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By S. S. Clarke

The sheet-ground deposits of the Tri-State district, because they are fairly uniform in thickness (7 to II ft.)—rather flat, with an easy dip to the west—and cover a large acreage, offered a problem of improved transportation. Heading advance in this kind of mine is very rapid and track haulage requires a larger crew of trackmen than do other mines, and a heavy investment in track. The use of some type of rubber-tired truck equipment for underground haulage was under consideration for some time by the Eagle-Picher Mining and Smelting Co. The nature of some ore bodies made them not adaptable to truck haulage and some mines that did offer possibilities were equipped with other types of haulage, and the expected savings by the use of trucks did not offset the loss of investment that would be incurred in discarding other equipment. The decision was made to sink a new shaft, designed for skip hoisting, on the Blue Goose ore body, and then develop truck equipment. Developing Tractors and Trailers A study was made of internal-combustion engines for power units in the tractors, but the extreme difficulty of controlling air circulation, and the heavy cost of drilling large churn drill holes for blower installations, led to the adoption of a tractor with a storage battery. The storage-battery tractor that was on the market for mine haulage was impossible to obtain because all were needed for defense work. Several standard electric-truck chassis were purchased and rebuilt in the Eagle-Picher shops. The steering wheel was shortened and set vertical, permitting the driver's seat to be moved forward, and the frame was shortened enough to take the battery box and the fifth wheel support that carried the gooseneck of the dump trailer. The fifth wheel, or turntable, is of the ball and socket type. Tig. I shows the ball fastened to the gooseneck. The socket is split to permit introduction of the ball and is held by through-bolts, and the socket is welded to a ½-in. plate, which in turn is attached to the tractor frame. Smaller wheels and tires were put on to reduce over-all height. If it had been necessary, the spring: could have been reversed, or taken off entirely, to gain additional clearance. The tractors are equipped with 48-cell, 23-plate, 400-A.H. batteries. With the initial tonnage and mileage, the batteries were about 50 per cent down at the end of a shift. Tractors have truck tires 7.50 by 20, 8-ply, front and rear, the rear being dud. After four months of use, the .tires are beginning to show signs of failure.* In

Jan 1, 1943

By S. C. Jones, J. A. Davis

This paper describes displacement mechanisms of micel-lur solution slugs, displuced by a thickened water "mobility buffer", in a glass micromodel and in consolidated Berea sandstone cores. Colored motion pictures, taken through a microscope, recorded interactions on a pore-to-pore bash in the micromodel. Production and differential pressure data from consolidated cores previously waterflooded to residua1 oil saturation revealed displacement mechani.sm.s macroscopically. Oil is displaced miscibly, while interstitia1 water is displaced immiscibly by the injected micellar. solution slug. Part of this water is dispersed and carried near the lead ing edge of the .slug, and some may remain as a residual saturation. Thickened water displaces any residual water. It also emulsifies with the. slug material, thereby displacing it. The displacement sequence forms banks of oil and water with relatively constant fractional flows, saturations, and mobilities. The mobility of oil and water flowing ahead of the .slug defines .slug viscosity required for stable displacement. Equations are presented that describe. several aspecis of the displacement flow behavior: oil breakthrough, oil cut during production of the stabilized bank, breakthrough of banked water and the resulting decreased fractional flow of oil, fluid mobilities, etc. Breakthroughs, calculated from the equations, are compared with experimental data. Slugs of micellar sollution, ranging from 1 to 5 percent PV, produced unusually high oil recoveries (up to 100 percent) in previously waterflooded consolidated cores. Investigation of the effect of mobility buffer size on oil recovery indicated that recovery is reduced in 3-in. X 4-ft cores for mobility buffer sizes .smaller than 50 percent of the pore volume. Introduction A new oil recovery process,* conceived in 1962, utilizes a slug of micellar solution' that has sufficient viscosity to avoid an unfavorable mobility ratio with respect to the oil and water it displaces. These micellar solutions are composed of hydrocarbons, water, and surfactants, and may contain electrolytes and co-surfactants. The micellar solution is displaced by a "mobility buffer". The mobility buffer is a bank of reduced-mobility fluid, preferably an aqueous polymer solution. The entire system is displaced by unthickened water. This paper describes some of the microscopic and macroscopic displacement mechanisms observed during laboratory development of the fluids for a potential field test. Microscopic phenomena were observed photographically in a specially constructed glass flow model and macroscopic data were taken from constant rate, tertiary displacements conducted in Berea sandstone cores. Qualitative Description of Micellar Solution Flooding This section qualitatively describes displacement mechanisms that occur when a slug of micellar solution is injected into a previously waterflooded core. The slug completely removes oil from the portion of core it contacts, and banks oil in the process. The oil saturation in this bank and the flowing WOR reach definite, fixed values that are determined by water and oil mobilities and slug characteristics. This zone of constant WOR and oil saturation will be called "stabilized oil bank" in the following discussion. When this stabilized bank reaches the outlet end of the core, the oil cut jumps from zero to a constant value equal to the fractional flow of oil. Saline water has little or no solubility in the slug. Because of this. the slug displaces in-place brine immiscibly. During slug injection, most of the brine is displaced ahead of the slug, some is carried along in the leading portion of the slug, and part of the brine may become immobile. The immobile, residual brine is later miscibly displaced by thickened water. Displacement of all the oil and only part of the water during slug injection results in a higher fractional flow of oil ahead of the slug than occurs when all the water is displaced after injection of thickened water is begun. The slug spontaneously emulsifies fresh (low-salt content) water. By emulsification, the slug can become infinitely diluted, and hence totally displaced, by thickened water. Fluid saturations in the core are shown schematically in Fig. 1. Parts A and B show conditions before and after the start of the thickened-water injection, respectively. Since thickened water displaces any residual brine by-

