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By J. H. Brophy, S. J. Michalik

A constitution diagram for the system Mo-Ir has been determined. The maximum solubility of iridium in molybdenum is 16 at. pct at 2110ºC and decreases to less than 5 at. pct at 1500°C. The solubility of molybdenum in iridium is 22 at. pct. Three intermediate phases appear in the system: 8 MoJr, having the p-tungsten structure; a phase, a cornplex tetragonal structure; and the hcp ? phase. Metallography, melting point determinations, X-ray diffraction and fluorescence, and electron micro-probe unalyses were employed in establishing the diagram. PREVIOUS to the present investigation, the intermediate phases in the Mo-Ir system were identified, but no detailed account of the phase diagram has been reported in the literature. Raub1 investigated alloys of Mo-Ir over an extensive range of composition between the temperatures of 800º and 1600°C. The in-termetallic compound MosIr was found to exist with nearly pure molybdenum, as the solubility of iridium in molybdenum was not detectable parametrically in this temperature range. MO3Ir was found to be iso-morphic with a ß-tungsten type structure, having a parameter of 4.959Å. An intermediate hcp phase, designated as the ? phase, ranged in composition from 52 to 78.5 at. pct at 800ºC, and from 41 to 78.5 at. pct Ir at 1200°C. Parameters noted for the ? phase were as follows: at 42.7 at. pct Ir, a = 2.771i0, c = 4.4366, c/a = 1.601; at 78.5 at. pct Ir, a = 2.736A, c = 4.378A, c/a = 1.600. Molybdenum was found to be soluble in iridium up to 16.5 at. pct Mo (83.5 at. pct Irj, with the parameter of iridium increasing to 3.845A at the solubility limit. Knapton,2 who investigated alloys between 15 and 85 at. pct Ir, essentially agreed with Raub's data, but, in addition, found a a phase in as-melted alloys near 25 at. pcto Ir. The oaphase lattice parameters were a = 9.64Å, c = 4.96Å, c/a = 0.515. The a phase was replaced by the 8 -tungsten phase on annealing at 1600°C. Knapton concluded that the a was stable only at elevated temperatures, and placed the composition of the a phase at approximately 30 at. pct Ir. The intermetallic compound Mo3Ir, with a lattice parameter of 4.965A, was included among the 8-tungsten structures reported by ~eller.' Matthias and Corenzwit,4 and Bucke15 studied the superconducting nature of MosIr, and reported a superconducting transition temperature of 8.$K. The present investigation describes the phase relationships in the Mo-Ir alloy system determined by melting point measurements, X-ray diffraction and fluorescence, and metallography. EXPERIMENTAL PROCEDURES Alloys for the determination of the phase diagram were prepared from powders. Commercial 99.9 pct Mo from Fansteel Metallurgical Corp. and 99.9 pct Ir powder from J. Bishop and Co. Platinum Works were used. The powders were weighed to nominal compositions, mixed, and then pressed, without binder, into compacts weighing 4 to 6 g. These were presintered in uacuo between 1200' and 1400°C for 1 hr, to reduce the degree of spattering during subsequent arc-melting. The compacts were arc-melted in a nonconsumable tungsten electrode furnace six times on alternate sides on a water-cooled copper hearth in an atmosphere of zirconium-getter ed argon at 500 mm of mercury pressure. In almost all cases, this procedure yielded buttons of satisfactory homogeneity. The composition of all melted buttons were confirmed by X-ray fluorescent analysis using the experimentally determined ratio of the iridium La1 line intensity to that of the molybdenum Ka1 line as a function of composition. In this determination four alloys analyzed by wet chemical methods were used as standards. An uncertainty range of ±1 at. pct has been attributed to all indicated compositions. All heat treatments and solidus measurements were carried out in tantalum resistance heating elements in vacuum conditions of 10-4 to 10-5 mm of mercury. A detailed account of this procedure has been reported by Schwarzkopf and Brophy.8 In the heat treatment and solidus measurements of iridium-rich alloys (50 at. pct Ir or greater), a tungsten lining was inserted into the tantalum resistance heating element because of a eutectic reaction which occurs between iridium and tantalum at 1948ºc.7 Heat treatments and solidus measurements carried out at compositions less than 40 at. pct Ir both with and without tungsten linings within the resistance

Jan 1, 1963

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By Harold A. Krueger

FACILITIES and mining operations of the National Lead Co., St. Louis Smelting and Refining Division, near Fredericktown, Mo., are situated in a famous mining area. Copper, lead, nickel, and cobalt have been mined here for more than 100 years, work having been started on a high sulphide copper outcrop in 1847. Lamotte sandstone is characterized by differential compaction on a rigorously eroded pre-Cambrian surface. The Bonneterre formation was therefore a good host for minerals not generally found in mineable quantities in these midwestern areas. Unusually complex minerals, however, make beneficiation difficult, and because of irregular ore thicknesses and elevations many engineers and operators have not attempted to mine the property. Others have tried who failed. This paper deals with economic, efficient, and competitive methods of mining these highly irregular orebodies, as compared to the open-stope, room-and-pillar methods normally used for horizontal-bedded lead deposits. For the purpose of this study it should be understood that the ore is found in two distinctly different types of occurrences, one to be designated as basin ore and the other as contact ore. Mining of basin ore is complicated by many faults, fractures, cross faults, and breaks. Contact ore is complex because it is found on flanks or slopes of pre-Cambrian knobs or highs. The dip of the mining floor for the latter type varies between 18" and 45". Occurrences of both types of ore are complicated by water courses or solution channels which carry unconsolidated shale, lime, sand, and dolomite. This material is also found between the bedding planes of the members of the Bonneterre formation. The water found where there are fractures, faults, and channels makes it very fluid and tacky, see Fig. 1, particularly after it has been blasted and handled by loading and hauling machines. Much of the ore can be wadded and thrown without dispersing. During early operations by the Buckeye Copper Co. in 1861 and the North American Lead Co. from 1900 to 1910, conventional narrow-gage railroad and side dump mine cars were used with hand shoveling. The complications of mining the contact ore, the only type attempted at this time, can be appreciated when it is realized that operators were obliged to use mules for haulage. Haulageways constructed on these slopes were of necessity similar to wagon trails or goat trails up the side of a mountain. In other words, it was merely a matter of going from side to side of the strike length of the slope, gaining a little in elevation on each shuttle trip. Production totaled only one to two tons per manshift. A few years later, about 1913, the property was purchased by combined Canadian interests known as the Missouri Cobalt Co., and the use of trolley locomotives was initiated. Between 1900 and 1928 a land agent using churn and diamond drilling methods prospected scattered sections of the area. In 1928 the first property was purchased by the present company, then operating as the St. Louis Smelting and Refining Co. Check drilling and prospecting was carried out by the company at various times between 1928 and 1939 to correlate the erratic mineralization. Much information about both types of orebodies was accumulated, but it was still questionable as to whether money should be invested to work these occurrences. In anticipation of high lead and copper prices, about the time World War II started, it was decided to develop and bring into production some of this ore. In 1942 No. 1 shaft was put down on the largest basin-type orebody and in 1943 No. 2 shaft was put down on contact-type ore. Operations were expanded when No. 3 shaft was completed in 1943, and progressed further in 1948, when National Lead Co. dewatered and opened No. 5 and 6 mines, old workings of the North American Lead Co. and the Missouri Cobalt Co. Because of the differential compaction of Lamotte sandstone over the pre-Cambrian porphyry, in some instances mineable thicknesses of basin-type ore occurred 20 to 30 ft above the sand. This is the exception rather than the rule, since most of the mineralization starts at the sand and is variable in thickness. The ore was attacked, therefore, by development drifts and crosscuts at the lowest possible elevation, where the ore immediately overlying the Lamotte sandstone could be drained and made accessible for mining. It was planned to connect to the drifts and crosscuts with raises to mine ore deposited 20 to 30 ft higher. The higher orebodies were thus mined as slusher levels. Slusher hoists were used to drag the ore into the raises, which were made into hoppers. The ore was then loaded into 32x32-in. ore cans, hauled to the shaft by battery locomotives, and hoisted by the conventional Tri-State method. The rate of efficiency was 5 to 6 tons per manshift underground. The contact-type ore was attacked in a similar way, except that the orebodies were not nearly so wide, so that they were more flexible for slusher loading into cans. This advantage was offset, however, by haulage complexities, since the railroad was constructed on steep slopes. Through experience and ingenuity, many improvements were made in mining both types of ores. The two levels, so-called, in the basin-type ore-bodies were connected as previously planned, more efficient locomotives replaced the older ones, and a

