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By Henry Louis

IN the discussion on the paper on subsidence by R. V. Norris and H. W. Montz (Teohnical Publication No. 153), H. N. Eavenson has been good enough to quote some of my views regarding the phenomena of subsidence; he has, of course, only mentioned these briefly, and as he has not represented them quite correctly, it is evident that I have failed to make them clear. Seeing that he, and apparently some of the others who took part in the discussion, appear to hold views different from my own, I think it advisable to put my opinions down as definitely and as clearly as possible, because it appears to me at least possible that some of the differences above referred to may be differences of language and definition rather than of fact. The statements I propose to submit refer to the simplest cases, namely where an area in a practically horizontal bed has been entirely worked out. My experience has been confined to beds of bituminous coal and of ironstone, but D I do not at the moment see why there should be any difference in principle between the working of a bed of anthracite and one of bituminous coal, provided that the other conditions are the same. I wish to emphasize that I am considering the conditions of complete working out of the mineral bed with the roof allowed to come down and form goaf in accordance with the general British practice; 1 am excluding, therefore, methods of hydraulic, pneumatic, and hand stowage where extraneous material is brought in to hold up the roof. I am also excluding the case of bord and pillar working where pillars of small size are deliberately left in, apparently with the idea that these will subsequently crush and let the roof down gradually. Such a method of mining has long been discontinued in this country, indeed in Europe generally, and I venture to predict that the time will come when such a recklessly wasteful method will no longer be tolerated even in the United States. Further, I am referring entirely td static conditions after the mineral bed has been worked out and the ground has come to rest; the phenomena that occur during the course of extraction are much more complicated, and these I am not discussing here. Under these conditions, I have

Jan 1, 1929

By F. F. Schmidt, H. R. Ogden, E. S. Bartlett

Continuing tantalum alloy development studies have been concerned with a more detailed investigation of promising binary, ternary, and more complex tantalum alloys containing Groups IV-A, V-A, VI -A. VII-A, and VII-A metal additions. A primary objective of these studies was to improve the elevated temperature strength of tantalum with minimum degradation of fabricability and low-temperature ductility by alloying. Several alloys were found to exhibit hot tensile strengths in the excess of 25, 15, and 10 ksi at 2700°, 3000°, and 3500°F, respectively, combined with excellent cryogenic toughness and ductility. Particularly outstanding in this respect are alloys in the system Ta-Mo-W. Ternary parameter curves describing hot-strength, low-temperature ductility, recrystallization temperature, weld ductility, and fabricability are presented for the Ta-Cb-V, Ta-Hf-W, Ta-Mo-Hf, Ta-Mo-V, Ta-Mo-W, and Ta-V-W systems. At present, only four metals qualify for large scale structural applications at temperatures above about 1800°F. These are tungsten, tantalum, molybdenum, and columbium. Research and development work on these refractory metals has progressed to the point where efforts are now primarily directed toward the improvement of high- and low-temperature mechanical properties through alloying. High melting point combined with excellent low-temperature ductility, fabricability, toughness, weldability, and high solubility and tolerance for both substitutional and interstitial solutes are important attributes of tantalum. Conversely, its high density and relative scarcity compared with other refractory metals are the major deterrents to the use of tantalum. Early studies in the Battelle alloy development program established base line data on the mechanical and metallurgical properties of high-purity tantalum.' These studies were extended to include a preliminary investigation of the effects of various alloying elements on the established base properties.2,3 Recent studies involved a more detailed investigation of promising binary alloys containing Cb, Hf, Mo, Os, Re, Ru, and W, and ternary Ta-Cb-V, Ta-Hf-W, Ta-Mo-Hf, Ta-Mo-V, Ta-Mo-W, and Ta-V-W alloys. Results of major importance obtained are given in this article. EXPERIMENTAL PROCEDURES Electron-beam-melted tantalum (Wah Chang) and high-purity alloying elements were consolidated as 150-g ingots by multiple nonconsumable arc melting under a partial atmosphere of helium. Homogeneity of the button ingots was checked radiographically. Alloy buttons measured about 1-1/2 to 1-3/4 in. in diam by about 0.35 and 0.50 in. thick. Nominal alloy compositions were checked on each alloy button by weighing after arc melting. In most instances, weight losses were within 3 pct of the total weight charged. Subsequent chemical analyses on a number of selected alloys showed interstitial contents to range as shown below: Range of Interstitial Content, ppm 0 C N H 25-75 50-100 10-50 1-5 Metallic alloying additions were found to agree well with nominal compositions based on weight change data. Fabrication directly from the cast button to sheet was conducted at low (75° to 900°F), intermediate (1800° to 2000°F), or high (3000°F) temperatures, depending upon alloy content and hardness. Evacuated stainless steel or molybdenum packs were used for fabrication at intermediate and high temperatures, respectively. The temperature required to complete recrystal-lization in one hour was determined by hardness measurements and metallographic observations (longitudinal test sections) after vacuum annealing at 180°F intervals in the range 2010° to 2910°F. All mechanical property evaluations were conducted on recrystallized sheet, nominally 0.04 in. thick. Weldability of the various alloys was screened using manual TIG welding techniques. Bend ductility (bend axis perpendicular to weld axis) was the criterion used for evaluation. Tensile specimens employed 0.200 by 1 in. gage lengths, and bend test specimens measured about 1 by 1/4 in. Low-temperature tensile tests were conducted using a crosshead speed of 0.02 in. per min up to the point of yielding, and 0.05 in. per min to fracture. High-temperature tensile tests were conducted in vacuum using a crosshead speed of 0.01 in. per min