Jan 1, 1969

By L. W. Rowe, S. S. Cole

A laboratory bench scale unit is described whereby finely divided chlorinatable residues are held for a short period by a restraining bed of a coarse-grained ore of comparable composition to permit 'flash' chlorination of the high surface area powder. Results on chlorinating a finely divicled titania residue using rutile sand as the restraining bed showed that 92 pct utilization of chlorine and equivalent conversion of TiO2 in the fine residue to Ticl4 were attained with a dust loss of about 3 to 4 pct and an equivalent amount of ch1om'nation of the titanium values in the restraining bed. THE chlorination of coarse grain titaniferous ores; i.e., 20 to 150 mesh sizes in fluidized beds is practiced commercially.' One effective method found for the chlorination of a very fine material in a fluidized bed is the "flash" chlorination technique.' By charging the finely divided charge material into the bottom of an established bed of optimum sized rutile, ilmenite, titanium beneficiate, or even an inert material such as sand, chlorination occurs before the dust can pass through the bed. The major requirement is that the bed be less chlorinatable than the charge. In this process, the charge consists of a finely divided (between one and 50 µ) titania residue obtained from a titaniferous ore beneficiation process. The charge is blended with a 20 pct wt basis addition of 20 to 60 .mesh size low volatile coke (to minimize hydrogen chloride formation) and fed to the bottom of the bed with the chlorine gas stream. A ball check device is used to permit entry of the charge and prevent the bed from discharging when the gas flow is cut off. The coarse pieces of coke serve to reduce compaction in the charge mixture and permit sat]sfactory feeding without plugging or jamming. Feed and gas rates are adjusted to provide for complete reaction of the charge material, with 90 to 95 pct chlorine efficiency, and maintain a gas flow in the range of 0.4 to 0.6 ft per sec. The equipment consisted of a 3-in. diam mullite tube enclosed in an electric furnace. A bow-shaped 3/4 in. Saran tube fed the charge from an air lock type of feed hopper (to prevent the escape of the chlorine gas into feed bin) to the bottom of the bed. The charge rate was set by providing a small awiliary hopper below the feed hopper where a weighed portion of the charge could be discharged in a specified time interval into the main chlorine stream. The chlorine rate was measured by a flowmeter. The reaction tube was filled with a bed of Australian rutile ore, this restraining bed serving two purposes; i.e., it rapidly heated the fresh charge to the reaction temperature and retained the particles in the bed for a sufficient time for selective chlorination to proceed. Due to the fine particle size and open, porous structure of the charge, this reaction was almost instantaneous, and, with proper adjustment of feed and gas rates, very little fine material passed through the bed without reacting and only a very small fraction of the restraining bed was chlorinated. A typical reaction procedure was to fill the bench scale fluidized bed unit with a 1 kg charge of Australian rutile ore analyzing 95.5 pct TiO2, 0.6 pct FeO, a beach sand deposit 40 to 150 mesh size. When fluidized, the bed was approximately 2 ft in depth. The ore was fluidized with a 1:l mixture of chlorine and nitrogen at a rate of 0.40 ft per sec. This gas mixture was necessitated by the comparatively short bed depth; a larger unit, with a deeper bed, could be fluidized with chlorine alone without significant free chlorine loss. Feed rate was 8.0 g per min of beneficiated ilmenite analyzing 85.0 pct TiO2, 2.6 pct FeO plus 1.6 g per min of 20 to 60 mesh petroleum coke. The

Jan 1, 1962

By G. H. Wilson

INTRODUCED in 1946 to serve a need in power-plant operation, closed-circuit TV has been used by well over 200 organizations in approximately 25 different industries. Known as industrial television, or simply ITV, it can be described as a private system wherein the television signal is restricted in distribution, usually by confinement within coaxial cable that directly connects the TV camera to one or several monitors, Figs. 1, 2. The picture is continuous and transmission is instantaneous, permitting an observer to see an operation that may be too distant, too inaccessible, or too dangerous to be viewed directly. Destructive testing or the machining of high explosives can now be conducted hundreds of feet away by personnel who still have close control through the eyes of the TV camera. It is also possible for one man to control operations formerly requiring the co-ordinated efforts of several workers. For example, at a large midwestern cement plant conveyance of limestone from primary crusher to raw mill and loading into five storage bins once necessitated the work of two men, one having little to do but prevent spilling of material by manually moving the tripper on the belt conveyor as occasion required. TV cameras mounted on the tripper now provide bin level indication to the conveyor operator at the crusher position so he is able to control the entire loading operation remotely, Fig. 3. By means of a switch, the picture from either camera is alternately available on a single viewer, or monitor, Fig. 4. Each camera is mounted on the tripper by means of a simple adjustable support and looks down into the bin, which is identified by the number of cross members on the vertical rod. Each associated power unit is located on a platform above the camera, Fig. 5. This centralized control by means of TV often has produced superior results, and in many instances saving in operating costs has been sufficient to write off equipment costs within six months to a year. Where a key portion of a process may be enclosed or otherwise inaccessible, TV again reduces the likelihood of mistakes and permits closer control by making available to the operator valuable information he might otherwise never possess. An example of this can be found at a strip mine where the coal seam lies 50 ft or more below the overburden, which is removed by a large wheel shovel. From his centrally located position the shove1 operator was unable to judge accurately to what extent the wheel buckets engaged the earth. His chief indication of efficiency was the amount of overburden on the belt conveyor as it passed his control point 75 ft from the wheel. Now, two television cameras mounted on the tip of the boom permit the operator to view the wheel from each side and provide him with a close-up view of the buckets so that he can take immediate and continuous advantage of their capacity, quickly compensating for ground irregularities and avoiding obstructions, Fig. 6. While the word television conjures up visions of highly complex and intricate apparatus such as that employed in modern TV studios and transmitting stations, the term industrial television should indicate compact, straightforward equipment. Most present-day ITV systems contain fewer than 25 tubes including camera and picture tubes. The average home television receiver alone requires at least that many tubes. Equipment like that illustrated in Fig. 1 contains only 17 tubes, of which 3 are in the camera. It can operate continuously and dependably, without protection, in any temperature from 0" to 150°F. It consumes less current than a toaster and weighs under 140 lb. Camera and monitor may be separated by 1500 to 2000 ft and by greater distance with additional amplification. This equipment is designed to withstand vibrations up to 21/16 in. and will operate successfully under more severe conditions of vibration and heat when suitable enclosures are provided. Any number of cameras may be switched to a single monitor, and any number of monitors, within reason, used simultaneously. Two types of applications in the mining industry have already been described. A third under serious consideration by several organizations will make use of ITV for remote observation of conveyor transfer points at copper concentrating plants so that evidence of belt breakdown and plugging of transfer chutes can be spotted immediately and costly overflow of material avoided. A television camera will soon be installed to view a trough conveyor near the exit of an iron-ore crusher to indicate clogging of the crusher as evidenced by reduction or absence of material on the