Jan 1, 1954

By M. R. Tek, F. F. Craig, J. O. Wilkes, C. S. Goddin

The waterflood performance of a water-wet, stratified system with crossflow is computed by a finite difference procedure. The effects of five dimensionless parameters on tile oil displacement efficiency, water saturation con-tour.7 and crossflow rates are evaluated in the absence of gravity forces. Crossflow due to viscous and capillary forces is shown to exert a significant effect on oil recovery in a field-scale model of a two-layered. water-wet sandstone reservoir. The crossflow is at a maximum in the vicinity of the front advancing in the more permeable layer. Under favorable mobility ratio conditions, the comparted oil recovery with crossflow always is interrnerliate between that predicted for a uniform reservoir and that for a layered reservoir with no crossflow. INTRODUCTION The important erects of reservoir heterogeneity on waterflood performance are commanding increased attention in the technical literature. Much of this attention is centered on two categories of layered reservoirs: those in which layers are non-communicating and those in which crossflow of fluids occurs between the layers. In the first category, the reservoir is assumed to consist of discrete layers, each uniform within itself and differing from the others only in such properties as thickness, porosity and absolute permeability. The performance within each layer is calculated by one-dimensional flow theory, and the performance of the total reservoir is obtained by summing individual layer performances. Capillary and gravity effects usually are not considered. Representative publications dealing with thi5 type of reservoir are those of Stiles,' Dykstra and Parsons,' Hiatt,3 Warren and Cosgrove' and Higgins and Leighton." Prediction of performance for reservoirs in the second category is considerably more difficult since viscous, capillary and gravitational forces all play important roles in causing crossflow between layers. A number of authors have investigated the simpler problem of two-dimensional displacement flow in a stratified system with a mobility ratio of unity and negligible capillary and gravity effects.'; Others have considered two-dimensional, non-steady-state flow of a single, slightly compressible fluid in a stratified reservoir. A limited number of laboratory oil displacement tests in layered models with crossflow have been reported. Miscible floods (with resultant zero capillary forces) in layered five-spot models were conducted by Dyes and Braun," who studied the effect of mobility ratio with zero gravity forces, and by Craig et al. 12 who studied the effect of gravity forces at constant mobility ratio. Waterfloods in layered five-spot models (with cross-tlow due to capillary, viscous and gravity forces) were conducted by Gaucher and Lindley,"' who showed the effect of gravity forces in causing underrunning of the injected water and by Carpenter, Bail and Bobek, 14 who demonstrated the reliability of Rapoport's" dimension-less parameters for scaling layered systems. Waterfloods in rectangular layered models were conducted by Richardson and Perkins."' who investigated the effect of velocity at constant mobility ratio and with zero gravity forces, and by Hutchinson," who studied the effects of varying mobility, layer permeability and layer thickness ratios. The differential equations which rigorously describe waterflooding in a heterogenous porous medium are non-linear and do not facilitate analytical solution. By using finite difference approximations it is possible to obtain a solution to any desired degree of accuracy. Such a solution, using an alternating direction implicit procedure (ADJP), is described by Douglas, Peaceman and Rachford.18 In the present study, a computer program using ADJP explores systematically the effects of important parameters on waterflood performance of a two-dimensional, two-layered, field-scale model of a water-wet sandstone system. Particular attention is given to evaluation of the water saturation contours and crossflow rates at the interface between layers to gain improved understanding of the crossflow mechanism. PROCEDURE BASIC: FLOW EQUATIONS The basic flow equations for two-dimensional, two-phase, immiscible, incon~pressible flow in a porous medium are:

Jan 1, 1967

By Kiyoshi Azuma, Hiroshi Kametani

The variation of the dissolution behavior of a ferric oxide with calcining temperature has been investigated. Samples were prepared by thermal decomposition of ferric hydroxide, nitrate, oxalate, and sulfate at low temperature, followed by the calcination in the temperature range between 600" and 1200°C. The samples of eight series and a fine crystalline sample of hematite were dissolved in 1 N hydrochloric acid at 55.2°C and the results are represented on double-log graphs for convenience. It is confirmed that all dissolution courses follouj either the accelerated process or the parabolic process except in the special case of the crystalline hematite which dissolced in accordance with the uniform dissolution of a particle. Examinations of the physical properties of the oxide powders revealed that the surface area measured by the permeability method is strikingly relevant to the dissolution behavior of the oxide. In the previous paper,' detailed data were presented on the effect of the kind of acid, the solution temperature, and the concentration of acid on the dissolution of two ferric oxides. It was also shown that these sam ples dissolved in strikingly different ways. The present investigation was carried out on the dissolution of various calcined samples prepared from various ferri salts by various methods to ascertain the course of dissolution. Pryor and Evans2 pointed out a change of the dissolution rate at around 700°C for a series of calcined ferric oxides prepared from the hydroxide. Several papers374 reported also the dissolution of ferric oxide samples. It seems, however, that a systematic account of the relationship between the dissolution behavior and physical properties of the oxide has not yet been given. This paper presents the variation of the dissolution of the oxide in relation to the calcining temperature and the change of physical properties of the calcines. EXPERIMENTAL Raw materials were prepared by precalcination of ferric hydroxide, thermal decomposition of ferric nitrate, oxalate, and sulfate, and aerial oxidation of ferric chloride vapor, at as low a temperature as possible. The products were crushed, ground, if necessary, and sieved with a 100-mesh Tylor screen prior to calcination, after which the specimens were dissolved in acid solution. The following is a detailed description of the preparation of the samples. Sample H. About 500 g of ferric chloride (guaranteed reagent) were dissolved in 5 liters of deionized water and filtered. Ferric hydroxide was precipitated by addition of the minimum amount of ammonium hydroxide solution, and the precipitate was washed continuously till chloride ion was not detected by silver nitrate solution, and then filtered. The filter cake was dried at 120°C for a week and ground, and the -100 mesh portion was used. Sample S. Ferric sulfate (guaranteed reagent) was pyrolytically decomposed in a crucible at 700°C for 24 hr and the product was sieved. In this case the following calcination was carried out at temperatures over 700°C. Sample B. Commercial ferric oxide (guaranteed reagent). About 15 kg of ferric nitrate were decomposed in a furnace maintained at 800°C for 2 hr. The actual temperature of the decomposition was not measured. The product was crushed and sieved, and the -100 mesh portion was used. Sample N. About 50 g of ferric nitrate (guaranteed reagent) were decomposed in a beaker in a sand bath until a red-brown dense solid was produced. This product was crushed and sieved, and subjected to complete decomposition at 500°C. The precalcined product was again sieved and used. Sample N2.5. Since the decomposition temperature was not controlled for sample AT, a different sample was prepared in a temperature-controlled furnace. The subscript represents the decomposition at 250°C. The product was treated in the same manner as sample N. Sample Nc. Under atmospheric pressure it is prac-tically inevitable that ferric nitrate hydrate melts to form a brown liquid at about 50°C before pyrolysis. For this reason, the salt was first slowly heated under reduced pressure (about 10-3 mm Hg measured in a trap refrigerated by dry ice-alcohol) to achieve dehydration without melting. About 5 hr were required for the dehydration and the partial decomposition. Then the temperature was elevated to 500° C in air for complete decomposition. The relatively porous product was sieved and used. Sample Ov. About 200 g of ferric oxalate hydrate (extra pure) were dehydrated under reduced pressure (as described above) followed by thermal decomposition at 500°C for 6 hr in air. The decomposition of this salt was accompanied by liberation of carbon monoxide, by which the ferric salt was initially reduced to a black powder. The powder changed in turn into brown ferric oxide as the gas liberation decreased and reoxidation predominated. The product consisted of sparkling fine particles passing through a 100-mesh screen. However it was ground and sieved as for the other samples. Sample D. Commercial fine powder for magnetic tape purposes. The preparation was as follows.5 Ferric chloride vapor and preheated excess air were mixed and passed into a reaction tube where oxidation took place at 450°C. The fine powder formed was collected in a cottrell chamber. The product was vacuum-degassed at 450°C for 1 hr and sieved.

Jan 1, 1969

By V. S. Gupt, W. H. Somerton

Rock may undergo great changes in physical properties when heated to high temperatures and then cooled, The temperature and intensity of reactions causing rock alterntiorl.s can he controlled by introducing certain chemicals during heat treament. Three typical outcrop sandsone samples were saturuted with common salt solutions, then heated to several maximum temperatures. After cooling, it was found that per-rrreuhilities had increased much more for salt-saturated samples than for samples not saturated with salt but heated to the same temperatures. This was only true, however. for samples heated above the melting points of the particrrlar salts. Potassium chloritle was particularly effective with Bandera sandstone. Samples saturated with potassium chlorirlr. solution and heated to 900C showed an 11-fold increa.se in perrrieahility. Samples without potasrirdrn chloritle but heated to the .same temperature. showed only a 2.7-jold incresre in permecahility. In application, it seems possible that injecting chemicalc into the formations from a wellbore followed by applying intensive borehole heating might promote reactions which would greatly improve permeability of the formalions. INTRODUCTION Great changes in the physical properties of rocks which have been heated to temperatures in the range of 600 to 800C have been reported earlier. Permeability increases of 50 per cent, and equivalent decreases in sonic velocity and breaking strength have been observed. Although there might be some suppression of certain reactions responsiblc for these changes when rock samples are heated under simulated reservoir pressure conditions. recent work has shown that these are more than offset by the increased importance of other reactions at high pressures. The reactions considered responsible for alteration of rock properties by heating include differential thermal expansion. dehydration, phase changes and dissociation of mineral constituents. Ceramists have long known that temperatures at which reactions occur can be raised or lowered by the presence of certain impurities in the sys. tcnl. Thus the possibility exists that by saturating rocks with appropriate solutions. desired reactions might be accelerated and unwanted reactions might be suppressec! when the rock is heated. Purpose of the present work was to investigate the effects of several common salts on the thermal alteration of sandstones. Types of reactions which might occur are reviewed and the findings of a number of thernlochemical alteration tests are presented. THEKMOCHEMICAL ALTERATION OF SANDSTONE MINERALS The most abundant constituent of most sandstones is quartz. Quartz can exist in at least eight different forms, but in the temperature range of immediate interest (to 900C) only two forms are of importance—alpha quartz below 573C and beta quartz above this temperature. Although the alpha-beta inversion of quartz is very important in thermal alteration of rocks. it is not particularly important here because the inversion temperature cannot be changed significantly by the presence of impurities or by the application of pressure. The transition temperature of quartz to tridynlite is 867C. but the transition is very sluggish. Tridymite is considered to be a metastable transition phase between two stable phases: quartz?cristobalite. However. the transition temperature for tridymite to cristobalite is 1,470C.4 The quartz-tridymite-cristobalite transition can be accelerated substantially by the presence of fluxing agents. Finely divided calcium, magnesium and titanium oxides accelerate the conversion while A12O may retard it. A Mixture of NaCl and CsCl reduces the temperature for cristobalite formation by as much as 420°C. Feldspars constitute the second most important mineral in sandstones. The melting points of feldspars range between 1,000 and 1,700°C depending upon the variety. No information available indicated any important role of fluxing agents in the thermal alteration of this group of minerals. Carbonate minerals are subject to dissociation at temperatures within the range of the present investigation. The dissociation of magnesite can start at a temperature of 373°C, but the reaction is sluggish and might not occur untll a temperature of 500°C or ,higher is reached. Dolomite dissociates in two stages at 500°C and 890°C, whereas calcite dissociation temperature is about 885C. Because CO2 is released in the dissociation of carbonates, ail such reactions are somewhat dependent upon CO2 partial pressure. In the absence of CO: in the surrounding atmosphere, the dissociation of calcite starts at 500°C.However. when 1 atm of CO surrounds the sample, the dissociation does not start until 900C. The other carbonates are apparently much less sensitive to a change in partial pressure of CO2. The aragonite calcite transformation can occur an!. where within the temperature range of 357 to 488°C. depending upon the presence of impurities such as barium. strontium. lead and perhaps zinc. The differential thermal analysis (DTA) curves of both magnesitc and dolomite vary with the presence of 'impuritics. The presence of iron in particular seems to affect the magnesitc DTA curve