Jan 1, 1963

By Ralph W. Smith

DEVELOPMENT of the lime industry- in Ste. Genevieve County began in a crude way in 1840. According to information furnished by the Missouri Bureau of Geology, in the early days small vertical kilns built of stone were used, and wood was the fuel. Modern development began in 1904 when one company introduced steel shells lined with firebrick. Within a few years two more companies erected similar plants, and in 1925 a fourth company began operations. Thus there are today four lime companies at Ste. Genevieve, all operating at high capacity and with a gradually increasing production. GEOLOGY The limestone of Ste. Genevieve County, which has the chemical composition and purity required for lime burning, is the upper 80 or 90 ft. of Spergen formation. The Missouri Bureau of Geology, in its geological map of Ste. Genevieve County, gives the Spergen formation a total thickness of 160 ft., but only the upper portion is pure enough for lime burning, with an average raw-stone analysis of 98.7 per cent CaC03; 0.35 Si02; 0.3 R203; 0.35 MgCO3; 0.3 S03. The Spergen, one member of the Meramac group, Iowa series, of the Mississippian system, underlies an area approximately 12 miles long north and south, by 5 miles wide east and west (Fig. 1) in Ste. Genevieve County. The entire eastern two-thirds of Ste. Genevieve County is sedimentary, the formations dipping gently toward the east on an average grade of about 4 per cent, or 2 ½ º. Thus the outcrops of the various sedimentaries run approximately north and south. A line running west from Ste. Genevieve crosses the outcrops of the sedimentaries, ranging from the top of the Mississippian to the Ozarkian geological system in a space of 12 miles. Thus with the gentle dip of 2 ½ º the 160-ft. thickness of the Spergen formation stretches out in its outcrop so that it is approximately one mile

Jan 1, 1938

By A. Stanley, J. C. Mead

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus % in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9 1/2 in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9 1/2 in. bed sinters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite recovery). These tests have proved successful in increasing ilmenite recovery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was plus from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By Joseph C. Mead, Alan Stanley

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus 5/8 in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9½ in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9½ in. bed sin-ters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite re-covery). These tests have proved successful in increasing ilmenite re-covery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was secured from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By G. K. Burgess

