mining operations are constantly faced with the problem of flooding from aquifers that overlie the orebodies; the magnitude of the problem is evident from the fact that the present groundwater inflow to these three pits is about 110 million mt /yr.Under such conditions, safe operations and pit wall stability are of great importance, and this has been accomplished very effectively by the appropriate use of seepage barriers and drainage systems. At KMA, the most important items used in the system of groundwater control are standard pumping wells, raise wells that drain into a 20 km to 30 km long underground drift located just outside the perimeter of the pit, horizontal wells driven through the pit walls, and horizontal drains. This paper describes the mining and engineering technology that have been developed at KMA to control the groundwater systems and protect the mines from flooding."

Jan 1, 1992

By William E. Foreman

Editor's Note: For purposes of uniformity throughout this book, Coal Preparation, 5th Edition, the editors have advocated the use of the U.S.A. Sieve Series of testing sieves (see Table 6-2) as the testing sieves of modem choice. The adoption and use of the U.S.A. Sieve Series and its forerunner, the Tyler Standard Screen Scale, has for over seventy-five years introduced uniformity into reports; has simplified the comparison of results; and has provided control for operation of modem coal preparation circuits. While a full explanation of the differences, and more importantly the similarities of these two measuring systems cannot be stated here, it is felt that a brief clarification of terms is in order. The difference between the U.S.A. Sieve Series and the Tyler Standard Screen Scale Sieve Series is the identification method. Tyler Screen Scale sieves are identified by mesh designation with the width of opening varying by the square root of two, or 1.414 times the width of opening in the previous sieve. The U.S.A. Sieves are identified by openings in millimeters or microns and also arbitrary numbers which do not necessarily mean the mesh count; this series uses width of opening variations in the order of approximately the ratio of the fourth root of two, or 1.189. Tyler Standard Screen Scale and U.S.A. Sieves can be used interchangeably, and to avoid confusion the nameplate of all Tyler Screen Scale sieves show the U.S.A. Number, while all U.S.A. Sieves show Tyler Screen Scale mesh equivalent numbers on the nameplate. INTRODUCTION Sizing is the process of separating mixed mineral particles into groups of particles all of the same size, or into groups in which all particles range between definite maximum and minimum sizes. This chapter applies the modem mechanical screen and the liquid-solid cyclone (hydrocyclone) to the process of sizing solid particles as applied to coal preparation. Screening is a mechanical process that separates particles by size. Modem coal preparation is made possible by efficient sizing. Sizing of coal is defined in terms of the opening through which an individual particle will or will not pass, or as two openings, the larger of which passes the particle and the smaller of which retains the particle. Screening accomplishes sizing by passing coal of a wide range of sizes through a series of screens, each of a decreasing size. The individual screen discharges are then directed into different screen products for subsequent sale or processing. Screening in modem preparation plants is done for the following reasons: 1. Removal of fine coal before a size reduction step. 2. Size separation of the coal into a coarse and fine fraction for sized products or for processing in coarse and fine coal circuits. 3. Recovery of heavy media from heavy media coal cleaning circuits.

Jan 1, 1991

By Errol Ladner

The application of large-diameter bore holes in underground operations at Heath Steele was first tried in February, 1975. Since that time, 90,000 feet of 6-l/2-inch holes have been drilled, primarily in open blasthole stoping, for a total of 1.4 million tons, 37% of underground production. This paper deals with a review of drilling with conventional percussion longhole drills, a practice which is still utilized at Heath Steele, and the changeover to accommodate the use of larger blastholes in underground operations. The discussion will include layouts for 2-inch-diameter blasthole stope blocks, 6-inch-diameter stope blocks, trials for I. T.H. (in-the-hole) stope rooms, and drill pat/ems, drilling and performance. The complexity involved in the development, drilling and blasting of 2-inch blastholes is overcome by the over-all simplicity of drilling and blasting in I. T.H. stopes. This has made the practice of open blasthole stoping more economical and efficient. The future indicates an increase in the company's I. T.H. program. Investigations are being undertaken regarding longer holes to decrease development costs. A VCR stope is also being developed for trial.

