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By John E. Conley

The slate industry of the United States has shown a marked decline in value of products made annually since the peak year 1925, although there has been moderate improvement over the lean years 1932 to 1934. In 1940, the value of all slate products sold was approximately 5¾ million dollars for a total output of 473,450 short tons. The quantity sold was a little less and the value a great deal lower than the annual averages for the past two decades. Enormous quantities of waste materials result from the quarrying, splitting and milling of slate. Obviously, utilization of such waste material is desirable, not only from the viewpoint of economical operation but also for the purpose of conserving natural resources. One possible use is to produce expanded or puffed lightweight aggregate for use in concretes. Such concretes are strong enough for specific uses, and the advantage of lightness could be obtained in such structures as the walls and roofs of buildings, suspension-bridge roadways and various other applications where weight is a factor. Conceivably, these uses could consume an appreciable annual tonnage. Use of Waste Slates for Making Aggregates The lightweight-aggregate industry has been discussed elsewhere4,5 and need not be covered in detail in this paper. The materials commonly utilized are naturally occurring aggregates such as volcanic cinder, tuff and pumice; by-product materials, such as ordinary cinders and air-cooled slags; and the specially manufactured products, including Haydite, Pottsco, Gravelite, Cel-seal, Superock and Waylite. Slates, however, have not been given the consideration they merit, especially in the United States. In the British Isles, some work has been done3 on certain Welsh slates by the Building Research Station to stimulate the use of these wastes. Requirements for Raw Materials As pointed out by Hughes,4 the lightweight-aggregate industry is segregated in regions adjacent to large cities in order to localize the market and avoid the necessity of long freight haulage. Waste slate from the deposits of eastern Pennsylvania, northern Maryland, and Virginia would meet the basic requirement of ally low-priced structural raw material. One apparent advantage is attained when slate waste, normally discarded, is utilized in the manufacture of a by-product. Mining or quarrying costs usually have been assessed against the main product. Some additional grinding might be desirable, but this operation would be relatively cheap, as most slates grind easily. It is believed that virtually all slates, when properly heated, will expand to some degree. However, types that will expand from 3 to 7 times their original volume are preferred. Fig, I illustratcs the expansion of a Pennsylvania slate. The cube originally

Jan 1, 1942

By Alexander Sattman, Anton Homatsch

Jan 1, 1894

By D. E. Baldwin

A method has been devetoped for predicting the produced concentration profile for a miscible slug in a five-spot pattern. The technique consists of dividing the five-spot into radial etements and applying an approximate radial solution of the dispersion equation to each element. To test the method, laboratory experiments were conducted in five-spot models. The models were 2-ft X 2-ft X 2-in. Berea sandstone slabs simulating one quarter of a five-spot. An aqueous system was used with tritiated water as the tracer fluid. Effiuent concentration profiles were obtained for slug sizes ranging from 0.7 to 10.2 per cent of the pore volume. Agreement between predicted and experimental profiles was excellent. A prediction was also made for a field tracer test. For this case the predictive technique was modified to account for stratification by use of a layer cake model. INTRODUCTION Dispersion in porous media is of growing interest to the petroleum industry because of the increasing importance of secondary recovery operations. One of the controlling factors in the recovery of oil by miscible displacement or by the use of waterflood additives is the degree of mixing between the fluids of interest. Consequently, one of the prerequisites of any recovery prediction is an adequate method of predicting dispersion. Also, any quantitative interpretation of waterflood tracer tests is dependent upon the ability to determine the dispersion of the tracer as it flows through the reservoir. During the past few years there have been reported in the literature the results of a large number of investigations, both theoretical and experimental, concerning dispersion in porous media. (An excellent review of these results has been presented by Perkins and Johnston." Only a few of these investigations, however, deal with non-lin-ear flow systems. Raimondi et al.9 have presented an approximate solution of the dispersion equation for radial, diverging flow. Lau et al. 7 and Bentsen 1 have presented laboratory data for radial systems and have shown that the dispersion can be adequately predicted by Raimondi's solution. Bentsen has also presented a small amount of data for a radial, converging system. Brigham and Smith' have applied Raimondi's solution, with approximating assumptions, to a five-spot pattern to predict the behavior of a field tracer test. The afore-mentioned investigations have been concerned almost entirely with radial diverging flow and have neglected the effect of converging flow. For this reason the present investigation was undertaken to develop a method of predicting dispersion in a five-spot pattern including the effects of both diverging and converging flow. The method which was developed consisted of dividing the five-spot into diverging-converging radial elements and applying an approximate solution of the disperson equation to each element. DEVELOPMENT OF PREDICTION METHOD The development of the prediction method consisted of two phases: construction of a radial model of a five-spot pattern and derivation of equations to describe the dispersion in the radial model. FIVE-SPOT MODEL A radial model which simulated the flow characteristics of a five-spot pattern was constructed by dividing the five-spot into diverging-converging radial elements (Fig. 1). Because of symmetry the model represented only an octant of a complete five-spot. Each element had an included angle of l° (making a total of 45 elements) and a radius r, such that the sum of the individual element areas was equal to the area of the five-spot octant. The normal five-spot fractional flow behavior was determined by averaging the experimental five-spot data of Caudle and Witte,' Dyes, Caudle and Erickson 5 and Fay and Prats. 4 The resulting curve is shown as a solid line in Fig. 2. To match this curve with the radial model it was