Jan 1, 1955

By R. L. Slobod, C. Gatlin

This .study defines the basic mechanism of the mis-cible displacerrzent of oil and writer from porous Medici by various water-driven alcohol .slugs. Three distinct alcohol slug processes were .studied. Considerable data concerning the quanity of alcohol required for oil recovery were also obtained. All data were obtained in a 1-in. diameter, 100-ft long, unconsolidated core. The porosity of this system was 35 per cent. and the permeability was approxitrlately 4 darcies. Total core pore volume wax 5,716 cc. All displacements were conducted at a constant injection rate of 5 to 6 cc//min, which corresponds to a frontal advance of 5 ro 6 ft/hr. The first portion of this paper is concerned with 111 use of one alcohol—isopropyl—a\. the slug material. Isopropyl alcohol (IPA) is completely miscible with both oil rind water; however, miscibility of the threc-component systems, oil-water-IPA, requires a relutively. high concentration of IPA. Hence, the displocement is not of the misoible type unless the IPA concentration is maintained above some critical value. A slug of IPA equal to only 13.5 per cent of the pore volume was jound to be sufficient to obtain complete recover?, of residua1 naphtha. In later studies two distinct process variations were developed. The firsr of these utilized methyl alcohol (MA) and IPA as slug materials. It was shown that methyl alcohol may he substituted for IPA at the front and rear of the slug with no loss of oil recovery. A .slug of 4 per cent MA-4 per cent IPA-4 per cent MA was sufficient for complere oil recovery. Because MA is considerably cheaper than IPA, this represents an important step toward economic application. A .second process variation used normal butyl alcohol (nBA) and MA as the composite slug, the nBA segnzent being injected first. This technique requires the smallest total slug size (approximately 10 per cent) of all processes studied. The high cost of nBA, however, precludes commercial application. It is possible that this basic process, subject to changes of alcohol type, may lead to a commrercicil process. INTRODUCTION Within the last 10 years, considerable study has been devoted to the general mechanics of miscible phase dis- placements in porous media. These studies have. in general, dealt with two-component displacernent. All reported field trials to date have been modified gas drives of one type or another. These utilize either lean gas at high pressure, rich gas at lower pressure or an LPG slug. The irreducible water content of the reservoir remains immobile during these processes. Although these techniques hold great promise. certain drawbacks do exist. Poor areal sweep efficiency, inherent in any displacement having a highly unfavorable mobility ratio, is one disadvantage. Furthermore, the miscibility of all processes which use gas is dependent on pressure. These pressure limitations may prohibit application to shallow reservoirs. There are also many areas where large quantities of LPG and natural gas are not readily available. These factors emphasize the need for improved techniques. The alcohol slug process is also a miscible-phase displacement process. It differs from the previously mentioned two-component techniques, however, in that both the reservoir oil and water are displaced by the slug. This behavior is a consequence of the miscibility of certain oil-water-alcohol systems. In the simplest case, a relatively small volume (slug) of an alcohol (such as isopropyl) is injected into the system. Water is then used to drive the slug through the medium. Thus. three components—-alcohol. oil and connate water-—exist at the front side of the slug. Miscibility is obtained at some alcohol concentration, this being dependent on the solubility of the particular system. Miscibility is maintained within the slug until the alcohol concentration falls below that value required to maintain miscibility. When miscibility is lost, the process reverts to a water flood. Certain inherent benefits of this technique are apparent. High pressures are not necessary to obtain miscibility in these liquid systems. Furthermore, water is a more desirable driving agent than gas because of the improved mobility ratio and the attendant improvement in areal sweep efficiency. The main disadvantage is, of course, the relatively high cost of alcohols. The use of an alcohol slug to recover reservoir oil is not a new idea." ' However, this process has received only limited study, presumably due to the lack of industrial interest caused by the seemingly prohibitive cost of alcohols. The main purpose of this study was to investigate the mechanism of the displacement of oil and water by various alcohol slugs and to develop process modifications aimed toward possible commercial application.

By J. E. Flinn, S. A. Duran

The steady-state creep behavior of poly crystalline cad mi inn was studied over a temperature range of (1.56 to 0.94 Tm. Two distinct mechanisms were found to occur over this temperature range. They were described by: where and represerqt the minimum strain rates corresponding to the low- and high-temperature regions, respectirely. The two regions of constant acti11ation energy were connected by a transition region where the strain rate was controlled by both mechanisms acting in parallel. At temperatures below a transition temperature of about 0.7 Tm the agreement between the activation energy value for creep and that for self-diffiision suggests a rate-controlling mechanism of dislocation climb. For cadwzium, steady-state creep at temperatures above 0. 7 Tm appears to be controlled by another mechanism, perhaps involving the behavior of dislocation jogs. FRENKEL et al.1 studied the high-temperature creep of polycrystalline cadmium and reported an activation energy of 21 kcal per mole for the 0.5 Tm < T < 0.8 Tm range. Based on observations of creep rate at only two temperatures, a value of 22.1 kcal per mole was determined by Medbury. These two investigations were for the purpose of showing agreement between the activation energy for creep and that for self-diffusion, reported3 as 18.2 and 19.1 kcal per mole, respectively, for diffusion parallel and perpendicular to the hexagonal axis. Gilman4 investigated prismatic glide in single crystals of cadmium over a higher-temperature range of 0.72 to 0.93 Tm, and found an activation energy of 29 kcal per mole. He also reported5 an activation energy higher than that of self-diffusion for prismatic glide in zinc single crystals deformed at temperatures near the melting point. This value was in good agreement with those found for an equivalent temperature range by Flinn and Munson6 and by Tegart and sherby7 for polycrystalline zinc. These two independent studies also disclosed at lower temperatures another value of activation energy near that for self-diffusion. It would be expected from the creep results on zinc and single-crystal cadmium that creep studies on polycrystalline cadmium, extended to temperatures near the melting point, might yield an activation-energy value higher than the 22 kcal per mole value found in earlier studies. The purpose of this paper is to report the steady-creep behavior of polycrystalline cadmium over a temperature range of approximately 0.5 to 0.9 Tm EXPERIMENTAL METHOD The cadmium used in this study was obtained in the form of as-cast rods, 0.5 in. diam, through the courtesy of the Bunker Hill Mining Co. The material was of 99.995 pct purity, as determined by spectro-chemical analysis. The creep specimens, which were 0.250 in. diam by 0.400 in. long and annealed at 300°C for 45 min to produce a stable average grain diameter 0.25 mm, were tested in compression using an apparatus similar to that described by Sherby.8 The specimen temperature was controlled to within ±0.5°C with the help of appropriate constant-temperature baths. The applied stress was maintained within 1.0 pct of the desired value by the additions of lead shot at fixed strain increments. No barreling was observed over the strains encountered during testing. Isothermal creep tests9 were used in the study with only a few differential temperature tests10 run for comparison purposes. Steady-state creep data were obtained over a temperature range of 60 to 287°C (0.56 to 0.94 Tm) at five stress levels ranging from 28.1 to 140.6 kg per sq cm. RESULTS The minimum or steady-state creep rate may be described by an equation of the following form:" where i is the minimum strain rate, S is the structure factor, F is a stress function, Qc is the energy of activation, T is the absolute temperature, and R is the gas constant. The minimum strain rates obtained in this study for cadmium were recorded on a semilogarithm plot as a function of the reciprocal absolute temperatures for the various stress levels, as shown in Fig. 1. This figure shows a characteristic transitional behavior" with a parallel interaction of two mechanisms. It is obvious that the activation energies corresponding to the individual processes are insensitive to stress because the curves are parallel. The discrete activation-energies values for the low- and high-temperature regions for the various stress levels are reported in Table I, and were determined by the least-mean-square method. For the low-temperature region, an activation energy of 20.7 ± 0.6 kcal per mole was obtained, and for the

