Search Documents

By Priscilla P. Nelson, Francis L. C. Fong

Introduction Fracture mechanics is the study of the behavior of cracks and their modes of propagation. During cracking, energy is consumed in creating new surface area. The more energy a material absorbs, the higher is its fracture resistance. This fracture energy is provided either by work done externally, by the release of elastic strain energy within the cracked body, or by a combination of these two effects. Fracture resistance is measured in terms of either the stress intensity factor, K, or the energy release rate (also referred to as crack driving force), G. For linearly elastic materials, these two quantities are related by the following equation: [ ] where [u] is Poisson's ratio and E is the modulus of elasticity. Both K and G are direct functions of the applied force and the crack length. The process of rock cutting with full-face tunnel boring machines (TBMs) involves the indentation of a rock surface by rolling disc tools. Excavation occurs as fractures, extending from the locus of chip indentation, propagate between concentric disc grooves and liberate rock chips from the intact rock face. Since disc cutting is achieved by a fracture process, it is logical to anticipate that differences in cuttability of a variety of rocks should be related to the variation of fracture material properties among the rocks. Such a relation has been demonstrated between the performance of full-face TBMs and the critical crack driving force for excavation in relatively high strength, brittle, low porosity, isotropic rock (Nelson et al., 1985). The objective of this paper is to report on the results of a continued investigation of the relationship between fracture material properties and the effectiveness of disc cutting. Additional reliable data on full-face TBM performance is difficult to obtain, so that the results of disc linear cutter tests performed by the Transport and Road Research Laboratory (TRRL) (U.K.) are utilized in this study. Fracture Material Property Evaluation If a crack tip is subjected to a perfectly plain strain condition, many materials will behave in a brittle manner, and the material resistance to fracture is constant during crack growth. This resistance is referred to as the plain strain fracture toughness, believed to be a material property. The standard testing procedure for this parameter is given in the American Society for Testing and Materials (ASTM) test designation E399. To ensure plain strain conditions, acceptable specimen geometries are specified. Unless these geometric specifications are adhered to, the measured fracture resistance is regarded as the apparent fracture toughness, and a subscript Q is used, as [ ]. The methodology adopted by ASTM E399 is within the framework of Linear Elastic Fracture Mechanics (LEFM). This approach assumes that material resistance against cracking is developed instantaneously, without initial stable or subcritical crack extension. Prior to crack advancement in rock, however, much of the crack driving energy is absorbed in a localized volume in front of the crack tip, referred to as the plasticity or microcracking zone. This capability for energy absorption substantially increases material resistance, and a crack may be arrested unless an increased crack driving force is applied. The LEFM or plain strain assumptions would underestimate the actual fracture resistance of such a material. Such a conservative approximation may be desirable in the design of structures against fatigue or brittle failure. In mining or tunneling, however, an underestimation of crack resistance can lead to the selection of underrated excavation equipment and a decrease in production efficiency. The development of fracture resistance in rock is best described by the material resistance curve, or R-curve, which is a plot of fracture resistance as a function of crack extension. Typically, peak resistance is only reached after a finite amount of subcritical crack growth. The

Jan 1, 1986

By Sidney Katell

Covering the period from January 1961 through December 1961, this report is a continuation of the series that began with the publication of Information Circular 7516 in October 1949. It is also the third report in the series to be prepared as a cooperative effort by the Bureau of Mines and the American Association of Cost Engineers. The cooperative agreement between the two groups provides for the publication of an annual cost bibliography with the work load involved in the preparation shared equally. The abstracted articles, concerned with all facets of cost engineering in the field of chemical and petroleum plants and related facilities, came from the following sources: 1. A.A.C.E. Bulletin. 2. Bureau of Mines Information Circulars. 3. Bureau of Mines Reports of Investigation. 4. Chemical Engineering. 5. Chemical Engineering Progress. 6. Chemical Week. 7. Coal Age. 8. Coal Utilization. 9. Cost Engineering. 10. Industrial and Engineering Chemistry. 11. Nucleonics 12. Oil and Gas Journal. 13. Petroleum Engineer for Management. 14. Petroleum Refiner. 15. World Oil.

Jan 1, 1962

By Richard L. Gordon

Minerals long have been important commodities in international trade. As an inevitable result, the governments of the world have employed a wide variety of programs that affect the flow of trade. Roughly speaking, we may distinguish between the desire of more developed countries (MDCs) like the United States, western Europe, and Japan to control imports of minerals and the concern of less developed countries (LDCs), like India, about exploiting and exporting their minerals. The terminology used to describe different countries and the classification of specific countries into groups is unsettled and often changing. The poorest countries have been characterized as backward, undeveloped, underdeveloped, and developing. The problem seems to be in selecting a term that is both accurate and polite. Less developed seems, for reasons outlined by Hagen [pp. 51-57], the most convenient term. The borderline between an LDC and an MDC is always open to dispute. Objective measures are difficult to develop because of deficiencies in basic data, and enormous conceptual problems arise in making valid comparisons from even good data [see Hagen, pp. 7-28]. Thus, policies of the MDCs can be evaluated chiefly by using a few basic arguments for protectionist attitudes towards international trade. In dealing with the LDCs, a more far-ranging evaluation that covers general economic policy goals, the role of foreign investment, and the indirect effects of trade is required. This duality breaks down when trying to deal with the depressed areas in an MDC; here the issues are more akin to those associated with problems in an LDC. Similarly, variants of arguments used in LDCs appear in debates about foreign investments in MDCs. The whole discussion of minerals in LDCs, moreover, has obvious implications for MDCs like Canada and Australia. They, too, must face problems about the best utilization of their resources. THE RESTRICTION OF MINERAL TRADE BY DEVELOPED COUNTRIES[a] At least to some extent, the world's leading industrial nations have intervened to control the inflow of minerals. Such regulation is particularly pronounced for energy. Both the United States and western Europe employ