Jan 1, 1967

By James T. Gow, Oscar E. Harder, Anton de S. Brasunas

This paper deals with the results obtained and the techniques employed in determining: 1. Liquidus and solidus temperatures of the HH and HT type heat-resistant alloys. 2. The relation of true temperatures (thermocouple) to apparent temperatures (optical pyrometer) for the molten HH and HT type alloys. This work was the outgrowth of needing to know something definite about these temperatures in order to study the effect of pouring temperatures, shake-out times, and heat-treatments on the properties of some heat-resistant alloys. The information developed is thought to be of interest to industry for guidance in foundry practices and the industrial heat applications of the alloys. This paper is based upon work that has been done for the Alloy Casting Institute at Battellc Memorial Institute. Liquidus and Solidus Temperatures of HH and HT Type Alloys Experimenta1 Procedures Temperature-time curves were obtained during the cooling and solidification of several HH and HT type alloys. The metal was melted in an Ajax high-frequency, induction furnace and poured into a core-sand mold for cooling. The temperature measurements were obtained by means of a platinum, platinum-rhodium thermocouple contained within a fused silica sheath, which was extended through the side wall of the dry core-sand mold to a point at about the center of the mold cavity, which was 3 in. in diameter by 5 in. deep. The cold junction of the thermocouple was kept in an ice bottle, and the temperature was recorded automatically on a direct-reading potentiometer recorder. Through a switching arrangement, the temperature recorded was checked occasionally with a portable Leeds and North-rup potentiometer indicator. Less than a 5°F. variation in the two instruments was always found. This is within the limit of accuracy of readings. The thermocouple used was checked for accuracy against a standard couple, both before use on this work and after completion of the work, by the instrument laboratory. Temperature-time Cooling Curves Fig. I is an enlarged reproduction of a part of one of the cooling curves over the temperature range of 2600' to 2200°F. for an alloy of the HT type with 0.45 per cent carbon, 16 per cent chromium, and 36 per cent nickel. This is a typically shaped cooling curve as obtained for the alloys. It will be noted that two temperatures "re shown on Fig. I for the liquidus temperature. It is thought likely that the higher temperature, A, at the point at

Jan 1, 1945

By R. M. Brick, M. A. Williamson

CERTAIN features of the response of copper and brass to deformation and recrystallization remain obscure. The textures obtained on rolled sheet are listed by Schmid and Boas1 as: No adequate explanation has been advanced for the presence of the secondary deformation texture of copper and its absence in brass or for the origin of the cubic recrystallization texture of copper.2 A recent publication3 attempted to rationalize the above empirically described annealing texture of brass on the basis of retention of the (110) [112] rolling orientation plus twinning about the central octahedral poles of the usual secondary (110) [115] orientation. The difference in the textures of annealed copper as compared to 70-30 brass is also shown by the difference in the position of ears for drawn copper, at o° and 90°, as compared with drawn brasses, at 45°, for all compositions between 90-10 and 70-30.4 Data on individual crystals are as yet too few to permit of an analysis of these textures comparable to that of mild steel as made by Barrett and Levenson.5 EXPERITMENTAL METHODS The materials studied included copper specimens cut from the central, equiaxed crystal portion of a tough-pitch wirebar; brass made from this copper and high-purity (99.99 per cent) zinc including alloys of 7 per. cent, 15 per cent, 24 per cent, 29 per cent and 36 per cent Zn and large single crystals of 70-30 brass. Sections ¼ in. thick of the cast polycrystalline copper and brasses were reduced 40 per cent by rolling, recrystallized to a common grain size of 0.045 mm. and then rolled directly to foil 0.002 in. thick. Hardness readings employing both the Vickers and Rockwell machines were taken at intervals during the reduction. For comparison, similar specimens were cross-rolled-i.e., turned through 90 for each pass-and comparable hardness data were secured. Finally sections from the single crystals of brass were cut to obtain the following initial orientations; these sections were rolled, hardness data secured and a complete micrographic record of structural changes obtained:

