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By C. Chu

An investigation has been made of the radial heat-wave process using a mathematical model in two-dimensional cylindrical coordinates. This model considers combustion, convection and conduction inside the reservoir, but only conduction in the bounding formations. From a study of the general features of the process, an important phenomenon has been revealed, namely, the feedback of heat into the reservoir on the trailing edge of the heat wave. The effects of various process variables on the performance characteristics of the process have also been investigated. It was found that up to the time when the combustion front reaches a given point, the per cent heat loss, provided it is not higher than 40 per cent, is approximately directly proportional to the square root of thermal conductivity arid fuel content, but inversely proportional to the square root of gas-injection rate and oxygen concentration. The effecr of reservoir thickness is more pronounced, since halving the thickness doubles the per cent heat loss. The most decisive factor in determining the center-plane peak temperature is the fuel content of the reservoir. Within the temperature range investigated, doubling the fuel content doubles the peak temperature in the early stage, but the rate of decline of the peak temperature is high. Reservoir thickness is also a very influential factor. The peak temperature is lowered when the thickness is reduced; however, the effect of thickness becomes less pronounced when the thickness is high. Reduction of oxygen concentration increases the peak temperature in the early stage but lowers it afterwards because of the higher rate of decline of the peak temperature. Increase in gas injection rate or decredse in thermal conductivity geives a higher peak temperature which stays high for a longer period. The propagation range of the heat wave is chiefly governed by the fuel content of the reservoir. An increase of 0.2 1b/cu ft in the fuel content increases the propagation range by 100 per cent. The propagation range is more than doubled by doubling the gas injection rate, or reservoir thickness, or by reducing the thermal conductivity by 50 per cent. Comparatively, oxygen concentration has less effect on the propagation range. INTRODUCTION Several investigators have conducted theoretical studies of a radial heat wave. Vogel and Krueger1 studied a system with a moving cylindrical heat source of constant temper- ature, considering conduction in the radial direction only. Ramey2 included conduction in the vertical direction in his studies. Bailey and Larkin2 attacked a more general problem where initial well heating, vertical heat losses and arbitrary frontal velocities were included. In all these studies, however, conduction was considered to be the only means of heat transfer. Bailey and Larkin in a later paper included the effects of convection in a study involving both linear and radial geometries. Vertical heat losses were neglected in the radial case. Katz5 studied a similar problem in a one-dimensional radial model, using a heat-loss coefficient to account for vertical heat losses. Selig and Couch6 mployed a cylindrical model and investigated two limiting cases. In one case they considered no heat loss from the reservoir whereas in the other they assumed a constant temperature at the interface between the reservoir and its bounding formations. Thomas' studied a more general case but assumed a permeable bounding formation so that the convection effect is not confined to the reservoir. In the present work a more realistic and more generalized model is used. It involves a two-dimensional cylindrical system with combustion, convection and conduction inside the reservoir, but only conduction in the bounding formations. The purpose is to establish the temperature distribution both inside and outside the reservoir, to study the general features of the radial heat wave process, and to investigate the effects of various process variables on the performance characteristics of the process. THEORY We fist consider a circular porous reservoir of thickness H extending vertically from y = — H/2 to y = + H/2. The reservoir extends from a well bore radius r, to an external radius re. A stream of oxygen-containing gas is introduced into the reservoir through the wellbore. The oxygen-containing gas reacts with the fuel contained in the reservoir and forms a combustion front wherever the prevailing temperature can support the combustion. It is here assumed that this combustion front constitutes a cylindrical surface source of heat having an infinitesimal thickness in the radial direction and extending vertically throughout the whole thickness H. This is designated as Region I. We next consider a Region II corresponding to the upper and lower formations bounding the reservoir, extending from y = — - to y = — H/2 and from y = + H/2 to y = + m. Since r, is very small, we may assume that the two bounding formations have the same dimensions, symmetric with respect to the center plane of the reservoir. In this way, we may take the upper half of the system alone into consideration. In contrast with

By R. L. Manegold

This paper outlines the reasons for and the application of suspended magnets above belt conveyors for the purpose of removing tramp iron to protect equipment. The design of a magnet strong enough to extract iron through a distance of 24 in. is described along with the proper method of application in the field. Advantages of the suspended magnet in contrast to the conventional magnetic pulley are also discussed. IN RECENT years there has been a decided trend toward bringing ores and coal out of open-pit and underground mines by long, sloping single-stage belt conveyors. Because the high investment cost of haulage conveyors dictates maximum speed and full loading for economy, conveyor belts are running faster than ever before, with a greater burden on the belts. In addition, larger blocks of ore are being mined and conveyed. Rougher use of large mining equipment generates more tramp iron; mechanical loading rejects nothing. Therefore the extraction of tramp iron and other deleterious materials for the protection of haulage conveyors at transfer points, chutes, and idlers is becoming an increasingly difficult job. In our experience, the best method of protecting long haulage conveyor belts is to extract the tramp iron, at least the large tramp iron, ahead of the main haulage conveyor or series of conveyors. It is most practical to collect the tramp iron at the earliest source possible, usually directly after the primary car unloading dump or after the primary crushers, which are ordinarily not affected by tramp iron, see Fig. 1. The coal or ore from its primary source is fed by chutes or feeders onto a buffer or shock belt conveyor, a short, ruggedly built conveyor system that can be considered somewhat expendable. Its primary function is to bring the ore up to speed for transmission to the haulage conveyor. There is no assured method of protecting this conveyor from tramp iron damage at its feed end. Likewise, there is a possibility, although remote if the magnet applied to this belt is installed correctly, that the tramp iron extraction process will damage the shock belt, which is, however, short and comparatively inexpensive. Magnetic head pulleys with their maximum effective range of about 8 or 10 in. cannot be considered for large installations, even though they are desirable because of the continuous and automatic tramp iron removal feature. Further, the most dangerous form of tramp iron with respect to potential damage to a haulage conveyor is the long rod, roof bolt, or rail segment, which is not extracted satisfactorily because of the limited tangential contact of such an iron shape with the magnet. Magnetic head pulleys are also undesirable because the dribble or fines from the head pulley discharge at the same point as the tramp iron. Screens must be used to separate the fines from the magnetics, resulting in additional expense and loss of head room. Magnetic detectors have been used with success, but they only indicate the presence of iron, and the conveyor belt must be stopped for its manual removal. The stoppage of any conveyor means that all preceding equipment must be stopped or a cascade of material on the idle conveyor will result. Occasionally combinations of detectors and magnetic head pulleys are used. The detector precedes the magnetic head pulley and detects only large pieces of tramp iron, while the magnetic pulley subsequently takes out the smaller ferrous metals. Suspended magnets overcome the disadvantages of magnetic head pulleys, detectors or a combination of both. There are several magnetic extraction problems involved in applying a suspended magnet at this