Jan 1, 1976

By Kenneth K. Humphreys, F. R. Zachar, A. G. Gilbert

INTRODUCTION by F. R. ZACHAR and A. G. GILBERT Coal producers are not only faced with the problems of mining coal, but must also decide to what extent they can invest capital and what operating and maintenance expense they can afford, after mining, to market their product in sufficient volume and at a reasonable profit. Regardless of whether the operator is contemplating a new operation or merely considering expanding present facilities, market requirements, ability to make a product meeting requirements, probable costs and possible future changes in raw coal or finished product must be considered. In addition, such items as flow sheet, structural designs, storage facilities, etc, must be examined individually and collectively. In estimating preparation costs to meet predetermined market specifications, the operator must calculate the economics of present preparation practices to anticipate the increase to total production which results when the quality of the finished product is increased. The consumers of coal are continually seeking to obtain the best and most adaptable product for their money. Regardless of whether coal is used for coking, industrial requirements, electric power generation, domestic fuel or for other purposes, it will be scrutinized from such standpoints as delivered cost, suitability for utilization in certain equipment, ease of ash disposal, adaptability to material handling system plans and its potential for generating dust problems. Utilities, due to large capacity volume requirements, are generally in a position to select coal that will result in the lowest over-all cost per million Btu. Freight costs, paid by the utility, are important and are given utmost consideration when a coal is under consideration. Steel companies, whether producing or purchasing coking coal, are continually seeking coal that will result in lowering the cost of pig iron production. Since it is usually necessary to clean to lower specific gravities, say 1.40, to turn out a product meeting metallurgical requirements, it follows that a high ash "middling" product is created. These low quality middlings do not command a good price. Therefore, unless the metallurgical coal can command sufficient realization to carry or subsidize the middlings, the operator may be ahead economically to prepare steam coal at a higher specific gravity. While the "unit-train" rate reductions have been helpful to the economy of the coal industry, the establishment of these rates has made mandatory the installation of storage and rapid car loading facilities at mines and plants shipping utility coal (see Chapter 15). A study of the economics

Jan 1, 1968

By A. Romero-Juarez

This paper is concerned with the effect of propping-agent concentration on flow capacity of a fracture in the case in which there is embedment of the propping agent. Previous published studies have shown definitely that there is a relationship between fracture flow capacity and prop ping-agent concentration, and it has been shown that the theoretical results are confirmed by laboratory experiments. The problem of directly finding sand concentration for maximum flow capacity of a sand-propped fracture is solved, and formulas and charts are given to obtain this concentration under various conditions of effective overburden pressure, medium sand-grain diameter and rock properties. It is shown that, for the same effective overburden pressure and the same rock characteristics, optimum sand concentration in pounds per gallon does not depend on medium sand diameter. For conditions met in bydraulic fracturing operations, it is found that sand concentration in grains per square inch for maximum flow capacity varies within a wide range of values; this indicates the convenience of using data of fracturing pressure and rock characteristics for calculating sand concentration in order to achieve the best results in fracture treatments. INTRODUCTION It has been pointed out in the published literature1 that one important factor controls the success of a hydraulic fracturing operation — the propping of the fracture. The main effect of the propping agent is to hold the fracture open by means of the reaction forces that oppose the pressure due to the overburden. It is assumed in this paper that the propping agent is sand, that the grains are spherical and uniform in size, and that this sand is distributed in a monolayer in the fracture. The existence of a sand concentration value for maximum fracture flow capacity has been recognized since the publication of one of the first studies on the subject.2 In the reference cited, flow capacity was assumed to be proportional to the cube of the fracture volume per square inch not filled by sand. The obtained experimental curves relating to flow capacity and sand concentration had the same appearance as the curves obtained with the afore-mentioned assumption. Recent work3 has shown that fracture permeability, in the case of a monolayer, can be calculated accurately by means of a modified Kozeny-Cacman equation; thus, it can be shown that fracture pqmeability is a function of free volume in the fracture and of the rock and sand wet surfaces. The flow capacity of a fracture is dependent on the permeability of the fracture and on the fracture width. If all excessive quantity of sand is present in the fracture, its width may be large but its permeability i:; low due to the greater flow resistance; too few grains of sand per square inch are present in the fracture, sand embedment in the rock is greater and the width becomes small enough to appreciably decrease the rate of flow. To determine optimum conditions of flow, therefore, it is necessary to know the dependence of fracture width on density or concentration of propping agent. This relation is given in a study4 in which the approach to embedment is similar to the one made in a penetration hardness test of metal. With the equacions presented by Darin and Huitt,3 the number of sand grains per square inch that results in maxirnum fracture capacity can be determined graphically by plotting fracture capacity vs sand concentration. Since the relationship between fracture capacity and sand concentration may be considered known, one is led to formulate the analytical problem of directly determining sand concentration for maximum fracture flow capacity, in order to study the effect of the various quantities that come into the process. The solution of this problem appears to be interesting both from the viewpoint of fluid mechanics and from the practical viewpoint of oil production in hydraulic fracturing treatments. EQUATION FOR FRACTURE FLOW CAPACITY It can be shown3 that the fracture flow capacity of a monolayer is given by the following expression,

