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By Louis A. Panek

INTRODUCTION AND SUMMARY With respect to the geomechanics aspects, the primary technical objectives in mining by an undercut-cave method are to achieve a controlled, sustained caving of the mineral body and to remove the fragmented ore with a minimum of dilution from surrounding unmineralized rock, while maintaining stability or control of the rock structure around the access openings. The structurally ideal ore body probably does not exist. If the rock mass is so weak as to cave on a short span, it may tend to be too sticky for easy drawing and handling, or create special problems in regard to support of the access openings. If the ore is strong, caving may be difficult and the caved fragments may be of such a large size as to require special equipment for transferring the ore from the caved zone. Given enough time and money to generate data and conduct trials, engineering and ingenuity can devise an appropriate combination of mine layout, sequence of extraction, and mechanical equipment to achieve a technically successful caving extraction operation to meet the foregoing requirements in many types of deposits. The large capital investment involved, however, reduces the freedom to make major changes once the mine development is well under- way, and the penalties for failure to accurately anticipate operating conditions militate against selecting any but the most obvious candidates for mining by an undercut-cave method. The demonstrated capability to extract large, deep, economically marginal deposits by this low-cost, high-volume method of mining provides an incentive to develop a rationale for predicting the cavability and stability characteristics of a deposit prior to mining, so that the undercut-cave method may be extended to a much wider range of mineral deposit characteristics. The ultimate goal is to establish as explicitly as possible the quantitative interrelationships between the measured rock-mass characteristics, the caving span, the size distribution of caved ore fragments, and the sizes and locations of stable access openings. Lacking an understanding of these relation- ships, a designer may readily change some factor in the wrong direction (e.g., excessively reduce the distance between the extraction level and the undercut to increase the convenience of operations for the undercutting crew, increasing the frequency of repairs to the extraction-level support system) or create unnecessary problems elsewhere in the system by introducing a design change that can achieve only minimal improvement in the factor of direct interest (e.g., unnecessarily complicate the ore-transfer system by changing the orientations of the openings, with- out succeeding in the objective of improving the ground support conditions). Although successful predesign is the prime objective, subsequent modifications in mine lay- out and sequence of extraction operations are inevitable. In developing the modified solution, systematic experimentation based on an understanding of the underlying structural relationships, coupled with monitoring measurements of selected diagnostic structural-behavior parameters, can achieve an acceptable solution in a minimum number of steps, which is far superior to the typical operational trial-and-error approach, in view of the cost of implementing each successive change. Since a drill-core sample of ore rock from a successful undercut-cave operation may exhibit a uniaxial crushing strength in excess of 100 MPa, the caving of such a rock mass is now commonly believed to be ascribable to the presence of discontinuities such as joints or fractures throughout the ore body. An essential part of the present investigation was therefore to characterize the natural discontinuities at each of the test sites by measuring their attitudes and spacings. The term "fracture" is used herein in a general sense to include any planar discontinuity without implication as to its suggested mode of origin. Most, but not all, of the fractures are properly termed joints. As a point of departure we may consider the possibility that the rock mass is transected by three families of joints, each family possessing a distinct orientation, such that parallelepipeds of intact rock are delineated by the jointing. Even if cementation is absent between adjoining parallelepipeds, the undercutting of the rock mass will not necessarily initiate sustained caving--owing to the all-around confinement, an arch may tend to stabilize over the undercut unless prevented from doing so by the failures of key blocks of intact rock. Thus, although the jointing can be assumed to weaken the rock mass, creating preferred directions of

Jan 1, 1981

By Martin Mayers

FIVE-SIXTHS of all the coal that is mined in the United States is burned, without previous treatment other than screening, for the produc-tion of heat and power, so that its value is fixed by its suitability for this purpose. Nevertheless, the characteristics of individual coals that determine their fitness for use in particular pieces of combustion equip-ment are not yet well known. This condition is due not only to the small extent of our knowledge concerning coal itself, but also to our compara-tive ignorance of the nature of the combustion process. The following paper is an analysis of the combustion process as it occurs in industrial furnaces, based largely on the results of laboratory research on combus-tion rates. From this analysis it is possible to determine the characteris-tics of the fuel that limit the capacity of a given piece of combustion equipment. Limitations imposed by other factors than the attainable speed of combustion will not be considered, since they are subject to a different type of analysis involving the balancing of maintenance costs against fixed charges. In order to observe the relations among the various factors that affect the speed of combustion, representative particles of fuel will be followed through two types of furnaces. A pulverized-coal furnace will be con-sidered first, because the independence of the individual particles facili-tates the analysis of the effect of furnace conditions. The second example will be a grate-fired furnace in which ignition of the fuel takes place on the underfeed principle. This includes both underfeed and chain grate stokers,(8)? but does not include hand-fired furnaces and some overfeed stokers. The furnace will be considered to be large enough, the tempera-ture high enough, and enough air provided, to burn the fuel completely, without, for the moment, specifying these quantities more completely.