Jan 1, 1941

By O. D. Gonzalez, H. A. Wriedt

Alloys ranging from pure iron to pure nickel were saturated with nitrogen gas at 918°, 999°, and 1217°C and analyzed. The solubility of nitrogen at 1-atm pressure was obtained as a function of nickel content for the whole range of iron-nickel alloys and the temperature dependence of solubility is shown for the experimental alloy compositions. The heat of solution and the entropy of solution of nitrogen in the alloys as func-tiom of composition up to 41 pct Ni are given. 1 HE acquisition of more exact thermodynamic information for systems composed of iron with various alloying elements continues to interest ferrous metallurgists. The interstitial alloying elements, especially carbon, hydrogen, and nitrogen, comprise a field of particularly intense study. A recent investigation of carbon solubility in Fe-Ni alloys held at 1000°C in gases of fixed carbon activity showed an unpredicted minimum in the solubility near 72 pct Ni.1 The present investigation was undertaken to see if a similar anomaly in solubility existed for nitrogen in Fe-Ni alloys. The solubility of nitrogen in ? iron at known nitrogen activities has been measured several times.&apos;-&apos; Corresponding information for the solubility of nitrogen in solid Ni and Fe-Ni alloys is not available. Juza and sachsze9 reported that metallic nickel dissolved 0.07 pct N in equilibrium with Ni3N at 445°C. This result was regarded as dubious by Turkdogan and Ignatowicz10 who found less than 0.0004 pct N in nickel specimens equilibrated at 600°C with ammonia-hydrogen mixtures having fugacities of N2 gas up to 35,000 atm. Such fragmentary reports indicate only that nitrogen gas has a very much lower solubility in nickel than in iron. MATERIALS A purified iron from Battelle Memorial Institute was used for solubility measurements. This grade contained more than 99.95 pct Fe, the highest content of any measured individual impurity being 50 ppm for each of As, C, Cu, and Pb. The nickel and iron-nickel alloys, supplied by the International Nickel Co., contained 0.02 to 0.04 pct C, 0.17 to 0.19 pct Mn, 0.04 to 0.08 pct Si, < 0.01 pct P, <0.03 pct S, and <0.20 pct Co as principal impurities. The Fe-0.59 pct Mn alloy was of zone-melted quality, the highest concentration of any individual impurity being 15 ppm Al. The "prepurified" grade of nitrogen gas from Matheson Co. had a nominal nitrogen content of 99.996 pct. An "electrolytic" grade of hydrogen gas from National Cylinder Gas. CO. in which the principal impurity is water vapor was admixed with the nitrogen to prevent oxidation of these specimens. This grade was preferred to a purer grade because the presence of some oxygen helps in suppressing contamination of the specimens, for instance, by silicon from the porcelain furnace tube. EXPERIMENTAL PROCEDURE The specimens were equilibrated in vertical tube furnaces through which a 99 pct N2-1 pct H2 gas was passed at ambient pressure. For the 915", 918", and 999°C runs, a wound-resistance furnace was used, and for the 1217°C runs, a furnace with a tubular silicon carbide element was used. The temperature control system on both furnaces was similar: a Weston "Celectray" controller with a modulator, responding to the electromotive force of a thermocouple with its hot junction in the furnace, switched the furnace current between high and low settings. The temperature distributions in the working zones of the furnaces were determined with calibrated Pt/Pt-10 pct Rh thermocouples. The overall uncertainty in temperature was estimated to be ±3° at 915° to 999°C and ±5° at 1217°C . Consideration of the small temperature coefficients of the equilibria studied, together with the error limits in the method of analyzing for nitrogen, showed that further refinement in temperature uniformity would give no improvement in resultant accuracy. The reactive gas was generated by mixing together "prepurified" nitrogen and "electrolytic" hydrogen directly from cylinders commercially supplied. The flow rates of nitrogen and of hydrogen were fixed by setting with bleeders the pressure drops across the individual calibrated flow meters of constant resistance. For any given alloy, the solubility of nitrogen at 1-atm pressure (% N) was obtained from the analyzed nitrogen content of equilibrated specimens (% Na)according to Sieverts&apos; law, thus

Jan 1, 1962

By T. R. Cass, M. R. Achter

A technique for measuring the creep and rupture strength of nickel specimens bonded by sintered oxide layers has been developed for the investigation of the role of grain-boundary oxide in the oxidation strengthening of nickel during creep at 817 °C. Prom the inverse dependence of the rate of bonding on the thickness of the starting oxide, it is suggested that the rate of diffusion of Ni++ through NiO is the controlling process at the start of sintering. Cornparisons of the creep and rupture strength and of the microstructure of sintered couples with conventionul creep specimens suggest rupture in nickel is prevented by load-bearikzg, ikltergranular layers of oxide. A number of recent investigations have shown that at high temperature and low stress the creep-rupture strength of nickel and nickel-base alloys may be much greater in air than in an inert gas or vacuum.&apos;-7 For example, at 817°C and 2500 psi, nickep failed in vacuum in 48 hr but not in air after 4000 hr, and a Ni-Cr-A1 alloy5 at 1038°C and 3000 psi failed in vacuum in 4 hr but not in air after 731 hr. Fig. 1 is a typical creep curve in air for materials experiencing this pronounced environmental strengthening. A common feature of these curves is the entry into a third stage of accelerating creep, concurrent with the for-