Jan 1, 1953

By R. L. Manegold

This paper outlines the reasons for and the application of suspended magnets above belt conveyors for the purpose of removing tramp iron to protect equipment. The design of a magnet strong enough to extract iron through a distance of 24 in. is described along with the proper method of application in the field. Advantages of the suspended magnet in contrast to the conventional magnetic pulley are also discussed. IN RECENT years there has been a decided trend toward bringing ores and coal out of open-pit and underground mines by long, sloping single-stage belt conveyors. Because the high investment cost of haulage conveyors dictates maximum speed and full loading for economy, conveyor belts are running faster than ever before, with a greater burden on the belts. In addition, larger blocks of ore are being mined and conveyed. Rougher use of large mining equipment generates more tramp iron; mechanical loading rejects nothing. Therefore the extraction of tramp iron and other deleterious materials for the protection of haulage conveyors at transfer points, chutes, and idlers is becoming an increasingly difficult job. In our experience, the best method of protecting long haulage conveyor belts is to extract the tramp iron, at least the large tramp iron, ahead of the main haulage conveyor or series of conveyors. It is most practical to collect the tramp iron at the earliest source possible, usually directly after the primary car unloading dump or after the primary crushers, which are ordinarily not affected by tramp iron, see Fig. 1. The coal or ore from its primary source is fed by chutes or feeders onto a buffer or shock belt conveyor, a short, ruggedly built conveyor system that can be considered somewhat expendable. Its primary function is to bring the ore up to speed for transmission to the haulage conveyor. There is no assured method of protecting this conveyor from tramp iron damage at its feed end. Likewise, there is a possibility, although remote if the magnet applied to this belt is installed correctly, that the tramp iron extraction process will damage the shock belt, which is, however, short and comparatively inexpensive. Magnetic head pulleys with their maximum effective range of about 8 or 10 in. cannot be considered for large installations, even though they are desirable because of the continuous and automatic tramp iron removal feature. Further, the most dangerous form of tramp iron with respect to potential damage to a haulage conveyor is the long rod, roof bolt, or rail segment, which is not extracted satisfactorily because of the limited tangential contact of such an iron shape with the magnet. Magnetic head pulleys are also undesirable because the dribble or fines from the head pulley discharge at the same point as the tramp iron. Screens must be used to separate the fines from the magnetics, resulting in additional expense and loss of head room. Magnetic detectors have been used with success, but they only indicate the presence of iron, and the conveyor belt must be stopped for its manual removal. The stoppage of any conveyor means that all preceding equipment must be stopped or a cascade of material on the idle conveyor will result. Occasionally combinations of detectors and magnetic head pulleys are used. The detector precedes the magnetic head pulley and detects only large pieces of tramp iron, while the magnetic pulley subsequently takes out the smaller ferrous metals. Suspended magnets overcome the disadvantages of magnetic head pulleys, detectors or a combination of both. There are several magnetic extraction problems involved in applying a suspended magnet at this

Jan 1, 1953

By Irving M. Symonds

Compania Minera de Penoles, S. A., at its Avalos unit operates a lead-zinc concentrator having a capacity of about 600 short tons per day. Lead and zinc concentrates are made by flotation methods. The concentrator is at Terminal de Provi-dencia, in the state of Zacatecas, some 70 miles south of Saltillo, Coahuila. The Penoles-Avalos railroad operating between the plant and Estacion Avalos, a distance of 6 miles, makes connections with the Coahuila and Zacatecas railroad, which in turn connects with the Nacionales de Mexico at Saltillo. The three railroads are each of different gauge, requiring the transfer of all supplies and concentrates at each junction. The mill operates on company ores, on ores from leased mines, and as a custom mill. The custom business is very small. The lead concentrates are shipped to company smelters at Monterrey, Nuevo Leon and Torreon, Coahuila. Since the European war, all zinc concentrates have been shipped to the United States. Ore Treated The major source of ore supply is a group of four mines owned or leased by the company, some three miles south of the mill in the mountains, from which the ore is delivered to the mill by a gravity aerial tramway. Custom ore comes in by rail. The ores occur in chimneys. The upper levels contained oxidized siliceous ores of lead, copper, silver and gold, with varying amounts of zinc. These ores were direct-smelted. In the first two years of operation a good deal of the tailing loss was caused by the oxidized content of the ores milled. Much of the ore in the transition zones between the oxides and sulphides was not mined. The ores now being milled are clean, hard sulphides in a limestone gangue, with only occasional contamination of oxides. The lead and zinc sulphides are of coarse structure and are freed even above flotation sizes. The galena is usually "steel" colored, easily and quickly floated. When there is any oxidation, it is usually on the coarsest crystals. The zinc minerals vary a good deal in color; a resin sphalerite predominates but there are varying shades of brown marmatites. On an average, the zinc sulphides contain 64 per cent zinc. The only common copper mineral is chalcopy-rite. Little is known of the occurrence of gold and silver. Waste minerals are mainly limestone and pyrite, both of which are barren. The ores vary from 20 to 35 per cent pyrite. Mill heads for the year 1940 and the first eight months of 1941 are given in Table I.

