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By Graham Bright

Metal mines were developed long before coal mines and the early lighting of underground workings was effected by torches and candles. The early coal mines were outcrop workings and little trouble was experienced from gas, but as the mines became deeper and shaft mines were opened it became apparent that an open-fame lamp was exceedingly dangerous where gas occurred. No satisfactory way of lighting gaseous coal mines was found until Sir Humphrey Davy brought out his famous safety lamp in 1815. The introduction of the Davy lamp was a great step forward but it was unsatisfactory from a light standpoint because the gauze absorbed a large percentage of the light given out by the flame. Clanny corrected this by the introduction of a cylindrical glass lens, which not only permitted a free distribution of the light from the flame but added greatly to the safety by eliminating the possibility of the flame coming in contact with the gauze and by protecting the flame against air currents. A very complete story of the flame safety lamp since its inception is given in the U. S. Bureau of Mines Bulletin No. 227, by J. W. Paul, I;. C. Ilsley and E. J. Gleim, published in 1924. A little more recent publication, known as Safety in Mines Research Board Paper No. 4, entitled "The Light Power of Flame Safety Lamps," by R. V. Wheeler and D. W. Woodhead, was published in England in 1927. The flame safety lamp was the only lamp available for gaseous mines for about 100 years. Although the incandescent electric lamp has been in use over 50 years, it was not found safe to use this lamp from wired circuits in gaseous locations. The storage battery was advocated shortly after the development of the incandescent lamp but until the introduction of the tungsten filament in the miniature mine-lamp bulb, the battery lamp was not a real possibility because of the excessive weight of the battery necessary to produce a sufficient amount of light. The actual development of the electric-battery type of lamp took place largely between 1910 and 1915. In England and Europe, the developmenb was almost entirely along the

Jan 1, 1936

By R. W. Gurry

Thermodynamic calculations indicate that boron is a better deoxidizer than silicon but probably is not quite as effective as aluminum. Boron should, therefore, be readily oxidized out of the open-hearth bath, which implies that it is not recovered in significant concentration from any scrap charged; also that, if added in the ladle, it tends to oxidize out during teeming unless protected by the presence of a better deoxidizing agent. These conclusions do not apply to liquid cast iron, in which the oxygen content is much lower than in liquid steel. The possibility of conserving certain strategic alloying elements widely used in steelmaking, notably manganese, chromium and molybdenum, by the use of boron, which in small concentration appears to be extraordinarily effective in promoting hardenability, raises a question as to what extent any boron present in the scrap charged into the open hearth is oxidized out during melting and subsequent refining. This question is significant because the optimum concentration of boron in the finished steel, which is fairly critical, would be very difficult to attain if the boron in the charge were largely retained in the metal. Moreover, unless boron is substantially eliminated in the open-hearth process its use on a large scale would eventually make it impossible to produce boron-free steel except by restricting the charge to virgin material. There is at present, unfortunately, virtually no satisfactory direct evidence to provide an answer to this question; but there are indications from practice that boron is oxidized out in the open hearth in the production of medium or low-carbon steel, although it may be retained if the carbon content remains fairly high; that is, if the oxygen is kept fairly low. In the absence of direct data, it is necessary to rely upon the indications of indirect evidence, the best approach now available being a thermodynamic calculation of the deoxidizing power of boron in liquid steel by the general method used so successfully by Chipman1 in his study of the reactions in the open-hearth process. The thermal data for boron, upon which such a calculation must be based, are reasonably satisfactory; the only quantities not directly determined are the entropy of melting of boron, the activity coefficient of boron in liquid steel, and the activity coefficient of B2O3in an open-hearth slag; but these values may be estimated with ample accuracy. The results of this calculation, which is presented later, are given in Table I, together with corresponding data for other deoxidizing elements. This table shows the calculated residual content of oxygen in equilibrium with different residual concentrations of the deoxidizer in the metal;

Jan 1, 1944

By R. W. Gurry

Thermodynamic calculations indicate that boron is a better deoxidizer than silicon but probably is not quite as effective as aluminum. Boron should, therefore, be readily oxidized out of the open-hearth bath, which implies that it is not recovered in significant concentration from any scrap charged; also that, if added in the ladle, it tends to oxidize out during teeming unless protected by the presence of a better deoxidizing agent. These conclusions do not apply to liquid cast iron, in which the oxygen content is much lower than in liquid steel. The possibility of conserving certain strategic alloying elements widely used in steelmaking, notably manganese, chromium and molybdenum, by the use of boron, which in small concentration appears to be extraordinarily effective in promoting hardenability, raises a question as to what extent any boron present in the scrap charged into the open hearth is oxidized out during melting and subsequent refining. This question is significant because the optimum concentration of boron in the finished steel, which is fairly critical, would be very difficult to attain if the boron in the charge were largely retained in the metal. Moreover, unless boron is substantially eliminated in the open-hearth process its use on a large scale would eventually make it impossible to produce boron-free steel except by restricting the charge to virgin material. There is at present, unfortunately, virtually no satisfactory direct evidence to provide an answer to this question; but there are indications from practice that boron is oxidized out in the open hearth in the production of medium or low-carbon steel, although it may be retained if the carbon content remains fairly high; that is, if the oxygen is kept fairly low. In the absence of direct data, it is necessary to rely upon the indications of indirect evidence, the best approach now available being a thermodynamic calculation of the deoxidizing power of boron in liquid steel by the general method used so successfully by Chipman1 in his study of the reactions in the open-hearth process. The thermal data for boron, upon which such a calculation must be based, are reasonably satisfactory; the only quantities not directly determined are the entropy of melting of boron, the activity coefficient of boron in liquid steel, and the activity coefficient of B2O3in an open-hearth slag; but these values may be estimated with ample accuracy. The results of this calculation, which is presented later, are given in Table I, together with corresponding data for other deoxidizing elements. This table shows the calculated residual content of oxygen in equilibrium with different residual concentrations of the deoxidizer in the metal;