Jan 1, 1962

By Robert E. Brand, Strathmore R. B. Cooke

For a study of the bonding that occurs in magnetite pellets during oxidation, cubes of magnetite, hematite, and quartz were prepared, each with one surface polished and nearly optically plane. These faces were fired together under pressure and at various temperatures to determine whether bonding would occur. Sections of surfaces that bonded were examined microscopically. An attempt was made to measure the strength effected by this treatment. EARLIER work1,2 has shown that the strength developed by fired magnetite pellets containing silica or silicates is due to three separate effects. Two of these are associated with the formation of slag in the interior of each pellet at temperatures above 1000°C, where the atmosphere is substantially though temporarily neutral. The third is independent of the silica content and is responsible for the development of a hard shell around the pellet during the initial stages of firing, a condition which permits the pellet to assume the increasing load of the overlying burden and prevents extrusion of the slagged and semi-liquid interior. Pellets made from hematite concentrates do not show these effects unless specific additives are incorporated, and it is these differences in behavior which in large part account for the difficulty in producing satisfactory pellets from such concentrates. Elsewhere it has been stated that "the most important factor responsible for the development of strength in fired magnetite pellets containing no additive is the oxidation of magnetite to hematite and the subsequent recrystallization and grain growth of this constituent."&apos; The object of the investigation reported in the present paper was to study in detail the mechanism described above and to evaluate its contribution to the strength of the fired pellet compared with other possible mechanisms, excluding slag bonding. This was effected by placing together polished faces of two cubes prepared from specific minerals and then heating them under constant pressure at different temperatures both in air and in nitrogen. The materials used were magnetite, hematite, and quartz. The magnetite was coarsely crystalline material from western Montana, the hematite was Itabira ore from Brazil, and the quartz was of optical grade but of unknown origin. A high-grade Lebanon (Pennsylvania) magnetite concentrate was used to prepare briquets for the investigation of bonding strength. Analyses of the iron oxide materials are given in Table I. The specimens for furnace treatment were prepared with a carborundum cut-off wheel, which cut the minerals into cubes approximating % in. along each edge. Care was taken to make the sides as parallel as possible. The edges of the faces to be polished were beveled, and the cubes were mounted in brass rings with sealing wax, this being essential to prevent the specimens from rocking during polishing. To prevent excessive friction, the wax was recessed so that only the brass ring and the face of each specimen would be in contact with the polishing lap. Each specimen was then ground with No. 600 carborundum to produce a level surface before polishing. The Graton-Vanderwilt process was not used for polishing because of the long time required by this method, but several other methods were tried. Cloth laps using levigated alumina or chromium oxide did not produce the desired polish. Satisfactory results were obtained by means of a pitch lap: which was used for all the mineral cubes, with levigated alumina as the polishing medium. The lap was operated in a machine similar to those used in the preparation of single optical surfaces. It was run at low speeds and moderate pressures, and polishing was stopped occasionally to prevent overheating of the lap. A good polish with flat surfaces could be obtained in about 5 min. After polishing the sealing wax was removed either in hot water or in a solvent such as turpentine.

Jan 1, 1955

By Henry F. Cooper, Lee Burford, John C. Thompson

In the past year or two, considerable effort has been expended to calculate the spherical wave propagation phenomena associated with explosions in a "hard rock" medium (Godfrey, 1969; McKay and Godfrey, 1969; Burford, et al, 1970). This effort was stimulated primarily by the recognition that early calculations predicted peak particle displacements which seriouly disagreed with field data. In most of the calculations mentioned above, rock was modeled as an inelastic continuum (see for example Isenberg, et al, 1970 for some recent results using "classical" elastic-plastic material property models). However, several investigators (Handin, et al, 1967; Cherry, et al, 1967) have shown that the simple classical yield surface models cannot simultaneously account for the observed brittle failures of intact materials under compression, extension and torsion. One approach to more correct modeling of dynamic rock behavior is to make the inelastic model more and more sophisticated. For ex- ample, recent experiments at the University of Utah (Swanson, 1969) have suggested possible modifications to the yield surface model that could better account for the ductile behavior of intact granite sam- ples under various states of stress. The approach followed in this paper proceeds in an entirely different direction. We want the physics which is built into our conception of rock to be as simple as possible while at the same time reproducing the most important features of rock&apos;s behavior. Therefore, we abandon the requirement that rock be treated strictly as a continuum. Instead we consider a "discontinuous" model in which one of the most significant characteristics of rock is treated..