Jan 1, 1966

By C. R. Kyle, R. L. Perrine

A transparent model of a reservoir has been used to study some characteristics of instability in miscible displacement. The linear model dimensions are 1/4 in. x 91/2 in. x 10 ft. The model is packed with spherical glass beads. Thee groups of experiments have been performed. The objectives were to measure, respectively, the rate of growth of the transition zone, flow velocity variations in the transition zone, and the effect of a graded viscosity zone on stability. All experiments were compared, where possible, with theory. The transition zone grew almost linearly with distance traveled, regardless of flow rate, where adverse viscosity ratios were used. At very adverse viscosity ratios and large distances traveled, some reduction below linear growth was noted. This could be attributed to the effects of mixing. General similarity in flow pattern was commonly shown in different tests, although exceptions occurred at the higher viscosity ratios. Injection of tracer spots permitted direct experimental measurement of velocity variations. Statistical methods were used to compare observed velocity variations with recently published theories. Surprisingly good agreement with some theoretical predictions was obtained. In the experiments using an initially graded viscosity zone, macroscopic stability was reached before predicted by theory. Under conditions theoretically considered to be slightly unstable, any fingering tendency was too slight to become observable within a 10-ft long system. The general correlation of all experiments with published theory was good, and suggests that theoretical predictions can be extended to reservoir problems. INTRODUCTION Early laboratory studies of miscible displacement demonstrated the high recovery efficiency possible with this method. Linear laboratory models showed that, under some circumstances, all in-place hydrocarbons could be recovered using as little solvent as 5 per cent of the pore volume. These results led to a series of pilot field projects. Under realistic conditions, however, other laboratory studies showed that a small solvent bank may not recover much more oil than a conventional gas drive. The cause of this lack of efficiency is unstable viscous fingering. The size of the solvent bank required to give optimum recovery is, therefore, a question of great economic significance. The object of the experiments described in this paper was the quantitative measurement of unstable flow parameters, the verification of present theories on the miscible displacement process, and correlation of these theories with ex peri mental observation. Of the theories which have been advanced, those by Dougherty, Koval and Perrine show promise of predicting reservoir performance with reasonable accuracy.l-7 The present experiments were correlated, where possible, with the theories of Koval and Perrine. Dougherty's theory was not used because of the additional computational difficulty it presented and the belief that no further information would result. EXPERIMENTAL EQUIPMENT A transparent reservoir model, consisting of a vertical reinforced plexiglass column, was used for the experiments. The column was packed with glass beads, which served as the porous medium. To induce a known velocity disturbance in the flow, a double cosine wave restriction in the transverse cross-section was placed at the top of the column. To mark the flow field and provide a means of measuring the local fluid velocity, a dye-spot injection system was provided. The column was 10 ft high by 1/2 in. wide by 1/4 in. deep. The cosine restriction was placed 16 in. from the top of the column. The glass beads were smooth and spherical and had a mean diameter of 0.0102 cm. Liquids entered the top of the column through a manifold allowing three points of entry (left, center, and right, or any combination of these). A positive-displacement pump with variable-speed drive was used to control liquid flow rates. The pump transferred viscous oil from a reservoir to an accumulator, which displaced the test liquid into

Jan 1, 1966

By E. S. Bumps, M. Baeyert, W. F. Craig

SOME time ago during a study of impact properties of tempered martensite,1 it was postulated that the consistently good ductility of tempered martensite might be caused by its relatively small and peculiarly shaped ferrite grains. The fer-rite grains of tempered martensite have approximately the same size and shape as the martensite "needles." Thus they form an interlocking mass of needle-shaped grains quite different from equiaxed or lamellar ferrite grain structures. When the common mechanical test methods are applied to steel, variations are often observed in the ductility of specimens that have closely similar hardness and tensile strength values. The ductility so measured appears to be structure dependent. When steel from the same heat has been heat treated to produce different structures with the same hardness, the elongation and reduction of area values from the tensile test and the transition temperature determined by the notched-bar impact test vary according to whether pearlite, tempered martensite, or other structural constituents were produced by the heat treatment. It has been widely recognized that tempered martensite gives a consistently good performance, when tempered to the same hardness as many other structures with which it has been compared. In recent years the isothermal transformation of austenite to specific structural products and the quantitative evaluation of the character of these products with respect to their nature and response to deformation has received considerable attention. The objective of the present study was to pursue somewhat further the dependence of ductility upon structure; specifically, it was desired to ascertain whether ferrite grain structure, including both shape and size of the grains, can account for the consistently good performance of tempered martensite in the notched-bar impact test. It was thought that a simple experiment would indicate whether the ferrite grain structure plays any part in the good ductility exhibited by tempered martensite in contrast to other steel structures with different types of ferrite grains. By determining the impact transition temperature, it was proposed to compare spheroidites having similar carbide particle size and spacing but obtained in such a manner that their ferrite grain structures would be very different. Spheroidite obtained by tempering martensite, with its small, needle-shaped grains, was to be compared with spheroidite from pearlite. If the latter is produced by sub-critical annealing, the ferrite grains correspond to the pearlite colonies. Thus, if the pearlite was not too coarse, the ferrite grains of spheroidite from pearlite are equiaxed in contrast to the needle-shaped grains of spheroidite from martensite. It was thought that the ferrite grain structure of spheroidite from martensite might depend to some extent upon the grain size of the prior austenite. The austenite grain boundaries limit the maximum attainable size of the martensite needles and thus of the ferrite grains in the derived spheroidite. In order to evaluate any possible influence of prior austehite grain size, spheroidites were to be prepared from martensites that had been formed from fine-grain austenite and also from coarsened austenite. As the carbide particle size and distribution were to be essentially alike in the various spheroidites, the difference would be in the ferrite grain size and shape. Thus any marked difference in transition temperature could be attributable to the character of the ferrite grain structure. There are certain considerations in assuming that these spheroidites would be equivalent in all respects except ferrite grain structure, and an attempt was made to take them into account. One of the considerations was the choice of the carbon content of the steel. An approximately eutectoid steel was selected for two reasons. First, the pearlitic structure would contain no proeutectoid ferrite which might complicate the picture by producing a non-uniform ferrite grain structure in the resulting spheroidite. Then, too, the high-carbon content would inhibit ferrite grain growth during the sub-critical treatment. Another factor to be taken into account was the choice of an alloying element to assure a martensitic structure throughout on quenching the impact specimens. Nickel was chosen, because it is a common alloying element and resides in the ferrite both upon its formation from austenite and throughout tempering. The formation of alloy carbides, or even a large solubility of the alloying element in cementite, would have complicated the interpretation by changing the composition of the ferrite .during spheroid-ization. The possibility of temper brittleness was minimized insofar as possible by using a tempering temperature as high as consistent with the 1 pct of nickel in the steel, namely, 1150°F. While it certainly is not claimed that no difference other than ferrite grain structure could exist between the spheroidites, nevertheless, reasonable precaution has been exercised within the limits of steel metallurgy. It is believed that any large difference in transition temperatures would reflect the difference in ferrite grain structure and that relatively good ductility in the spheroidites from mar-