Jan 1, 1934

By Sidney Cornell

THE technical- advantages of adding oxygen to air and producer gas, or using it as a reactive agent, producing 400 B.t.u. gas instead of. the present 150 B.t.u., with higher flame temperatures and a reducing flame possible, with a smaller volume of gases passing through a furnace, will result in quicker melting, sharp working, and the production of a higher quality steel. Probably electric-furnace quality steel can be produced in an open-hearth furnace using oxygen. A modern, standard, 100-ton open-hearth furnace, with all equipment, costs from $350,000 to $500,000; there are about 1000 open-hearth furnaces in this country. When considering the use of oxygen in the open-hearth furnace, it is necessary to study: (1) The use of oxygen in the present furnaces, such that modification of existing furnaces and equipment cost the minimum, although the full advantages of oxygen are not thus obtainable. (2) The use of oxygen in a furnace designed for its use, producing quality and quantity steel at the lowest possible cost. This entails the junking of approximately one-half the value in existing equipment-writing off $250,000,000. The basis of such an estimate is the present cost of manufacturing open-hearth steel. Objection may be taken to some of the unit prices here given, but while some companies obtain their ores for much less, the prices quoted are -market price. They are used to show the maximum economy obtained by oxygen at that point.

Jan 11, 1924

By Jed H. Mosgrave

One of the most interesting of all the subjects required of persons studying the different facets of coal mining is coal mine gases. Some mine gases have been a real problem since the very beginning of underground mining and in the early days there was insufficient knowledge to really cope with them. The history of mining shows that until the last 50 years, work was done primarily by the hand methods which are primitive by today's standards. Miners had to contend with gases without the aid of enclosed lighting, permissible mine machinery, sophisticated testing devices, and high pressure ventilating fans. In early days the miners were not educated as to the different mine gases nor did they have the technical equipment to determine the actual composition of the deadly atmospheres they encountered. They simply applied a descriptive term to each gas or mixture of gases depending on its effect on a person or the surrounding atmosphere. The gases or mixtures of gases were then referred to as "damps" which was probably derived from the German word "dampf," meaning vapors or gases. The more common damps encountered were as follows: Blackdamp A condition of the atmosphere where a miner's light would be extinguished and in which breathing would be difficult, sometimes even resulting in suffocation. An atmosphere deficient in oxygen and which could contain a combination of any of the mine gases. Firedamp Flammable mixtures of methane and air which would either burn or explode when ignited. White Damp A mixture of gases containing carbon monoxide.

Jan 1, 1973

By H. DeWitt Smith, W. H. Sirdevan

The mine operated by the United Verde Copper Co. is situated near Jerome, Ariz., on the eastern flank of the Black Hills, at an elevation of approximately 5500 ft. (1676 m.) above sea level. The mine and the town of Jerome are located on a steep hillside, which slopes rapidly to the valley of the Verde River. The smelter town of Clarkdale is situated in the Verde valley 4.1 mi. (6.5 km.) distant in an air line from Jerome, and at an elevation of 3560 ft. The two towns are connected by the standard-gage railroad of the Verde Tunnel & Smelter R. R., a subsidiary of the United Verde Copper Co. This road is 11 mi. long and has a compensated 4 per cent. grade. At Clarkdale, it connects with the Verde Valley branch of the Santa Fe R. R. The narrow-gage (36-in.) line of the United Verde & Pacific Ry. was, until 1919, Jerome's only rail connection with a main trunk line. This road connects with a branch line of the Santa Fe system at Jerome Junction, after traversing 26 mi. of mountainous country. From the time of its location, in 1876, until its purchase by Senator W. A. Clark, of Montana, in 1889, the United Verde mine was worked on a small scale and shipments of high-grade gold-silver ore were made from the surface workings. Until 1894, when the narrow-gage railroad to Jerome Junction was completed, all supplies for the mine and smelter

Jan 1, 1922

By Prof John E. Church

It is a well-known fact that under similar conditions a ton of pig iron can be made from any ore with less fuel when charcoal is used than when coke or anthracite is employed for heating. The cause of this superiority is entirely undiscovered, and no explanation has yet been offered which has won even temporary acceptance. Such as they are the hypotheses heretofore presented are of two general kinds, and they come from those two divisions of the working force which are so often antagonistic in all professions, the scientific and the practical men. The theory of the scientific men is drawn from the chemical reactions which are known to go on in the furnace, and it is well expressed in a recent paper by Prof. Akerman, of the Stockholm (Sweden) School of Mines. It makes the superiority of charcoal as a fuel in the blast furnace to depend upon its quick and thorough reduction of carbonic acid. This property belongs to charcoal in virtue of its porous structure. Its cells have a diameter of only one twenty-four hundredth of an inch, and one grain of charcoal has been computed to have a surface of two and four-tenths square feet. According to the theory in question, carbonic acid is formed by the first impact of the air upon the fuel; but in charcoal furnaces the acid is immediately reduced to oxide by taking the necessary amount of carbon into combination, and the gas reaches its highest reducing power even in the near vicinity of the tuyeres. Coke, being harder, does not produce so strong a reducing gas in that part of the furnace. Carbonic acid is formed as in the case of charcoal, and remains longer before its reduction to oxide is completed. Anthracite is the