Jan 1, 1971

By B. H. Bergstrom, D. D. Crabtree, C. L. Sollenberger

Experimental data is given for several natural minerals on the load required for the fracture of single specimens over a very wide size range. The experimental results are compared with published data, such as for coal specimens from 30 to 64 in. in size. Correlations of the specific energy (per unit mass) required for fracture with specimen size are given as functions of specimen shape and the experimentally observed load-size relationships. These correlations are discussed with reference to comminution theory. Comminution, through compression or impact, can result only if sufficient force, or energy, is applied to a specimen of a given size. The relation of specimen size to force requirements has not been adequately considered previously. For single particle fracture, it was previously shown 1 that the specific energy (energy, e, per unit mass, M) at fracture of single glass spheres was inversely proportional to the product size modulus, k, as indicated in the equation: e/M = C/k [1] This was found true not only for numerous glass specimens of the same nominal size but was true for all nominal sizes tested from 1/8 to 1 in. Also, it was observed that the smaller spheres required a greater specific energy at fracture than the larger spheres and consequently had a smaller product size modulus. However, the relationship giving the required fracture energy as a function of specimen size was not further elaborated upon. This paper will extend the development of the feed size effect on specific fracture energy for brittle materials. SPECIFIC ENERGY-FEED SlZE RELATIONSHIPS FROM PREVIOUS DATA For the particular glass spheres tested in the previously described experiments,&apos; it can be shown, as on Fig. 1, that the specific energy varied inversely as the 4/3 power approximately of the feed size, F, in the span of nominal sizes ranging from 1/8 to 1 in. The load at which a 3-in. diam glass sphere (a crystal gazer&apos;s ball) was broken was within the range predicted by an extrapolation of this energy-feed size relationship. The product size distribution from this

Jan 1, 1963

By E. C. Wright

THE bessemer process is nearly one hundred years old. William Kelly, the American inventor, was able to demonstrate that he had accomplished the pneumatic purification of molten pig iron as early as 1847; Henry Bessemer&apos;s English patent was granted in 1855. The application of this steelmaking process since then and the present position of bessemer steel are concisely discussed on pages 355 and 356 of the fifth. edition of "The . Making, Shaping and Treating of Steel." The great expansion of the automotive, machine-tool, chemical and oil industries after 1910 opened up a large market for special steels, whereas the output of steel before that time was mainly for structural purposes. The need for building more blast-furnace capacity complementary to bessemer operations, combined with the availability of the necessary amounts of steel scrap, both contributed to the extension of open-hearth plants at the expense of bessemer capacity. The ability of the open-hearth furnaces to consume virtually all types of raw materials as far as phosphorus content is concerned, and the need for selecting low-phosphorus ores of rapidly decreasing sources for the bessemer, also played a part in the reduction of bessemer steelmaking operations and the increase in basic open-hearth plants. As a consequence, the proportion of bessemer steel made in this country decreased from 65.7 per cent of the total in 1900 to 37.1 per cent in 1910, and only 6.8 per cent of the total in 1937. The metallurgical concepts existent between 1910 and 1920 caused the writing of many specifications of very low phosphorus and sulphur content, which definitely barred bessemer steel from many commodities. Now sulphur is being added to many open-hearth and bessemer steels in amounts far exceeding the maximum sulphur content of good bessemer steel. Rephosphorizing is also widely practiced for several specific applications. Even today, the chemical-analysis limits of many important steel specifications have an arbitrary and archaic tinge. The old standard "0.30 .to 0.60 per cent manganese" might be mentioned in this connection as similar to the rigid requirements for maximum sulphur and phosphorus that were established at least 25 years ago. The literature is full of vague and complex discussions of the effects of nitrogen and oxygen where these elements existed in steel in amounts less than 0.02 per cent. Many failures have been attributed to these elements without reasonable proof on which to base the conclusion. Methods of determining oxygen in steel require such complicated equipment that only a few research laboratories have the required apparatus. Even so, the accuracy of oxygen determinations in steel has been only recently reproducible. Up to the present time, no one has succeeded in obtaining a sample of molten steel that accurately indicates the full. oxygen content of the metal in the furnace or ladle because the molten sample

Jan 1, 1944

By R. G. Clay, C. F. Seaman

THE mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 05 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that time, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to he used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered, the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. ORIGINAL EQUIPMENT AND INSTALLATION A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would he available at each shaft for places where there were no regular blasting stations and could he used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots and the up position would connect it to the line from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1942