Jan 1, 1951

By T. R. A. Davey

IN 1947 the author became interested in the fundamental aspects of the desilverizing of lead by zinc, conducted some experimental work, and searched the technical literature for all available fundamental data. Since then a revival of interest in the subject in Europe resulted in the appearance of quite a number of papers. It became evident that it would be more profitable to collect together and examine thoroughly the results of various workers, than to attempt to duplicate the experimental determinations. There are many inconsistencies in the various publications, and it is opportune to review at this time the present status of knowledge on the Ag-Pb-Zn system. There is also a need for a clear description, in fundamental terms, of the various desilverizing procedures. This paper is presented in four sections: 1—There is an historical review of the origins of the Parkes process, of the results of many attempts to find a satisfactory fundamental explanation for the phenomena, and of the modifications proposed to date. 2—A diagram of the Ag-Pb-Zn system is presented. This is believed to be free of obvious inconsistencies or theoretical impossibilities, although thermodynamic analysis subsequently may reveal errors. 3—The fundamental bases of the various desilverizing procedures, which have been used up to the present day, are described; and a new method is suggested for desilverizing a continuous flow of softened bullion in which the bullion is stirred at a low temperature in two stages producing desilverized lead at least as low in silver as that from the Williams continuous process and a crust which, on liquation, yields a very high-silver Ag-Zn alloy. 4—A suggestion is made for the revival of de-golding practice, following a recently published account which does not seem to have attracted the attention it deserves. The terms "eutectic trough" and "peritectic fold" as used in this paper are synonymous with "line of binary eutectic crystallization" and "line of binary peritectic crystallization" as used by Masing.' The German literature on ternary and higher systems is rather extensive and a fairly general system of nomenclature has arisen, whereas in English usage the corresponding terms are not as well established. For this reason the meanings of terms used in this paper, together with the equivalent German terms, are given as follows: 1—Eutectic trough—eutektische rinne: line at which a liquid precipitates two solids S1 and S2 simultaneously. If the composition of a liquid which is cooling reaches this line, it then follows the course of this line until a eutectic point is reached, or until all the liquid is exhausted. The tangent to the eutec-tic trough cuts the line joining S1S2. 2—Peritectic fold—peritektische rinne: line at which a solid S1 and a liquid L transform into another solid S2. If the composition of a liquid which is precipitating S1 reaches the line, on further cooling only S2 is precipitated. The liquid composition moves from one phase region (L + S1) into the other (L + S2), and does not follow the course of the boundary. The tangent to the peritectic fold cuts the line S1S2 produced nearer S,. 3—Liquid miscibility gap, or conjugate solution region—mischungslucke: the region within which two liquid phases coexist in equilibrium over a certain range of temperature. A system whose composition is represented by a point in this region comprises one liquid at high temperature; then as the temperature is progressively reduced, two liquids, one liquid and one solid, one liquid and two solids, and finally three solids. 4—Liquid miscibility gap boundary—begrenzung der flussigen mischungsliicke: the line along which the surface of the miscibility gap dome, considered as a solid model, intersects the surrounding liquidus surfaces. 5—Tie lines—konoden: lines joining points representing the compositions of two liquids, a liquid and a solid, or two solids, in equilibrium. In binary systems the only tie lines customarily drawn are those through invariant points, e.g., through the eutectics of the Pb-Zn and Ag-Pb systems, or the various peritectics of the Ag-Zn system, as in Figs. 1 to 3. In ternary systems it is desirable to draw sufficient tie lines to indicate the slopes of all possible tie lines. 6—Ternary eutectic point—ternares eutektikum: point at which liquid transforms isothermally to three solids, S1, S2, and S Such a point can lie only within the triangle 7—Invariant peritectic (transformation) point— nonvariante peritektische umsetzungspunkt: (a) — On the miscibility gap boundary, the point at which two liquids and two solids react isothermally so that L, + S, + L, + S2. (b)—On the eutectic trough, the point at which a liquid and three solids react iso-thermally so that L + S, + S2 + S3. Such a point must lie on that side of the line joining S,S which is further from S,. (c)—A further possibility, not found in this ternary system, is that the point is at the intersection of two peritectic folds when the reaction concerned is L + S, + S, + S Historical Introduction Karsten discovered in 1842 that silver and gold may be separated from lead by the addition of zinc.2 Ten years later Parkes used this fact to develop the well known desilverizing process which bears his

Jan 1, 1955

By O. Hoover, M. Herman, V. V. Damiano

The Jetter and borie' teclznique of determining textures using a spherical specimen has been applied to tlze study of compression-rolled beryllium sheet. Snzall spheres the order of 1 mm in diam cut from the beryllium sheet were autotnatically rotated about tz41o axes using the G.E. single-crystal goniometer. Quantitative pole figures were obtained without tke need to apply absorption corrections. Compression-rolled beryllium exhibited peak intensities ,for (0002) planes of positions tilted 10 deg to the rolling plane and a near random distribution of (1010) planes about the nornal to the rolling plane. TECHNIQUES for determining textures of rolled sheet material are amply described in the literature. The techniques are found to be variations of two basic methods. One due to Decker, Asp, and arker, referred to as the transmission method, utilizes a thin-sheet specimen in which the X-ray beam enters the specimen from one side and the intensity of the beam which emerges from the opposite side is measured. The second method due to chulz,3 referred to as the reflection method, utilizes a thick specimen and the intensity of the beam emerging from the same side is measured. The two rotations of the specimen in the beam are designated a and 8. In order to completely determine the texture of sheet material, it is generally necessary to use a combination of the two methods. The calculations involved in correcting the raw X-ray data for absorption effects and the combining of the data obtained by the two methods are very laborious and time consuming. To avoid the intensity corrections which arise as a result of the changing diffraction volume and path length within the sample other methods have been proposed. The Norton method utilizes a cylindrically shaped specimen cut from the sheet material. Since the rods have rotational symmetry, the absorption correction is constant for rotations about the sheet texture. Jetter and Borie' employed a spherical specimen to analyze the fiber texture of extruded aluminum rods. The spheres were rotated rapidly about the fiber axis to include a large number of grains in the X-ray beam and changes in intensity with respect to tilts of the fiber axis were measured. The absorption correction was constant for all angles and was neglected. The Jetter and Borie' technique finds excellent ap- plication to very fine-grained low-absorbing metals in which the entire sphere volume can contribute to the diffraction volume. In the case of low-absorbing metals, however, serious limitations on specimen thickness occur as demonstrated by Braggs due to de-focussing effects. Peak shifts may occur which negate the assumption that integrated intensities are proportional to peak intensities. These limitations in sphere size to the order of 0.5 to 1 mm for beryllium require that the grain size be sufficiently small to include a large enough statistical sample. The present paper describes the application of spherical specimens less than 1 mm in diam to the quantitative determination of pole figures for compression-rolled beryllium sheet having a grain size the order of 10 p. EXPERIMENTAL 1) Specimen Preparation. Two techniques for spark-machining beryllium spheres were tried. One involved the use of a hollow cylinder as a cutting tool. The tool was fed into the rotating cylindrical specimen as shown in Fig. l(a). The hollow cylinder was carefully aligned such that the axis of the cylinder and the axis of the specimen lay in the same plane and were 90 deg to each other. As the hollow cylinder was fed into the rotating cylindrical specimen, a spherical shape was formed as shown in Fig. 1. Alignment was very critical. Slight misalignment resulted in the formation of a barrel-shaped specimen instead of a sphere. A second technique involved the use of a cutting wheel shaped as shown in Fig. 2 with a groove of the desired radius. A section of the sheet specimen was first turned into a cylinder on the left part of the cutting wheel. It was then shifted to the right and a spherical specimen was turned as shown in Fig. 2. The axis of the cylinder lay in the plane of the sheet. Flats corresponding to the rolling plane of the sheet were used to grip the specimen during the machining operation and these served to identify the rolling plane of the sphere. 2) Rotation of Spec=. The spherical specimen is shown mounted on the G.E. single-crystal goniometer in Fig. 3. The knob A of the goniometer shown in Fig. 3 rotates the specimen about the pedestal axis. These angles have been designated as @ angles. The knob B rotates the specimen about an axis perpendicular to the pedestal axis. These angles have been designated as p angles. A device was made to automatically drive the single-crystal goniometer by means of two flexible shafts connected to the A and B knobs as shown in Fig. 3. The motor system was designed to rotate the knob A, thus rotating the specimen through angles of $I while the B knob remained stationary. After one complete

Jan 1, 1967

By Emerson F. Heald

A systematic study was made of new and old data on chemical activities in Fe-Pt alloys at elevated ternperatuves. Experimental results may be expressed in terms of the excess free energy using Least-squares analysis of the data gave the following values for the constants: for the temperature range 1130° to 1350°C, and tentatively to 155O°C, B = -3.326564 and C = 0.221051; for the temperature range 650" to 850°C, B = -2.555690, C = 1.762735, and D =0.097196. In the study of iron-containing silicate systems, it is sometimes desirable to have a direct experimental measure of the activity of iron in the system. The well-known solubility of iron in platinum, often a headache in experimental work on iron compounds under reducing conditions, can be used to advantage in this respect. If the activity of iron in an Fe-Pt alloy in equilibrium with the silicate is known as a function of the composition of the alloy, chemical analysis of the alloy will give a knowledge of the activity of iron in all of the phases in the system. The present study was undertaken in order to elucidate the characteristics of Fe-Pt alloys as iron activity indicators. This work is intended to tie together some previous work, which may be summarized as follows. Larson and Chipman1 determined the activity of iron in Fe-Pt alloys at 1550°C by equilibrating platinum metal with calcium oxide-iron oxide-silica melts of known iron activity. Compositions of the resulting alloys were determined by chemical analysis. A similar study was carried out by Taylor and ~uan,' who worked at 1300°C. They brought the Fe-Pt alloys into equilibrium with iron oxide under conditions of known partial pressure of oxygen, and thus, from the work of Darken and ~urr~,~ conditions of known iron activity. Compositions were determined indirectly, by following the change in weight of the sample. Sundaresen et el* used the electromotive force of cells in which the alloy formed one electrode in order to measure the activity of iron in the alloy at 650" and 850°C. These temperatures were chosen to be above and below the first-order phase transition which takes place upon the ordering of Fe3Pt and the second-order transition which occurs upon the ordering of FePt3. EXPERIMENTAL 1) High Temperatures. The starting materials used were thin platinum foil, about 0.002 mm thick, and Fisher Certified reagent ferric oxide, Fez03, which had been heated for 24 hr at 1000°C. An intimate mixture of 80-mesh Fez03 and platinum platelets was placed in a thin platinum foil envelope. The latter was suspended from thin platinum wires in the hot zone of a vertical-tube, platinum-wound furnace of the type described by Muan and ~sborn.~ A capillary gas mixer similar to that used by Darken and Gurry3 was used to prepare a precisely known mixture of carbon dioxide and hydrogen, which was allowed to flow upward through the furnace tube. The partial pressure of oxygen in contact with the sample was thereby fixed at a value which was calculated from the charts prepared by porter? Temperatures were measured with a Pt-10 pct Rh-in-platinum thermocouple, which was calibrated using the melting points of gold (1062 .@C) and diopside, CaMgSizOB (1391.5"C). Temperature control was maintained to within i3"C with a Geophysical Laboratory proportional controller, using the furnace resistance as the sensing element. Samples were quenched by passing a small current through the platinum suspension wires, allowing the sample to drop into a bath of dibutyl phthalate at the bottom of the furnace tube. Prior to chemical analysis the samples were washed with acetone and dried. 2) Chemical Analysis. It proved possible, in almost all cases, to separate the Pt-Fe platelets physically from particles of iron oxide. The platelets were dissolved in a small volume of aqua regia, evaporated to dryness, and redissolved to 0.1 M HC1. In order to determine iron in the platinum alloy potentiometrically, it is necessary first to remove the platinum. A 10-cm column of Amberlite IR-120 cation exchange resin in the hydrogen form provided separation quickly and quantitatively: The mixture of iron and platinum in 0.1 M HC1 was added to the top of the column, and washed with about 100 ml of 0.1 M HC1. Under these conditions, the iron, principally in the form of cations such as FeC1" and FeCl;, is held quantitatively in the uppermost centimeter of the column. The platinum, in the form of anions such as PtC&- , is washed through without being adsorbed. After a qualitative test with stannous chloride indicated all of the platinum was removed, the iron was