Jan 1, 1943

By Irving M. Symonds

Compania Minera de Penoles, S. A., at its Avalos unit operates a lead-zinc concentrator having a capacity of about 600 short tons per day. Lead and zinc concentrates are made by flotation methods. The concentrator is at Terminal de Provi-dencia, in the state of Zacatecas, some 70 miles south of Saltillo, Coahuila. The Penoles-Avalos railroad operating between the plant and Estacion Avalos, a distance of 6 miles, makes connections with the Coahuila and Zacatecas railroad, which in turn connects with the Nacionales de Mexico at Saltillo. The three railroads are each of different gauge, requiring the transfer of all supplies and concentrates at each junction. The mill operates on company ores, on ores from leased mines, and as a custom mill. The custom business is very small. The lead concentrates are shipped to company smelters at Monterrey, Nuevo Leon and Torreon, Coahuila. Since the European war, all zinc concentrates have been shipped to the United States. Ore Treated The major source of ore supply is a group of four mines owned or leased by the company, some three miles south of the mill in the mountains, from which the ore is delivered to the mill by a gravity aerial tramway. Custom ore comes in by rail. The ores occur in chimneys. The upper levels contained oxidized siliceous ores of lead, copper, silver and gold, with varying amounts of zinc. These ores were direct-smelted. In the first two years of operation a good deal of the tailing loss was caused by the oxidized content of the ores milled. Much of the ore in the transition zones between the oxides and sulphides was not mined. The ores now being milled are clean, hard sulphides in a limestone gangue, with only occasional contamination of oxides. The lead and zinc sulphides are of coarse structure and are freed even above flotation sizes. The galena is usually "steel" colored, easily and quickly floated. When there is any oxidation, it is usually on the coarsest crystals. The zinc minerals vary a good deal in color; a resin sphalerite predominates but there are varying shades of brown marmatites. On an average, the zinc sulphides contain 64 per cent zinc. The only common copper mineral is chalcopy-rite. Little is known of the occurrence of gold and silver. Waste minerals are mainly limestone and pyrite, both of which are barren. The ores vary from 20 to 35 per cent pyrite. Mill heads for the year 1940 and the first eight months of 1941 are given in Table I.

Jan 1, 1943

By F. Morgan, M. Muskat, R. G. Russell

Core experiments with radioactive tracers are reported in which an artificial radioactive material, radiovanadium, was used in aqueous solution in a study of the mobility of connate water. Sandstone cores saturated with active water were first oil-flooded to simulate virgin sand conditions in which the pores are filled with oil and connate water. They were then subjected to flooding with inactive water. Measurements were made of the activity of various portions of the effluent as well as the activity of the core after the flooding process was completed. In additional experiments the distribution of activity along the core was measured during the flooding process. In still other experiments direct flooding tests were made on cores fully saturated with active water. In all cases it was found that after injecting into the core a volume of inactive water equal to a pore volume, the effluent contained more than 80 per cent of the active content originally in the core. It is felt that these results imply a high degree of mobility of the connate water when subjected to flooding, although at least a partial contribution of the effect of tracer transfer within the core cannot be definitely excluded. The implications of this result with respect to water-flooding, core analysis and electric logging are discussed. A better approach to determining the connate water saturation of oil-producing rocks, by measuring the fluid content of core samples at the surface, is by the use of crude oil or oil-base muds as the coring fluid. Although water flooding is of considerable importance in the recovery of ail, comparatively little is known with certainty about the role played by the connate water during the operations. Two possibilities exist. Either the connate water is retained in the formation while the flood water sweeps past, or it is pushed along to a greater or lesser degree by the flood. Thus far no explicit recognition in terms of quantitative studies appears to have been given to the displacement of connate waters by the intruding extraneous waters, although such a possibility might be important in determining the efficiency of the oil -displacement process in water-flooding operations, especially in argillaceous rocks. On the other hand, it is a common observation that the first water produced with the oil as a result of water-flooding is definitely saline.*.' Similar questions regarding the mobility of the interstitial water arise in the quantitative description of the advance of edge-water drives. In the interpretation of the water saturations determined from core analyses, when coring with water-base muds, it is generally assumed that the drilling fluid merely adds to the connate water without displacing the

Jan 1, 1947

By M. Muskat, F. Morgan, R. G. Russell

Core experiments with radioactive tracers are reported in which an artificial radioactive material, radiovanadium, was used in aqueous solution in a study of the mobility of connate water. Sandstone cores saturated with active water were first oil-flooded to simulate virgin sand conditions in which the pores are filled with oil and connate water. They were then subjected to flooding with inactive water. Measurements were made of the activity of various portions of the effluent as well as the activity of the core after the flooding process was completed. In additional experiments the distribution of activity along the core was measured during the flooding process. In still other experiments direct flooding tests were made on cores fully saturated with active water. In all cases it was found that after injecting into the core a volume of inactive water equal to a pore volume, the effluent contained more than 80 per cent of the active content originally in the core. It is felt that these results imply a high degree of mobility of the connate water when subjected to flooding, although at least a partial contribution of the effect of tracer transfer within the core cannot be definitely excluded. The implications of this result with respect to water-flooding, core analysis and electric logging are discussed. A better approach to determining the connate water saturation of oil-producing rocks, by measuring the fluid content of core samples at the surface, is by the use of crude oil or oil-base muds as the coring fluid. Although water flooding is of considerable importance in the recovery of ail, comparatively little is known with certainty about the role played by the connate water during the operations. Two possibilities exist. Either the connate water is retained in the formation while the flood water sweeps past, or it is pushed along to a greater or lesser degree by the flood. Thus far no explicit recognition in terms of quantitative studies appears to have been given to the displacement of connate waters by the intruding extraneous waters, although such a possibility might be important in determining the efficiency of the oil -displacement process in water-flooding operations, especially in argillaceous rocks. On the other hand, it is a common observation that the first water produced with the oil as a result of water-flooding is definitely saline.*.' Similar questions regarding the mobility of the interstitial water arise in the quantitative description of the advance of edge-water drives. In the interpretation of the water saturations determined from core analyses, when coring with water-base muds, it is generally assumed that the drilling fluid merely adds to the connate water without displacing the