Jan 1, 1967

By M. C. Chang, J. Dasher

The Crucible mine in Pennsylvania, operating on Pittsburgh seam coal, is rated at 5000 tpd. The washing plant, built in 1943, is rated at about 400 tph, using hydroseparator boxes to wash the coarse coal (4 x 5/16 in.) and hydrotator units for the fine coal (5/16 x 0). It is generally acknowledged that gravity concentration equipment cannot be used efficiently to bene-ficiate coal finer than about 65 mesh. Before the present flotation circuit was installed, Crucible recovered its fine coal slurry, predominantly —24 mesh material, by passing it through a battery of five 14-in. and one hundred and ten 3-in. Hey1 & Patterson cyclones in parallel with a 75-ft Dorr thickener. About two thirds went to the cyclones and one third to the thickener. Underflows from cyclones and thickener were filtered, respectively, in one 12-disk Eimco and two 4-disk Oliver filters fitted with stainless steel cloth. The resulting cakes were added to the coarse metallurgical coal. The filtrates, overflow from the 3-in. cyclones, and thickener overflow were recycled to the coarse coal plant. The three major problems encountered in this system became acute when the mine switched from track-mounted machinery to the more economical off-track equipment. The quality and uniformity of product declined, too much slime accumulated in the circuit, and waste pond disposal became highly impracticable. Lowering of Quality and Uniformity: Since the filter cake was not beneficiated in any way, addition of this material, which contained about 18 pct ash. would certainly lower the quality of metallurgical coal. Difficulties of handling the high-ash, high-moisture cake were compounded by the lack of adequate blending facilities at the mine, making it impossible to maintain uniform feed for the coke ovens. For a two-month test period, to verify the source of trouble, all filter cake was excluded from the metallurgical coal. The excluded filter cake was stocked in the yard. Slime Buildup: The moisture retained in the coarse coal products and the filter cake provided daily the only means of slime bleed from the circuit. As this bleed was not enough to offset the incoming fines in the raw coal feed, there was excessive buildup of slimes. To operate at all, it was necessary to purge the system each weekend by pumping all the solids from the thickener to a waste pond. The slimes built up during the week progressively increased the apparent viscosity and density of the circulating plant water, so that hydroseparators and hydrotator units had to be constantly adjusted. Uniformity, yield, and quality of the metallurgical coal suffered in consequence, and the filter cake ash and moisture contents increased from Monday to Friday. Limited Waste Disposal: Location of the plant does not permit unlimited waste pond area. Maintenance of a waste pond for the purpose of emptying the thickener during weekends became a costly burden that could not be tolerated. HOW THE PROBLEMS WERE SOLVED The flowsheet (Fig. 1) chosen to overcome these difficulties is considered the one most economical to install and operate. In a preliminary change made before the flotation circuit was installed, the cyclones and thickener were first operated in series, the overflow from the 3-in. cyclones, after flocculation, being discharged into the thickener and the thick-

Jan 1, 1960

By J. G. Hadley

IN THE early stages of the depression the Utah Copper Co. realized that an unemployment problem would he created which demanded an intelligent and sympathetic solution. The company recognized that as many men as possible must be kept at work on a basis which would ultimately permit of the most efficient reabsorption of displaced labor. The company therefore determined to distribute available work among the greatest number of men and to that end adopted a system of rotating employees on a part-time basis. An employer requires a certain number of days&apos; or shifts&apos; work during the month. Obviously one employee working thirty days or shifts a month is equal to two employees each working fifteen. In principle, therefore, where two employees equally fitted to do certain work are available the work could be divided between them on an equitable basis and twice as many persons would benefit. A rotation plan on this basis was adopted by the Utah Copper, Co. on Jan. 30, 1930. Some variations of the orig¬inal plan have been made from time to time, the reasons for which will be discussed.

Jan 1, 1932

By Glenn C. Waterman

THE diamond drill is a very important tool in exploration and development testing and its use is increasing. In almost all cases results of diamond drilling are analyzed on the basis of grade and tons. A proper evaluation of core and sludge assays is important if drilling results are to be acceptable as a basis for geologic and engineering appraisal. The relatively wide variation in assay averages as calculated by various well-known combining methods indicates that the engineering choice of a method may affect the outcome of the drilling in terms of ore and waste. The problem of combining assay results from core and sludge samples has been discussed many times in conference and in the literature.&apos;" Most writers agree that the field of disagreement in methods is large and that the engineer on the job must consider features unique to his drilling, pick one of several combining methods, and depart from the rules when abnormal results come in. All the discussion to date can be summed up by the admission that as yet there is no fairly simple, generally acceptable combining method that is practicable over a wide range of drilling conditions, ground conditions, and ore occurrence.

Jan 1, 1955

By Douglas Bunting

In the anthracite fields of Pennsylvania, mine fires occur with more or less regularity and their existence is an ever-present hazard in coal mining. In all probability 90 per cent. of the mine fires can be ascribed directly or indirectly to the ordinary miner&apos;s open lamp. Other causes may be smoking, electrical installations, gas explosions, gas feeders, and the communication of fire to the outcrops from ash dumps, culm banks, timber and brush fires. Fires in pump rooms, engine rooms, and barns have been largely eliminated as a result of the Act of Assembly, approved June 15, 1911, which provides that all inside buildings, including engine rooms, pump rooms, barns, etc., shall be constructed of incombustible materials. Smoking of pipes, cigars, and cigarettes should be prohibited in the mines and too much care cannot be exercised in preventing fires by enforcing all rules designed for that purpose. Mine fires that involve underground fighting are invariably dangerous, and all fires, irrespective of the method pursued in extinguishing them, involve property loss and expense. Each fire generally presents a more or less original problem, dependent on the particular conditions. Generally mine fires may be divided into two classes: Those occurring in inaccessible workings and those occurring in accessible workings. The methods pursued in extinguishing or controlling mine fires are as follows: (I) direct attack, (2) smothering by sealing, (3) flushing with silt or other solids, (4) flooding with water, (5) cutting off, (6) surrounding with incombustible materials, (7) digging out by stripping from the surface. Most fires arc extinguished in their incipiency by direct attack and failure to extinguish by this method is due to late discovery, lack of water, and facilities for fighting, or improper procedure in the initial attack. Should the method first pursued be unsuccessful, careful consideration should be given in determining next the method to be adopted; then the work should be carried on as rapidly as possible and under the supervision of the most competent persons. Materials and supplies should be furnished without delay and in sufficient quantities for the work to be accomplished.

Jan 1, 1922

By H. R. Hose

The total crude fluorspar reserves in Western Europe, including the USSR and the Soviet sphere, Thetotalare estimated to exceed 17 million metric tons. The total fluorspar production in 1951 in Western Europe, excluding the USSR and Soviet sphere, amounted to 371,109 metric tons. The chief fluorspar-producing countries in Europe, excluding Russia and the Soviet sphere, are France, Western Germany, Italy, Spain, and the United Kingdom. TOTAL crude fluorspar reserves in western Europe, containing more than 35 pct CaF2, are estimated at 12 million metric tons, while reserves in the USSR and Soviet sphere may exceed 5 million metric tons. The bulk of these reserves are concentrated in the USSR, Spain, United Kingdom, Western Germany, and Italy. Since published estimates of the tonnages of individual mines and explored deposits are not available for most countries in Europe, the estimates given in Table I for individual countries are necessarily approximate, as they have been compiled from various sources and modified and combined according to the author&apos;s judgment. The majority of fluorspar deposits in Europe belong to the late Paleozoic metallogenetic epochs and have been found within or along the borders of the Hercynian massifs. Deposits in the Central and Esterel massifs of France, those of Western and Eastern Germany, Sardinia, the Catalonian coastal range and the Sierra de los Santos in Spain, and those of Devonshire and Cornwall belong to the Hercynian metallogenetic epochs. The deposits in southern Norway and Sweden belong to the middle Paleozoic or later, those of northern England to the late Paleozoic or later, those of northern Spain and Greece to the late Mesozoic, and those of northern and probably central Italy and Switzerland to the early Tertiary. Fluorspar occurs typically in veins, lenses, and stringers, but also in bedded replacement deposits and pockets, and may be colorless to white, yellow, green, pink, brown, blue, or purple to almost black. Except for the colorless variety, the only kind which can be used for the optical industry, the color of the fluorspar does not affect its ultimate use. Crystalline cubic or octahedral fluorite of CaF2 composition, with 48.9 pct fluorine and 51.1 pct calcium, has a specific gravity of 3.18. The ore in place, however, usually averages 3.0 metric tons per cubic meter owing to the quantity of other minerals asso- ciated with the fluorspar in the veins and the physical state of the fluorspar, which may vary from sugary to massive crystalline or microscopic inter-growths with barite or quartz. Barite of 4.5 sp gr, quartz of 2.66 sp gr, and chert and calcite are commonly associated with fluorspar. Other minerals usually present in lesser amounts are celestite, wad, silver, lead, zinc, copper, iron, and uranium minerals. In some cavity fillings the rhythmic deposition of the fluorspar, barite, and quartz into successive crusts may make it possible to handpick the crude ore on belts or to separate the fluorspar from other minerals while the actual mining is carried on. Where metallurgical grade lump spar is required for foundries, separation of quartz and barite may be made also by jigs and heavy media, and in some mines the lead and zinc sulphides may be separated and recovered by vibrating tables. However, except in the case of unusually high grade deposits, acid grade fluorspar is obtained by flotation. By this process fluorspar deposits grading down to 35 pct CaF2 in the crude ore can be worked at a profit when fluorspar, lead, zinc, and silver prices are high, as the sulphides of these ores can be separated from the fluorspar by flotation and concentrated.