By Kenneth H. Johnston

This publication describes a thermal-recovery project and two waterflood projects in Carter, Garvin, and Stephens Counties, Okla., that will be visited during the 1966 annual tour by the Bureau of Mines in cooperation with the Kansas-Oklahoma Waterflood Association. Among the unique features of the projects arc (1) the movement of heavy (15° to 28° API gravity) crude oil through the reservoir and into the well bore with energy furnished by burning a portion of the crude oil therein; (2) the use of an inverted nine-spot pattern with dually completed injection wells, to produce oil from deep (8,200 to 10,200 feet) reservoirs; and (3) the flooding of five sand lenses separately by using controlled water injections through multizone completed injection wells. Discussion of the individual projects includes information on their early history, method of completing and operating wells, source and treatment of injection fluids, oil production decline curves, and the results obtained from the development.

Jan 1, 1966

By Jon G. Bredeson

A survey of electromagnetic noise reports was performed and a summary of the results are given. Industrial concerns were surveyed for techniques to minimize the effects of electromagnetic noise and are presented. The common types of communication links used in mines today are presented. A mathematical development was presented which determines the probability of bit errors versus distance for various types of communication links designed for in mine use. The probability of undetected errors for the protocols of the communication links is given together with burst error detection.

Jan 1, 1981

By John A. Kiefer

Ground fault and overcurrent protection criteria for coal mine a.c. distribution systems were examined under a Bureau of Mines funded investigation. The report begins with a literature review which discusses relaying topics including instrument transformers, ground-fault pickup methods, static relaying, and comparison of relaying techniques. A chapter on ground-fault relaying provides recommendations for maximum and minimum pickup levels as a function of maximum ground current and power system line-to-ground capacitance. Research in the area of phase overcurrent protection has resulted in a complete set of procedural recommendations for the selection and setting of overcurrent relays and current transformers, molded-case circuit breakers, and distribution system fuses in coal mine power systems. The final chapter, which discusses relay system maintenance, introduces a unique test set which was developed by KETRON and used to test the performance of relaying systems in coal mines.

Jan 1, 1980

By C. Perez-Rosales

A simplified method for evaluating specific surface of porous media is described. The theory takes as its point of departure a previous statistical method. Due to its simplicity, it is considered that the proposed method has certain advantages over methods currently used. By making some simplifications, the procedure for estimating specific surface is reduced to superposing, on an enlarged reproduction of a section of the sample to be analyzed, an evenly spaced grid and counting the number of intersections of the grid lines with the perimeter of pores. An indirect test of the theory is presented and the advantages of the method are discussed. INTRODUCTION The specific surface of a porous medium, defined as the surface area of the pores per unit bulk volume, is an important parameter with regard to flow of fluids through porous materials. In particular, it has long been recognized that specific surface is closely related to permeability. Since the internal surface of any porous material is extremely complex, the measurement of specific surface can only be accomplished by indirect means. It is considered that the best method' in current use for determining specific surface is the statistical method of Chalkley, Cornfield and Parka who originally applied it for estimating volume-surface ratios of morphologic components, such as cells or nuclei. This method will be outlined later. A method which has been used extensively in determining specific surface is based on the adsorption of a vapor by the internal surface of porous media. However, this method does not yield the surface pertinent to fluid flow, since it includes the tiny molecular interstices which do not influence the permeability of the porous material.' This method has been employed by Brooks and Purcell measure internal surface area of sedimentary rocks. The Kozeny equation, which relates the specific surface to the permeability of a porous medium, has found wide use in determination of specific surface. However, the KOzeny equation is not strictly correct. Large discrepancies have been found between theory and experiments.' Consequently. the specific surface estimated by this method is subject to serious uncertainties. Finally, there are other methods for specific surface measurements which are based upon a variety of physical and chemical properties: the heat conduction by a gas, ionic adsorption phenomena and direct chemical reactions of the surface with other substances. This paper describes a simplified statistical method for determining specific surface of porous materials. It is considered that, due to its simplicity, it has certain advantages over the methods presently used. For its validity, it is required that the porous structure of the material under study be of a spatially random nature. This is not a serious limitation since the naturally occurring porous media have this characteristic. THEORY The method proposed in this paper takes as its point of departure the interesting result obtained by Chalkley et al The method shows that if a line of length r is dropped repeatedly at random over a section of a porous medium, and counts are made of the number of times the two end points fall within pore areas (denoted by h for "hits") and the number of times the line intersects the perimeter of pores (denoted by c for "cuts"), then in a very large number of throws where S is the internal surface of the sample (surface of pores) and V, the volume of the pores (Fig. 1). Since the specific surface S, and the porosity 4 are given, respectively, by S. = Sl V, and 4 = V,l V,, where V, is the total volume of the sample, it follows from Eq. 1