Jan 1, 1947

By R. M. Hardgrove

Pulverized coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the dificulties incident to the burning of coal. In a. cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverizeti in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. Process of Burning Cement Clinkers The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output, This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,' in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. Wet vs. Dry Plants The study presented here covers 88 plantsl of which 41 are wet and 47 dry. Of

Jan 1, 1942

By R. M. Hardgrove

Pulverized coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the dificulties incident to the burning of coal. In a. cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverizeti in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. Process of Burning Cement Clinkers The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output, This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,' in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. Wet vs. Dry Plants The study presented here covers 88 plantsl of which 41 are wet and 47 dry. Of

Jan 1, 1942

By J. F. Pierce, S. M. Enterline

Since January 1959 a zinc refiner of novel design has been in operation at Blackwell, Okla., producing 99.995+ zinc from the output of the Blackwell horizontal retort smelter. The refiner is a continuously operated, pyro-metallurgical unit with principal dimensions horizonhl permitting one floor operation. It is very ruggedly constructed. Furthermore it is not subjected to deterioration from high concentrations of iron in the feed metal because the boiling units are separate from the rectification columns and constructed with silicon carbide brick arches through which the heat is transferred radiantly to the zinc. A zinc refiner of novel design has been in operation since January 1959 at the Blackwell, Okla. plant of the Blackwell Zinc Co., Inc., a wholly owned subsidiary of American Metal Climax, Inc. Developed by the company's technical and operating staffs, the new refiner produces special high-grade zinc (nominally 99.995+ pct Zn) from the prime western1 slab output of the Blackwell horizontal retort smelter. The refiner is a continuously operated, pyromet-allurgical, high-capacity unit employing fractional distillation to separate zinc from its impurities, consisting principally of lead, cadmium, and iron. In prime western slab, these elements may vary in concentration from 0.005 to 1.6 pct. Table I shows the impurities usually encountered in the Blackwell prime western metal fed to the refiner together with their boiling points and vapor pressures at 907oC. The commercial pyrometallurgical smelting processes for extracting zinc from ore include distillation from retorts and a limited degree of refining can be accomplished by redistillation in similar retorts. To achieve special high-grade purity, an impractical number of batch redistillations would be required, however, with an accompanying loss of much metal. A continuously operated and integrated unit is required in which fractional vaporization and condensation can take place in a large number of stages. The first commercial continuously operated unit was developed by the New Jersey Zinc Co.4 In the separation of any two substances the degree of separation is limited if a mixture exists having a constant boiling point. The Zn-cd5 and Zn-pb6 systems contain no constant boiling points and hence these metals may be separated to any degree of purity. Basically, the pyrometallurgical refining of zinc involves three steps: a) iron, copper, and other metals having no appreciable vapor pressures at 907ºC concentrate in the bath when the impure zinc is boiled, b) lead, bismuth, antimony, and other metals having boiling points above but closer to that of zinc are selectively condensed from the vapor boiled from the bath, and c) cadmium and other metals having lower boiling points than zinc are fractionally distilled by reboiling and subsequent condensation of the zinc from the vapor. The boiling and condensation of zinc must be conducted in vessels constructed of materials which are refractory at the temperature involved (907 ºC) and it is highly desirable that they also be resistant to attack by the various metals. This requirement dictates construction from relatively small, simply shaped pieces which limits the pressures which may be employed to avoid leakage through the numerous joints. Stresses induced must be limited essentially to compressive ones and suitable provision must be made to permit thermal expansion of the components in order to minimize these stresses. Where heat is to be added for boiling or removed for condensation, refractory material through which the heat flows must have a high thermal conductivity. Silicon carbide is highly conductive but is subject to attack by iron and is expensive compared to other commonly employed refractories. A hard, dense aluminous fireclay is well suited to contain these metals where high thermal conductivity is not required. Zinc will penetrate the pores of fireclay brick but only to a limited depth in an exterior wall and has little effect other than an increase in weight and conductivity. The design objectives of the Amax refiner were: a) to achieve a long life by heavy, rugged construction and by employing silicon carbide refractories only where required for heat conductivity and not in contact with liquids of high iron concentrations, b) to simplify operation, manning and control by essentially horizontal construction, and c) to obtain high capacity in a single unit. DESCRIPTION Flow of metal through the Amax refiner is shown diagrammatically in Fig. 1 and schematically in Fig. 2.