THE question of the allotropy of iron, in spite of a vast amount of experimental work and perhaps an even greater amount of theorizing, is not yet settled. That there is a definite transformation in iron near 900,° the A3 point, is generally recognized, as well as the fact that the temperature of this transformation is lowered by the addition of carbon and metallic elements. On heating, the transformation Ac3 is always found at a higher temperature than the transformation on cooling, Ar3. Whether the A3 transformation is sharp like the melting point of a pure substance, or extends over a considerable range of temperatures, embracing perhaps as a lower limit the A2 change, appears to be still an open question for pure iron. The nature and identity of the A2 transformation, it would appear from recent publications, has not yet been satisfactorily settled. In the present paper a critical, historical summary is given of the experimental investigations of the location of A2 and A.3 in pure iron, together with brief mention of some of the theoretical aspects of the subject, and an account is given of a series ?of experiments carried out with several samples of very pure iron which were studied by two methods of thermal analysis in the range from 5000 C. to 1,000° C., and which it is believed furnish a contribution of considerable interest as to the location and nature of the A3 and A2 transformations. Great attention has been given to the details of experimental manipulation and to the preparation of the samples, as it was soon found that some of the discrepancies noted in the work of many experimenters may be traced largely to lack of precautions which we have found essential. For example, to anticipate somewhat, it is necessary to provide a suitable practically gas-free furnace, the heating and cooling of which may be regulated to a nicety. To render a minimum the deforming effects on the cooling and heating curves of the poor conduction of the sample, it must be of a form and mass to pass on its heat rapidly and completely to the thermocouple; and finally the sample itself must be freed from gases which, if present, further mask the definiteness of the heating and cooling curves. We, have also varied within wide limits the several factors which may influence the determination of the exact location and shape of the transformation ranges, especially the mass and preparation of sample, rate of heating and cooling, design of furnace; and finally, as nearly as it is experimentally possible, we have worked with a substance approaching pure iron, free from occluded gases and other impurities, and contained within a vacuum.

Jan 10, 1913

By P. P. Reichert, L. S. Gournay, N. T. Burdine

An investigation of equivalent pore entry radii in typical samples of petroleum reservoir rock and the pore volume associated with each value of pore entry radius has been made. Theoretical discussions together with experimental procedures for obtaining pore entry radii and the distribution of pore volume with pore entry radii are presented. Experimental results on a number of core samples, along with typical distribution curves, are shown. Data on the per cent of the pore volume filled by mercury at a pressure of 1500 psi are included. The results of re-runs of samples, made possible by regenerative apparatus, show the repeatability of the data and indicate the amount of physical change in the samples by mercury penetration. Theoretical equations for calculating permeability from pore size distribution data were derived and the results of such calculations are compared with measured gas permeability data. The effect of the shape, surface area, and weight of the sample on the pore size distribution of reservoir rocks was investigated experimentally. Mercury capillary pressure curves are compared with those obtained by the porous diaphragm method using pas to displace water. Standard core analysis techniques, in general, lead to quantities which are statistical averages of the varying properties of the samples under examination. Although such statistical information has real value in predicting the gross performance of porous bodies. it fails to provide fundamental information concerning the properties of the porous medium itself or any processes which may be occurring in it. For example, a gas permeability measurement on a core sample is an indirect way of determining an average pore radius for that particular sample. Since there are many combinations of pore radii that will give the same radius. and, hence, the same permeability, no information is obtained on the pore size distribution. Furthermore, similar values in permeability do not imply similarity in other properties of a porous medium. For instance, it is possible to have two core samples with identical permeability which would have different residual oil contents at the end of an air-oil drive, or even different amounts of interstitial water. It has been recognized for some time by the petroleum industry that a determination of pore size distribution for porous reservoir rocks offered promise of increased understanding of fundamental flow processes in the porous matrix. and therefore of petroleum res- ervoir performance in general. There have recently been published several methods for determination of pore size distribution in porous materials. A report is given here of the modifications in equipment and technique required to adapt one of these methods to measurements on petroleum reservoir core samples. Although only a few correlations or applications of the method are available for presentation at this time, the technique is brought to the attention of the industry so that the opinions and experience of others with this or similar techniques may be exchanged. It seems possible that the method will have wide application in the future. METHOD There are three general methods which may be employed to measure the pore size distribution in naturally occurring rock formations. One method1 is based on experimentally determined adsorption isotherms for calculating the surface area of the porous medium. Knowing the surface area and the pore volume, a rough estimate of the pore radius can be made. Since this method depends on the thickness and uniformity of the adsorbed layer, the validity of the results would be highly questionable. A second method utilizes capillary pressure data obtained by static dis-