Jan 1, 1879

By John A. Church

IT is a well-known fact that under similar conditions a ton of pig iron can be made from any ore with less fuel when charcoal is used than when coke or anthracite is employed for heating. The cause of this superiority is entirely undiscovered, and no explanation has yet been offered which has won even temporary acceptance. Such as they are the hypotheses heretofore presented are of two general kinds, and they come from those two divisions of the working force which are so often antagonistic in all professions, the scientific and the practical men. The theory of the scientific men is drawn from the chemical reactions which are known to go on in the furnace, and it is well expressed in a recent paper by Prof. Akerman, of the Stockholm (Sweden) School of Mines. It makes the superiority of charcoal as a fuel in the blast furnace to depend upon its quick and thorough reduction of carbonic acid. This property belongs to charcoal in virtue of its porous structure. Its cells have a diameter of only one twenty-four hundredth of an inch, and one grain of charcoal has been computed to have a surface of two and four-tenths square feet. According to the theory in question, carbonic acid is formed by the first impact of the air upon the fuel; but in charcoal furnaces the acid is immediately reduced to oxide by taking the necessary amount of carbon into combination, and the gas reaches its highest reducing power even in the near vicinity of the tuyeres. Coke, being harder, does not produce so strong a reducing gas in that part of. the furnace. Carbonic acid is formed as in the case of charcoal, and remains longer before its reduction to oxide is completed. Anthracite is the

Jan 1, 1879

By T. L. Joseph

BLAST furnaces are most efficient thermally when the C02 in the top gas is highest. Oxygen introduced in the air blast is converted to CO in the combustion zones. The extent to which CO, generated in front of the tuyères, is converted later to C02 depends upon the ratio of ore to coke in the charge and the degree to which C02, formed by reduction of iron oxides with CO, reverts to CO by the secondary reaction C + C02 = 2C0. About 300 to 400 lb. of carbon1 per ton of pig iron is normally involved in the solution-loss reaction or direct reduction of FeO with carbon, which produces the same over-all results; that is, lower C02 in the top gas. Riche has recently reported a reduction in fuel consumption at slower rates of blowing on two furnaces in spite of higher heat losses per ton of iron. Lower fuel consumption at slow operating rates was due to more complete reduction of the ore that resulted in higher C02 in the top gas and consequently more efficient use of carbon. Kinney3 has reported substantial savings in fuel as a result of crushing and sizing the ore charge. Improvements in practice brought about by sintering magnetite concentrates, flue dust and fine ores have been due to the establishment of conditions favoring more complete reduction of the ore in the upper part of the furnace. The extent to which oxides and fluctuations in composition in pig iron affect the quality of steel and finished product needs further study. There is evidence that more thorough preheating and reduction of the ore in the upper part of the furnace tend to give a better quality of iron. Bogatzkii4 found from laboratory work that the amount of oxygen enter-