Jan 1, 1963

By M. R. Tek, M. L. Katz

The flow of fluids in stratified porous systems, in which adjacent layers of the system are in communication over part or all of their common interface, in many instances involves crossflow between layers. A theoretical study has been made of the unsteady-state flow of slightly compressible fluids during depletion of bounded stratified porous systems. Consideration has been restricted to single-phase flow, with negligible capillary and gravity effects. Analytical expressions have been derived for pressure distribution and for (luxes at the producing face in two-layer systems with complete fluid communication between layers. These expressions occur in the form of double Fourier and Fourier-Bessel series. The method of extending the solution to any number of layers is demonstrated by deriving the equation for pressure distribution in a three-layer system. On the basis of the exact results for two-layer systems, certain generalizations regarding the behavior of stratified systems in general are presented. The performance of stratified systems, in terms of cumulative flux as a function of time, is shown to lie at all times between bounds established from single-layer theory. The upper bound is given by treating the system as a single layer with arithmetic average physical properties. The corresponding lower bound is the summation of the fluxes from each layer as treated individually. Variation in performance of stratified systems between the upper and lower bounds is determined by the extent to which crossflow occurs in the system. The principal parameters affecting the amount of crossflow are the vertical permeability of the system, the ratio of system thickness to "radius of drainage" and, for radial systems, the ratio of system thickness to interior radius. At very early times, the system performance is near the lower bound. As time and the radius of drainage increase, the per- formance tends toward the upper bound. INTRODUCTION One of the principal problems confronting the reservoir engineer today is the incorporation of the effect of reservoir heterogeneity into predictions of reservoir performance. The first part of this problem is the characterization of the actual reservoir heterogeneity in a mathematically useful manner.1, 2 It is then necessary to calculate the effect of this heterogeneity on the system performance. It is frequently possible to estimate this effect by comparison with the known behavior of an idealized system. One such idealized system is the stratified or "layer-cake" model. Many studies have been made of the behavior of stratified systems, from both a theoretical and experimental standpoint. The great majority of the theoretical studies have been concerned with systems of separated layers, with no fluid crossflow. Among the more noteworthy of such theoretical papers is that recently published by Lefkovits, et al,3 in which a comprehensive analysis was presented for the behavior of bounded stratified systems which are unconnected except at the wellbore. Three recent references have dealt with stratified systems including the effect of crossflow. Jacquard4 has presented mathematical derivations pertaining to depletion of two layer systems producing at constant rate. A doctoral dissertation by Katz5 contains analytical solutions and tables of results for pressure distribution and fluid flux during constant-terminal-pressure depletion of two-layer and three-layer systems with crossflow, together with tables of eigenvalues necessary in the computations. The dissertation also presents results from experimental studies of stratified systems using a heat-transfer analog model. This paper is largely based on the material in this dissertation. A third paper has been presented by Russell and Prats6 since completion of the work resented herein, containing mathematical derivations for and discussion of the behavior

By E. P. Pfleider, O. Rellensmann

For many years mine surveyors and exploration engineers have sought an accurate means of transferring meridian underground by using the gyro-compass. These efforts have generally failed, either because they were inaccurate or because the size and cost of the unit were prohibitive. During the past ten years, however, great strides have been made. Autonetics in California has developed a comparatively light-weight (100-lb) gyro-theodolite, and the Institute for Mine Surveying at Clausthal Mining Academy in Germany has developed an instrument that is accurate to within 20" to 30" of arc. Concurrently, Minneapolis-Honeywell has produced an astounding gyro, the size of a salt shaker, that weighs less than 0.5 lb and compares favorably in performance with the 50-lb gyros of conventional design. This unit, designed for missile guidance, should be adaptable for surveying small boreholes. Given the proper interest and approach from the engineering fraternity, excellent gyrocompasses could probably be available for accurate underground survey work at nominal cost within the next three years. Principles of the Gyro-Compass: The gyrocompass, used for 50 years as a navigational instrument in ships and aircraft, senses the turning axis of the earth and points out true polar north. It does this by employing the earth's rotation, the force of gravity, and the precession characteristics of a gyro. For further details the reader is referred to several excellent publications'"3 covering the principles and application of the gyroscope. The heart of the gyro-compass is the gyroscope, or rapidly spinning wheel, which rigidly maintains its angular orientation in inertial space. Free gyros employ universal joints to allow the wheel axle to maintain its rigid spatial orientation independent of base motion disturbances. Free gyros are independent of the earth's turning and are not north-seeking. Only as they might be initially oriented parallel to the earth's turning axis could any north indication be obtained. In practice, the free gyro tends to drift off its set position and lose its accuracy. The Sperry-Sun Well Surveying Co. instrument (5M in O. D.) is of this free gyro type. It is counterbalanced, however, against the rotational effects of the earth, so that orientation of the spinning axis will be maintained throughout the survey.

Jan 1, 1960

By R. D. Satterley

The Sherwood mine is an iron-ore mine owned and operated by the Inland Steel CO. in the Iron River district of the Menominee Range in Michigan. The property consists of an 80-acre tract in the village of Mineral Hills, about a mile north of the city of Iron River. The Ore mined is a reddish brown, nonbessemer hematite and limonite. The ore bodies lie in a large syncline having an east-west strike and are associated with a rock structure that is folded in a complex manner. The two 40-acre tracts run north and south and the shaft is at about the center of the north forty. The shaft is 1286 ft. deep, with the principal operating level at a depth of 1000 ft. from surface. A new level at a depth of 1200 ft. is in process of development. The ore body at the 1000 ft. level is east and west at approximately the line between the north and south forties and at this elevation is from 50 to 250 ft. wide and extends above the level at varying heights of from 50 to 250 feet. The 1000-ft. haulage drift extends south from the shaft, through the east-west ore body to the slates on the south side of the body. The ore body was intercepted in the drifting at about goo ft. from the shaft and was about 200 ft. wide at that point. A crosscut was extended to the east and west property lines along the south boundary of the ore body. This east-west ore body is laid out in areas go ft. wide for stoping with pillars 40 ft. wide. With this plan a stope is approximately 80 by 200 ft. and has a pillar on both sides that is approximately 40 by 200 ft. The ore has a height of from 50 to 250 ft. above the elevation of the level, therefore the height of the stopes varies in accordance with the height of the ore body at each location. From the crosscut, two main raises, each 6 by 7 ft., are put up into each 80 ft. stope area (Fig. 1). One of the main raises serves as a manway and the other as an ore raise, through which the ore is transferred from the stope to the tramming equipment. The original layout as for scraping of the ore from under the stopes into the ore raises. The ore raise is put halfway (east and west) along the 80-ft. stope area and usually is in iron formation. The manway and ore raise are on approximately 15-ft. centers. These main raises are Put up to a height of from 22 to 30 ft. above the level floor, then from the toll of the ore raise a scraper drift (8 X 8 ft.) is driven across the width of the ore body, which is in the neighborhood of 200 ft. From the scraper drift, mill raises are put up at 12½-ft. centers, staggered on opposite sides of the drift, and these are extended 23 ft. above the elevation of the scraper-drift floor. The sublevel connecting the top of these mill raises is called a mill "sub." Other sublevels are developed parallel to the scraper and mill subs at 25-ft. vertical intervals above the mill sub. The mill sub and the other subs above the mill sub are known as sublevels or dog drifts, and are 4 ft. wide and 6 ft. high. These dog drifts