Jan 1, 1950

By C. M. Means

CONSIDERING the hazard involved in mining operations, statistics show, that a very small percentage of accidents is chargeable to electricity. These accidents do represent quite a large percentage of those that are preventable and they are the direct result of the introduction, for purely economical reasons, of a dangerous element. The introduction of electricity in mines should decrease the hazard, and in no case should electricity be applied to the mechanical operation of equipment if by so doing the dangers incidental to mining are increased. The greater number of accidents are the result of persons coming in contact with exposed conductors at potentials varying from 250 to 500 volts direct current. The use of alternating current at higher potentials than these is quite common in large mines, but it is very rare that an accident happens from this source. This is explained by the fact that where these high voltages are warranted, proper precautions are taken, and the installations are directed by those who fully appreciate the dangers incidental to the work. Direct current is in much more general use than alternating current because of its adaptability to haulage locomotives and the low cost of installation in small operations. We are, however, coming to a more extended application of alternating current for the operation of mechanical devices used in connection with mining. Our underground direct-current wiring system is an evolution from the surface trolley and feeder systems. This is probably due to the fact that our earlier successful applications were electric locomotives, which required the use of a trolley wire. The trolley current has been almost banished from industrial establishments on the surface because of the hazard involved, yet we use it indiscriminately underground, where the dangers from accidental contact or fire are infinitely greater. This does not mean that the underground trolley systems should be eliminated, but it does imply the proper safeguarding of such equipment as is necessarily a part of the trolley systems.

Jan 4, 1914

By Charles O. Burgess

The improvement in the quality of cast iron by means of alloy additions has been a metallurgical accomplishment of recent years. Work in this field has shown that chromium is one of the most effective of such additions but, in general, the influence of chromium has been studied only in connection with other alloying additions, with irons other than foundry irons, or with relatively high percentages of chromium. The present study was initiated for the purpose of determining the influence of small additions of chromium (0.25 to 1.00 per cent) on the quality and resultant properties of cast irons of various foundry grades. The effect of small chromium additions on cast iron was investigated with particular reference to casting characteristics, hardness vs. section, strength, machinability, heat resistance and structure. Material Tested Three separate types of cast iron were investigated as to the effect of small chromium additions. The first series of these irons was prepared from remelted foundry cast iron at the Union Carbide and Carbon Research Laboratories, and the other two series consisted of cupola melts made under usual operating conditions. The first series was made by melting 30-lb. charges of foundry scrap iron containing 2.86 per cent carbon, 2.72 silicon, 0.50 manganese, 0.062 sulfur and 0.794 phosphorus in a carbon resistance furnace. The chromium was added as high-carbon ferrochrome just before pouring. The irons were cast at approximately 1250" C. (2282" F.). The analysis and physical properties of these irons are given in Table 1, test 1. The second set of cast irons was made in a cupola in accordance with normal practice, the charge containing approximately 42 per cent steel scrap and the chromium additions being made to the ladle. Analysis and tests of these heats are shown in Table 1, test 2. A small amount of copper was present in heat C of this series. Further data on the effect of chromium on strength were obtained from a third series of 13 cupola heats containing from 0.06 to 0.50 per cent chromium and made without steel additions. The heats were all

Jan 1, 1933

By M. W. Ditto

THe conversion of the beehive coke plants, in this country, to byproduct plants has been slow, because the coal supplies were near the centers of the steel industry. With the growth of this industry, especially with its development around Chicago, it became necessary to transport large tonnages of coal from the eastern districts and then convert it into coke. The losses due to transportation costs were partly offset by the value of the byproducts recovered. To save the transportation costs, it was desirable that the coal deposits of Indiana and Illinois be utilized. This coal had always been classified as non-coking; it was also considered unsuitable for the metallurgical field because of the high ash and sulfur contents. The conditions, however, were promising enough to start experimental and development work, which crystallized in the design of the Roberts coke oven. The fundamental features of most coke ovens, with respect to the application of heating gas and the recovery of byproducts, are the same and, in the last few years, the tendency has been toward the better application of heating to the walls, higher thermal efficiency, and, by the control of heating conditions, the increase of byproduct yields. Structural features also have been improved and the use of high-grade refractory material has allowed the use of higher temperatures with the resulting higher rated capacity per oven per day. Great improvements have been made in both recuperators and regenerators, particularly with reference to the individualizing of each oven with respect to its adjacent oven, so that each can be operated as a unit if desired. The highest development of the flue type of oven has been applying individual regenerators for the recovery of waste heat. After some research work, it was found that the high-volatile Illinois coals could be utilized for coke by obtaining the better application and control of the heating conditions in the oven. Most of this high-volatile coal that had been used gave a good coke structure under certain conditions. The fact that coke was occasionally made from this coal, demonstrated the possibiity of making coke from it at all times pro-