Jan 1, 1936

By C. R. Sandberg, C. A. Connally, N. Stein

This paper describes the results of volumetric efficiency tests on oil well pumps handling gas oil mixtures. The work was performed in a large scale, above ground unit wherein test conditions could be accurately controlled and measured. The main variables studied were gas/oil ratio (including gas from solution and free gas mixed with oil), pump compression ratio, pump stroke length, pump speed, and clearance volume between the valves at their closest approach. Results are presented for two different pumps and for oils of two viscosities. Relatively small amounts of gas entering the pump resulted in large decreases in volumetric efficiency. Under conditions where the pump was operating at reduced efficiency because of the presence of gas, it was found that variation in the clearance volume between the standing and traveling valves had a considerable effect on pump efficiency level. This effect of the valve clearance volume was found to be significantly altered by the viscosity of the oil used in the tests. The effects on pump efficiency of the other variables studied were found to be relatively small over the range of conditions utilized. INTRODUCTION The production of oil by pumping is often hampered by low volumetric efficiency. A direct increase in lifting costs results from low volumetric efficiency. An indirect increase in lifting costs, probably greater than the direct increase, results from additional wear and tear on pumping equipment and from the down-time necessary for the repairs which can be traced to low-efficiency operation. Both increases in lifting costs tend to reduce economically recoverable oil. A number of different factors can contribute to low pump efficiency. A known basic cause of low efficiency is the presence of free gas in the pumped fluid. Pump volumetric efficiency is calculated only on the basis of liquid pumped and because any free gas pumped is discounted, this volume of free gas would represent a loss of pump efficiency. However, gas also causes a reduction in pump efficiency because it is a highly compressible fluid. It is known that pumps some- times "gas lock" because of excessive gas-to-liquid ratios in the pump barrel. Little is known of the role of gas compressibility in the intermediate case where the pump is operating at low efficiency. The opinion exists, however, that oil-well pumps tend to operate at higher efficiency with long stroke lengths at low speeds, but no quantitative studies of these pumping variables have been reported. It was believed that a much better understanding of the variables which control pump volumetric efficiency could be obtained and that possibly some suggestions as to the methods for increasing efficiency might be found from a study of the operation of pumps handling gas under closely controlled conditions. Previous investigators have studied the effects on pump efficiency of such factors as oil viscosity, oil temperature, slippage of oil. past pump plungers, pump submergence, valve size and spacing, pressure above pump plunger and fluid vapor pressure. However, none of these published investigations were conducted with pumps being subjected to large amounts of gas such as might be the case in a pumping well, nor did any of the investigations study the effect of variation in stroke length or pump speed. A large-scale teat unit was therefore constructed for studying the operation of pumps handling gas and for evaluating effects of such variables as pump stroke length and pump speed. PROCEDURE AND EQUIPMENT A schematic diagram of the pump testing equipment is given in Fig. 1. A 45-ft length of 6-in. casing is mounted vertically in a 65-ft tower. Sight ports are mounted in the casing at intervals near the location of the pump intake and the liquid level in the casing. These sight ports are fitted with Lucite windows sealed by neoprene "0" rings. The Lucite windows are machined to conform to the I.D. of the casing so that no obstruction to flow is present along the casing wall. The casing is fitted with a tubing head and 2-in. tubing is hung inside the casing. Pumps are seated in a shoe attached to the 2-in. tubing. A 1-in. polish rod is attacked directly to the pump without any intervening sucker rods. The top of the polish rod is attached to the weight carrier, which contains a number of weights to be used to force the polish rod in against tubing pressure on the down-stroke. This is necessary because a long string of sucker

Jan 1, 1953

By S. Luszcz, R. W. Vook, Fred Witt

In situ X-ray investigations were made on polycrys-talline silver films deposited by vacuum evaporation on (111) CaF2 and (100) NaCl single-crystal substrates at 80°K. The films were evaporated and annealed in an X-ray diffractometer attachment having a residual gas pressure of 2 x lo-' Torr. All measurements were made without exposing the films to the atmosphere. Measurements were made on the films in the as-deposited state and after various annealing treatments. The intrinsic stacking and twin fault densities, the magnitudes of the uniform and nonuniform strains, and the crystallite sizes were determined. In addition the textures in the films were measured qualitatively. The results obtained for the as-deposited films on single-crystal substrates are in substantial agreement with previously reported results for silver films deposited on glass. Intrinsic stacking and twin faults, as well as uniform and nonuniform strains, were present in these films. During the various annealing treatments (up to 350°C) the faults and nonuniform strains annealed out. Considerable grain growth and texture changes occurred also. The effects were much greater for the NaCl substrate than for the CaFz substrate. The relative magnitudes of the grain growth in the variously oriented grains could be explained qualitatively in terms of the thermal strains and strain energies introduced into the differently oriented grains during the initial, irreversible anneal. These strains were due to the different thermal expansion coefficients of the film and substrate. X-RAY diffraction measurements on evaporated films deposited on substrates at low temperature have the advantage that many of the imperfections introduced into the film during deposition are "frozen in". Thus, the influence of a very important experimental variable, substrate temperature, on the imperfection structure of evaporated metal films may be studied. Moreover, the effects of annealing such films makes possible the study of thermally activated recovery processes in these films. The present study was designed to determine the influence of single-crystal substrates on the resultant film structure relative to the previous results obtained using glass substrates.' To this end great care was taken to keep the experimental variables the same in the two cases. Different experimental conditions would, of course, result in films having different physical properties. Again the initial substrate temperature was in the neighborhood of 80°K and the films were subsequently annealed to 350°C. The pure metal silver was chosen for evaporation, primarily because of its relatively low stacking fault energy and consequent high fault density in the as-deposited state. The silver films were formed by evaporation onto air-cleaved {ill} CaF, and (100) NaCl surfaces cooled to 80°K in an X-ray diffractometer attachment2 having a base residual gas pressure of 2 X l0-' Torr. The films were not exposed to the atmosphere until all of the X-ray data had been recorded. In this way one of the most important experimental variables, environment, could be well-controlled and reproduced. X-ray measurements were made at the temperature of deposition and included determinations of the diffraction line peak positions, line shapes, and integrated intensities. The peak position measurements were used to determine the intrinsic stacking fault densities and the average uniform strain in the film. The shapes of the diffraction lines provided information on the twin fault density, true crystallite size, and average nonuniform strain. The preferred orientation in the film was determined qualitatively from the integrated intensities. I) EXPERIMENTAL PROCEDURE The evaporator attachmentZ was charged with 99.999 pct Ag pellets positioned in a tantalum filament which had been outgassed previously at l0-8 Torr. The CaFz and NaCl single crystals were cleaved in air and then placed in position in the chamber so that their cleavage surfaces were on the diffractometer axis. The chamber was prepumped using a sorption pump, sealed off, and then baked at 150°C for 24 hr. The ion pump operated during the bakeout cycle. The substrate was then heated to 500° C by means of an auxiliary heater and kept hot until the rest of the chamber was cooled slowly to room temperature. This bakeout procedure consistently resulted in an ultimate pressure in the low lo-' Torr range. The substrate was then cooled down on 80°K. Its temperature was monitored by a thermocouple wedged into the rear of the copper substrate holder. The diffracted intensity and peak position of the 111, 222, and 333 CaF, lines were measured prior to evaporation. Nickel-filtered, pulse-height-discriminated copper radiation was used. Similar measurements were made for the 200 and 400 lines from NaC1. These measurements were used as a lattice parameter check and to determine the thickness of the evaporated silver films from the attenuation of the substrate lines. The evaporation rates were approximately 3A per sec for both films while the maximum pressures during evaporation were 3 x lo- ' and 7 x 10"8 Torr for the CaF, and NaCl cases, respectively. The film thickness was measured by the attenuation of the CaF, and NaCl substrate lines and by at optical interference method. Values of 1700 and 1500A, respectively, were obtained for the silver