Jan 1, 1968

By C. L. Sollenbeger, R. B. Greenwalt

PERHAPS the most widely used dispersants or gangue depressants in nonmetallic flotation are sodium silicates, which vary in silica-to-soda ratio from 1 to 3.75. Typical manufactured silicates in order of decreasing solubility and increasing amounts of silica are Metso, silica-to-soda ratio of 1.00; D, 2.00; RU, 2.40; K, 2.90; N, 3.22; and S-35, 3.75.* References in flotation literature1,2 to the use of sodium silicates are often weak because they fail to mention the type of silicate used. Metso and silicate N have occasionally been mentioned, but when the type of silicate is not mentioned, it is usually assumed to be N, the cheapest of the soluble silicates and the one recommended by sodium silicate manufacturers as a flotation agent. In the All is-Chalmers Research Laboratories a systematic study was made of the effect of different alkali-silica ratios on the concentration by flotation of two scheelite ores. One of these was a high grade ore from the Sang Dong mine in Korea. The effect of such factors as pH; addition agents; and conditioning time, temperature, and pulp density on the flotation efficiency of this ore have been described previously. The other ore was a low grade ore from Getchell Mines Inc., Nevada. The mineralogy and techniques of concentrating this ore have been described by Kunze. Hereafter these ores will be referred to as the Korean and Nevada ores. Experiments were made with both to determine the effect of three factors—-type of silicate, concentration of silicate, and pH of the pulp—on recovery and grade of tungsten in a rougher concentrate. Average WO, content of the Korean ore was 1.50 pct and of the Nevada ore 0.27 pct. The predominant tungsten mineral in both ores was scheelite, which was accompanied by a small amount of powellite. The powellite and scheelite were finely disseminated through both ores and required a —200 mesh grind for liberation. Major gangue minerals in the Korean ore, in decreasing order of abundance, were amphi-boles, quartz, biotite, garnet, fluorite, and calcite. Bulk sulfides composed about 3 pct of the total weight. Gangue in the Nevada ore, in descending order of abundance, was garnet, alpha quartz, calcite, phlogopite, wollastonite, and amphiboles. Sulfide minerals were 3 to 4 pct of total weight. Batch flotation experiments were made with 500-g samples of ore, each sample wet-ground to 90 pct passing 200 mesh. The finely ground ore was floated in a Fagergren batch cell at 25 pct solids. The natural pH of the Nevada ore was 8.9 and of the Korean ore, 8.5. The D, RU, K, N, and S-35 sodium silicates were obtained in colloidal dispersions with varying amounts of water. The most alkaline, Metso, was in dry powdered form. For convenience in addition, 5 pct solutions by weight were prepared from each of the silicates, on the basis of dry sodium silicate dissolved in the correct amount of distilled water. Chemical analyses of the various silicates are given in Table I, together with the pH of the 5 pct solutions. A preliminary bulk sulfide float was made with secondary butyl xanthate as the collector and pine oil as the frother. The WO] analysis of the sulfide concentrate was nearly 1 pct for the Korean ore and about 0.1 pct for the Nevada ore. The tungsten contained in the sulfide concentrate constituted about 3 pct of the total tungsten in each ore. No effort was made to recover these tungsten values. The scheelite was floated with oleic acid. Adjustments in pH were made with sulfuric acid or sodium carbonate. A 1 pct solution of 85 pct Aerosol OT was sprayed on the froth and sides of the cell during the scheelite float to aid in dispersing the minerals and to decrease the entrapment of gangue particles. Six tests were planned for each of the six types of silicate in which concentrations of 1, 2, and 4 1b of silicate per ton of dry ore were investigated at both 6.5 and 10 pH. All tests were made at room temperature. The performance of each silicate was judged from the grade and recovery of WO, in the scheelite rougher concentrate. Tungsten recovery was calculated on the basis of the scheelite remaining in the ore after the preliminary sulfide float. Testing of each silicate at three levels of concentration and two levels of pH required 36 tests with each scheelite ore. Variance analyses were performed on the concentrate grades and recoveries to determine whether or not the type of sodium silicate, the concentration of sodium silicate, or the pH significantly affected recovery or grade. Results Concentrate Grade: A variance analysis of the concentrate grades for the Korean ore showed that concentration of the silicate and pH of the ore pulp were major factors in producing a high grade concentrate. Also, the silica- to-so da ratio was important as an interaction with pH. The concentrate grade vs silica-to-soda ratio is plotted in Fig. 1. The curves show that the concentrate grade improved with an increase in concentration of sodium silicate and also

Jan 1, 1959

By J. R. Szedon, T. L. Chu, G. A. Gruber

Amorphous adherent filnzs of silicon dioxide have been deposited on silicon substrates by the oxidation of silane at temperatures ranging from 650 to 1050C. Various diluents (argon, nitrogen, hydrogen) were used to suppress the formation of SiO2 in the gas phase. Deposition rates of the oxide were determined over the temperature range in question as functions of' re-actant flow rates. Etch rate studies were used for a cursory comparison of structural properties of deposited and thermally grown oxides. From electrical evaluation of metal-insulator-silicon capacitors it was determined that the interface charge density of deposited films is similar go that of dry-oxygen-grown films in the 850° to 1050 C temperature range. Deposited films exhibit several ionic instability effects which differ in detail from those reported for thermal oxides. Stable passivating films of silicon nitride over deposited oxides appear to be practical for use in silicon planar device fabrication. Such films can be prepared under temperature conditions which have less effect on substrate impurity distributions than in the case of grown oxides. AMORPHOUS silicon dioxide (silica) is compatible with silicon in electrical properties and is the most widely used dielectric in silicon devices at present. Silica films can be prepared by the oxidation of silicon or deposited on silicon or other substrate surfaces by chemical reactions or vacuum techniques. The ability of thermally grown silicon dioxide films to passivate silicon surfaces forms one of the practical bases of the planar device technology. Properly produced and treated films of grown SiO 2 can have low densities of interface charge (-1 X 10" charges per sq cm) and can be stable as regards fast migrating ionic sgecies. 1 To maintain these properties, even with an otherwise hermetically sealed ambient, the Sia layers must be at least l000 A thick. Such thicknesses require oxidation in dry oxygen for periods of 7.8 hr at 900°C or 2 hr at 1000°C. Although oxidation in steam or wet oxygen can reduce these times to 17 and 5 min, the resulting oxides must be annealed to produce acceptable levels of interface charge., Oxidation or annealing involving moderate to high temperatures for extended periods of time can be undesirable. Under some conditions, there can be changes in the distribution of impurities within the underlying substrate. A chemical deposition technique using gaseous am-bients is particularly attractive and flexible for preparing oxide films. With a wide range of deposition rates available, films can be produced under condi- tions of time and temperature less detrimental to impurity distributions in the silicon than in the case of thermal oxidation. The pyrolysis of alkoxysilanes, the hydrolysis of silicon halides, and various modifications of these reactions are most commonly used for the deposition of silica films.3 Silica films obtained in this manner are likely to be contaminated by the by-products of the reaction, organic impurities, or hydrogen halides. The use of the oxidation of silane for the deposition process has been reported recently.4 The deposition of silica films on single-crystal silicon substrates by the oxidation of silane in a gas flow system has been studied in this work. The deposition variables studied were the crystallographic orientation of the substrate surface, the substrate temperature, and the nature of the diluent gas. The electrical charge behavior of Si-SiO2-A1 structures prepared under various conditions was investigated by capacitance-voltage (C-V) measurements of metal-insulator-semiconductor (MIS) capacitors. The experimental approaches and results are discussed in this paper. 1) DEPOSITION OF SILICA FILMS The overall reaction for the oxidation of silane is: The equilibrium constants of this reaction in the temperature range 500° to 1500°K, calculated from the JANAF thermochemical data,= are shown in Fig. 1. In addition to the large equilibrium constants, the oxidation of silane is also kinetically feasible at room temperature and above. However, the strong reactivity of silane toward oxygen tends to promote the nucleation of silica in the gas phase through homogeneous reactions, and the deposition of this silica on the substrate would yield nonadherent material. The formation of silica in the gas phase can be reduced by using low partial pressures of the reactants. Argon, hydrogen, and nitrogen were used as diluents in this work. 1.1) Experimental. The deposition of silica films by the oxidation of silane was carried out in a gas flow system using an apparatus shown schematically in Fig. 2. Appropriate flow meters and valves were used to control the flow of various reactants, i.e., argon, hydrogen, nitrogen, oxygen, and silane. Semiconductor-grade silane, argon of 99.999 pct minimum purity, oxygen of 99.95 pct minimum purity, and nitrogen of 99.997 pct minimum purity, all purchased from the Matheson Co., were used without further purification. In several instances, a silicon nitride film was deposited over the silica film. This was achieved by the nitridation of silane with ammonia using anhydrous ammonia of better than 99.99 pct purity supplied by the Matheson CO.' The reactant mixture of the desired composition was passed through a Millipore filter into a horizontal water-cooled fused silica tube of 55 mm