Jan 1, 1944

By D. H. Davis

HAND picking was the earliest form of coal preparation, first practiced to improve the outward appearance of the coal being loaded and to remove any pieces that might appear objectionable to the buyer. It is still extensively practiced on the plus 1-in. sizes and particularly on the plus 4-in. and larger sizes. The percentage of ash reduction that can be obtained by hand picking depends entirely upon the amount of free impurity occurring in the coal. In a particularly dirty seam, where a considerable quantity of rock occurs with the coal as mined, hand picking produces sizable reductions in ash. If the coal is prepared clean at the working face hand picking is a final check; in some cases, on the smaller sizes, it is practiced only to show that an effort is being made to prepare a clean product. Because of the variety and different methods of handling coal, the early plants were not always adapted for e5cient hand picking. At most mines an attempt is made to - remove the visual impurities at one or more of the following places in the surface plant: (1) in the tipple during screening; (2) on picking band, belt, or shaking conveyors; (3) during the loading of the coal as it goes into the railroad car. Whatever the physical arrangements of the plant may be, the usual practice is to run the coal over the screen, picking table or boom at such a rate that each piece of impurity is exposed to the view of the picker. Coal should be uniformly sized and spread out on the table so as not to exceed one particle in depth. The flow should be even in tonnage and the speed of travel should be from 30 to 60 ft. per minute and not exceed 80 ft. per minute. The dirtier the coal, the slower the speed required for a fixed number of pickers. Actually, picking capacity depends on the number of particles per minute that can be picked out and rejected. The percentage of impurity removed is determined by this and the average weight of piece rejected. Handling a mixture of sizes from nut to the largest lump over the same picking table or conveyor is not conducive to efficient separation by hand picking because the lump hides the smaller impurities. Picking tables should not exceed 5 ft. in width, so that a picker on either side will be able to remove impurities from the center of the table. Guards usually are placed at the side of the screen, boom or table, to allow the pickers to stand close enough to remove the impurities without too long a reach.

Jan 1, 1950

By N. Plint, E. T. Pinkney

Copper segregation developments within the Anglo American Group are described with particular reference to the work of the TORCO Project Team. Some radical differences have been adopted in the mechan-ical approach as compared with methods used by previous workers. Most important among these are the following: (1) The heating unit is a fluidized-bed roaster. (2) A moving bed segregation chamber is used in which a packed bed of heated ore, to which reagents have been added, moves slowly downwards under gravity and is discharged at a controlled rate by means of a suitable valve mechanism. (3) The quantities of reagents used (sodium chloride and some form of carbon) are only a small fraction of those recommended by other investigators. (4) Retention times of only 2 min have been satisfactory for the complete segregation of many types of ore. Project TORCO is the code name given by Anglo American Corp. of South Africa, Ltd., to the investigation undertaken in Zambia to develop a plant capable of treating large tonnages of refractory copper-bearing ores by the segregation process. The process has now been under investigation for six years (1960-66) and during that period work has progressed from bench-scale laboratory tests through small-scale pilot-plant work to a 500-tpd commercial pilot unit. Test work has been carried out mainly on the ores of the Zambian Copperbelt and results have been encouraging. Recoveries on the order of 90% have been obtained on all Copperbelt ores tested with high concentrate grades. PRELIMINARY INVESTIGATION Laboratory scale segregation test work on Nchanga banded sandstone ore and oxidized ore from Kansan-shi was originally conducted by Professor Maurice Rey. SociCtC Miniere et Mktallurgique de Pefiarroya, in 1957. Both ores responded favorably to the process. In November, 1960, the consulting engineers of Anglo American Corp. of South Africa, Ltd., initiated preliminary investigations on the application of the segregation process to Copperbelt ores. In early laboratory work, samples were heated in a closed, externally heated tube furnace for one hour at temperatures ranging from 800 to 830°C. The free moisture content of the ore feed was 4% and reagent additions were 0.5% salt and 1.0% coal. Both salt and coal were pre-crushed to -48 mesh Tyler.* All samples except an ore from the Bancroft mine responded well to the process and recoveries of over 90% were obtained. The presence of dolomite was responsible for the relatively poor recoveries from the Bancroft ore, as tests on one of the other ores with additions of calcite or dolomite confirmed. Recoveries dropped from 94 to 45% Grades of all the rougher flotation concentrates produced were low at 10-15% copper. The laboratory investigation was then followed up on a pilot-plant scale. An attempt was made to carry out the segregation process in a coal-fired 5-ft-diam fluidized-bed roaster. The feed consisted of a banded sandstone from Nchanga to which 0.5% sodium chloride had been added. Reducing conditions were maintained in the roaster by the addition of coal in excess of that required for heating the feed. The excess coal and added salt constituted the segregation reagents and the temperature of the bed was held at approximately 750°C. Recoveries of copper and grades of concentrates were low. The reasons for lack of success were: (1) The reaction is a gaseous one and a fluidized-bed roaster is therefore unsuitable as a segregation unit. (2) Coal in excess of that required for heating was introduced into the fluidized bed and under these conditions some of the copper would be reduced in situ. The term "reduced in situ" is used here to describe any copper mineral reduced to the metallic state in its original site within the ore particles. If this copper is to be recovered by subsequent froth flotation it must first be liberated from the gangue by grinding or some other method which, in practice, is usually not feasible on ores of this type.