By L. D. Jaffe

From the limited experimental data in the literature, preliminary values were derived for the thermal diffusivity of titanium alloys and for the quenching severity of various mediums used in heat treating them. The thermal diffusivity, under quenching conditions, was found to be approximately 0.006 sq in. per sec. The quenching severity H for brine is approximately 8, for water 4, for oil 0.9, for air 0.08, and for argon 0.02 in. -' The quenching severity of the Jominy-Boegehold end quench water jet against titanium alloys appeared effectively infinite. Using these results, graphs were prepared showing ideal round sizes and Jominy positions having cooled conditions equivalent to those at the center and surface of titanium alloy rounds, plates, and sheets quenched in various media. THE importance of proper quenching of steel parts has long been recognized, and much quantitative work has been done on the subject. Recently, it has become evident that for titanium alloys, too, the rate of cooling from high temperature is important in obtaining high strength and hardness. Maximum hardness and strength may not be attained in the as-quenched condition but rather after a subsequent tempering or aging treatment. Nevertheless, if cooling from temperatures near the ß transus is too slow, for the hardenability of the alloy used, a soft a-ß structure will be produced which will not harden on subsequent aging.'-" The quenching of parts of practical interest can usually be approximated by simple shapes, such as cylinders (rounds) and plates (or sheets) of various thickness and combinations of these. A standard that has been used extensively for comparison purposes is the center of an ideal round: a cylinder quenched in a medium that instantly cools its surface to room temperature." For steel, considerable practical use has been made of the Jominy-Boegehold end quench hardenability specimen,".' and its value for titanium alloys is becoming evident. It therefore appeared worthwhile to attempt comparison of cooling conditions in titanium alloy rounds and plates of various sizes quenched in various media with cooling conditions in ideal round and Jominy quenches. These comparisons are based upon measurements reported in the literature. Quenching Severity of Jominy Water Jet and of Brine As one starting point, there were available the data of Phillips and Tobin8 for positions on a Jominy specimen having the same hardness as the centers of round bars of various size quenched in 5 pct NaOH brine, shown in Table I. The Jominy speci- men is a round bar, normally 1 in. diam and 4 in. long, quenched by a water jet at one end and air cooled on the other surfaces.' After quenching, hardness is measured vs distance from the quenched end on longitudinal flats ground 0.015 to 0.10 in. deep. The data of Table I were based on a subsized specimen of 3/4 in. diam. Since the positions tested are close to the jet quenched end, cooling there may be taken as due solely to the jet, and air cooling can be neglected. These positions are then equivalent to various locations near one surface of an 8 in. plate jet quenched on both sides. Assuming that the thermal diffusivity is constant and that Newton's law of cooling holds, temperature vs time curves at corresponding positions in Jominy specimen and round bar then may be calculated for various assumed quenching severities of the water jet and of the brine. Cooling curves are generally taken to be metallurgically equivalent when they show the same time to reach a temperature halfway between the initial (heat treating) temperature and the final (quenching medium) temperature.5 On this basis, using Russell's tables,9 upplemented by heat flow tables derived in similar fashion by the author for positions near the surface of severely quenched plates, it was found that the experimentally observed equivalence, Table I, was best matched if the quenching severity H of the Jominy water jet was taken as infinite (H = co). This means that the jet quenched surface should be considered as instantly cooled to the temperature of the water