Jan 1, 1969

By H. C. Chao, L. H. Van Vlack

Small MnS inclusions with known crystallographic orientations were placed inside powder compacts of low-carbon steel. After the metal was axially campressed with negligible end friction, the deformstions for the metal and the inclusions were compared. The MnS inclusions deformed more when the [100] direction was aligned with the compression axis than when the [111] direction was parallel to this axis. The deformations of the inclusions in the two principal radial directions were equal for each of the above orientations. Inclusions with [110] compression alignments did not deform with radial symmetry. The relative deformation of the inclusion and metal was closely dependent upon the relatiue hardness of the two phases. The relative deformation of the two phases was not sensitive to the rate of deformation. RECENT studies by the authors1.&apos; suggested that the plastic deformation of MnS in steel would probably be highly sensitive to the orientation of the inclusions and to the temperature of the metal. This paper reports an investigation of these factors upon MnS behavior within steel. Manganese sulfide (MnS) possesses an NaCl-type structure and typically has extensive (l10) {110} slip as a separate (noninclusion) crystal.&apos; A secondary slip system, ( l 10) { l l l}, has also been observed where the major slip system is restricted. In general, MnS inclusions must be rated as a highly deformable second phase.3 The amount of sulfide deformation varies, however, with several composition and processing factors. Some of these have been only partially assigned. For example, it is known that minor amounts (<0.01 pct) of silicon within free-machining steels will increase the amount of MnS deformation,4 but the mechanism of the added deformation can only be surmised at the present. Manganese sulfide and steel have sufficiently comparable deformation characteristics so that slip which is started in steel may be continued through the sulfide inclusions and back into the steel if the crystal orientations are favorable.5 A more detailed discussion of previous work on the plastic deformation of NaC1-type crystals and on the plastic deformation of inclusions within a metal is given in Chao&apos;s work.6 EXPERIMENTAL PROCEDURE The manganese sulfide which was used in this study was prepared by previously described methods.&apos; Single crystals of MnS, both as cleavage cubes and as spheres, were oriented within steel powder compacts so that the desired crystal directions were parallel to the direction of axial compression. A four-stage hydrostatic compaction procedure was used and involved the following steps. In the first stage part of the powder was placed in a metal die 1 in. in diameter with a thick (1 in. OD, 5/8 in. ID) rubber liner which had one end plugged. The steel powder was hand-rammed, making it as dense as possible before placing a carefully sized MnS crystal (either as a sphere or as a cube) near the center. The crystal was oriented with the chosen direction vertical; viz., [001], [011], or [111], with the aid of a X10 microscope. A pair of tungsten wire threads 0.020 in. in diameter was inserted along the side of this (&apos;core compact" to locate the desired plane after the compression tests. After the crystal was positioned in the center of the die, more powder was added and carefully rammed by hand. The die was then capped with a rubber plug of the same hardness and thickness as that of the liner. The whole assembly as shown in Fig. 1 was compacted by a ram load of 54,000 lb (about 70,000 psi). In the second stage a smaller, 3/4-in, rubber-lined die was used to give a stress of approximately 120,000 psi. The above process was repeated with the initial compact serving as a core for a larger compact. The final product after sintering was a cylinder 1 cm long and 1 cm in diameter, having a density of 7.54 g per cu cm. This was close to the theoretical density since the metal contained a non-metallic phase. There was no evidence of MnS deformation during the hydrostatic compaction or subsequent sintering. Elevated-temperature hardness data were obtained by procedures previously described.2 Compression tests for inclusion deformation utilized the cylinders which were described above. The critical problem in these tests was the lubri-

Jan 1, 1965

By H. L. Riley, B. W. Gandrud

AS far as the present writers know, this system of flotation has not been used elsewhere in this country, but in the last couple of years it has been introduced, with minor variations, at one plant in England and one in Wales.&apos; The system has been described and discussed in a number of publications.2-5 The following is quoted from an abstract of the latest of these,5 a paper presented at an International Conference on Industrial Combustion in 1952. On the basis of experience to date with the commercial plants, it is believed that the kerosene-flotation process incorporates all the necessary elements to make it greatly superior to anything else now available for treating of fines in wet processes of coal preparation. Additional study and investigation are still needed, however, to determine if certain phases of the process can be improved to such an extent as to make it generally satisfactory and acceptable to the industry. Further improvements will be needed with respect to the capacities of the flotation cells and the reagent consumption. The situation referred to above explains why an investigation is being made of the possibilities of achieving better cell capacities. Results obtained from this investigation, which is still in progress, are believed significant with regard to both cell capacity in general and the relation of cell design to cell capacity in particular. Commercial equipment now being used in a laboratory-type investigation should have performance characteristics similar to those of the larger machines. Equipment and Procedures: All flotation tests have been made in a standard Denver sub-A 24x24-in. unit cell of 12-cu ft volume. Cell modifications to make it more suitable for the tests were an adjustable front-wall section for varying cell depth and a perforated scraper-drag assembly for removal of the float product. There is also an apron dry-coal feeder, a gravity-feed water supply, reagent feeders, and a centrifugal pump that feeds the mixture of coal, water, and reagents into the flotation cell. A wattmeter connected into the drive-motor circuit records the power requirements of the impeller throughout each run. Dry coal, water, and reagents are all fed through a pan-type intake to the feed pump. A Sturtevant blower was set up to furnish air for supercharging. A centrifugal pump with a garbage-can intake provides for disposal of refuse flow to an outside settling tank. Figs. 1 and 2 show the flotation cell; Fig. 2 also illustrates the blower for supercharging. For purposes of this investigation, the percentage by weight of the feed coal recovered in the float product under a standard set of conditions has been considered as the criterion of cell capacity. The authors realize that such a criterion may be somewhat unorthodox, as the term cell capacity is usually understood to refer to feed input and ordinarily takes into account the ash analyses of the float product and refuse. However, the word capacity is flexible enough so that Webster gives one definition as maximum output, a definition which seems to justify, at least partly, acceptance of the above criterion. It has been the authors&apos; experience in the Birmingham district that the ash-reduction efficiency of the coal-flotation process is generally satisfactory and that the only real problem is to increase the rate of float recovery so that the feed rate to any given bank of cells can be increased without undue loss of coal in the refuse. Originally it was planned to operate the flotation cell to simulate continuous operation during sampling periods. It was assumed that operating for reasonable time with feed coal, water, and reagents turned on would stabilize conditions so that the weight of float coal discharged during a fixed time interval would be an accurate measure of the rate at which the coal was being floated. It developed, however, that this supposition was erroneous. The float coal, caught for fixed time intervals and weighed, gave widely varying results in duplicate runs. Efforts to correct this trouble failed, and it was decided to try to operate on a batch-test basis, whereby all the float coal produced during a run on a known weight of feed coal would be caught in tubs, dewatered, and weighed. This method gives consistent and reproducible results, with total float product weight rarely varying by more than 3 or 4 pct on duplicate runs. The standard test procedure is as follows: A 132-lb sample of dry feed coal is weighed and placed in the feed hopper. The feeder is adjusted for a rate of 800 lb per hr. Feed water and reagents are turned on, and the feed and refuse pumps are started. One minute later the impeller is started. Six minutes are allowed for the cell to fill up with the water-reagent mixture. The feed of dry coal is started at the end of this 6-min period. One minute later the float-coal removal drag is started. The float coal is caught in one tub for the first 6 min after the flow of feed coal starts. Tubs are then changed, and the float coal is caught in a second tub until the feed coal runs out, when the tubs are again interchanged to catch the float coal for the remainder of the run in the first tub. The cell is kept running for 3 min with the water and reagents on after the feed stops to allow residual float coal to be removed. At the end of a test the wet float coal in both tubs is weighed and the total weight recorded. The product in the second tub is used for moisture determination and screen-size analyses. When the