Jan 1, 1979

By J. Peck, M. J. Scobie, M. D. Pollitt

"IntroductionOne aspect of current research in the Canadian Centre for Automation and Robotics in Mining (CCARM) involves the automatic identification of different rock types based on particular drill performance parameter responses. In this paper, the four drill performance parameters of torque, pulldown pressure, rotary speed and penetration rate, were measured from an instrumented rotary blasthole drill during field investigations undertaken at the Fording River Mine, Elkford, British Columbia. The variations in these parameters were subsequently correlated with the corresponding rock types using the results from gamma, neutron and core logs in the same or adjacent boreholes. Work to date has sought to understand the relationships between drilling performance parameter variation and rock properties. Using the results of these studies, the present research is aimed at developing techniques whereby classical statistical pattern recognition methods are applied to drill data to permit the discrimination of different rock units. A monitoring system incorporating such a capability would greatly enhance the quality and amount of geological information on a per borehole basis while drilling. This data in tum could be used economically and directly towards explosive loading, geological exploration and reserve estimation. Background and PhilosophyThis work is an extension of equipment monitoring studies at the Fording River Mine. At this site over a two-year period, 5180 m of data was acquired on an instrumented Bucyrus-Erie 60R drill during routine production blasthole drilling. Throughout these studies, a 311mm diameter S8M Security tungsten-carbide tricone bit was used. The mine, located in southeast British Columbia, is a large-scale, open pit coal operation. The means of excavation revolves around a drill-blast-muck cycle and employs a standard shovel-truck system of transport. Four main rock types are present: sandstone, siltstone, mudstone and coal, with average compressive strengths ranging from 10 MPa to 150MPa respectively (peck 1989). However, based on the field studies, very few differences in terms of strength and compositional properties were seen between the sandstones and siltstones at the site. Therefore, they are referred to as sandstone/siltstone in all further discussions."

Jan 1, 1991

By Charles R. Stewart, William W. Owens

Reservoir performance predictions based on laboratory core test data assume that fluid flow is laminar for the laboratory test. A study has been made to determine the validity of this assumption for laboratory tests on various types of porosity found in producing limestone formation. Data are presented which show that turbulence and slippage can occur during laboratory tests on hetero-geneous-type porosity limestones, thus causing serious errors in measured single-phase permeabilities and two-phase relative permeability characteristics. In single-phase flow tests it is possible to eliminate turbulence and correct for slippage or to eliminate both factors by controlling test conditions. It is not always possible to control test conditions and thereby eliminate turbulence and slippage in two-phase .flow tests. A correction method is presented which can be used to calculate the true two-phase laminar flow relative permeability characteristic even though furbulence and slippage exist. .INTRODUCTION It is customary to make use of Darcy's law and modifications of this law, together with laboratory data on formation core samples to predict the performance of producing reservoirs. Such predictions are based on an assumption that fluid flow is in the laminar or streamline region for the laboratory test. It was the purpose of this inves- tigation to determine the extent to which turbulent flow may occur in laboratory fluid flow tests on hetero-geneous porosity limestones. Considering that turbulent flow conditions might exist in some laboratory fluid flow tests, additional emphasis was placed on the development of a method to correct for turbulence when laminar flow conditions could not be attained. FLUID FLOW CONCEPTS FOR POROUS MEDIA The Influence of Pore Geometry on Fluid Flow One of the more important factors influencing fluid flow in porous media is the geometry of the pore space which includes such characteristics of the pores as size, shape, distribution, roughness, uniformity, etc. In general, oil- and gas-producing formations can be divided into two broad types on the basis of pore geometry. One has been called sandstone-type porosity media, which is characterized by a small range in pore size, uniformity in shape of the pores, smooth pore surfaces and a regular and uniform distribution of pores. The other type has been called heterogeneous porosity media and is usually limited to the dolomites and limestones. This type is characterized by a wide variation in the size, shape, and distribution of the pores and rough, irregular pore surfaces. It is therefore apparent that conditions are much more favorable for turbulent flow* in heterogeneous-type porosity media than in sandstone-type porosity media. Interrelationship Between Turbulence and Gas Slippage In studying the problem of turbulent flow in laboratory tests on porous media, it is necessary to be aware of the interrelationship between slippage and turbulence for gas flow. As a result of slippage or the Klinken-berg effecta, apparent perrneabilities to gas are greater than the true value because there is no stationary layer of gas in contact with the walls of the flow channels. Gas slippage decreases as the mean free path of the gas molecules decreases. Since the mean free path of any gas decreases with increasing density, increases in static pressure result in lower apparent gas permeabilities. However, a reduction in gas permeability can also be due to turbulence. Therefore, in studying only turbulent flow in porous media, it is necessary to hold gas density, and slippage, constant or to reduce slippage to a negligible value by operating at high static pressures. Presentation of Laminar and Turbulent Flow Data A graphical relationship between permeability and a pseudo-Reynolds number, N,, will be used to show the two types of fluid flow, i.e., laminar and turbulent. The usual graphical method for such a description has been the use of friction factor-Reynolds number charts4. On such a logarithmic diagram, the laminar region appears as a straight line having a slope of 45 degrees. As the friction factor decreases and the Reynolds number increases, the turbulent region is reached and appears as a deviation from the 45-degree slope line. However, in petroleum engineering literature resistance of por-