Jan 1, 1969

By John P. Cosgro

The increasing use of internal furnace-boilers for power-plants at mines (doubtless due to the facility with which they may be installed by reason of their portability; the fact that they require no masonry setting, cast-iron fronts, buck-stays, etc., and practically no foundation; and the successful maintenance of their claim to equal and even superior economy, as compared with boilers of other types1) warrants this description of a method of remedying an injury which may occur, to such a furnace, even though it satisfies the requirements of the formula adopted by the U. S. Board of Supervising Inspectors of Steam Vessels.2 I refer to the sagging and deformation which may result from overloading the boilers for a considerable period—say, several days. This is sometimes unavoidable under the exacting conditions of continuous day-and-night service, and (in spite of all rules and precautions) the occasional careless or unskillful work of attendants. Such a trouble is by no means unknown on shipboard, where this type of boiler is so largely employed.

Jan 1, 1906

By F. S., Mc Nicholas

A practical discussion of the theory of A block caving is presented which applies particularly to the physical conditions of the Climax orebody although the conditions are sufficiently characteristic of caving ore-bodies to provide valuable information applicable elsewhere.

Jan 1, 1950

By William G. Kegel

Aside from having a usable product and good mining conditions, the greatest asset for a profitable coal mining organization is an effective mine maintenance program. The first step in acquiring this is the presence of good mechanics-men who have a desire to be mechanics and who want to know more about the design and control of the machines in their charge. The knowledge and ability to be found in a good mechanic is the result of ambition, curiosity, and determination. With these qualities as a start, the aspiring mechanic can launch his career if he begins learning all he can about mechanics, electricity, and hydraulics. The path is rough, but the goal is rewarding. Mining machinery probably absorbs as much or more abuse than any other heavy equipment in use today. And when one considers the conditions under which much of the repair work must be done, it can be seen that the mine operator needs the best of men and tools to perform the service. MOTORS AND POWER SUPPLY Mining machines, from their inception, have utilized the electric motor; and down through the years of mechanization, this power source has undergone many changes to better adapt it to the job. The first motors used were large, heavy-framed, open d-c motors (direct current), using a series or shunt winding. A motor having one or more field coils connected in series and themselves in series with the armature, is called a series motor (Fig. 1). The series motor had the advantage of quickly developing starting torque, and it was, and still is, the best motor for locomotive or shuttle car tramming service. The series motor, however, speeds up as the load is reduced and, if not properly controlled, can overspeed. The shunt motor, as its name implies, has one or more field coils in