Jan 1, 1956

By B. C. Raynes

IN the early part of 1952, under the auspices of the U. S. Atomic Energy Commission, Horizons Inc. undertook an investigation dealing with the preparation of high purity thorium metal in order to develop an economically competitive production process for this metal based on a recovery of the metal by a fused salt electrolysis, since a survey of the literature of thorium chemistry indicated that electrolytic methods had not received much attention, despite the consideration that electrochemical production methods often offer attractive commercial potentialities. Thorium metal has been prepared by methods which fall into four general categories: 1) reduction of halides or double halides; 2) reduction of the oxide; 3) thermal decomposition of halides; and 4) electrolytic methods. Some of the important results obtained by various investigators since about 1900 in recorded attempts to prepare pure thorium metal are summarized in Table I. A bibliography of pertinent literature is included in this paper. Thorium is strongly electropositive and possesses a great affinity for oxygen and nitrogen. As a result the prospects for the production of high purity thorium from aqueous media or oxygen bearing compounds are remote. The most promising and profitable area of investigation for production of thorium metal lies in fused salt electrolysis. Thorium has been produced previously by this technique from halide salts in fused alkali chloride eutectic mixtures. Until recently, however, the lack of known large uses for thorium has inhibited the investigation of these methods. The basic problem in an investigation into fused salt production methods for thorium is to develop a suitable fused salt system for the economic preparation of thorium. The first portion of the work in these laboratories was a small scale investigation of the electrolysis of potassium thorium fluoride, KThF5, under a variety of conditions. High purity thorium was prepared in this manner, but the process did not appear to offer the simplest and most attractive continuous commercial operation. The work reported here relates to an all-chloride system employing thorium chloride as the cell feed. ThCL4 was produced as an anhydrous material stabilized in molten sodium chloride. This product, further diluted in NaCl, was subjected to electrolysis. The electrolytic process has been studied in cells capable of holding up to 50 lb total salt charge. Ductile, pure thorium metal has been reproducibly obtained in this operation. Considerable experimentation was carried out in the preparation of ThCL as a cell feed free from moisture, oxides, and other impurities. Thorium was cathodically deposited as a coarsely crystalline powder interspersed with residual bath salts. Aqueous procedures for separating and recovering the metal were developed. The thorium metal powder obtained has been evaluated with respect to its hardness, chemical purity, and certain other physical properties. Preparation of ThC1,—Thorium chloride, and the chlorides of some other metals (aluminum, beryllium, tantalum, zirconium, etc.) cannot be obtained from aqueous solution by direct heating and evaporation. Evaporation and heating of a solution of thorium chloride yields a hydrated salt or partially hydrated salt, and finally Tho,. Anhydrous thorium chloride is obtained by forming a hydrated ammonium chloride complex salt in acidified water solution, dehydrating this compound, and finally decomposing the double salt and subliming the ammonium chloride component. Various methods for the preparation of anhydrous thorium chloride, in accordance with this concept, were investigated as indicated in Table I1 from hydrated thorium chloride (ThCl, . 3H2O), and from thorium oxycarbonate (ThOCO,). In the method found most satisfactory for this work, thorium oxycarbonate was dissolved in an excess of concentrated HC1 to give a clear yellow solution. At least 2 mol NH,Cl were added to the solution per mol of contained thorium. Ammonium thorium chloride hydrate was crystallized from solution by cooling and saturation with HC1. The resultant complex, (NH,Cl),ThCl;xH,O, was dehydrated by evaporation at 130° to 150°C and thoroughly dried at about 250°C. The dried ammanium thorium chloride was then mixed with sodium chloride in proportions to form NaThCL or to produce a ThCL4-NaCl mixture with a given thorium content. The ammonium chloride was sublimed off at temperatures above 500°C in an inert atmosphere furnace to give an anhydrous chloride source electrolyte, essentially free of Tho, and of NKC1.

Jan 1, 1958

By J. Geertsma

The relationship between porosity and the speed of propagation of acoustic waves in fluid-saturated porous rocks as measured by the Sonic log and by ultrasonic tecbniques is analyzed. Biot&apos;s continuum theory 1 is used to explain the difference in acoustic wave propagation between a dry and a liquid-saturated porous material. The porosity is a variable in this theory. However, the acoustic wave propagation in the dry rock depends too on porosity, and this dependence is not predicted by the theory. Frequently in dry sandstones, a nearly linear relationship between reciprocal acoustic wave velocity and porosity is observed in the low-porosity range. The physics behind this behavior is outlined. An empirical relationship of the form, 1/v = A + F, applies accordingly for many porous dry rocks, provided the porosity is the only variable. The presence of a liquid in the pores changes the value of B, and this change is found to be in agreement with the Biot theory. The time-average relation introduced some years ago2 results in an equation of the same type — 1/V = F/Vf + (1 - F)/Vr — but is not based on a sound physical picture. Still, this relation sometimes predicts approximately correct A and B values. Carbonate rocks with their complicated pore structures sometimes show a different relationship between wave velocity and porosity, unfavorable for log interpretation. Examples are presented. The simultaneous presence of calcite, dolomite and anhydrite, with their different grain densities and matrix compressibilities, complicates acoustic-log interpretation in carbonate rocks still further. Other complicating effects of acoustic-log interpretation are discussed. They are related to the influence of shale streaks and natural fractures on the average wave velocity observed by the logging tool and to the effect of adsorption phenomena on wave propagation in unstressed rocks particularly in sandstones. INTRODUCTION The velocity of propagation of sound waves in porous sediments as a function of depth is a quantity frequently measured. Velocity loggers of various d-sign may be used, for which velocity is defined as the. distance between a wave generator and a wave detector (one-receiver system) or the distance between two wave detectors (two-receiver system), divided by the shortest time required for a vibrational pulse to cross this distance. Velocity loggers ale used to assist in seismic prospecting, to differentiate between the various types of sedimentary rock layers, and also to determine rock porosity and pore fluid content. This paper is especially concerned with the relation between velocity and porosity and the effect of the pore fluid. Most velocity-log interpretation in terms of porosity is based on a formula advocated by Wyllie, et al,2 suggested earlier by Hughes and Jones,3 and known as the "time-average relation". This formula app1ies for a model consisting of a layered system of parallel alternating slices of a solid and a liquid, crossed by the wave path perpendicular to the interfacrbs. This must be an unsuitable model for the derivation of wave propagation properties of liquid-saturated porous media. It suggests that only rock matrix and fluid properties influence the wave velocity. The surprising fact, nevertheless, is that applications of this formula to clean sandstones under specific conditions (sufficient effective stress) results in porosity predictions which are close to reality. Other authors, such as Gassmann,4 White and Sengbush,5 Brandt,6 and Hicks and Berry,7 make use of more realistic models to arrive at a wave velocity formula for porous sediments. Their models consist of various liquid-saturated packings of frictionless spherical grains and are, therefore, also highly idealized. A number of experimental results obtained in the laboratory can be better explained with these model theories than by the time-average relation. A study of a general character, not depending a priori on a special model of the porous structure, was still lacking.