Jan 1, 1956

By A. M. Adair

The effect of cathodically charging zone-refined iron with hydrogen and the resultant yield stress-grain size relationship has been investigated by tensile testing at -31°C. The resulting Petch plots (using a least-squares analysis) reveal a significant reduction in the slope and an increase in the intercept for specimens cathodically charged with hydrogen, regardless of whether the hydrogen is present during testing or is removed by evacuation. Other than surface blistering, the major effect of cathodically charging this material is to create occasional grain boundary cracks and thus presumably a large number of free dislocations at grain boundaries. These results are consistent with the hypothesis that the slope of the Petch plot is related to the stress required to activate dislocation sources at or near pain boundaries. While there is no significant effect of the presence of hydrogen on the lower yield stress of this material, the ductility is considerably reduced. After hydrogen removal by evacuation, the ductility recovers substantially and almost complete recovery of ductility is experienced by the smaller grain size specimens. THE influence of hydrogen on the mechanical properties of ferritic materials has received much attention throughout the history of the metallurgy profession. This interest has stemmed from the obvious importance of preventing the catastrophic failure of parts in service, and from the desire to determine what mechanism is responsible for the observed behavior. The various mechanisms proposed for hydrogen em-brittlement have been discussed in detail by Tetel-man,'who concludes that a combination of the applied stress and hydrogen pressure within voids best accounts for the many experimental observations reported. From a quite reasonable correlation of the strain aging and delayed-failure behavior of a variety of materials,2-7 Troiano and coworkers are of the opinion that the delayed-failure mechanism (slow strain rate embrittlement as well) is better explained as a lattice effect; solute elements diffuse to regions of high triaxial stress, form uncondensed atmospheres there, and effect a reduction in the cohesive strength of the lattice. Since hydrogen yield points8,9 have been observed in ferritic materials, and since the hydrogen embrittlement process may be related to strain aging, one further contribution to embrittlement could be hydrogen-dislocation interactions which make crack blunting by plastic deformation a more difficult process. The etch" relationship offers a means of comparing the relative ease of initiating and propagating plastic deformation in metals and alloys. In the present study, this relationship was used to compare the behavior of zone-refined iron in three conditions: uncharged, cathodically charged with hydrogen, and cathodically charged with hydrogen and evacuated. These Petch plots and the observed ductility parameters allowed a separation of the effects of the very presence of hydrogen and the act of cathodically charging hydrogen into this material. MATERIALS AND EXPERIMENTAL PROCEDURE An ingot of zone-refined iron was obtained from Battelle Memorial Institute which contained the following (typical) interstitial impurities in ppm: carbon, 10 * 5; nitrogen, 1; oxygen, 22; hydrogen, 0.9. The ingot was cold-reduced to a 1/2-in.-sq cross section, recrystallized 1 hr at 575°C, and then straight-rolled to a thickness of about 0.105 in. Tensile specimens were milled from the strip having a 1-in. gage length

Jan 1, 1967

By Harlan W. Nelson, James B. Purdy

A study of data obtained during laboratory tests to determine the suitability of bituminous coals for use in residential underfeed stokers of the clinkering type has led to the following general conclusions: 1. Performance data obtained during the test furnish useful and practical information upon which to judge the performance of stoker coals. 2. Conclusions derived from the results of evaluation tests are in general agreement with use experience of the coals in the field. The method is of practical use in this regard. 3. Few direct relationships were found between data furnished by the evaluation test and determinations by standard laboratory tests. This is particularly true with regard to coke formation and clinker formation, two of the most important factors in determining the degree of satisfaction to be obtained in applications of the small underfeed stoker. It has not been the intention in this paper to prescribe this evaluation test as the best method for the evaluation of coals for use in residential stokers. That there are other test methods having similar objectives serves to point out the necessity for full-scale laboratory tests. At the present time, a standard test procedure for the evaluation of coals for use in residential stokers is not available. For the past several years a joint committee, appointed from the membership of the Residential Stoker Committee of Bitulninous Coal Research, Inc., and from the Engineer- ing and Research Committee of The Stoker Manufacturers Association, has been working to develop a standard procedure for testing and evaluating bituminous stoker coals. This procedure has been completed to the extent that it has been drawn up in tentative form. It is now planned to install the equipment in several laboratories and to run check tests, using portions of identical lots of coal, to determine the degree of reproducibility offered by the method. Completion of this standard test procedure should greatly facilitate the evaluation of coals for use in residential stokers. Acknowledgments The cooperation of coal operators sponsoring the coal evaluation tests, in granting permission to use data and results from their tests, is gratefully acknowledged. Acknowledgment is also made to Carroll F. Hardy, Chief Engineer, Appalachian Coals, Inc., for his interest and timely suggestions on the preparation of this paper. Special thanks are due to Ralph Sherman, Assistant Director, Battelle Memorial Institute, who was largely responsible for development of the evaluation procedure, for helpful suggestions and advice. References I. H. A. Sherman: The Evaluation of Coal for Use in Domestic Stokers. Univ. of Ill., Eng. Experiment Station, Circular No. 39 (1939). 2. . Q. Shotts: The Relation of Free- swelling Indexes to Other Characteristics of Some Alabama Domestic Stoker Coals. TP 2314, Coal Tech., (Feb. 1948); Trans. AIME (1948) 177, 502. 3. . J. Helfinstine and C. C. Boley: Correlation of Domestic Stoker Combustion with Laboratory Tests and Types of Fuels 11. Combustion Tests and Preparation Studies of Representative Illinois Coals. Ill. State Geol. Survey, Re1. of Zntlestigations No. 120 (19467 ) p. 35. DISCUSSION E. R. Kaiser*—AS a former fuel engineer assigned to residential stokers at Battelle, the writer has read with considerable interest the paper by Messrs. I'urdy and Nelson. It is gratifying to note that earlier test techniques have been developed and that criteria found important in the years 1935 to 1938 are still important in judging the performance of stoker coals for residential heating. The authors have plotted a number of the primary coal data singly against performance data of the stoker in an effort to establish relationships. Unfortunately, wide scattering of points usually resulted, which did not illustrate what experience generally indicates to be the case. For example, Fig 3 indicates almost no trend in the percentage of released ash converted to clinker with change in ash-softening temperature. Other factors, such as fuel bed temperatures and zones of heat release must have influenced the results. In the present state of our knowledge, we cannot explain all of the reasons why suitable stoker coals perform satisfactorily, nor why one satisfactory coal may be better than another. It would therefore