Jan 1, 1969

By H. L. Hartman

The drop tester has proved an invaluable tool for the investigation of percussion drilling in the laboratory in "slow motion". It has allowed the process of rock penetration by impact to be studied a single blow at a time. In the present work, the relation of cratering to indexing has been determined under conditions simulating those in the bottom of the hole at atmospheric pressure. Indexed blows form craters which are influenced by others adjacent to them. Two unique effects are involved: (1) the provision of additional free faces in proximity to the point of impact and (2) the creation of subsurface damage by the previous blow(s). Both have a pronounced influence on the volume of rock removed per blow, which governs the rate of penetration in actual drilling. The emphasis in this study was placed on determining effects of index distance and impact energy on crater volume when blows were struck on a previorusly "drilled" or damaged surface, a situation most representative of down-hole conditions. The results indicated that (I) the optimum index distance on a damaged surface is greater than on a fresh, undamaged rock surface; (2) the volume of rock broken at optimum index distance is about the same for both surface conditions with die-shaped chisels but tends to be greater on a cratered surface with wedges; (3) the optimum index distance and maximum crater volume are proportional to the energy level; (4) at a given energy level, the volume of rock broken by dies generally exceeds that by wedges; and (5) maximum crater volume varies, approximately, inversely with width of die but is nearly independent of included angle of wedge. INTRODUCTION This is the third of a series l,2 of papers reporting on one phase of a continuing program of drilling research being conducted in the Rock Mechanics Laboratory at The Pennsylvania State U. The phase under consideration is crater geometry — blow energy relations in percussion drilling. While the previous papers have discussed mainly craters produced by single blows, indexed-blow studies are the subject of this paper. The basic mechanism of drilling under study here — that of penetration of a chisel-shaped bit subjected to impact — is the one employed in probably 90 per cent of all rock drilling in mining, percussion drilling. It also forms the fundamental action in nearly all oil well drilling with the so-called rotary (roller-bit) method. Here, the action is more complex, however, and may in actuality comprise rotary-percussion drilling. Certainly, the down-hole air and mud hammer drills, which superimpose percussion on rotation, belong in this latter category. In any event, impact blows constitute one of the principal mechanisms in the vast majority of drilling, mining or petroleum, and it is essential in drilling research to study percussion in all its basic aspects. Percussion drilling can be simulated in the laboratory by means of an impact drop tester. Striking one blow at a time with chisel-shaped bits, the drop tester permits study of most of the parameters in the drilling process except blow frequency and the interrelationship of certain variables. Single-blow craters, the simplest to produce, allow the determination of basic relationships, such as volume vs blow energy; while indexed craters most closely resemble the action in real drilling and provide data for the establishment of optimum operating criteria. Indexed-blow craters are those which are influenced by others in proximity. The simplest case to visualize, and the one studied in detail here, is that in which a given crater is formed adjacent and parallel to a previous crater, such that the second crater breaks into the first. A more complex situation usually exists downhole, where indexing is achieved by rotation of a radial-winged or rolling-cutter bit. But, admittedly, the phenomenon of indexing is more likely to be understood eventually if it is first studied in its simplest form, that related to parallel craters. Limited research has been conducted on indexed blows. The most complete study was that by Simon, 3,4 who proposed universal indexing curves to relate the important variables. His work, however, was conducted entirely with rows of craters formed on a

By D. Harrington

FROM 1870 to date, the coal production of Utah has been somewhat less than 85,000,000 tons. There have been at least three major disasters with total fatalities about 380; or approximately. 4.4 persons killed by mine explosions per million tons produced, a record that is decidedly bad. Even during the periods when there have been no disasters, such as the years 1914 to 1920, inclusive, the fatalities .in ordinary working ran about 5.4 per million tons produced, against 3.7 per, million tons. for the whole United States for the same years. The five-year period following the war, 1918 to 1922 inclusive, showed 4.7 fatalities per million tons, hence the fatality record in the coal mines of Utah has uniformly been in excess of the average of the coal mines of the entire country. Notwithstanding this poor record, Utah has given to the coal-mining world many excellent safety methods and practices and the Utah. coal mines in general take greater measures toward safe mining than do the coal mines of any other State; if practices and methods allowed and pursued in coal mines of other States were tried in Utah coal mines, there would be a succession of disasters. Utah's principal producing coal mines, operated by about eighteen companies and numbering less than forty, have comparatively thick seams. Very little coal under 5 ft. thick is being worked and practically none under 4 ft.; much, if not most, of the coal comes from seams in excess of 8 ft., some running as high as 30 ft. In thick seams in which coal is left on floor and roof, any dust found or made is entirely combustible. In addition, the thick seams induce the use of large capacity cars, handling which in rooms and other workings on fairly heavy grades (up to 10 or .12 per cent.) tends to cause haulage accidents. The thick seams give high ribs and roof, in which it is difficult to prevent slabs of coal or rock from dropping on workers and also difficult to keep free of gas and of coal dust. The coal is high volatile, running usually over 40 per cent and in some instances nearly to 50 per cent; hence is highly flammable. Most of the coals have very high resin content, the resin usually being in dust form, hence almost as highly flammable as black powder. At .least one-third of the mines give off methane and several give off other