By J. E. Udd, L. B. Geller

"Responding to requests by the Canadian mining industry and the mine regulatory authorities, the Canada Centre for Mineral and Energy Technology (CANMET) undertook a major research project on evaluation of wire rope testing in 1986. The objectives were to document, analyze, and clarify - in basic operational terms the reasons for ongoing mine shaft wire rope failures that occur despite mandatory non-destructive testing (NDT) with electromagnetic (EM) instruments on a routine basis.The project's investigative scope and depth was unprecedented. It could not have been successfully completed without the wholehearted cooperation of industry, research organizations, and mine-inspectorates at both national and international levels.The project's goals were attained. II was shown that, at present, the problem areas in Canada are oj a human , rather than of a hardware-related nature. Consequently, the principal recommendations are concerned with a more widespread, more uniform and thorough operator training and certification process at industrial and regulatory levels. Secondary suggestions concern areas of hardware-related improvements.In this paper an up-to-date general overview is given of the major project results. A n extensive reference list is also provided of papers that contain all relevant documentation in detail."

Jan 1, 1992

By Masatsugu Taniguchi

Onoda Cement Co., Tsukumi is one of the largest limestone producer in Japan. From mines called Suisho and Ganji, it supplies limestone for cement plant, lime manufacture, iron and steel industries and for many purposes. The Suisho mine is highly productive, though it's operations are governed by the utmost regard for its impact on the surrounding environment. However, it became difficult to meet additional expanding demand and to assure continuity of high rate production. Present production rate 1s 6 million tons per year. The Ganji mine, some 2 km south west of the Suisho mine, began first-stage development In mid- 1973 and in january I976 started production. The capacity of major facilities is 9 million tons per year. Full production is predicted in 1979. Mining and milling methods of these mines under strict environmental considerations are discussed.

Jan 1, 1976

By A. J. Barry

[This is the third report related to fundamental retail roof-bolting research in underground mines (1, 2)3/ and describes an investigation of the torque-load relationship for 3/ -inch, expansion-type mine roof bolts. In a previous investigation the relationship between torque and bolt load was established for 1-inch slotted-type bolts (1). The two types of bolts differ in diameter and method of tightening and consequently may exhibit different torque-load behavior. The 1-inch bolt is tightened by turning the nut, the 3/4-inch bolt by rotating the beaded bolt itself. Experience indicates that roof bolts should be tightened to moors satisfactory anchorage sand to minimise nag of tie roof over the mine opening. The generally accepted means for determining the bolt load is to see a torque wrench to measure he torque applied in tightening the bolt and, from a predetermined relationship between torque and molt load, to, determine the corresponding bolt load. Because frictional effects may vary between one type of shell assembly or bolt and others, 7 types of expansion shell units and 3 types of 3/4-inch-diameter, mild steel bolts were tested.]

Jan 1, 1954

By Remilton D, Raston GR

The low price of nickel which has prevailed since 1982 has dictated that unit production costs must be minimised. Early in the crisis, a decision was made within Western Mining Corporation to reduce unit costs by increased rates of production, not by a significant reduction in manpower. This paper details developments which have taken place at Kalgoorlie Nickel Smelter to enable the higher production rates to be achieved, or to increase energy efficiency. It is significant that most developments were achieved without the need for capital expenditure.

Jan 1, 1985

By E. M. De Souza

"A scientific testing program, sponsored by the Mining Industry Research Organization of Canada (MIROC), was undertaken on the surface and underground facilities of Inca's Crean Hill Mine to investigate the effectiveness of using dry drilling with a dust control system.The surface test work consisted of the monitoring of advance rates and bit life white a series of 165 mm holes, wet and dry, were drilled at both mine air (0.69MPa) and high pressure air (1.72MPa). The results, based on statistical evaluations of field data, indicated that penetration rates can be increased by an average of 24% when drilling dry with high air pressure and by approximately 27.5% when drilling dry with mine air as compared with wet drilling under the same conditions. Dry drilling also has the effect of reducing button wear by approximately 69% using high pressure air. The underground test work employed a dust collection system developed for use with large diameter holes (165 mm) and high pressure drilling. The testing program has indicated that an average increase in penetration rate of 35% can be achieved when dry drilling in rock (waste) and of 49% when dry drilling in ore with high air pressure. Although preliminary dust monitoring did not indicate any appreciable increase in dust concentration during dry drilling, further dust monitoring is still required.A comparative cost analysis has indicated the potential for reduction in drilling costs when dry drilling is utilized. Further testwork is to be conducted to develop a reliable dust collection system for underground production."