By D. L. Murphy

As the world&apos;s demand for iron ore increases and available sources are depleted, increasing attention will be centered on deposits in the Quebec and Labrador portions of the Labrador Trough. Improvements in the beneficiation of low grade iron ore stimulate exploration for deposits amenable to concentration. This study is centered on the Mt. Wright-Lake Carheil area, lying between Wabush Lake and Mt. Reed in Quebec. Described in detail are the general lithology, structural and economic geology, and mineralogy of the iron-formation. The iron ore province of Quebec and Baffin Island extends from the southern shore of Baffin Island southeasterly through the Labrador Trough to Matonipi Lake, approximately 185 miles northwest of Seven Islands. (See Fig. 1.) Future geologic field studies will very likely extend the limits of the province to include the ferruginous and associated rocks of the Lake Mistassini area. At present, only the deposits near Knob Lake and Lake Jeannine (Mt. Reed) have been brought into production although in the near future several other properties including those at Wabush Lake will commence operations. Within the next decade or so, as the world&apos;s demand for iron ore increases and available sources are depleted, this region will undoubtedly become a major mining district. Initial reports of iron-rich rocks in Labrador date back to the investigations undertaken by A. P. Low near the turn of the century. During the ensuing years the country was prospected chiefly for precious and nonferrous base metals because of its inaccessibility and the low tenor of the iron ore. In 1945 an evaluation of the direct shipping ores near Knob Lake indicated that sufficient reserves were present to justify the enormous capital expenditure required to bring the deposits into production. Initial open-pit mining began in 1954 following the construction of a townsite, a 360-mile long railway to Seven Islands, and an elaborate loading dock. Recent improvements in the beneficiation of low-grade iron ore have stimulated widespread exploration for deposits amenable to concentration, particularly along the southwest extension of the Trough. The Mt. Wright-Lake Carheil area, bounded by longitudes 66O45&apos; and 67 30&apos; west and latitudes 52 30&apos; and 52 45&apos; north, is a nearly rectangular tract of slightly less than 500 sq miles. It is in Saguenay County, 174 miles north-northwest of Seven Islands (see Fig. 1). Except for the periods of the spring breakup and the fall freeze up, the area is readily accessible by air.

Jan 1, 1962

By J. K. Stanley

Iron-cobalt alloys in the range of 35-50 pct cobalt are of interest in the electrical industry because they possess the highest magnetic saturation of any magnetic material known. l1,2The magnetic saturation induction of the 35 pct alloy is 24,200 gausses, compared to 21,600 for pure iron (Fig I). This difference is significant in electrical apparatus, whcre weight or space is at a premium. Saturation values higher than that for iron can be attained up to about 70 pct CO hut economics restrict the use to as low a cobalt content as possible without sacrifice of saturation. Fig 2 shows the iron-cobalt equilibrium diagram. Up to the present time, two factors have retarded the use of iron-cobalt alloys in the electrical industry for magnetic purposes.? One is the high cost of cobalt ($I.50 per pound) and the other ig the inherent brittleness of the binary iron-cobalt alloys. Straight iron-cobalt alloys up to 25 pct cobalt are ductile and easy to process to thin sheet or strip by conventional methods of hot and cold rolling, but the alloys containing higher percentages of cobalt, including those for which the saturation magnetization is a maximum (35 to 40 pct), are increasingly brittle with increasing cobalt, at least up to 50 pct, and are consequently difficult to process, either hot or cold. Work by White and Rah13 indicated that iron-cobalt alloys could be made ductile by alloying with vanadium, and after quenching the alloys (vanadium up to about 4 pct) were found to he colt1 rollable to very thin gauges. With this as a background, we resumed our work on these alloys with the objective of developing methods for economical commercial production of wide strip. For this purpose we have concentrated on the 35 pct alloy in order to obtain the highest possible saturation value. Experiments were carried out on many alloy additions to the iron-cobalt alloys and on the response of these compositions to heat treatments. This work finally culminated in a commercially producible iron-cobalt alloy which has been named Hiperco.* Materials StudIEd and Procedur&apos; Binary alloys of iron and cobalt in the range of 35 pct cobalt are difficult to hot work (by forging or rolling) without such additions as C, Mn, Si, V, Cr, Mo, W Or perhaps others. Another beneficial effect of such additions is that they increase to a greater or less extent the strangely low electrical resistivity of these alloys about 10 microohms per cm3

Jan 1, 1948

By J. K. Stanley

Iron-cobalt alloys in the range of 35-50 pct cobalt are of interest in the electrical industry because they possess the highest magnetic saturation of any magnetic material known. l1,2The magnetic saturation induction of the 35 pct alloy is 24,200 gausses, compared to 21,600 for pure iron (Fig I). This difference is significant in electrical apparatus, whcre weight or space is at a premium. Saturation values higher than that for iron can be attained up to about 70 pct CO hut economics restrict the use to as low a cobalt content as possible without sacrifice of saturation. Fig 2 shows the iron-cobalt equilibrium diagram. Up to the present time, two factors have retarded the use of iron-cobalt alloys in the electrical industry for magnetic purposes.? One is the high cost of cobalt ($I.50 per pound) and the other ig the inherent brittleness of the binary iron-cobalt alloys. Straight iron-cobalt alloys up to 25 pct cobalt are ductile and easy to process to thin sheet or strip by conventional methods of hot and cold rolling, but the alloys containing higher percentages of cobalt, including those for which the saturation magnetization is a maximum (35 to 40 pct), are increasingly brittle with increasing cobalt, at least up to 50 pct, and are consequently difficult to process, either hot or cold. Work by White and Rah13 indicated that iron-cobalt alloys could be made ductile by alloying with vanadium, and after quenching the alloys (vanadium up to about 4 pct) were found to he colt1 rollable to very thin gauges. With this as a background, we resumed our work on these alloys with the objective of developing methods for economical commercial production of wide strip. For this purpose we have concentrated on the 35 pct alloy in order to obtain the highest possible saturation value. Experiments were carried out on many alloy additions to the iron-cobalt alloys and on the response of these compositions to heat treatments. This work finally culminated in a commercially producible iron-cobalt alloy which has been named Hiperco.* Materials StudIEd and Procedur&apos; Binary alloys of iron and cobalt in the range of 35 pct cobalt are difficult to hot work (by forging or rolling) without such additions as C, Mn, Si, V, Cr, Mo, W Or perhaps others. Another beneficial effect of such additions is that they increase to a greater or less extent the strangely low electrical resistivity of these alloys about 10 microohms per cm3