Jan 8, 1925

By Earl G. Alexander, W. R. Runyan

The nature of silicon growths which form on amorphous silicon dioxide during epitaxial deposition in unmasked areas was investigated. Octahedral silicon crystals nucleate and grow on the oxide at pinholes, adjacent to foreign matter, and probably at oxygen-deficient sites. Generation of these oxygen-deficient sites on the oxide surface is shown to be thermody-namically favorable under usual deposition conditions. The conditions which affect saturation (SiCl4 concentration, flow rate, temperature) determine the number of sites on which silicon will nucleate and grow. Reaction product HCl is likely responsible for suppressing silicon growths on the oxide near exposed areas of silicon. Evidence is given that the oxide-silicon interaction occurs in hydrogen only and involves water as a transport agent. SINCE shortly after the first epitaxial layers of silicon and germanium were grown over large areas, efforts have been underway to grow selectively on specified portions of the surface. Such selective growth should be particularly applicable to monolithic integrated circuits. Usually the approach has been to use a mask to delineate the desired area,'-7 but the VLS methoda can also be used, and direct programmed atomic beam deposition has been suggested.9 The masking techniques may embody two forms: in one the mask is a material which can later be removed along with any growths attached to it; in the other, attempts are made to prevent growth on the mask, so that over a whole slice the only deposition that will occur is in the unmasked regions. The present work is concerned with only the latter case, and specifically with silicon. Joyce and Baldrey1 suggested the use of thermally grown silicon dioxide as a mask material and showed that for the proper choice of operating conditions (total flow, SiCl4 concentration, total deposition time, and temperature) spurious growth could be minimized and that if the deposition times were not too large, and the criteria not too stringent, a usable process occurred. Later Schnable et al 4 observed that after extended deposition times, during which most of the silicon dioxide became covered with growth, there was always a band of oxide around the exposed silicon area that was quite free of spurious silicon crystallites. This latter observation is particularly intriguing since it extends the hope that with proper precautions masking can be much more versatile than it presently is. EXPERIMENTAL PROCEDURE Silicon substrates oriented to within 1.5 deg of a (111) plane were prepared by etch polishing lapped

Jan 1, 1967

By D. H. Davis

HAND picking was the earliest form of coal preparation, first practiced to improve the outward appearance of the coal being loaded and to remove any pieces that might appear objection- able to the buyer. It is still extensively practiced on the plus 1-in. sizes, and particularly on the plus 4-in. and larger sizes. The percentage of ash reduction that can be obtained by hand picking depends entirely upon the amount of free impurity occurring in the coal. In a particularly dirty seam, where a considerable quantity of rock occurs with the coal as mined, hand picking produces sizable reductions in ash. If the coal is prepared clean at the working face hand picking is a final check; in some cases, on the smaller sizes, it is practiced only to show that an effort is being made to prepare a clean product. Because of the variety and different methods of handling coal, the early plants were not always adapted for efficient hand picking. At most mines an attempt is made to remove the visual impurities at one or more of the following places in the surface plant:2 (1) in the tipple during screening; (2) on picking band, belt, or shaking conveyors;. (3) during the loading of the coal as it goes into the railroad car. Whatever the physical arrangements of the plant may be, the usual practice is to run the coal over the screen, picking table or boom at such a rate that each-piece of impurity is exposed to the view of the picker. Coal should be uniformly sized and spread out on the table so as not to exceed one particle in depth. The flow should be even in tonnage and the speed of travel should be from 30 to 60 ft. per minute and not 'exceed 80 ft. per minute. The dirtier the coal, the slower the speed required for a fixed number of pickers. Actually, picking capacity depends on the number of particles per minute that can be picked out and-rejected. The percentage of impurity removed is determined by this and the average weight of piece rejected. Handling a mixture of sizes from nut to the largest lump over the same picking table or conveyor is not conducive to efficient separation by hand picking because the lump hides the smaller impurities. Picking tables should not exceed 5 ft, in width, so that a picker on either side will be able to remove impurities from the center of

Jan 1, 1943

By L. M. Pidgeon, P. G. Thornhill

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. lengtn of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of ¼ in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10o to 15oC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press Operating at a temperature of 140°C

Jan 1, 1958

By J. H. Wernick, D. Dorsi, K. E. Benson

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. iength of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10O to 15OC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press operating at a temperature of 146°C