Jan 1, 1950

By Robert W. Lawson

INTRODUCTION Prior to 1705, underground mining was limited to mineral deposits located above or a very small distance below the water table. The development of the steamdriven mining pump was motivated by the need to extract minerals from deposits considerably lower than the water table. Today, the majority of mine pumps are driven by electric motors, although some small utility pumps are driven by compressed air. Of the numerous types of pumps that might be applied to mine service, the two types used most frequently underground are the centrifugal and plunger pumps. Each of these types has its proper application. In most instances, plunger pumps are best for handling low water volumes against high heads, while centrifugal pumps effectively handle either low or high water volumes against low heads or high water volumes against high heads. This chapter describes pumps applied to underground mine service, where the mine service consists of mine drainage (removing mine water as it accumulates), mine dewatering (removing water from a flooded mine), and pumping for special applications such as hydraulic mining and hydraulic hoisting. CENTRIFUGAL AND PLUNGER PUMPS Modern mine pumps are classified as either centrifugal (kinetic) or plunger (positive-displacement) pumps. The following paragraphs describe each of these classifications. Centrifugal Pumps Centrifugal pumps may be grouped into various types, depending upon their various features: Number of Stages: In a single-stage pump, the total head is developed by one impeller, while a multi¬stage pump utilizes two or more impellers acting in one casing to develop the total head. Fig. I shows a cutaway view of a single-stage horizontal split-case double -section centrifugal pump. Fig. 2 shows a cutaway view of a six-stage medium-pressure split-case diffuser-type centrifugal pump. Type of Casing: A volute pump has a casing in the form of a spiral or volute. A circular-casing pump has a casing of constant cross section concentric with the impeller. A diffuser pump is equipped with a diffuser that converts velocity head to pressure head. Shaft Position: A horizontal pump normally has the shaft in a horizontal position, while a vertical pump normally has the shaft in a vertical position. Fig. 3 illustrates a vertical turbine-type slurry pump. Suction: A single-suction pump equips the first stage with a single-suction impeller, with fluid entering the impeller on only one side. A double-suction pump equips

Jan 1, 1982

By R. R. Angel

Drilling rate is a parameter that should be considered in determining the volume requirements for air and gas drilling. The use of past methods which ignore the effects of the solids content upon the pressure and velocity of the annulus flow stream can result in undercalculation of the required volurne by as much as 50 per cent. A vertical-flow equation is presented for determining volume requirements. This equation includes the effect of the solids that are transported up the annulus in the flow stream by incorporating the drilling rate as one of the parameters. The effect of down-hole temperature on required circulation rates is also ana-lvzed. A simple approximate method of determining volume requirements is presented. This method is more accurate than the methods used in the past. Hole cleaning difficulties are analyzed for a recent air drilling job where past methods indicated that - excess air was being used. Sample curves of calculated boitom-hole pressures are presented for air and gas drilling in several hole sizes. INTRODUCTION In certain areas the use of air or natural gas as a circulating medium for drilling oil and gas wells is becoming a common practice. Large increases in penetration rate and bit life are achieved through the use of these media in preference to mud or water. Drilling rates as high as 90 ft/hr have been obtained in shales. The importance of maintaining adequate circulation rate is generally recognized; however, much disagreement exists among drilling operators as to what constitutes "adequate" circulation rate. In quarry drilling, where annular velocities can be accurately determined, an annular velocity of 3,000 ft/min of standard air is required for best results in rocks having approximately the same density as those commonly penetrated in drilling oil or gas wells'. Although this standard air velocity has proven satisfactory for quarry drilling, some oil and gas well drilling operators believe that an equivalent annular velocity of more than 4,000 ft/min is required; others believe that as little as 2,000 ft/min . is sufficient. Much of this disagreement results from determining the required circulation rates with methods which fail to incorporate the drilled solids in an equation which is applicable to vertical flow. Hughes Tool Co. Bulletins No. 23' and 23-A3 present data for determining circulation rates based on the Weymouth formula. These data do not include the drilling rate as a parameter and, therefore, neglect the effect of the solids being transported up the annulus. In spite of this apparent defect and the fact that the Weymouth formula is not applicable to vertical flow, the Hughes data have well served the drilling industry in many areas and are important and timely contributions to the science of air and gas drilling. The Hughes data purposely omit a correction for increasing down-hole temperature. At slow drilling rates this effectively compensates for the use of a formula which is not valid for vertical flow; however, volumes determined by the Hughes method are not sufficient to support rapid drilling rates at moderate and great depths. For example, Phillips Petroleum Co.'s Cauthorn "D" No. l in the Vinegarone field for Val Verde County, Tex., was air drilled from 1,500 to 9,300 ft using a compressor delivering 1,400 cu ft/min. The 8 3/4-in hole was drilled with 5-in. drill pipe, and drilling rates as high as 90 ft/hr were obtained between 7,000 and 9,300 ft. No water or caving hole was encountered. At 7,728 ft it was necessary to wash-out 60 ft of cuttings to reach bottom after a trip to change bits. At 8,130 ft a twist-off occurred and the drill col-lars were stuck in drill cuttings. These difficulties indicate that the 1,400 cu ft/min of air was not sufficient to keep the hole clean. Hughes data indicates that 1,180 cu ft/min is sufficient to produce an annular ve-