By Hans Allenius, R. T. Hukki

This paper describes in quantitative terms the effect of sharpness of classification on the performance of the closed grinding circuit. The analysis is based on a large number of laboratory experiments designed to simulate the two most common industrial closed grinding systems. The experimental results show quantitatively the degree of improvement achievable by more efficient classification. In an earlier paper1 the senior author has presented a trend showing qualitative analysis of mill and classifier performance in tile closed grinding circuit. According to this analysis, the master key to a major improvement appeared to be effective removal of finished fine material from the classifier sands, or in other words, improved sharpness of classification, leading simultaneously to a substantial reduction of the circulating load. One purpose of this paper is to present the results of a quantitative investigation of the closed grinding circuit. Another is to set forth the essential pertinent variables revealed by the quantitative experiments carried out by the junior author. In the two most common systems of closed-circuit grinding, (1) feed to the circuit is introduced into the grinding unit, normally the ball mill, operating in closed circuit with the classifier; or (2) feed to the grinding circuit is introduced into the primary grinding unit, normally the rod mill, discharging into the classifier in closed circuit with the secondary grinding unit, normally the ball mill. In the following discussion, these two systems are analyzed separately in the indicated order. BALL MILL - CLASSIFIER CIRCUIT The Test Apparatus and Procedure: The test appara- tus included: (1) a F 195 mm x 220 mm laboratory ball mill rotated at a speed 77% of the theoretical critical speed, and a 7-kg batch of F 20 mm-F 50 mm steel balls; (2) a conventional 65-mesh test sieve and a Ro-Tap sieve shaker serving as a classifier; and (3) a Permaran instrument manufactured by Outo-kumpu Co. for surface area determinations by the permeability method. As no continuous closed-circuit experiments could be brought about on a laboratory scale, the process was broken down into alternating grinding and sizing steps in such a way that the new feed plus the returning sands always formed a batch of 1000 gm. For each selected grinding period and for each selected sharpness of classification the basic steps were repeated six times. Steady-state conditions were then reached. Crystalline vein quartz was used as a test material. Feed to the process consisted of -10-mesh fraction of this quartz crushed in rolls. This fraction included 15% of-65-mesh material. Experimental Results: The general flowsheet is shown in Fig. 1. Table I gives the essential data obtained representing steady-state conditions under the indicated set of variables. The table is based on 110 grinding experiments, 440 screen analyses and an equal number of specific surface area determinations. Fig. 2 shows the relationship between the cumulative net energy consumption and the cumulative number of mill rotations as evaluated by separate experiments. Note that this relationship is not linear but is instead represented by a slight curve. The data given in Table I are presented in graphical form as follows: Fig. 3 shows the produced -65-mesh material in grams per minute vs. time of grinding in minutes; the sharpnesses of classification at 65 mesh were 100%, 75% and 50%. It is clearly indicated that the highest-capacity figures call for relatively short grinding times and for the sharpest possible classification. Fig. 4 presents the specific surface area on the final -65-mesh product in square centimeters per gm vs. time of grinding in minutes. In conventional mineral dressing processes, a final product characterized

Jan 1, 1969

By R. A. Wagstaff

SINCE the time of the early Egyptians, the use of copper has been a boon to the life of most of the civilized world. Its use has been varied; in many connections, the art by which it attained its greatest usefulness has been lost. In this paper an endeavor will be made to take into consideration some of the many changes that have occurred during the last 40 years of copper-smelting operations. During that period the copper blast furnace has been almost forced out of existence by new methods of ore concentration-first, gravity concentration, and then flotation. These new methods of concentration introduced many new factors into the metallurgy of copper. The product was wet, it was fine and it was high grade. To overcome all these factors many changes have been necessary throughout the smelting agenda. New equipment has been necessary for transportation, unloading, sampling, bedding and moisturing in all smelting plants, especially the custom smelter. During the early days of "flotation," little was known about filtering or the use of an alkaline circuit; often the concentrate going to the smelter contained 20 to 25 per cent moisture. It is easy to imagine the difficulty arising with a product of this nature when it had to be shoveled by hand labor. To meet these conditions, mechanical unloading and sampling devices were soon developed. In many instances such good progress was made that actual costs for handling this unfavorable product were lower than those formerly incurred on the old product. Today most sampling of copper ores and concentrates is being done by mechanical sampling, pipe or augur sampling, or shovel sampling. Where concentrates are uniform in metal content, a great many of the plants now use pipe or augur sampling. For spotty ores, the old shovel sample with cone and quartering during the final stages is still in use. Where big tonnages are handled, mechanical sampling of concentrates is being done, with satisfactory results. To curtail slow and tedious cutting and mixing procedure, mechanical mixing boxes, splitters and cutters are used to cut down labor costs. MILLING As crude ores become scarcer, the use of milling operations increases. The few crude ores that are still smelted are used as fluxing types. To use these ores finer grinding has been found advantageous. This has made for better bedding and fluxing conditions, which in turn makes for faster smelting, lower fuel consumption and better metal recovery. The degree of fineness is more or less determined by the nature of the product. Most plants have set a standard of through ¾ -in. mesh screen for roaster- reverberatory smelting and through ¾ in. mesh screen for converter fluxes. For some refractory ores; even finer grinding has been found to be advantageous for good reverberatory smelting.