Jan 1, 1958

By W. R. Aufricht, E. H. Koepf

The theory, measurement, and application of capillary forces in reservoir rock have been treated ex-tensively in the various petroleum publications.1,2,3,4 These forces are of prime importance in oil production, and data on capillary properties are among the most important tools used in the study of reservoir mechanics. The principal application of measured capillary pressure data by the production and reservoir engineer is in the determination of the initial distribution and mobility of fluids in the reservoir. The distribution refers to the magnitude of the oil and interstitial water saturations under capillary equilibrium conditions and to their variation throughout the reservoir as a result of changes in rock properties and nearness to free water. The magnitude of the water (or oil) saturation, as well as the porosity, is required in volumetric calculations of oil in place. In low permeability reservoirs, where long transition zones are common, the magnitude of the oil content, even below the depth interval where commercial rates of flow of oil can be obtained, must be evaluated since this oil contains energy available to the production of the remainder of the reservoir. The mobility of the water phase is of particular importance in water-flood calculations since mobile water may condemn what would otherwise be a successful water flood. In reservoirs with predominantly small pores, high immobile water contents are possible. Conversely, reservoirs with predominantly large pores may contain mobile water even though the interstitial water contents are relatively low. This presentation reviews the distribution and mobility of interstitial water contents as determined from capillary pressure data and shows typical data for various types of carbonate reservoirs. A discussion of the variations found in different types of carbonate reservoirs is given. DISCUSSION Interpretation of Laboratory Data Fig. 1 is a typical laboratory capillary Pressure curve where water saturation as Per cent of Pore space is plotted vs capillary Pressure (psig) for an air-water system. This par- ticular curve was obtained by measurement on a sample of San Andres dolomite with intergranular-type porosity. The sample studied had a permeability of 10.0 md and a porosity of 18.8 per cent. The capillary pressure-saturation relationship illustrated is an inherent property of the rock (a function of the size and distribution of the pores) and represents the ability of this water-wet rock to retain water against the pressure differential exerted by a non-wetting fluid (oil or gas) to remove it. In the laboratory a known pressure differential is exerted from a pressure source; in the reservoir the pressure differential is controlled by the density difference between the hvdrocarbons and the water. The laboratory capillary pressure values may be converted into terms of equivalent height of water column using known values of reservoir oil and water density and measured or average values of interfacial tension

Jan 1, 1958

By W. Jolley

Decarburized iron and Fe-3.28 pct Ni alloys were impact and tension tested in the temperature range of ambient to 4°K. It was found that alloying with nickel improves the fracture properties of the ferrite matrix. The improvement in toughness of iron alloyed with nickel is attributed to the effect of the nickel addition on the temperature and strain rate dependence of the yield stress as well as on slip behavior. The Fe-Ni alloy exhibits wavy slip when strained at temperatures down to at least 50°K. This behavior contrasts with that of the unalloyed iron where a change It is a generally accepted fact that of the common alloying elements of steel only nickel and manganese improve both the room-temperature yield strength and the notch impact properties. It has been argued for a number of years that manganese refines the ferrite grain size, the pearlite spacing, and the grain boundary carbides in medium- and low-carbon steels and thereby increases resistance to brittle fracture.''' This hypothesis recently was confirmed for a 0.04 pct C iron containing manganese, but at the same time from wavy to planar slip occurs at about 120°K. The observed lower temperature and strain rate dependence of the lower yield stress in the Fe-Ni alloy as compared to the unalloyed iron may be attributed to this difference in slip behavior. In contrast to suggestions published previously, transmission electron microscopy showed that the uniform dislocation distributions and the lack of dislocation tangles which are produced by low-temperature straining of bcc materials cannot be attributed to the absence of cross slip at these temperatures. no increase in notch toughness was observed in a decarburized Fe-Mn alloy.3 It thus appears that manganese only increases the toughness of Fe-Mn-C alloys by refining grain boundary carbides and has no toughening effect in carbon-free iron. On the other hand, the mechanism whereby nickel improves toughness has not been established. Hodge et a1.4 have clearly shown that a 3.28 pct Ni addition lowers the impact transition temperature of 0.02 pct C steel approximately 35°C when the grain size is maintained constant. Recently, Wullaert and Tetelman6 stated that nickel refined the grain boundary carbides of an Fe-C alloy and consequently increased the toughness. They also observed the previously reporteda-'