Jan 1, 1973

By Richard Franchot

IN the inevitable conquest of the blast furnace by metallurgical science in the solution of the problem of how to make more and better iron or to burn less coke, or both, it is highly desirable first to find out why the efficiency of the furnace generally is no better than it is. The immediate need would seem to be a theory to account for the existing facts. Explanation for the quantitative results obtained in current practice has been strangely lacking. One reason for this may be found perhaps in the custom of think-ing of efficiency in terms of coke consumption. Before efficiency can be explained there should be a fairly exact knowledge of what it is. A statement of the pounds of coke used per ton of iron made may have value in indicating comparative results in smelting similar ores with similar cokes, but it tells very little of the efficiency of the operation. When the materials are different, one practice burning 1800 lb. may be less efficient than another using 2000 lb. A better measure is the pig and slag per unit of coke. What is needed is a specific description of efficiency as a' definite quantity having an exact meaning. Such description is afforded by expressing efficiencies in terms of the respective proportions of the fuel energy absorbed in and chargeable to the various actions known to be -taking place in the furnace. Concretely, the energy value of a pound of coke having 85 per cent carbon blown with air heated to 1200° F. is about 7650 lb.-cal. (13,770 B.t.u.). And the work done in smelting a pound of average iron from hematite, comprising reduction and melting with production of a half pound of free running slag from an equal weight of limestone,

Jan 9, 1926

By C. C. Miller, A. R. Belyea, C. H. Herty, E. B. Burkart

The removal of sulfur from steel has been studied by many investigators, but the quantitative relationships between the factors involved have not been determined. This is undoubtedly due to the number of variables encountered in practice, an attempt to study any one reaction or relation being affected by other reactions the relative importance of which cannot be estimated. The factors that affect the extent of removal of sulfur in the basic open-hearth process are generally considered to be: 1. Basicity of the slag: (a) Iron oxides in the slag; (b) carbon in the metal or carbide in the slag; (c) temperature. These factors affect the reactions: FeS + CaO = CaS + FeO FeS + CaO + C = CaS + CO + Fe. 2. Amount of sulfur in the gas. 3. Amount of manganese in the metal. The factors which affect the rate of removal of sulfur include, in addition to these: 4. Fluidity of the slag: (a) Normal changes in fluidity with temperature; (b) changes in fluidity because of the addition of special reagents such as fluorspar. 5. Agitation of the bath. Within the past two years, two papers have been presented dealing with the influence of sulfur in the gas on slag and metal. Nead' has shown that if a gas high in sulfur be used for heating, the metal and slag will absorb sulfur. Whitely2 concluded from his experiments that if a

Jan 1, 1925

By R. E. Zinkham, R. J. Goodwin

A mathematical study has been made of the amount of support a cement sheath could provide to casing cemented into the earth. Several assumptions were required to make the analysis, but only two of them are limiting: (I) the pipe must be completely surrounded with cement, and (2) any mud filter cake between the cement and formation has the same physical properties as either the cement or formation. The calculations showed that little support would be provided to the pipe before an unsupported cement sheath failed in tension; however, when the cement is confined between the pipe and wellbore and is loaded in compression, the pipe could receive a considerable amount of support. In fact, the theoretical results indicate the lower grades and larger sizes of pipe could have their working pressures doubled when reasonable compressive loads were applied to a surrounding cement sheath. These data are shown in six charts. Other down-hole conditions such as setting the cement under pressure, increased temperature and cement confinement all tend to increase the potential usefulness of the sheath. Because of size limitations, a laboratory program to verify the most important results of this mathematical study would be very difficult. However, small-scale field tests would be practicable. This paper shows that, if a solid cement sheath can be obtained in the field by either primary cementing or by repair after detection of flaws by surveys such as the new cement-bond logs, the use of this approach to reducing pipe costs merits further consideration. INTRODUCTION A modification in casing design practices is proposed which may either reduce the amount and grade of steel required to contain a specified internal pressure or permit the working pressure to be increased for a specified weight and grade of pipe. One of the more important considerations in casing design is its resistance to collapse; however, Bowers' and, more recently, O'Brien and Goins' have shown many casing programs are unnecessarily conservative in this respect, and they have indicated how savings can be made by designing for more realistic down-hole conditions. Earlier, Saye and Richardson howed that pipe costs could be reduced by considering the cement sheath as a part of the casing string when collapse resistance was being calculated. More recently, Rogers4 has raised the question as to whether a cement sheath might be considered in the design for burst resistance of the cemented casing. Calculations have been made for the increased burst resistance a cement sheath would provide for casing in a wellbore, and the results show that a sizable amount of support could be obtained in some instances. These data are presented in addition to a discussion of several other factors that are considered to affect the burst strength of pipe supported by cement. Two types of support are treated: Case I for tensile loading of the unconfined cement sheath, and Case for compressive loading of the confined cement sheath. ANALYTICAL TREATMENT AND RESULTS CASE I—TENSILE STRESSES IN AN UNCONFINED CEMENT SHEATH Conditions like this would most likely occur in a greatly enlarged portion of the hole where the cement was not in immediate contact with either the formation or a thin and hard mud cake. The mathematical analysis for this condition, as shown in the Appendix, rests on the following concepts. Pressure inside a unit length of pipe causes: (1) a tensile or tangential stress to be exerted over the longitudinal cross-sectional areas of the pipe and cement; and (2) an equal amount of strain in both the pipe and cement that is uniformly distributed over the wall thickness of each. This analysis was then used to make several calculations for a cement sheath around 51/2-in. OD pipe. The results are illustrated in Fig. 1, which shows that a tensile stress of 500 psi is imposed on a 5-in. thick sheath when the casing contains a pressure of only 1,450 psi. It also shows that a 10-in. thick sheath would be stressed to 500 psi in tension when the pipe contained a pressure of only 2,350 psi. Alternatively, if the stress analysis is made by means of the Lame thick-wall cylinder theory, the inner fibers of the 10-in. thick sheath will be stressed to 500 psi in tension when the pressure in the pipe is only 990 psi. This, of course, reveals that an unconfined sheath is of little support to the pipe in burst; however, an entirely different result is obtained when the cement is confined between the pipe and formation.