Jan 1, 1958

By James R. Aggson

This Bureau of Mines report describes floor heave ground control problems that have been encountered in a new underground coal mine in West Virginia. Previous experience in the coal seam, the results of physical property investigations, and in situ rock stress determinations are discussed. The apparent cause of the ground control problems is the existence of a high, biaxial, horizontal stress field coupled with a floor member that has time-dependent deformation characteristics, A theoretical analysis of the failure process is developed. The analysis is assisted by finite element techniques. Mine design recommendations based on this analysis are presented.

Jan 1, 1978

By Arthur C. Green

A group of 13 cities situated in Los Angeles and Orange counties in Southern California is engaged in constructing an aqueduct to carry water from the Colorado River at a point near Parker, Arizona, to a distribution point near Riverside, California. The aqueduct will have a total length of approximately 241 miles, and is designed to deliver 1605 cu. ft. of water per second, or a little more than one billion gallons per day. The aqueduct crosses a barren, desert waste and the project includes approximately 91 miles of 16-ft. tunnel. The large amount of tunnel driving that has been done in this country in the past few years has resulted in some remarkable developments in equipment and methods. The building of the Colorado River Aqueduct,, with its 91 miles of tunnels, has called on the best engineering knowledge in the country. New and better equipment has been designed for all phases of the work, and this equipment has been coordinated and adapted to the various conditions with a high degree of efficiency. The result has been that records for tunnel driving have been set under adverse natural conditions that exceed the estimates for progress under good conditions. All of the tunnels in the main aqueduct, except the Valverde and Bernasconi at the western end, have a finished diameter, inside the concrete lining, of 16 ft. Fig. 4 shows a section of the tunnel with the concrete lining in place. Fig. 5 illustrates the finished tunnel in three typical sections. The standard tunnel grade is 3.432 ft. per mile in the 16-ft. tunnels, but the Valverde and Bernasconi tunnels have a steeper grade and a finished diameter of 15 ft. 3 in. The rough bore of the large main aqueduct tunnels averages 12.95 solid cubic yards of material per foot of advance in timbered sections and 10.76 solid cubic yards per foot in unsupported sections. Tunnel Driving While the Conway mucker has been used on tunnel work for years, a larger and more powerful Conway was designed to meet the require-

Jan 1, 1937

By V. Tang

In this paper, Chinese economic situation are estimated and prospective developments predicted to assist those US firms who have been trying to do business in China, or are considering to pursue the China market.

Jan 1, 2010

INTRODUCTION The regional characteristics of porphyry copper deposits in South America southward from Pantanos and Pegadorcito, Columbia, will be summarized. The age of formation of deposits spans the period from the Permian to the present. Where possible, the general characteristics and the individual structural features of the deposits are integrated into the plate tectonic model. Individual deposits cited in this chapter, located in Fig. 1, are the best known and most fully described, and form the Andean copper belt (Peterson, 1958). Smaller porphyry copper deposits fall both east and west of the line of large, high grade deposits (e.g., Chuquicamata and Toquepala) that make up the spine of this belt. Most of those smaller deposits (e.g., Mi Vida [Koukharsky and Mirre, 19761) are omitted from Fig. 1 and Tables 1 and 2, however. Neither the Andean orogen nor the individual deposits within it are as well de- scribed in the literature as are their counter- parts in the Cordilleran orogen of North America. Therefore, generalizations relying on regional geologic mapping, geophysics, and isotopic geochemistry lack details pertinent to this type of deposit in the Cordilleran orogen. The South American porphyry copper province has been described by Hollister ( 1974) to include most of the Andean orogen excepting the tin province of Bolivia. Porphyry copper deposits have not been commonly found within the tin belt, although copper-tin (e.g., San Rafael, Peru) and copper-tungsten (e.g., Llamuco, Chile) de- posits are known. James (1971) shows the depth to the Moho to be appreciably greater where the tin belt exists, but disclaims any genetic significance from the coincident appearance of the tin-tungsten province and the inhibition of porphyry copper development where the sialic crust thickens. Dates on tin mineralization in this belt (Clark and Farrar, 1973) are similar to some dates for porphyry copper mineralization known else- where in the Andean orogen. It seems that some mechanism is needed to explain the geographic separation of porphyry copper- molybdenum and simultaneous tin-tungsten- molybdenum metallization within the same tectonic system and appearing at similar distances from the trench. The thickened crust of the tin belt is one of the few known differences. Tables 1 and 2 list some of the most important porphyry copper deposits explored to date, but do not include Antaminas, Peru. or a number of other large skarn deposits