Jan 1, 1946

By K. Brack, G. H. Schwuttke

Annealing properties of subszerface amorphous lavers produced through high-energy ion implantation in silicon are studied. The buried layers are produced through the implantation of ions (nitrogen), ranging in energy from 1.5 to 2 mev. X-ray interference patterns, transmission electron microscopy, and resistivity profiling are used to study the annealing characteristics of the ion damage. The annealing experiments indicate a low temperature (below 700°C) and a high temperature (above 700°C) region. Significant changes occur in the amorphous layer during the high-temperature anneal. Such changes are corre-lated with the re crystallization of the amorphous silicon and the formation of subsurface (buried) silicon-nitride films. TODAY'S main problems in the field of ion implantation are related to the accurate determination and prediction of 1) the distribution profiles of implanted ions, 2) the lattice sites occupied by the implanted ions, 3) the lattice damage produced through ion implantation, and 4) the annealing characteristics of damage centers in the lattice. This paper reports investigations concerned with the problems listed under 3) and 4). EXPERIMENTAL Our investigations cover the energy range of incident ions from 100 to 300 mev and from 1 to 2.5 mev. The emphasis of this study is on the energy range from 1.5 to 2 mev. The experiments are conducted with single charged nitrogen ions. To implant the ions a van de Graaff generator is used as described by Roosild et al.1 Accordingly, a gas containing the desired ion specie is passed through a thermome-chanical leak into a radio frequency activated source. The positive ions are driven into the van de Graaff with the help of a variable voltage probe. Emerging from the accelerator the ions drift into a magnetic analyzing system and here the desired ion specie is bent 90 deg into the exit port. The ion beam leaving the analyzer is defocused and drifts down a 4-ft long tube to hit the silicon target. At this position the 20 pamp ion beam has a circular cross-section of 2.1 cm. N2 is used as a source gas for nitrogen ions. The implantation target is silicon with zero dislocation density, 2 ohm-cm resistivity, (111) orientation, mechanically-chemically polished, and 1 mm thick. The target is mounted on a water-cooled heat sink and kept at room temperature. A fluence of 1015 to 1016 ions per sq cm is used. RESULTS 1) Silicon Perfection after Bombardment. High-energy ion bombardment of silicon has some striking effects on lattice perfection. Some results were reported in detail previously at the Santa Fe conference2 and are here briefly summarized for the benefit of the experiments described in the following. 1.1) Identification of Surface Films on Silicon. After bombardment all samples are found to be coated with surface films. The films on the silicon surface vary in thickness and color; they can be transparent, slightly brown, or opaque. The films are thicker and darker in the high-intensity area of the beam and they delineate the bombarded surface area of the crystal. The films produce electron diffraction patterns characteristic of carbon and of SiO2. Carbon is predominant. The presence of carbon in these films was confirmed by use of the electron microprobe. Formation of the films occurs independently of the ions used and is attributed to a contaminated vacuum of the high-voltage machine. The carbon is most likely the product of the pump oil which is cracked and polymerized under ion impact. The films stick tenaciously to the silicon surface and burn off in a low-temperature Bunsen flame. 1.2) Mechanical Perfection of the Silicon Surface. The mechanical perfection of the bombarded silicon surface was investigated through optical microscopy, electron microscopy in which the replica technique is used, and optical interferometry. No mechanical damage of the surface was visible after bombardment. However, if a bombarded sample is soaked for several minutes in hydrofluoric acid (HF), gas bubbles may develop in certain spots of the silicon surface. It is also noted that in these areas the surface film starts to peel off. Relatively large patches of film come off if the sample is soaked in HF during ultrasonic agitation. After HF treatment, pits may be present on the silicon surface. The pit dimensions are estimated to be as large as 50 µ. The pits appear in the region of most intense irradiation. 1.3) Lattice Perfection After Bombardment. No lattice damage is found on the silicon surface. Electron transmission micrographs and selected area diffraction patterns of the surface show no difference before and after bombardment. Measured approximately 2 µm down from the surface, the silicon lattice throughout this depth is of good perfection. Well-defined Laue spots and Kikuchi lines are obtained from the surface as well as from the indicated area below the surface. However, some radiation damage is dispersed in this top layer. A sharp boundary line separates this surface layer from a highly damaged layer which extends further downward into the silicon. Typical of this

Jan 1, 1970

By R. T. Hukki, A. M. Gaudin

Previous work on the principles of comminution has shown: (I) that the surface produced is proportional to work input (Rittinger law,1a-3); (2) that there is regularity to the weight distribution of the fragments according to size.4-10 In this paper it is brought out that the surface on the fragments is distributed in definite and simple relation to size. The experimental work has involved: I. Construction of an apparatus capable of measuring the surface of mineral fragments independently of the chemical composition of the fragments and independently of the site of the surface; 2. Refining the methods available for measurement of particle size; 3. Designing and building a crusher capable of operating on very small quantities of mineral, if necessary on single crystals, and in such fashion as to minimize handling losses of material. Complete experiments have been carried out with massive quartz and individual quartz crystals. partial experiments, lacking surface measurements, have been made with glass spheres, galena, sphalerite, fluo-rite, salt and coal. Introduction The existence of regularity in weight distribution with size was noted many years ago, especially by A. 0. Gates;' the subject was also studied in detail by one of us.6 In the latter paper it was shown that rocks can be divided into two classes: homogeneous rocks, which show no preferential breaking, and heterogeneous rocks, which show preferential breaking at the grain size of the rock. It was shown, furthermore, that the fragments obtained by rupturing a homogeneous rock occur (in the fine range) in such abundance as to permit the writing of a relatively simple equation to relate the weight retained on a screen to the screen opening. With screens of which the openings are in constant ratio, the weight retained, w, is related to the screen size x by the equation w = Cxm [I] In this equation C and m are experimental constants. A plot of weight retained against particle size is a straight line of slope m if the scales for x and w are both logarithmic. In the fine range, the slopes of the size-distribution graphs plotted according to the method described above vary within a limited range. An inspection of a number of size-distribution data by various experimenters indicates that the slope very seldom exceeds 1.0, but is usually greater than 0.7. More recently, schuhmann8 proposed a method of charting that has so many advantages of its own that it has been used primarily in this work. For fine particles Schuhmann has found that the data are represented by the equation (x)m k [2]