By David R. Maneval, W. E. Foreman, J. Richard Lucas

INTRODUCTION The objective of this chapter is to inform the industry, as well as the public, of the challenges in dealing with problems associated with air and water contamination by coal preparation plants. The need for an informed industry is most important. It is hoped that some contribution may be given to a more efficient and economical approach to the problems of plant waste contaminants at individual plants. The problem has many facets, and consideration should be given to specific areas before any significant problems develop at a given plant. It is then possible to have a reasoned approach before the pressures of an emergency environment force hasty and incomplete solutions. It will be necessary to anticipate these problems in the design of new preparation plants, and advance consideration should be given to all problems concerned with contaminants.

Jan 1, 1979

By Ray W. Arms

DRY cleaning, or pneumatic separation, is not, strictly speaking, a recent discovery. Among the archives of the Patent Office may be found many patents dating back as far as 1850 which cover early attempts to separate materials of varying specific gravity or of different shape by means of air. Hundreds of patents covering this art have been issued, which may be roughly classified into four general groups, as follows: 1. Stationary devices with pulsating air currents. The separating surface is usually riffled and air is supplied by bellows or compressors. This group also includes air jigs, which have been used rather extensively. 2. Stationary devices with continuous air currents. These submitted the material to a continuous current of air, either horizontal or vertical. Chaff is blown from wheat by such a device. 3. Reciprocating or vibrating devices with pulsating air. A small group in which the pulsating air is supplied by bellows and some motion provided in the separating surface to move the stratified material to various discharge points. 4. Reciprocating or vibrating devices with continuous air supply. This is by far the most important group and all recent developments have been along this line. All of the above groups involve the stratification of material by air and include none of the dry methods that use other principles, such as coefficient of friction, magnetism, etc. Air Stratification In all of the above groups of air separating devices, the same phenomena are found as exist in water concentrators; that is, free settling will be found in some and hindered settling in others, with a complex action on the reciprocating tables similar to the wet tables. The experimental value of the settling ratios of particles in air has never been determined but, following the formulas given in Richards'

Jan 1, 1924

By William B. Hughes

The rubber-sleeve core barrel was developed to improve core recovery from unconsolidated sands, where it is most difficult to obtain cores with conventional barrels. The use of a rubber-sleeve core retainer, together with other departures from clsual design, has required a considerable amount of testing to prove its op-erability. Tests run in shallow experimental holes and in drilling wells in the U. S. and Vemzuela have, to a large extent, proved the operation of the barrel. Internal mechanism troubles, such as sandbinding of the sleeve, have been largely overcome. Cores have been recovered from completely unconsolidated sands where conventional coring has almost been abandoned because of poor recovery. In addition to retaining the core until brought to the surface, the barrel gives a "packaged-ar-cut" core, which is coilverzient for handling and transportatiorz. In the initial work, drag blade cutters, roller cutters, and tungsten carbide insert cutters were usrd for cutting unconsolidated sands. Further bit head development is presently being conducted using diamond heads in unconsolidated sands where shaler, hard sands, and limestones are encountered in adjacent formations. Thir work should render the barrel capable of cratting and recovering most types of formations, possibly including broken and frac- tured fornlations. This development is being carried on in cooperation with a diamond bit manufacturer. INTRODUCTIO N For many years the oil industry has needed greater recovery of cores from unconsolidated formations. Conventional equipment has been used to core unconsolidated sands with little success. Specifically, a core barrel was needed which would recover unconsolidated sands in a condition that would give useful reservoir information. DEVELOPMENT OF 'THE RUBBER-SLEEVE CORE BARREL Basic Requirements of Unconsolidated Sand Corino The basic need was defined as a core barrel that would obtain a continuous unconsolidated core with sand grains in place as deposited in the formation. A preliminary survey of coring equipment and search of the literature pointed out the weaknesses of conventional core barrels. Laboratory tests using unconsolidated sand in a metal tube simulated actual sand movement into the core barrel during conventional coring. In conventional barrels the core sometimes crumbles and bridges, or fails under compressive column action, becomes oversize, and wedges against the walls of the inner barrel. The core barrel requirements established by the surveys and preliminary tests are as follows. 1. The core should be held to- gether as it is cut so it cannot crumble and bridge against the walls of the coring tool, care being taken not to permit damage from fluid circulation. 2. The core should be continuously supported as it is cut so it will not have to support itself in compression. 3. The core should be packaged in the barrel to retain an undisturbed condition during the trip out of the well. The packaging should be such that the core can be removed from the barrel and shipped to the core analysis laboratory without physical disturbance or alteration of fluid content by exposure to the atmosphere. Operation of the Rubber-Sleeve Core Barrel The rubber-sleeve core barrel (Fig. 1) meets the basic requirements for the recovery of unconsolidated sands.