Jan 1, 1948

By J. B. NEALEY

THE common method of shale mining, drilling, shooting and steam-shovel loading, is fast giving place to a comparatively new method which is purely mechanical. This machine is known as the shale planer and takes a cut from a very large face, 50 ft. or more in height. As shale deposits are usually stratified, the quality varying with the stratification, a very uniform product is obtained with this planer, for its "cut" is across the entire face, vertically. This planer was designed and built some 15 years ago by J. M. Powell, president of the Indiana Drain Tile Co., Brooklyn, Ind. Unevenness in mixing the material from the different layers in the shale deposit of this company was causing considerable fluctuation in the firing- shrinkage and color of the ware produced, thus lowering the quality. Some method of maintaining a uniform proportion was vital and this planer was therefore devised. Its capacity was approximately 175 tons daily: cut, mixed and loaded.

Jan 1, 1929

By AIME AIME

POSSIBILITY of a greatly increased manganese production from domestic source; is indicated by news that the Bureau of Mines has been successful in producing concentrates with high manganese content from the low-grade deposit, in the Las Vegas, Nev.) area. The first unit of the comprehensive group of pilot plants built by the Bureau of Mines at Boulder City, Nev., under a defense appropriation has started operations, demonstrating definitely the applicability of a process developed in the Bureau laboratories. About 12 1/2 lb. of manganese is needed for the production of every ton of steel, and substitution is difficult. The United States produces only about 3 per cent of its total needs of this ore. Only high-grade ore, such as the Russian ore, containing 18 per cent or more manganese, is generally used by American producers in making ferro¬manganese the form in which manganese is usually added in steel manu-

Jan 1, 1941

By Irving Cadoff, Kurt Komarek, Edward Miller

THE lead-telluride phase diagram as compiled by ansen&apos; has one intermetallic compound, PbTe, with solid-solubility limits extending from 22 to 45 wt pct Te (31.4 to 57.1 at. pct). Pelzel2 redetermined the solubility limits on the lead-rich side of the compound by measuring the relative areas of the two phases in samples annealed in the liquid + compound region, and obtained data that agreed with that of Hansen. Recent studies of the semiconducting properties of PbTe3 however, have indicated that addition of either lead or tellurium to PbTe does not produce samples having carrier concentrations greater than about 10 19 cm-3. Assuming that each impurity atom results in one ionized charge carrier, this carrier concentration corresponds to a solubility range of less than 0.04 at. pct. Fritts and Karrar4 have also observed two phases in samples containing from 0.15 to 3.0 pct excess Pb. To resolve the discrepancy between these data, the solid solubility limits of the compound PbTe were reinvestigated, using several other techniques. Ten 25-g samples with compositions of 1/2, 1, 3, 5, and 10 wt pct excess Pb and Te and two 100-g samples of 15 and 65 wt pct Te were prepared, sealed in evacuated quartz capsules, and melted at 950oC. The 1 pct excess Pb and Te and the 15 and 65 wt pct Te alloys were homogenized for 96 hr at 450°C; the remaining specimens were annealed 50°C below the solidus temperature reported in Hansen. The specimens were polished and etched with dilute HNO,. All the samples were two phase, the second phase appearing at grain boundaries or as spheroidized inclusions in the grains. In the case of the tellurium rich specimens, the second phase could be identified as a eutectic. Debye-Scherrer powder diffraction patterns of the i, 1, 3, 5, and 10 pct excess samples were obtained using CuKa filtered radiation. The films were indexed as the B1 or NaCl structure. The calculated lattice parameter of PbTe was the same for all the specimens, with an average of a0 = 6.460 ± 0.0005A. Lines due to excess lead or tellurium were very faint but could be clearly seen in all the films. The relative areas of the two phases in the 15 and 65 wt pct Te samples were corrected for the differences in density of Pb, Te, and PbTe, and taken as proportional to the relative volumes of the two phases. The corrected relative volumes agreed with a phase diagram having negligible solubility of both lead and tellurium in lead tellurium. The density of the two specimens was obtained to give a third independent determination of the phase limits. The specimens were broken into small pieces to eliminate any internal porosity, and adherent air bubbles were removed by vigorous shaking under water for several hours. The density of lead (11.34 g per cm3) and of tellurium (6.25 g per cm3) was taken from the International Critical Tables and that of stoichiometric PbTe (8.25 g per cm3) calculated from the lattice parameter. For the 15 wt pct Te samples, the density assuming no solubility was calculated to be 9.89 g per cm3; the observed density was 9.86 g per cm8. For the 65 wt pct Te alloy, the density calculated for the case of no solubility was 7.24 g per cm3; the measured density wag 7.22 g per cm3. These results are consistent with those of planimetry and X-ray. SUMMARY Data have been presented which show that the solid solubility of both lead and tellurium in PbTe is