Jan 1, 1956

By J. Leja, J. H. Schulman

FROTH flotation is usually effected by the addition of a collector agent and a frothing agent to an aqueous suspension of suitably comminuted mineral ores. The action of collectors is to adsorb onto the surfaces of minerals to be separated, sensitizing them to bubble adherence. The action of frothers has, in the past, been accepted as that of froth formation only, brought about by a lowering of the air/water interfacial tension. Substances capable of producing froth are classed1a,b according to their relative capacities for production of froth-volume and froth stability in the simple frother-water system. The purpose of this paper is to show that the surface active agents acting as frothers become effective only when there is a suitable degree of molecular interaction taking place between collector molecules and frother molecules at the air/water and solid/ water interfaces. Further, the discussion will demonstrate that the actual mechanism of adherence of an air bubble to a suitably collector-coated particle is due to the molecular interaction collector-frother. This leads to the formation of a continuous interfacial film of associated molecules, anchored to the mineral by polar groups of the collector, and enveloping the whole bubble. The tenacity of adhesion mineral-to-bubble results from the strength and the visco-elasticity of this mixed film. Some 20 years ago Christman2 postulated mutual dependence of collector and frother in effecting flotation. This view was, however, strongly opposed by Wark,3 who pointed out that an addition of frother had no effect on the value of contact angle once this was established in the solution of collector. More recent work by Taggart and Hassialis&apos; indicated that the presence of frother, namely, cresol, leads to the immediate establishment of a contact angle on sphalerite, partially coated with xanthate, whereas an air bubble fails to make contact in potassium ethyl xanthate solution alone, even after 60 min induction time. Wrobel5 raws attention to the selectivity of frothers in flotation. Many instances of antagonistic effects of certain mixtures of frothers (or collectors and frothers) on flotation froth have been known to flotation operators and have been reported in literature. Taggart6 and Cooke7 give several examples of incompatibility of certain ratios of frothers and collectors, e.g., oleate and long-chain sulphates, pine oil and soaps. Monolayer Penetration. Properties of insoluble films produced by molecules of surface active agents orientated at the air/liquid interface are conveniently studied by the Langmuir trough technique, described fully by Adam.&apos; Using the trough technique Schulman and Hughes" and Schulman et al.10a. b, c, d,e established the existence of molecular interactions occur- ring between certain types of surface active agents. Their experiments revealed the phenomenon of penetration of an insoluble monolayer (e.g., a film of a long-chain alcohol) by a soluble agent (e.g., sodium alkyl sulphate) injected into the substrate (water or salt solution). The degree of molecular interaction taking place on penetration is determined by changes in the surface pressure of the resulting film, changes of its surface potential and its mechanical properties (viscosity and rigidity). When the interaction takes place between both polar groups and both hydrophobic groups of the two participating amphipathic molecules a molecular complex is formed. Complexes formed on penetration of the monolayer at interfaces are not necessarily true chemical compounds: they are labile in solution, the activity and reactivity of individual components are greatly different from those of the molecularly associated complex, and on crystallization they usually separate out into components. However, when formed in the orientated state at interfaces they are found to be very stable, although some mixed films spread as monolayers of stoichiometric complexes can show further penetration by subsequent additions of the soluble component injected into the substrate.&apos;" The degree of association between two or more types of surface active agents is very sensitive even to small changes in electric (dipole) moment of the polar groups of the amphipathic molecules as influenced by magnitude and position of neighboring ions or dipoles, their size, concentration, and stereochemistry. In addition, the molecular association is greatly influenced by concentration and type of inorganic salts in the substrate, by its pH, and by temperature. The nonpolar groups of interacting molecules greatly affect the stability of molecular complexes. Progressive shortening of the aliphatic chain of one of the reacting molecules weakens (at an increasing rate) its tendency to form stable complexes. Similarly, introduction of a double bond of cis-form into one of the reacting chains, which changes the straight hydrocarbon chain into a kinked one, or introduction of a branched chain, reduces the stability of the associated complex. Monolayer Adsorption. Using the trough technique and injecting metal ions into the substrate (water or salt solution) underlying insoluble films of fatty acids, alkyl amines, and sulphates, Wolsten-holme and Schulman11a,b,e. &apos; and Thomas and Schulman" have established conditions, namely, pH, concentration. and steric factors, under which molecular interactions take place between the polar groups of the surface active agents and the metal ions. These interactions are marked by great changes in the solubility and mechanical properties of the monolayer of the agent; no surface pressure increases are observed as in monolayer penetration experiments. The results of these adsorption studies, correlated with flotation experiments, indicated that in the case of fatty acids and alkyl sulphates their adsorption onto minerals of base-metals takes place by a similar

Jan 1, 1955

By R. F. Dickerson, D. A. Vaughan, A. F. Gerds

ALTHOUGH the metallurgy of uranium has been under intensive study since the early 1940&apos;s, no systematic effort has been made to identify the non-metallic inclusions in uranium. Uranium carbide (UC), which is probably the most common inclusion found in graphite-melted metal, has been tentatively identified by previous investigators, but the other nonmetallic inclusions have received little attention. Since metallography is a valuable tool in metallurgical studies, the metallographic identification of the nonmetallic inclusions in uranium is important. Such an investigation has been completed and the identification of slag-type inclusions and of uranium monocarbide, uranium hydride, uranium dioxide, uranium monoxide, and uranium mononitride is described. Metallographic Preporation It is often possible to prepare specimens for metal-lographic examination equally well by several methods. The specimens which were examined in this work were prepared by one of two acceptable methods. For the convenience of the reader, both methods will be discussed in detail and will be referred to simply as Method I or Method II in the subsequent sections. For both Methods I and 11, specimens for microscopic examination usually were mounted either in bakelite or in Paraplex room temperature mounting plastic. Method I—Specimens were ground in a spray of water on a revolving disk covered successively with 120-, 240-, and 600-grit silicon carbide papers. It was necessary to perform the final grinding operation carefully on worn 600-grit paper to keep the scratches as fine as possible. After washing and drying, the specimens were polished for 3 to 4 min on a slow speed wheel (250 rpm) covered with a medium nap cloth. Diamet Hyprez Blue diamond polishing paste, Grade 00, 0 to 2 µ, was used as abrasive with kerosene as lubricant on the wheel. Specimens were washed thoroughly in alcohol and final polished electrolytically in an electrolyte composed of 1 part stock solution (118 g CrO, dissolved in 100 cm3 H2O) with 4 parts of glacial acetic acid. A stainless steel cathode was used. At an open circuit potential of 40 v dc, a polishing time of 2 sec retained inclusions well with the bath at room temperature. If additional etching was required to sharpen the interface between the metal and the inclusions, an electrolyte composed of 1 part stock solution (100 g CrO3 and 100 cm8 H20) and 18 parts glacial acetic acid was used at room temperature. Best results were obtained by etching for from 10 to 15 sec at 20 v dc in the open circuit. Surfaces obtained by this method are suitable for microscopic examination. However, if desired, they may be etched further with other chemicals. Method 11—Rough grinding was done on a wet 180- or 240-grit continuous grinding belt. The specimen was then ground by hand successively on 240-, 400-, and 600-grit silicon carbide papers in a stream of water. Final polishing was accomplished on a 4 in. high speed wheel (3400 rpm) covered with Forstmann&apos;s cloth. Linde B levigated alumina, suspended in a 1 volume pet chromic acid solution, was the abrasive. Specimens usually were polished in 5 min or less by this technique. Often the inclusions present in the metal were identified in the mechanically polished condition. When etching was required to outline inclusions more sharply, one of the two following methods was used. In the first method, the specimen is etched lightly while electropolishing in the chromic-acetic acid solution described above (1 part of stock solution to 4 parts of acetic acid). The electrolyte was refrigerated in a dry ice-ethyl alcohol bath and specimens were etched at 60 v dc on the open circuit for 2 or 3 cycles of 3 to 4 sec each. The second technique utilizes electrolytical etching at about 10 v dc (open circuit) in a 10 pet citric acid solution at room temperature. X-Ray Diffraction Technique The major problem in the identification of inclusions in metals by X-ray diffraction techniques is the extraction of a sufficient amount of each type of inclusion to obtain an X-ray diffraction pattern. In the present study, X-ray diffraction patterns were obtained from individual inclusions of the order of 10 µ diam. The polished and etched samples shown in the micrographs were examined at a magnification of X54 or XI00 with a binocular microscope. This allowed sufficient working distance to extract the inclusions with a needle probe for powder X-ray diffraction analysis. Friable inclusions such as MgF2, CaF2, UO2, and UH3 could be freed from the metal by probing the as-polished and etched surface. The fine particles then were picked up on the end of a Vistanex-coated glass rod (0.002 in. diam) which was held in a brass adapter made to fit the powder X-ray diffraction camera. The end of the glass rod was centered in the path of the X-ray beam. In the case of the UC, UO, and UN inclusions which are smaller in size, more metallic in appearance, and less friable than the other inclusions, it was necessary to etch the inclusion in relief before extraction. UN inclusions etched sufficiently in relief in the electrolytic polishing solution described in Methods I and II by increasing the polishing time. UN inclusions were relief etched by extending the

Jan 1, 1957

By Peter R. Drummond

THE final casting of refined copper has been re-J- stricted for generations by the following sequence of operations: Filling the reverberatory furnace, melting, skimming, blowing or flapping, and poling. The hoped-for 24 hr cycle, producing 300 tons or more, has been taken up largely with the necessary bat time-consuming tasks of cleaning the bath, sulphur elimination, and in turn removal of excess oxygen to produce tough-pitch copper. Incidental to comparatively slow melting under combustion gases, copper oxides react with the furnace lining, and the slag so-formed must be completely recycled. The three-phase arc furnace has eliminated some of the cycle stages, and telescoped the remainder into a continuous operation. Electrical energy, supplied to graphite electrodes enclosed in high grade refractories, rapidly melts copper cathodes and sustains a stream of metal, containing approximately 0.01 pct oxygen, without contamination from fuel. The arc was struck on the first large electric furnace for melting copper in the United States on April 13, 1949. The earliest use of this type of furnace was at Copper Cliff, Ont., in 1936, and an admirable description of their installation has been published? Copper, melted in the Baltimore furnace, is used to cast billets, and the installation differs somewhat from the Canadian, as will be described. The arc furnace is a heavy-duty, three-phase furnace, holding 50 tons, the general outline of which appears on Fig. 1. The steel shell is 15 ft ID with a bottom radius of 14 ft 2 in. The roof, separate and distinct from the body, consists of a 15-ft water-cooled, cast-steel ring of the same outside diameter as the furnace. The center line of the furnace lies 9 ft 6 in. from that of the trunnions, permitting a 5" backward tilt for skimming, and a 40" maximum nose tilt forward for complete draining. Normally, the furnace overflows by displacement, and the use of the forward tilt arrangement is restricted to covering charging delays. The charging slot, 3 ft 8 in. x 5 in., lies on the north center line, the tap hole on the south, and the 30x30 in. skim door 45" to the west of the slot. The original 20-in. graphite electrodes were replaced with 14 in. in December 1949. Three conventional direct current winch drives, governed by electrical controls, position each electrode which has individual mast supports and counterweights. An independent circulation supplies cooling water for the electrode glands, the roof ring, charge slot, and the skim door frame. Arc Furnace Refractories Hearth: Fused-in monolithic bottoms had been used in copper arc furnaces, installed prior to April 1949. These consisted of thin layers of periclase, successively fused in place over preliminary brick courses. Heat was obtained from the arc, using a T-like arrangement of broken electrodes resting directly on the periclase to be fused. The operation, taking weeks to perform, was very expensive. The chemically-bonded magnesite-brick bottom, installed at Baltimore, was the first of its kind and a radical departure from previous practice. It consists of a 1 to 6-in. layer of castable refractory laid on the steel shell, modifying it to a 12 ft 2 in. bottom radius. Two courses of 9x2 % -in. fireclay straights and keys follow. The third course is made of 9-in. magnesite blocks of special shape to form circles of an inverted arch. It was started by a four piece keystone with skew-backs forming the outer course. The fourth course also started on a central keystone, or button, of four 90" segments, 12 in. diam x 13 Vz in. deep, and continued with 13%-in. blocks. Skewbacks at the shell completed the course to produce a horizontal surface for the side walls with a single course of No. 2 arch fireclay against the steel. Dry chrome-magnesite cement was brushed over each course after laying, and a 1-in. expansion space between the brick and the shell was filled with the same mixture. The total bottom thickness, excluding the castable material, was 5 in. of clay plus 22% in. of chemically-bonded magnesite. Tap Hole: A 5-in. OD and 3-in. ID silicon-carbide tube constitutes the tap hole and is set tangential to the upper course of the furnace bottom. Molten metal fills the tube when the furnace is level and filled to capacity. Side Walls: The lining, against the shell, consists of a 9x4Y2x3 in. soldier course of fireclay, using straights and No. 1 arches to turn the circles. A second soldier course of 9x4&apos;/2x2&apos;/2-in. fireclay was laid in a somewhat similar fashion. Three courses of 13Y2x6x3 in. and 9x6~3 in. of final magnesite, laid flat, completed the lining, using Nos. 1 and 2 keys to turn the circles. Cardboard spacers were placed between every two bricks in horizontal courses, and a thin coat of chrome-magnesite cement filled the joints between the firebrick and magnesite. A sprung-arch spanned the skim door with jambs of suitable magnesite shapes. Charge Slot: The slot is 3 ft 8 in. wide x 5 in. high. A silicon-carbide sill of special shapes has a 30" slope to allow cathodes to slide easily into the bath. The original arch was flat, and composed of Nos. 1 and 2 wedge magnesite with a 6-ft radius. It projected 5 in. over the sill, and, being a flat arch, gave an 18 15/16-in. opening between the inner edge and the metal line. The whole assembly was later raised 9 in., and the flat arch replaced with an arch, the lower edge of which maintained the 5-in, width from the outer to inner edges as shown in Fig. 2. A water-cooled, cast-copper jacket protects the steel shell behind the slot.