By K. Prins

ONE of the more pressing problems faced by the coal industry today is the development of adequate means for meeting conservation laws, particularly those involving stream pollution, in various parts of the country. Discharge of dirty coal washery water into streams and rivers is almost universally frowned upon. Many states have enacted laws carrying heavy fines to curb the practice. The Prins stream-cleaner is one of the latest machines to enter the market. It is closely related to the cyclone thickener in principle. Eleven stream-cleaners are currently operating, ranging in size from 4 to 16 in. diam. In more recent installations the water enters directly in line and on a tangent with the impeller. The impeller consists of a vertical shaft up through a packing gland and bearings, and a V-belt pulley. The lower part is a tubing fastened to the shaft above, extending through the water intake compartment and provided with six vertical flat bars welded to the tubing. Portholes are situated in the upper end of the tubing, immediately below the point where shaft and tubing join together. The portholes are placed so that they are in open communication with the upper compartment of the stream-cleaner from which the processed water is discharged. The impeller is motor driven with a wide range vari-pitch drive employed between motor and impeller. The motor is mounted vertically, and the mounting provided with a vertical hinge allowing for needed adjustment of the wide range vari-pitch drive. The dirty coal washery water entering the machine under 20 lb psi pressure, flows from the compartment above the impeller, between the impeller blades, and is whirled around in the vertical section of the impeller enclosure. The speed of the impeller supplies centrifugal force and velocity required for separating suspended solids from water. The lower part of the machine consists of a cone, whose action is similar to other machines of the same type. The underflow discharge orifice is a cold rolled steel block machined to correspond with the cone angle and allows insertion of steel tubes of different diam. On 16 in. machines a 1 ¾ in. ID vertical discharge pipe is used. Provision is made for attaching a curved section of the same diam to the vertical pipe, to which, in turn, different lengths of horizontal pipe can be connected. Curved Pipe Advantageous It has been found that a curved pipe offers resistance to discharged material flow. In addition, the rotary motion of the underflow can be easily arrested in a curved pipe. Impeller speed of the 16 in. diam machine is provided from 400 to 800 rpm. A speed of about 474 rpm is suitable for maintenance of a constant underflow in coal slurry. In one installation 5x ¼ in. coal is cleaned in a Jeffrey Baum type washer at a 225 tons per hr rate. Washer installation is of the conventional type and a drag type sludge tank is used for water clarification purposes. Capacity of the water system, including the Baum washer, is about 40,000 gal. Before placing the stream-cleaner in operation, it was necessary to flush out the entire system every five days of two shift operations. The only time the system is drained now is for repair work on the sludge conveyor or the rig. The suction line of the stream-cleaner pump terminates in a number of small branch lines located at a depth of about 4 ft above the sludge conveyor. Each branch line extends the full width of the tank and is provided with four intake ports, each one with a funnel shaped inlet projecting downward. The arrangement provides an extensive pick-up area, for dirty water, and the inlets are arranged for a low rim velocity, preventing the taking in of coarse particles. The funnels are also arranged to extend up or down in the tank. They are set to pick up -60 mesh material exclusively.' The material is a high ash and high sulphur product and thus has to be disposed in the refuse conveyor. The underflow of the stream-cleaner is discharged on top of the washery refuse which is carried in a drag type horizontal conveyor, discharging into another refuse conveyor inclined at 30º with a short horizontal loading section. Some Disadvantages The impeller inherits certain disadvantages because of the nature of its construction. Additional moving parts make it subject to wear and maintenance costs. The advantage of being able to maintain a constant speed, however, to produce desired water velocity in the machine outweighs the drawbacks. Better separation between water and solids can be obtained by regulating time of residence of water through adjustment of valves in the intake and discharge lines. The amount of fines encountered during plant operation will vary because of higher or lower moisture in coal passing over fine coal vibrating screens. Even the amount of fines picked up by underground loading machines will be inconstant. Consequently, the percentage of solids will vary in water to be processed. The velocity in the feedlines to the slurry thickeners will fluctuate, with the required water velocity lacking. Another advantage advanced for the machine is its ability to operate on 15 to 25 lb line pressures at the water intake, reducing pump power required and pump maintenance.

Jan 1, 1952