Jan 1, 1992

By J. J. Taber

The turbulent flow drag coefficients, or friction factors, have been experimentally determined for the cut-tings normally encountered in drilling operations. The gas law and average drag coefficients for characteristically flat particles (limestones and Shales) and for angular to sub-rounded particles (sandstones) were used to extend Newton's equation to a more useful form. The resulting equations are expressions for the slip velocity of a particle as a function of the particle size and shape, the bottom-hole injection Pressure, and the Bottom-hole temperature. The velocities necessary to lift actual rock cuttings as observed in the laboratory were compared with velocities predicted from the derived equations. Results indicate that there is no single correct circulating velocity and that the commonly used 3,000 ft/min linear velocity may he sufficient to lift only very mall particles. The advantage, in terms of horsepower requirements, of cycling high pressure air to obtain lower compression ratios was shown. INTRODUCTION A promising departure from standard rotary hydraulic drilling is air or gas drilling. In many instances the rate of penetration and bit life increases resulting from this method have been very substantial. Further the application of air drilling in areas of lost circulation or water sensitive pay zones has been highly successful. Purpose For the drilling contractor, the practical question remains, "How much air pressure and volume should I have?' lt is necessary to have a reasonable knowledge of the air velocities and pressures required in air drilling operations for the proper design and most advantageous use of surface equipment and for adequate removal of cuttings from the borehole. Eased on experience, most operators agree that at least 3,000 ft/min linear velocity is necessary for satisfactory hole cleaning.' It was the purpose of this investigation to evaluate the drag coefficients, or friction factors, for the cuttings normally encountered in drilling operations, i.e., sandstones, limestones and shales, and to utilize the values obtained to develop an expression for the minimum velocities and pressures (hence the volumes) necessary to lift these cuttings. THEORY Several investigators have done extensive work on the ability of drilling mud to lift bit cuttings. They conclude that the ability of a drilling mud to lift cuttings is dependent on the density difference between the rock being drilled and the drilling mud, the cutting size and shape, the mud flow constants, and the flow state (laminar, transitional or turbulent) of the mud. However, drilling muds are not true fluids but have a variation of viscosity with velocity in laminar flow and a constant apparent viscosity in turbulent flow. A survey of the literature revealed that no experimental work had been done to determine the velocities necessary to lift actual rock cuttings using a compressible fluid such as air or gas. In 1953, Nicolson6 stated that in air drilling fluid mechanics the particles could be assumed spherical in shape and that ; constant drag coefficient of 0.5 could be used due to the highly turbulent flow of the circulating air. By equating the turbulent resistance on the particle to the gravitational force on the particle, he obtained the expression, ^ = 2.67^y ,.......(1) where V, is slip velocity of spherical particle, ft/sec; D is particle diameter, in.; p, is particle density, 1b/ft8; and p, is fluid density, lb/ft3. However, as stated earlier, most operators seem to agree that adequate hole cleaning can be obtained by circulating a sufficient volume of air to give a linear velocity of 3,000 ft/rnin in the annular space. Basic Equation The velocity with which a solid particle freely falls through a fluid will increase until the accelerating force, gravity, is equal to the resisting forces, buoyancy due to the fluid displaced by the particle and friction due to the relative motion of the particle through the fluid. When the accelerating and resisting forces become equal,

By D. J. McAdam, R. W. Mebs

THE technical cohesive strength of a metal means, not the interatomic forces, but the technically estimated resistance to fracture. An example of such resistance to fracture is the so-called "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture, According to the prevalent view,8 the technical cohesive strength of a metal, in any particular state as regards prior mechanical treatment and heat-treatment, is determined by a limiting value of the algebraically greatest principal stress.? In two previous papers by one of the authors,14,15 evidence was presented to indicate that the technical cohesive strength of a metal cannot be represented by a single stress value, but that it comprises an infinite number of values corresponding to the infinite number of possible combinations of the principal stresses. The technical cohesive strength thus comprises an infinite number of technical cohesion limits, each representing a specific stress combination at fracture. In the same papers, evidence was presented that plastic extension causes a continuous increase in the technical cohesive strength. Such a variation is not in accordance with the prevalent view that plastic extension first increases, then decreases, the technical cohesive strength.8 In this paper, as in the previous papers,14,15 the algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated S2. The technical cohesion limit under polarsymmetric tension (S1 = S2 = S3) will be termed the "disruptive stress," and the cohesion limit under unidirectional tension (S2 and S3 = o) will be termed the "severing stress." 14,15 The yield strength of a metal, according to the generally accepted theory, may be represented by a constant value of the shearing-strain energy, and hence by a constant value of the sum of the squares of the three principal stress differences. In the second of the previous papers,"' however, evidence was presented that the limiting yield stress decreases with increase in the volume stress, the algebraic average of the three principal stresses.* This influence of the volume stress becomes prominent only in the field of triaxial tension. In the two previous papers, 14,15 diagrams were presented to show the variation of the technical cohesion limit, yield stress, and ultimate stress with the combination of principal stresses and with plastic extension. Those diagrams and the resultant conclusions are derived from experimental data obtained by other investigators. They are based chiefly on results of tension tests

Jan 1, 1943

Jan 1, 1955