Jan 1, 1944

By William M. Newton, John E. Sherborne

The relation between the properties of electric logs and the fluid used in drilling a well is briefly discussed in this paper, and the fact that the resistivity of the drilling mud and its filtrate can appreciably affect such logs is shown. Apparatus was designed and a procedure established to determine the effect of various factors on the resistivity of drilling mud. Five muds commonly employed in California were tested and the following observations were made: I. The effect of raising the temperature from 80°F. to 180°F. is to decrease the resis-tiviiy of the mud or filtrate approximately 50 per cent. 2. The resistivity of the mud in most cases closely approximates that of its filtrate. 3, The change in the resistivities of muds caused by the addition of chemical is not the same function of the amount of chemical added for each mud. 4. The effect produced by increased sodium chloride content is to reduce markedly the resistivity of the mud and its filtrate. 5. Weighting materials, such as Baroid and limestone, tend to increase the resistivity of drilling mud. 6, Cement and counteracting reagents reduce mud resistivity. Introduction During recent years, the practice of "electrically logging9&apos; a geologic formation by means of a system of electrodes lowered in a well bore has provided data of inestimable value concerning the nature of the sediments penetrated. Thus, it is generally possible, by means of electric logs, to distinguish readily the differences between shales, sandstones, cherts, limestones, and other rock forms. The distinguishing characteristics recorded in this manner have made possible long-range and closely detailed mapping and visualization of deeply buried structures1-5 to a degree not to be obtained with a reasonable expenditure of money by any other means at present in use. Besides being of value in correlation work, electric logs have been found, in some instances, to be useful in estimating the nature of fluid contained within the pores of a rock. In special cases, even the relative permeabilities at different horizons have been determined.6 That electric logs should reflect in some measure such physical properties of rocks as porosity and permeability and, in addition, give rise to data pertaining to the fluid content of rock interstices, becomes evident when the conditions giving rise to the electrical phenomena measured are known. However, it is not the purpose of this paper to review the present status of electric logging; rather, the reader is referred to the papers cited on page 220 for the principles and practice involved. Importance of Drilling FLuid in Electric Logging Because the drilling fluid must be the medium through which the electric logging device is made to traverse the formations penetrated by a borehole, and since, in general, the drilling mud completely surrounds

Jan 1, 1942

By John E. Sherborne, William M. Newton

The relation between the properties of electric logs and the fluid used in drilling a well is briefly discussed in this paper, and the fact that the resistivity of the drilling mud and its filtrate can appreciably affect such logs is shown. Apparatus was designed and a procedure established to determine the effect of various factors on the resistivity of drilling mud. Five muds commonly employed in California were tested and the following observations were made: I. The effect of raising the temperature from 80°F. to 180°F. is to decrease the resis-tiviiy of the mud or filtrate approximately 50 per cent. 2. The resistivity of the mud in most cases closely approximates that of its filtrate. 3, The change in the resistivities of muds caused by the addition of chemical is not the same function of the amount of chemical added for each mud. 4. The effect produced by increased sodium chloride content is to reduce markedly the resistivity of the mud and its filtrate. 5. Weighting materials, such as Baroid and limestone, tend to increase the resistivity of drilling mud. 6, Cement and counteracting reagents reduce mud resistivity. Introduction During recent years, the practice of "electrically logging9&apos; a geologic formation by means of a system of electrodes lowered in a well bore has provided data of inestimable value concerning the nature of the sediments penetrated. Thus, it is generally possible, by means of electric logs, to distinguish readily the differences between shales, sandstones, cherts, limestones, and other rock forms. The distinguishing characteristics recorded in this manner have made possible long-range and closely detailed mapping and visualization of deeply buried structures1-5 to a degree not to be obtained with a reasonable expenditure of money by any other means at present in use. Besides being of value in correlation work, electric logs have been found, in some instances, to be useful in estimating the nature of fluid contained within the pores of a rock. In special cases, even the relative permeabilities at different horizons have been determined.6 That electric logs should reflect in some measure such physical properties of rocks as porosity and permeability and, in addition, give rise to data pertaining to the fluid content of rock interstices, becomes evident when the conditions giving rise to the electrical phenomena measured are known. However, it is not the purpose of this paper to review the present status of electric logging; rather, the reader is referred to the papers cited on page 220 for the principles and practice involved. Importance of Drilling FLuid in Electric Logging Because the drilling fluid must be the medium through which the electric logging device is made to traverse the formations penetrated by a borehole, and since, in general, the drilling mud completely surrounds

Jan 1, 1942

By E. F. Ketterer, D. B. Collyer, A. B. Wilder

AISI constructional alloy steels are a series of steels in which a numbering system signifies the chemical composition limits. SAE (Society of Automotive Engineers) and AISI (American Iron and Steel Institute) use basically the same numbering system. The steels contain, in part, alloy additions of chromium, molybdenum, nickel, and vanadium. They are used widely in the automotive industry but have not been adopted for high temperature service. The purpose of this investigation was to study changes in microstructure observed in typical AISI constructional alloy steels in the welded and un-welded conditions after exposure for 34,000 hr at 900" and 1050" and, in several instances, 1200°F. The tensile and creep rupture properties after exposure for 10,000 hr were also evaluated. The investigation is a continuation of the work previously discussed. Little was known concerning the existence of graphite in medium or low carbon steels before 1943. It is important to note that continued exposure at elevated temperature is required in order to develop graphite in steels. Interest in the presence of graphite in steel, and also in changes in the properties of steels after long service at elevated temperatures, developed after 1943 when a failure of a welded joint of high temperature piping was associated with the graphitization of a relatively low carbon content steel with 0.50 pct Mo,8,0 Numerous investigations have beeen made in recent years studying the relationship of this type of failure to changes in properties. Attempts have been made to evaluate the performance of a number of steels, particularly in the heat-affected zone of welds. It was the unfortunate failure of a weld-heated zone in high temperature piping in the steam-power industry which focused attention upon possible failures, and upon the need for improved steels for high temperature piping service. Investigation by the National Tube Div., U. S. Steel Corp., Lorain Works, has included a wide variety of steels in an attempt to learn the effects of numerous factors and elements upon the long time performance of steels, although many of the compositions studied will never be used at elevated temperatures.1,4-7 Included were the well established AISI steel series, which offered opportunity to test moderately high carbon content steels as well as observe the effect of nickel, chromium, molybdenum, etc., upon the properties of these well known steels. The properties of the AISI steels were also studied for comparison with the previously reported properties of the Cr-Mo steels, now used by the steam-power industry.5,7 Materials and Testing Procedure The AISI alloy steels in Table I were exposed for 10,000 and 34,000 hr at 900°, 1050°, and 1200°F.