Jan 1, 1947

By A. M. Gaudin, R. T. Hukki

Previous work on the principles of comminution has shown: (I) that the surface produced is proportional to work input (Rittinger law,1a-3); (2) that there is regularity to the weight distribution of the fragments according to size.4-10 In this paper it is brought out that the surface on the fragments is distributed in definite and simple relation to size. The experimental work has involved: I. Construction of an apparatus capable of measuring the surface of mineral fragments independently of the chemical composition of the fragments and independently of the site of the surface; 2. Refining the methods available for measurement of particle size; 3. Designing and building a crusher capable of operating on very small quantities of mineral, if necessary on single crystals, and in such fashion as to minimize handling losses of material. Complete experiments have been carried out with massive quartz and individual quartz crystals. partial experiments, lacking surface measurements, have been made with glass spheres, galena, sphalerite, fluo-rite, salt and coal. Introduction The existence of regularity in weight distribution with size was noted many years ago, especially by A. 0. Gates;' the subject was also studied in detail by one of us.6 In the latter paper it was shown that rocks can be divided into two classes: homogeneous rocks, which show no preferential breaking, and heterogeneous rocks, which show preferential breaking at the grain size of the rock. It was shown, furthermore, that the fragments obtained by rupturing a homogeneous rock occur (in the fine range) in such abundance as to permit the writing of a relatively simple equation to relate the weight retained on a screen to the screen opening. With screens of which the openings are in constant ratio, the weight retained, w, is related to the screen size x by the equation w = Cxm [I] In this equation C and m are experimental constants. A plot of weight retained against particle size is a straight line of slope m if the scales for x and w are both logarithmic. In the fine range, the slopes of the size-distribution graphs plotted according to the method described above vary within a limited range. An inspection of a number of size-distribution data by various experimenters indicates that the slope very seldom exceeds 1.0, but is usually greater than 0.7. More recently, schuhmann8 proposed a method of charting that has so many advantages of its own that it has been used primarily in this work. For fine particles Schuhmann has found that the data are represented by the equation (x)m k [2]

Jan 1, 1947

By C. H. Herty

Almost every metallurgist who has given the Howe Memorial Lecture has had a personal contact with the distinguished gentleman to whose honor this hour is devoted. Unfortunately for me, such personal contacts were not possible. However, in the preparation of a thesis on the open-hearth process, one of my earliest tasks was to make a survey of the literature, and it was from Professor Howe's Metallurgy of Iron and Steel that I drew a number of references to previous work which formed the foundation for my reading. In his volume there is such a complete exposition of the then current thought on the subject of slags and on their role in the elimination of phosphorus from steel that I feel no hesitation in discussing in this Howe Lecture the subject of "Slag Control." In the present era of close specification, the drive for uniformity of product, and the close inspection of the product and of the cost sheet, the utmost efforts must be made to control the steelmaking variables to the best advantage of the steel producer. One of the major variables is the slag in the furnace and its effect upon the composition and quality of the metal. What advantages are to be gained from slag control? In the broad sense the advantages must be tremendous, because the very foundations of good steel melting are based on "shaping up" the slag to a point where, with proper temperature, the heat is ready for the final additions. Every melter works toward these ends through his experience with innumerable heats of steel. The elimination of undesirable constituents such as phosphorus and sulphur is entirely a matter of slag composition and physical nature, a point well known for over half a century. The role played by the slag in assisting in proper heat transfer from roof and flame to the steel has been equally well known. In more recent years there has come the problem of oxygen control in the more accurate gauging of additions for making high-grade rimmed steel for the new strip mills, and in the deoxidizing additions for grain-size control and cleanliness of forging steels, both carbon and alloy. In this problem the oxidizing characteristics of the slag are of prime importance.

Jan 1, 1940

By C. H. Herty

Almost every metallurgist who has given the Howe Memorial Lecture has had a personal contact with the distinguished gentleman to whose honor this hour is devoted. Unfortunately for me, such personal contacts were not possible. However, in the preparation of a thesis on the open-hearth process, one of my earliest tasks was to make a survey of the literature, and it was from Professor Howe's Metallurgy of Iron and Steel that I drew a number of references to previous work which formed the foundation for my reading. In his volume there is such a complete exposition of the then current thought on the subject of slags and on their role in the elimination of phosphorus from steel that I feel no hesitation in discussing in this Howe Lecture the subject of "Slag Control." In the present era of close specification, the drive for uniformity of product, and the close inspection of the product and of the cost sheet, the utmost efforts must be made to control the steelmaking variables to the best advantage of the steel producer. One of the major variables is the slag in the furnace and its effect upon the composition and quality of the metal. What advantages are to be gained from slag control? In the broad sense the advantages must be tremendous, because the very foundations of good steel melting are based on "shaping up" the slag to a point where, with proper temperature, the heat is ready for the final additions. Every melter works toward these ends through his experience with innumerable heats of steel. The elimination of undesirable constituents such as phosphorus and sulphur is entirely a matter of slag composition and physical nature, a point well known for over half a century. The role played by the slag in assisting in proper heat transfer from roof and flame to the steel has been equally well known. In more recent years there has come the problem of oxygen control in the more accurate gauging of additions for making high-grade rimmed steel for the new strip mills, and in the deoxidizing additions for grain-size control and cleanliness of forging steels, both carbon and alloy. In this problem the oxidizing characteristics of the slag are of prime importance.