Jan 1, 1969

By L. N. Cosner, C. F. Austin, J. K. Pringle

Shock wave attenuation studies with three igneous rocks are reported: a spessartite elastic to low intensity stress waves, a diorite which is anelastic, and a porous scoria. The level of shock passing through a given thickness of rock has been determined by donor-receptor methods using high speed streak cameras. Test results include shock propagation velocities adjacent to the explosive, the dependence of test results upon test geometry, and the variation in shock wave attenuation in these three common igneous rock fabrics. The utilization of explosives in rock mediums involves a series of interrelated phenomena. These are the detonation process within the explosive, the transfer of energy from the explosion products to the surrounding rock medium in the form of sound or shock waves, and the transmission of the resulting waves through the rock medium to create damage in the rock mass. The subject of this paper is the investigation of the close-in phenomena that occur in the intensely shocked rock immediately adjacent to a detonating explosive charge. The specific aspects studied include propagation velocities, free surface velocities, pulse attenuations and the determination of the equations of state for the test rocks employed. In addition, a series of tests was conducted to indicate the dependence of the test results on the charge diameter utilized. The equations of state for rock materials have been investigated by a number of authors. Lombard, ',' Grine and Fowles,3,4 Zharkov, 5 Hughes and McQueen,6 Chabai,7,8 and Bass9 have published equation of state data on rock materials similar in structure or composition to those investigated in the study. Rinehart10 has prepared a voluminous tabulation of the available equation of state data from these publications and in addition has included data on many other rock and earth materials in his compilation. The reader is referred to the extensive tabulations and bibliography of Rinehart for classical studies involving the equations of state of metals and other solids and liquids. The problems of sympathetic detonation between drill holes in hard rock environments has received the attention of various authors as have the problems of explosive rock coupling close-in stress wave attenuation and explosive diameter effects. 11 to 18 In particular, the don or-receptor geometry of this report is very similar to the geometry of adjacent drill holes in rock. Most published pulse velocity and attenuation work in rock has been devoted to velocities at much lower amplitudes than those associated with close-in explosives phenomena and most attenuation studies involve transmission and attenuation across large spans of material. Excellent work in these areas has been published by the U. S. Bureau of Mines and by many individuals and groups engaged in geophysical research. SELECTION OF TEST MATERIALS A series of rock materials was desired for this test program that would be meaningful industrially, and that would also yield comparative results from a research viewpoint. Based on the availability of a series of related petro-fabrics, for which there is an abundance of physical and shock response data, three igneous rocks were selected as the test series. The applicability of shock wave studies in igneous rocks to industrial problems is readily demonstrated. Hardrock tunneling, quarrying and mining take place in igneous rocks throughout the world. From a research point of view, igneous rocks exhibit a wide variety of grain sizes and considerable diversity in composition, yet the common igneous rocks have a general similarity in their constituent mineral structures and in their grain-to-grain bonding mechanisms. The latter is based on the concept of rock formation through magma cooling and crystallization, though the degree of liquefaction in granitic plutonic masses is indeed a moot question. Three specific types of igneous rock were chosen for study. A coarse-grained granitic textured rock was desired that would serve as a representative of the widely abundant intrusive rocks of the world and

Jan 1, 1967

By Donald F. Irwin

Amid the violent economic changes of recent years that have affected Canadian gold-mining operations so deeply, the discovery and early developments of Kirkland Lake might easily be overlooked. There was more than one "false dawn" for the district before developments finally established the fact of a superb mineralized zone extending through the present operating properties. Never far from the railway, Kirkland Lake has had the benefit of direct rail service for many years. The pioneer discovery on which the first mill and cyanide plant was erected was the "Tough-Oakes," this property being on what is now the eastern end of the producing area; some time later, the "Lake Shore" was discovered at the present center of the district, and a small plant was built upon it also. Both of these early treatment plants were of pre-war design. Progress in gold metallurgy has caused a gradual change in cyanide practice since the early days at Kirkland Lake, which may be observed in the various plants now operating there. From the beginning, it was recognized that Kirkland Lake ores were more responsive to cyanidation than to alternative recovery methods, and that their mineralization contained gold in very finely divided form. Although sulfides and tellurides are present, the quantity is small, being in the order of 1 or 2 per cent by weight, so that gravity concentration is undesirable, since gold is not confined to the sulfides. Some complication is offered by the presence of tellurides, which have caused lengthy controversy over their importance in recovery methods, but which are now generally regarded as less harmful in the light of increasingly high recoveries being obtained. Nevertheless, finely divided gold intimately associated with sulfides in the presence of tellurides; and occurring in the hard Kirkland Lake porphyry ore, substantially free from natural slime-forming constituents, indicates fine grinding. This has been the universal practice although the degree of grinding has varied a little among the different mills and is considerably finer now than formerly. The natural condition as outlined caused the establishment of fine grinding at an early stage in the development of Kirkland Lake, during the period of valuation of gold at, $20.67 per ounce. With the arrival