Jan 1, 1958

By E. T. Turkdogan, J. H. Swisher

The solubility of oxygen in zone-refined iron was determined in the temperature range.from 881" to 1350°C. The solubility in a iron at 881°C ms found to be about 2 to 3 ppm; in y iron, the solubility was found to increase from about 2 to 3 ppm at 950°C to about 25 ppm at 1350°C. The permeability of oxygen in an iron -0.1 pct Al alloy was determined in the y iron range, using an internal oxidation technique. By combining the permeability and solubility data, the diffzisivity of oxygen in y and a iron was calculated. The oxygen diffusivity in solid iron may be s~immarized as follows: 8820 For 6 iron and IN a recent paper, Hepworth, Smith, and Turkdogan1 reported on the solubility, permeability, and diffusivity of oxygen in 6 iron, together with permeability data for oxygen in a iron. In the present investigation, similar measurements were made in the ? phase region. In addition, a solubility measurement was performed in the a phase region to permit calculation of the diffusivity of oxygen in a, iron. References to early work on this subject are given in the previous publication.&apos; EXPERIMENTAL Solubility Measurements. The oxygen solubility was determined by equilibrating cylindrical samples 0.3 in. diam by 14 in. long of zone-refined iron in water vapor-hydrogen gas mixtures. The zone-refined iron was prepared by B. F. Oliver of this laboratory, using an apparatus and technique described elsewhere.2&apos;3 Six zone-melting passes were used to achieve a total impurity level of about 30 ppm. Of this 30 ppm, the combined nickel and cobalt content was about 20 ppm, oxygen was 4 ppm, and all oxidizable impurities less than 1 pprn each. A vertical resistance furnace wound with molybdenum wire was used for the experiments. The temperature was measured before and after each experiment with a Pt/Pt-10 pct Rh thermocouple. In the gas train, flow rates of hydrogen and argon were measured with capillary flow meters, and the resulting mixtures were passed through a column containing 90 pct oxalic acid dihydrate and 10 pct anhydrous oxalic acid to obtain predetermined ratios of H2O to H2. The vapor pressure of H2O above this mixture as a function of temperature is well-known.4 The exit gas was ana- lyzed periodically for H2O, and good agreement (+3 pct) with the calculated composition was obtained. The zone-refined iron specimens were held in the furnace for a sufficient length of time for equilibration, e.g., 18 hr at 1350°C and 1 week at 881°C, then quenched in a brine solution. After removing the surface oxide from the samples by machining, duplicate analyses were obtained by a combined vacuum fusion-infrared method; oxygen analysis was reproducible within 2 ppm. Permeability Measurements. The permeability of oxygen in alloys containing about 0.1 pct A1 was determined by internal oxidation and measurement of the subscale thickness as a function of time. The experimental alloys were prepared by adding aluminum to electrolytic iron (grade 104A plastiron) that had previously been vacuum-carbon deoxidized. The resulting ingots contained about 20 ppm O, 100 pprn C, 40 pprn Si, 50 ppm Mo, 20 ppm P, 20 ppm S, and 20 pprn Zr as the principal impurities. After hot rolling the ingots to 1-in.-thick slabs, specimens were machined from the stock in the form of rectangular plates, 4 by 5 by 2 in. The general procedure for the permeability experiments was the same as those for the solubility measurements. The specimens were cooled in a reducing atmosphere rather than quenched, however, in order to maintain a clean surface for measurement of the sub-scale thickness. This measurement was made on a polished cross section of each specimen, using a microscope with a micrometer stage. The inclusions formed at the lowest temperature 1033°C were too small to be seen with an optical microscope. An electron micrograph showing the size and shape of individual particles is shown in Fig. 1. The dark band in the picture is a boundary between two subgrains. The subscale thickness in these samples was measured with an optical microscope after heavily etching in 2 pct nital, which gave contrast between the subscale and the unoxidized zone.

Jan 1, 1968

By H. I. Aaronson, C. Laird

The kinetics of thickening and of lengthening of ?&apos; plates in an Al-3.93 pct Cu alloy in the temperature range 203" to 300" C were determined by means of transmission electron microscopy. The rate of thickening was found to be less than that allowed by volume diffusion control at all temperatures, by amounts which increased with decreasing temperature, in agreement with the predictions of a general theory of precipitate morphology.1 Thickening was treated on the basis of the ledge mechanism. Ledges were deduced to spread across the broad faces of ?&apos; plates at volume dzjrfusion-controlled rates, as anticipated from the disordered structure of their edges. Lengthening of 8&apos; plates, on the other hand, took Place more rapidly than allowed by volume dzjrfusion. This occurred despite clear morPhological evidence of a bmrier to growth at the edges of these plates. It was concluded that the misfit dislocation structure comprising the barrier requires that lengthening take place by a jog mechanism. The tnisfit dislocations, however, also serve as diffusion short circuits, and allow high overall lengthening rates to be achieved. In Part I&apos; it was shown that, within the range of aging temperatures and times studied, the broad faces of 8&apos; plates formed in Al-4 pct Cu are fully coherent with the a, matrix. Virtually .all of the dislocations present in these faces were found to have developed as a result of plastic deformation in the a phase. Such dislocations are thus "intruders", rather than the more usual misfit-compensating variety. The edges of 8&apos; plates were confirmed, by extension of the earlier studies of Mat-suura and Koda,3 to be made up of edge-type misfit dislocations, in sessile orientation with respect to lengthening of the plates. These interfacial structures should cause 8&apos; plates to thicken and to lengthen at rates less than those allowed under the condition of volume diffusion control, such as would be expected if the interphase boundaries had disordered structures.&apos; The narrow width of 8&apos; plates, the reproducible crystallography of their broad faces, and the appearance of these plates in cross section as octagons rather than as circular discs2 provide qualitative support for these deductions. The present study of the rate of thickening and the rate of lengthening of 8&apos; plates was undertaken in order to examine them on a quantitative basis. I) THICKENING KINETICS OF THE BROAD FACES OF?&apos; PLATES A) Literature Review. The measurements now available on the thickening kinetics of single-phase precipitate plates consist of one plot of the thickening of a proeutectoid ferrite plate in an Fe-C alloy,&apos; showing (as predicted) thickening rates less than those allowed by volume diffusion control. B) Experimental Procedure. Details of the preparation of the 4-3.93 pct Cu alloy used in this study have been previously reported.4 As in Part 1,&apos; transmission electron microscopy was the observational tool employed. A general description of the apparatus and procedures of the electron microscopy studies is given in Section I of Part I. In thin foils, 0&apos; plates tend to form at and parallel to the foil surface.&apos; A direct investigation of the thickening process by means of hot-stage transmission electron microscopy was therefore not feasible. It was thus necessary to use the conventional method of aging individual bulk specimens for a wide range of different times at the various temperatures studied. In each specimen, the thicknesses of a number of plates were measured. Since thin foils prepared from "bulk-aged" material contain a large proportion of grains with orientations near (001) , it was relatively easy to find, near the edges of the foils, the characteristic multifold patterns of intersecting extinction contours which indicate regions where the foil is exactly at an (001) orientation. The thicknesses of large numbers of plates were measured along the (200) branches of the "stars" so that the 8&apos; plates were precisely parallel to the optical axis of the microscope. Wherever possible, intersecting extinction contours were adjusted with the parameter s > 0 to improve the visibility of the plates in bright-field illumination. These precautions, in combination with taking the measurements at the thinnest parts of the foils, minimized the errors in the measurement of the thickness of the plates resulting from inexact parallelism to the electron beam. Since the plates were very thin, it was not easy to measure their thickness on the photographs. The techniques of enlargement and of microdensitometry were employed to minimize errors from this part of the measurement. A further source of possible error, that the plates can appear thicker because of contrast associated with mismatch normal to the plane of the plate, was also considered. The images of the plates were usually thinner than those of dislocations, however, and no anomalous changes in apparent plate thickness were observed when regions of foil containing plates were tilted through various diffracting conditions. Any error from this cause must therefore be small. Other sources contributing errors were: a) the microdensitometer traces per se and the subjective estimates of their peak limits, and b) slight fluctuations in magnification associated with small changes in the current of the objective lens of the electron microscope. The overall error probably amounted to no more than 5 to 10 pct. In order to obtain readily interpretable data on thickening kinetics, it is essential that the diffusion fields of adjacent ?&apos; plates not be allowed to overlap. Calculations&apos;-&apos; showed that this condition is definitely not ful-

Jan 1, 1969