Jan 1, 1940

By W. S. Gatley, F. C. Appl

This paper presents a combined analytical and experimental study of chisel penetration vs time during chisel impact on rock, a problem of fundamental importance in improving the performance of roller-cone bits or percussion drilling toots. For a given force-time relationship between chisel and rock, the problem of determining the penetration (displacement) vs time of the chisel is formidable. This is so because the rock is a nonlinear system with distributed mass and distributed damping (friction, dissipation of energy due to rupture, etc.). Since the literature does not contain adaptable solutions, the rock behavior to impact was simulated approximately by an "equivalent" lumped system, that is, an "equivalent" mass, spring, dashpot system. With this assumption, an analytical solution was found for chisel penetration vs time due to a sinusoidal load between chisel and rock. From this solution were found curves, in terms of dimension-less variables, for the maximum depth of penetration vs the frequency of the sinusoidal loading and for the energy transfer vs frequency. The results of this analysis were used to predict the penetration rate of rotary rock bits us rotary speed. The curve indicated that an optimum speed exists. To verify this analysis, an experimental apparatus was constructed and used to apply a sinusoidal pulse to a chisel penetrating a rock specimen under atmospheric conditions. Strain gauges were mounted on the chisel shank and a velocity transducer was mounted between the chisel and the rock surface. The velocity was integrated electrically and picked up simultaneously with the strain gauge signal on an oscilloscope. Permanent records were made photographically to provide simultaneous records of force us time and penetration us time. In comparing the experimental results for limestone and dolomite with the theoretical results, good agreement was found in the frequency range of the experiments. Unfortunately, the inertia effect (peak penetration) indicated by the theory occurs at a frequency much higher than could be obtained experimentally with the apparatus constructed. A "rate-of-loading" effect is indicated theoretically, but has not yet been verified experimentally. INTRODUCTION The process of drilling with percussion tools or rotary rock bits is basically related to the transient response of rock to surface impact. Each time a bit tooth contacts the rock, high stresses are developed which result in penetration and rock removal. As the tooth moves on, stresses are relieved and a new cycle begins as the next tooth contacts the rock. Thus the drilling process, which consists of an endless succession of these cycles, can be studied in terms of a single cycle. It is apparent, therefore, that the study of single-chisel impact on rock is fundamentally important in improving the performance of roller-cone and percussive-type drills. Previous studies in this area have been conducted by Simon1 and Hartman2 by means of drop tests. In these tests a chisel was attached to a weight and allowed to fall, due to the force of gravity, so that the chisel was driven into a rock specimen upon impact. Strain gauges were attached to the chisel shank and the resulting force-vs-time curves were recorded photographically from an oscilloscope screen. The depth of penetration and crater dimensions were also measured. These tests have provided much valuable information but, as mentioned by the investigators, have not provided complete information on the effect of "rate of loading'! This is partly due to the fact that the chisel motion during drop tests is not a controlled motion which can be varied in form and frequency. Therefore, it seemed that additional information could be obtained by studying chisel impact under conditions where both the motion and the frequency of loading could be controlled. This paper presents a combined analytical and exper-

By C. J. Ramsburg

The construction and operation of by product coke-oven plants in America are essential to strong national defense and of the greatest importance to many widely diversified undertakings as well as to safety and health. Carbonization Process Coal used under a boiler or in a locomotive is consumed by combustion. Coal used in a by-product oven is distilled or carbonized, meaning that heat is applied to the coal, out of contact with air, causing a distillation that removes the volatile matter in the coal in the form of gas and vapor, leaving coke as a final product. From the gas and vapor distilled from the coal, many valuable substances can be secured, known as by-products. Hence the name "by-product oven." The simplest example of this principle can be shown by means of an ordinary clay pipe. If powdered coal is placed in the bowl of the pipe, the opening of the bowl is plugged with clay, and the bowl placed in a flame, the bowl will become hot, transferring the heat through the wall of the bowl to the coal within. When this coal attains a temperature of about 750°F., gas will begin to pour out of the pipe stem, and this gas may be lighted. A black liquid also will exude from the stem of the pipe, This is tar mixed with certain other products. What remains in the bowl after distillation is coke. A coke oven corresponds to the bowl of the pipe. A modern oven of the latest type is a refractory brick chamber approximately 42 ft. long and 14 ft. high, averaging 18 in. in width. The oven is filled with coal through closable charging holes in the oven top. Each end of the oven consists entirely of a self-sealing door, the removal of which permits the finished coke to be discharged by means of a power-operated pusher machine. The ovens are built in batteries, and the heating is effected by gas burned in vertical flues in the oven walls, each set of flues heating the walls of two adjacent ovens. Such ovens carbonize a charge of approximately 18 tons of coal in 18 hr., or have a daily throughput of 24 tons per oven, producing about 17 tons of coke. In times of need, the output may be increased by raising the temperature of flues, thus shortening the coking time. There is one great essential difference in principle between distillation in a clay pipe and in a coke oven. In the 18-in. oven, the distillation proceeds from two parallel walls held at a high temperature (over zooo°F.) and instead of all the coal being heated at once, the distillation is a progressive one creeping inward. The average travel of this heating is 35 in. per hour from each vertical wall, and as the volatile products are formed progressively they must push away from the interior toward the wall through the previously formed hot coke, then up the walls and Over the top of the coal charge, finally escaping through the outlet pipe. This results in a cracking of the components into many different and valuable substances not formed when the carbonizing temperature is low.

Jan 1, 1942