Jan 1, 1935

By W. M. McKewan

Magnetite pellets were reduced in flowing hydrogen at pressures up to 40 atm over a temperature range of 350° to 500°C. The rate of weight loss of oxygen per unit area of the reaction surface was found to be constant with time at each temperature and pressure. The reaction rate was found to be directly proportional to hydrogen pressure up to 1 atm and to approach a maximum rate at high pressures. The results can be explained by considering the reaction surface to be sparsely occupied by adsorbed hydrogen at low pressures and saturated at high pressures. PREVIOUS investigation1,2 have shown that the reduction of iron oxides in hydrogen is controlled at the reaction interface. Under fixed conditions of temperature, hydrogen pressure, and gas composition, the reduction rate is constant with time, per unit surface area of residual oxide, and is directly proportional to the hydrogen pressure up to one atmosphere. The reduction rate of a sphere of iron oxide can be described3 by the following equation which takes into account the changing reaction surface area: where ro and do are the initial radius and density of the sphere; t is time; R is the fractional reduction; and R, is the reduction rate constant with units mass per area per time. The quantityis actually the fractional thickness of the reduced layer in terms of fractional reduction R. It was found in a previous investigation2 of the reduction of magnetite pellets in H2-H,O-N, mixtures, that the reaction rate was directly proportional to the hydrogen partial pressure up to 1 atm at a constant ratio of water vapor to hydrogen. Water vapor poisoned the oxide surface by an oxidizing reaction and markedly slowed the reduction. The enthalpy of activation was found to be + 13,600 cal per mole. It was also found that the magnetite reduced to meta-stable wüstite before proceeding to iron metal. The following equation was derived from absolute reaction-rate theory4,8 to expfain the experimental data: where Ro is the reduction rate in mg cm-2 min-'; KO contains the conversion units; Ph2 and PH2O are the hydrogen and water vapor partial pressures in atmospheres; Ke is the equilibrium constant for the Fe,O,/FeO equilibrium; Kp is the equilibrium constant for the poisoning reaction of water vapor; L is the total number of active sites; k and h are Boltzmann's and Planck's constants; and AF is the free energy of activation. Tenenbaum zind Joseph5 studied the reduction of iron ore by hydrogen at pressures over 1 atm. They showed that increasing the hydrogen pressure materially increased the rate of reduction. This is in accordance with the work of Diepschlag,6 who found that the rate of reduction of iron ores by either carbon monoxide or hydrogen was much greater at higher pressures. He used pressures as high as 7 atm. In order to further understand the mechanism of the reduction of iron oxide by hydrogen it was decided to study the effect of increasing the hydrogen pressure on rebduction rates of magnetite pellets. EXPERIMENTAL PROCEDURE The dense magnetite pellets used in these experiments were made in the following manner. Reagent-grade ferric oxide was moistened with water and hand-rolled into spherical pellets. The pellets were heated slowly to 550°C in an atmosphere of 10 pct H2-90 pct CO, and held for 1 hr. They were then heated slowly to 1370°C in an atmosphere of 2 pct H2-98 pct CO, then cooled slowly in the same atmosphere. The sintered pellets were crystalline magnetite with an apparent density of about 4.9 gm per cm3. They were about 0.9 cm in diam. The porosity of the pellets, which was discontinuous in nature, was akrout 6 pct. The pellets were suspended from a quartz spring balance in a vertical tube furnace. The equipment is shown in Fig. 1. Essentially the furnace consists of a 12-in. OD stainless steel outer shell and a 3-in. ID inconel inner shell. The kanthal wound 22 in. long, 1 1/2, in. ID alumina reaction tube is inside the inconel inner shell. Prepurified hydrogen sweeps the reaction tube to remove the water vapor formed during the reaction. The hydrogen is static in the rest of the furnace. The sample is placed at the bottom of the furnace in a nickel wire mesh basket suspended by nickel wire from the quartz spring. The furnace is then sealed, evacuated, and refilled with argon several times to remove all traces of oxygen. It is then evacuated, filled with

Jan 1, 1962

By B. M. Larsen, T. E. Brower

IN an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats in different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper.2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, ?[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus permits more definite conclusions to be drawn. This excess concentration of oxygen, ?[0], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration ?[O] in liquid steel varies from about 0.009 to 0.035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following : [ ] as in a reasonably steady state-that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour-?[O] returns from

Jan 1, 1946

By Konrad Stern

The terms asphalt and bitumen have been applied interchangeably, although incorrectly on occasion, for centuries to various bituminous materials that were found in nature. Native bituminous materials of only the solid types, deposits of which are found on or near the surface of the ground, together with the principal asphaltic type petroleum bitumens will be considered in this chapter. The international dictionary compiled by the "Permanent International Association of Road Congresses" still gives the full name as "Asphaltic Bitumen" but of recent years the word "Bitumen" has been recognized as most appropriate and is defined by the American Society for Testing Materials as follows: "Mixtures of hydrocarbons of natural or pyrogenous origin, or combinations of both, frequently accompanied by their nonmetallic derivatives, which may be gaseous, liquid, semisolid or solid and which are completely soluble in carbon disulfide." Asphaltic pyrobitumens, as defined by Abraham,2 is a species of pyrobitumen, including dark-colored, comparatively hard and nonvolatile solids; composed of hydrocarbons, substantially free from oxygenated bodies; sometimes associated with mineral matter, the nonmineral constituents being infusible and largely insoluble in carbon disulfide. Deposits of native bituminous materials are found scattered over the entire world, but many are relatively unimportant as far as commerce is concerned. No attempt has been made to cover all of the known deposits. The principal bituminous materials of commercial interest are: 1. Native asphalts, solid or semisolid, associated with mineral matter such as Trinidad Lake asphalt, Rock asphalt and Selenitza, Boeton and Iraq asphalts. 2. Native Asphaltites, such as gilsonite, grahamite and glance pitch, conspicuous by their hardness, brittleness and comparatively high softening point. 3. Asphaltic bitumens obtained from nonasphaltic, semi-asphaltic and asphaltic crude petroleum by distillation, blowing with air and the cracking of residual oils. 4. Asphaltic pyrobitumens of which wurtzilite and elaterite are of chief interest industrially as they depolymerize upon heating, becoming fusible and soluble in contrast to their original properties in these respects.2 5. Mineral waxes, such as ozokerite, characterized by their high crystallizable paraffine content. Native Lake Asphalt At Brighton situated in the southwestern part of the Island of Trinidad is one of the largest deposits of native asphalt in the world.3 The lake is oval in shape, covering an area of 114 acres and having a depth of 285 ft at its center. The material forming this deposit is an emulsion of water and bitumen with a suspension of mineral matter consisting of impalpably fine clay dispersed by nature in the asphalt, and small quantities of silica and vegetable matter. The entire mass of asphalt is in slow but constant motion as evidenced by the migration of islands of vegetation from one place to an-