Jan 1, 1978

By K. D. Jacob, W. L. Hill

Fluorine recovery in the United States has been restricted chiefly to manufacture of ordinary superphosphate and wet-process phosphoric acid. However, there is an expanding use of fluorine by industry in the form of hydrofluoric acid. Improved methods for recovering fluorine from phosphate rock would yield an adequate supply of fluosilicates and a substantial contribution to the hydrofluoric acid requirement. THE bulk of natural phosphates is comprised of calcium phosphates, which are usually apatites;' calcium aluminum phosphates such as pseudowavel-1ite; and aluminum phosphates, which occur in extensive deposits, notably on Grand Connetable Island." Nearly all deposits of current commercial value are apatite, often with admixed material of other classes. In the trade coarsely crystalline material, mainly fluorapatite of igneous origin, moves as apatite, whereas the fine-grained material of sedimentary origin is marketed as phosphate rock. The latter classification includes the phosphate shales in western North America. Although several genetic varieties of phosphate rock are recognized, most known reserves are of marine origin. Calcium phosphate is concentrated from Florida land pebble and Florida hard rock deposits and from Tennessee brown rock deposits. Phosphate Rock as a Fluorine Carrier: More than two decades ago Reynolds and Jacob surveyed the fluorine content of commercial phosphate rock from various parts of the world. Later these samples were re-analyzed by a new and more reliable method of determining flourine. The revised results with interim additions of new samples, published a few years ago in a brief summary," are shown in Figs. 1-3. Through the points representing the respective varieties of rock medial straight lines are drawn. In some instances the lines are well defined by the points; in others, where scattering renders the position of the best line uncertain, the greater weight is given points near the end of the covered range. Wherever distinct varieties are not recognized among regional deposits the custom is to identify the sample with the geographical location of the deposit. Although all rock varieties cannot be classified with certainty, the plotted results exhibit three distinct types of rock in which the F to P2O5 ratio diminishes, remains sensibly constant, and increases, respectively, as the grade of rock increases. Existence of two types of Florida land pebble, Fig. 1, is supported by additional results recently supplied by a producer." Generally speaking, phosphate rock from continental deposits carries considerably more fluorine, and apatite and insular phosphates less, in proportion to the phosphorus content than is required by the formula for fluorapatite. Exceptions to this are apatite from Canada and certain samples of Florida waste-pond phosphate. The manner in which excess fluorine is held and the nature of the carbonate in carbonate apatite are discussed in recent articles. High-grade calcium phosphates with F to P2O5 ratios less than 0.05 have been found in Curacao, Christmas Island," Mexic,13 and elsewhere. The calcium aluminum phosphates and aluminum phosphates aften carry notable amounts of fluorine. The F to P2O, ratio for one sample of pseudowavel-lite concentrate was 0.05." Aluminum phosphate from Grand Connetable Island is very low in fluorine, whereas some aluminum phosphate minerals" carry as much fluorine as highly fluorinated phosphate rock. Fluorine in Domestic Phosphate: The curves shown in the figures permit interpolation of the F to P2O5 ratio corresponding to the midpoint of selected ranges in grade of any one of the several varieties of phosphate. Multiplication of the interpolated ratio by the mid-range grade (P2O5) yields a figure for the percentage of fluorine in this grade. Then, with the use of the mid-range grade, the amount of fluorine held in phosphate reserves can be estimated, see Table I. Accordingly, on Jan. 1, 1950, the fluorine complement of known domestic reserves exceeded 420,000,000 long tons of the element. A previous figure," deduced in 1940 by a simpler procedure

Jan 1, 1955

By Waldemar Lindgren, Clyde P. Ross

To the miner, as well as to the geologist, the eastern part of Cuba is a most interesting region. Here we find, in contrast to the moderate relief predominating elsewhere in the island, an imposing mountain range, the Sierra Maestra, extending east and west parallel to the coast, its precipitous front facing the blue Caribbean Sea. In a geological sense this range is largely an unknown land, the only well-explored region being that in the vicinity of Santiago. The range also contains the most important mineral deposits of the island. They comprise, first, a series of iron deposits yielding a partly hematitized magnetite with low content of phosphorus; second, a remarkable copper-bearing vein at El Cobre; and third, various manganese deposits. Probably more deposits will be found, for the larger and western part of the range is as yet little explored and its slopes are covered by a thick tropical jungle. The following notes are based on a short visit by the senior author in January and February of 1914 to the mines at Daiquiri, El Cobre, and Mayari. For many courtesies and great assistance he is deeply obliged to Charles F. Rand, President of the Spanish-American Iron CO.; to George W. Pfeiffer, General Manager, and to the several members of his staff. Geological Features of the Sierra Maestra For our knowledge of the geology of the Sierra Maestra we are indebted to reconnaissance work by Fernandez de Castro, R. T. Hill, Mr. C. Hayes, T. W. Vaughan, and A. C. Spencer. The summary by the latter three geologists furnishes the best guide to the region.' The range is undoubtedly outlined by a great east-west dislocation. Its southern slope drops abruptly to the sea, while on the north side a much gentler declivity leads down to the rolling plateau of Tertiary limestone which occupies much of the adjacent part of the island. Viewed

Jan 1, 1916