Jan 1, 1952

By D. H. Yardley

GEOCHEMICAL investigation of trace elements in surface materials was begun near Ely, Minn., in 1953 along the basal contact of Duluth gabbro with Giants Range granite (Fig. 1). This article presents data on the distribution of copper and nickel in till and in stream sediments in the area and proposes an explanation for the types of distribution found. The Duluth gabbro, one of the world&apos;s largest basic intrusives, intrudes rocks which range in age from Keewatin to middle Keweenawan. Within the test area the gabbro is in contact with granite except for short sections where it is in contact with remnants of iron formation. Sulfide mineralization occurs within the gabbro, near and parallel to the basal contact for a distance of several miles. Schwartz and Davidson&apos; have described the geologic setting of the mineralization. The sulfides, believed to be syngenetic, include chalcopyrite, cubanite, pentlandite, pyrrhotite, and minor amounts of bornite. They occur disseminated in the silicates and as small interstitial masses. The ratio of copper to nickel is about 3.8:1, based on 66 chemical analyses of rock samples from various outcrops (Ref. 1, p. 702, and Ref. 2). Test Procedures: With specified exceptions, all nickel and copper tests were made by the chromo-graph method,:&apos; which measures the intensity of a colored spot formed by a reaction between the metal being determined and special reagent paper. The intensity is then compared to the intensity of spots prepared from samples of known metal content. Details of the test procedure are outlined in another article (Ref. 4, pp. 77,78). All soil samples tested in this investigation to date have been weighed on an analytical balance. However, a volumetric scoop designed to provide about 0.1 g of soil adds to the speed and ease of testing and has been found to give satisfactory results (Ref. 5, p. 531, and Ref. 19). The size of the samples used for the tests was 0.1 g. Whenever such small samples are used there is some question as to whether they are representative of the several grams in the field sample. Many repeat tests of the samples used in this investigation demonstrated that results can be reproduced within the limits of accuracy of the method without formal mixing beyond that inherent in screening the soil fractions. Furthermore, the 0.1 g is probably as representative of the field sample as the field sample is of its area of influence. Hawkes and Lakin (Ref. 6, p. 291), who considered the general problem, compared ground and quartered bulk samples of 500 g with 5-g grab samples. They concluded that "there is no significant loss in accuracy of data by substituting grab samples for bulk samples." The term soil implies somewhat different things to the geologist, engineer, and soil scientist.&apos; For convenience the term as used in this article refers to unconsolidated material (the mantle) overlying bed rock. Sampling Procedure: Samples were taken at 100-ft intervals along north-south traverse lines across the gabbro-granite contact. The soil (till) samples were taken at an average depth of 1 ft, which was below the high-humus surface layer and into clean till. Samples taken at 1-ft intervals down to ledge showed as high a metal content at 1 ft below the air-surface as at greater depths and in two instances were slightly higher. The till at 1-ft depth did not appear to differ from material at greater depths. Total depth to bedrock has been tested at only a few points and where measured varied from 1 to 10 ft. Aerial Distribution Contours and Profiles: Plotting of copper, nickel, and cobalt content in contour form (Fig. 2) shows that anomalous amounts of these metal ions occur in till over and closely adjacent to mineralized areas of the gabbro. Contouring nickel content alone, or the copper content, outlines the same target area. Contours of the copper content provide a more distinct anomaly than nickel because of the higher copper concentration. The traverses are rather widely spread for interpolation; however, drilling has confirmed the target area essentially as shown.

Jan 1, 1959

By D. S. Gleason, D. E. Severson, D. R. Skidmore

Pittsburg and Midway Coal Co. has modified the German Pott-Broche process, on which patents date back to 1927, to produce on a bench scale liquid products by solution hydrogenation of coal. A continuing program of lignite solution-hydro gena-tion experiments is directed toward investigating coal solution reactions, determining favorable conditions for the solution refining of lignite by the Pott-Broche process, and investigating some of the uses for the de-ashed product obtained from lignite The German Pott-Broche process1" on which patents date back to 1927, has been modified by the Pittsburg and Midway Coal Co., a Gulf Oil subsidiary, to produce on a bench scale liquid products by solution -hydrogena-tion of coal." The objectives of the present effort are to investigate coal solution reactions, to determine favorable conditions for the solution refining of lignite by the Pott-Broche process, and to investigate some of the uses for the de-ashed product obtained from lignite. This paper is a summary of results to date in a continuing program of lignite solution-hydrogenation experiments. The coal solution reaction program has several principal aims. The first of these is to determine whether lignite can be successfully dissolved in solvents that might be practical for commercial development. The second object is to determine whether the solvents function after successive cycles of use, recovery, and reuse. It seems necessary to the economics of a potential commercial process that the solvent be recycled. Third, it is desired to learn something about the distribution of the ash constituents between cake and filtrate. The extent of ash removal is important. The nature and quantity of mineral matter passing through the filter may determine end-use marketability. For certain use applications, trace quantities of certain minerals can be objectionable, e.g., titanium and vanadium must be very low in electrode carbon for aluminum production. The Solution Reaction The coal solution Process involves an extremely complex system of chemical reactions. An initial solvent such as anthracene oil is a mixture of hundreds of different compounds with a boiling range of roughly 500" to 750°F at atmospheric pressure. The coal macro-molecule is broken down by thermal decomposition and solvent action into myriads of different compounds, some the same as those comprising the solvent. This similarity in structures opens up the possibility of production and subsequent recovery of solvent. Some solvent is inevitably lost by reaction. Regeneration of solvent was not a problem in the early German Pott-Broche plant. The coal refinery was an integral part of a petroleum refinery complex and replacement solvent was readily available. A coal refinery using lignite, however, might be isolated from other hydrocarbon processing facilities and the regenerability of solvent could be vital to the economic success of the venture. Several structural features of the solvent molecules have been cited as important to the coal solution process.&apos;. The first of these is aromaticity of the material, the second, ability to transfer hydrogen to another molecule, as for example the ability of tetralin to transfer hydrogen and be converted to naphthalene. Finally, the presence of hydroxyl groups on aromatic rings within the molecule, i.e., phenolic character, seems beneficial. Mixtures of pure compounds have been tried by various investigators. Mixtures of o-cresol, a phenolic substance, and tetralin were found to dissolve bituminous coal better than either substance alone.3 This maximum solubility was not found with lignite." Hydrogen contributes to the reaction by hydro-genolysis and by combining with free radicals and molecular "loose ends" to stabilize the compounds formed in coal depolymerization. High boiling point, and correspondingly high molecular weight, seems to be a property which improves solution potential for coal with a given type of compound.&apos; The maceral components of the coal appear to have an important bearing on its ease of solution. The fusain portion is quite inert to solvent action, but the an-thraxylon material dissolves quite readily.3 The hydrogenation reaction can be improved by the use of a catalyst; commercial hydrogenation catalysts having been found effective. Although cost is involved in the use of catalyst and catalyst recovery, the resulting saving in time and perhaps lowered temperature or pressure might justify their use in the solution refining process and decrease the total process costs. Apparatus and Procedure The coal solution runs were made in a 1-gal stainless steel stirred autoclave. The autoclave was provided with thermocouple wells and a transducer to permit continuous recording of temperature and pressure. The autoclave stirrer was magnetically driven, eliminating the leakage inherent with a rotating pressure seal. For runs in which a catalyst was used, the catalyst in the form of beads was placed in a wire mesh container mounted on the stirrer shaft. A control system programmed the heatup and reaction cycle. The permissible heating rate was 5°F per min because of the need to minimize thermal stress in the autoclave body. The temperature was raised at that rate until the reaction temperature was attained and then the temperature was held constant for the desired length of time. The maximum temperature seldom exceeded the average run temperature by more than 15°F.

Jan 1, 1971