Jan 1, 1960

By J. E. Burke, Y. G. Shiau

SEVERAL attempts have been made to account for the fact that grains in a fully recrystallized metal will coarsen on annealing Two fundamentally different hypotheses have been advanced; with several variations of each. One school considers that the cause of grain growth is the surface energy of the grain boundary The theory of Jeffries1 that grain size and grain size contrast control growth is an example The second school considers the primary cause to be a difference in the lattice energy on either side of the grain boundary The more perfect grains are considered to grow at the expense of their less perfect neighbors, with a decrease in free energy, and a consequent increase in the stability of the system. Burgers2 presents the case for the strain or lattice imperfection theory of grain growth quite completely The latter theory is attractive, since there is no doubt that differences in lattice energy or lattice perfection can cause grain boundary migration A recrystallization nucleus grows for this reason as was beautifully demonstrated by van Arkel and Ploos van Amstel3 Although it is generally stated that grain growth does not precede recrystallization, several workers4,5 have shown that small strains may induce a single crystal in an aggregate to grow, or at least may cause some' grain boundary migration The origin and nature of the imperfections that are considered to be responsible for normal grain growth have never been clearly described, but it is generally agreed that they are a consequence of the difficulty of atomic reorientation in the solid state, and that they do not seriously differ from the type of imperfection that-can be introduced by mechanical deformation An object of the present work was to determine whether the introduction of mechanical deformation would cause grain growth in a specimen which would not otherwise show it under the conditions used It was also planned to determine the effect of such deformation the rate of growth in a specimen which would show - growth in the unstrained condition Recently Maddigan and Blank6 have indicated that slightly strained specimens of alpha brass can undergo considerable growth prior to recrystallization if annealed at low temperatures French7 has reported that at deformations less than 17 pct it is very difficult to detect the beginning of recrystallization in the same material. It was therefore planned to broaden this work to include a complete study of the microscopic behavior of alpha brass under conditions of temperature, time, grain size and deformation such that recrystallization would not occur or would occur to only a small extent EXPERIMENTAL PROCEDURE The brass was prepared by melting cathode copper and commercially pure zinc

Jan 1, 1947

By C. S. Smith, K. K. Ikeuye

In an earlier paper1 one of the authors called attention to the significance of the relative free energies of grain boundaries and interphase boundaries in alloys in determining the shape and distribution of micro-constituents. The ratio of the interphase and inter-crystalline boundary energies (that is, —?12, where ?12 is the energy of the interface between a crystal of phase 1 and a crystal of phase 2, and ?11 is the energy of the boundary between two crystals of phase 1) will be denoted by F. It is this ratio that determines, by the simple triangle of forces relation, the dihedral angle, ?, of phase 2 where it adjoins a boundary between two grains of phase 1. On a metallographic sample various angles are actually obtained because of the random orientation of the plane of observation, but the statistically most frequent angle is the true one. These basic principles were discussed in detail in the earlier paper. The experiments described below show the effect of both composition and temperature on ? and T for some alloys containing a liquid phase, and provide data on the rate of approach to the equilibrium angles in both solid and liquid phases. All the alloys were melted in a graphite crucible and cast into ingots 3/4 by 11/8 in. cross-section. Some of the aluminum alloys consisted of two immiscible liquids from which uniform castings could not be obtained. The amount of liquid phase varied locally to a considerable extent and the analyses were meaningless. Nevertheless, the composition of the liquid would not differ from that in an alloy of the intended gross composition, and the synthetic composition of the alloys was therefore used in considering the surface energy results. The ingots were given initial deformation by cold-rolling to 50 pct reduction, annealed, and given a second 50 pct reduction prior to cutting into small samples for heat treatment at the times and temperatures subsequently reported. The resulting micro-samples were carefully prepared metallographically and examined at magnifications of 650, 1280, or 2500. The objectives used were 21X, 0.40 NA (dry); 41X, 0.65 NA (dry); and 80X, 1.40 NA (oil emulsion). The appropriate angles were measured by rotation of the stage of a Bausch and Lomb metallograph to align the appropriate boundaries successively with a cross-hair in a micrometer eyepiece. About 250 angles were measured on each phase. Though the angles were read to 1°, the accuracy is probably not above 3°, because of uncertainty in setting the cross-hair tangent to the sometimes curved boundary. The corners of very small particles, particularly in early stages of adjustment after cold-rolling, were difficult to measure and the results were slightly dependent on the magnification used (compare Fig 1-C and 1-D). With particles of larger size and more nearly equilibrium shape, magnification made no difference (vide Fig 1-F and 1-G). The measured angles were grouped in 10° intervals (5° if ? was less than 20°) and plotted in bar

Jan 1, 1950