Jan 1, 1961

By J. J. Pfeiffer, A. A. Oming, J. W. Myers

The method used to process washery refuse for use as a building material aggregate is described. Results of studies made in investigating this process are summarized. The Bureau of Mines, in cooperation with two commercial coal producers, studied the feasibility of using coal washery refuse in manufacturing lightweight aggregate for building materials. Lightweight aggregate is used in making masonry products of lower bulk density than can be produced from conventional aggregate such as sand, gravel, and crushed rock. Among the advantages of lightweight masonry are reduced deadweight loads which reduce materials cost, lower cost of construction labor, and improved thermal and sound insulation. A process for making lightweight concrete aggregate from coal-washery refuse appeared attractive because of the availability of raw material which contained sufficient fuel to carry out the processing. Vast quantities of such refuse have accumulated throughout the coal-producing areas of the U.S. Generally it is stored in unsightly piles, creating a problem of air pollution and using space that could be devoted to other purposes. Conversion of the refuse to lightweight aggregate could remedy these conditions and effectively utilize a material normally considered waste. The first part of the program was conducted in cooperation with the Truax-Traer Coal Co. to determine the feasibility of utilizing refuse from the company&apos;s coal preparation plant at Ceredo, W. Va., in manufacturing lightweight aggregate. The aggregate was to be used in making cement blocks comparable to those known in the building trades as cinder blocks. This project was completed in 1954. Based on the results of this work, the cooperator began commercial development and formed the Trulite Corp. for this purpose. A plant capable of handling 120 tpd of refuse was constructed for approximately $120,000. It began operating on June 1, 1955.&apos; In 1959 a similar study was made, under a cooperative agreement with the Carbon Fuel Co. of Charleston, W. Va., to rrscertain the suitability of refuse from its plant for conversion to lightweight aggregate. Under this program various modifications of equipment were made, and the investigation of process variables begun in the previous project was extended. PROCESSING METHOD The formation of lightweight aggregate from coal-washery refuse js governed, in part, by the fusion characteristics of the mineral constituents of the refuse when heated above their softening temperatures. The heat required for this purpose is supplied by the combustihle matter in the refuse. Some of this combustible matter is coal with high ash content, but most of it is carbonaceous shale. Whereas the density of aggregate prepared from clay or shale depends largely on expansion, the density of aggregate made from refuse is reduced by burning out the combustible matter and, to a lesser extent, by expansion of the residue in the fuel bed. A chain-grate stoker with an updraft air system was chosen for the conversion process in preference to a downdraft sintering machine or a rotary kiln. In down-draft sintering, which operates on the overfeed burning principle, the volatile matter from the refuse would distill off without burning, resulting in inefficient combustion and interfering with proper operation of the equipment. The rotary kiln is designed primarily for processing material that lacks, or is low in, combustihle matter with the heat of reaction supplied from external sources. Washery refuse, which is relatively high in carbonaceous material, could burn at the wall of the kiln causing clinker formation with resultant decrease in throughput and increase in production cost. With an up-draft air stream on a chain or traveling-grate stoker, the volatile matter would be carried into the ignition and burning zones where the heat could be utilized more efficiently. The combustion procedure in this type of unit may be considered as consisting of four periods: 1) initial ignition of the top of the bed by furnace radiation, 2) ignition travel through the bed until the whole bed is ignited, 3) burnout of the combustible material with high air rates to fuse and expand the residue, 4) a cooling period. Investigation of the process required a study of the variables that affect both

Jan 1, 1962

By FREDERICK LAISY

A NUMBER of new plants for the treatment of copper ores were completed or under construction during the year. Among these may be mentioned the plants of the International Nickel Co., those of the Hudson Bay Mining and Smelting Co. and of the Roan Antelope Copper Co. Three copper re- fineries have been constructed, one at El Paso, Tex., and the other two in Canada. Two more are projected, one in France and one in England. The International Nickel Co.&apos;s plant consists of an 8000-ton concentrator and a smelting plant containing thirty 10-hearth Herreshoff roasting furnaces, five reverberatory furnaces, and eight Smith-Pierce converter5 . The smelting plant is estimated to have a capacity of 5000 tons of ore or concentrate per day. The roasters are superimposed on the reverbratory furnace plant. As a consequence of this arrangement, the furnace building is unusually high. being 183 ft. from yard level to highest roof elevation.

Jan 1, 1931