Jan 1, 1923

By M. W. Ditto

THe conversion of the beehive coke plants, in this country, to byproduct plants has been slow, because the coal supplies were near the centers of the steel industry. With the growth of this industry, especially with its development around Chicago, it became necessary to transport large tonnages of coal from the eastern districts and then convert it into coke. The losses due to transportation costs were partly offset by the value of the byproducts recovered. To save the transportation costs, it was desirable that the coal deposits of Indiana and Illinois be utilized. This coal had always been classified as non-coking; it was also considered unsuitable for the metallurgical field because of the high ash and sulfur contents. The conditions, however, were promising enough to start experimental and development work, which crystallized in the design of the Roberts coke oven. The fundamental features of most coke ovens, with respect to the application of heating gas and the recovery of byproducts, are the same and, in the last few years, the tendency has been toward the better application of heating to the walls, higher thermal efficiency, and, by the control of heating conditions, the increase of byproduct yields. Structural features also have been improved and the use of high-grade refractory material has allowed the use of higher temperatures with the resulting higher rated capacity per oven per day. Great improvements have been made in both recuperators and regenerators, particularly with reference to the individualizing of each oven with respect to its adjacent oven, so that each can be operated as a unit if desired. The highest development of the flue type of oven has been applying individual regenerators for the recovery of waste heat. After some research work, it was found that the high-volatile Illinois coals could be utilized for coke by obtaining the better application and control of the heating conditions in the oven. Most of this high-volatile coal that had been used gave a good coke structure under certain conditions. The fact that coke was occasionally made from this coal, demonstrated the possibiity of making coke from it at all times pro-

Jan 1, 1923

By H. M. Strong

Nickel and carbon form a metastable nickel-nickel carbide eutectic system which was ohsen)able by freezing latent-heat arvests at pressures 210 khars. The eutectic freezing temperature Is pressure had a slope of -0.9°Cper kbar. At 50 khars, the nickel-nickel carbide eutectic was at 1297° * 5°C and 0.34 mole fraction Ni3C. The composition of the eutectic increased at an initial rate of 0.004 mol fraction per kbar. The carbon content of the a phase also increased with pressure. Seveval illustrations of nickel carbide microstruc-tures are given and data on the lattice expansion of nickel in terms of carbon content are included. IRON, cobalt, and nickel form a variety of carbides with compositions varying between MeC and MC. Of the three, the carbides of iron are most easily recovered. The nickel carbides are so unstable at temperatures above 900°C that their occurrence in Ni-C alloys is unusual. Under sufficient pressure, however, the temperature range for the stability of the ferrous metal carbides is likely to be increased, or their decomposition suppressed (e.g., see Hilliard' on FesC), providing an opportunity to study the unstable carbides more closely. This was found to be true for nickel carbide which, under pressure, formed a eutectic with nickel. This paper describes the conditions of formation of nickel-nickel carbide eutectic compositions and their subsequent recovery. Despite the comparative instability of the nickel carbides, something is already known about them through various chemical reactions at temperatures not exceeding a few hundred degrees, and from their recovery by the very rapid quenching of molten Ni-C alloys from temperatures of about 1450°C or above. At least four carbides of nickel have been reported: A discussion of the state of knowledge about Ni3C is given by ansen.' The studies of Morrogh and williams7 who obtained Widmanstatten nickel carbide structures and of sidorenko8 who first obtained nickel carbide eutectic structures are also of interest. Sidorenko used short heating times for molten nickel in a graphite crucible followed by very rapid quenching to obtain the eutectic structures. Long heating times produced only the nickel-graphite eutectic probably due to the presence of graphite crystals in the melt which nucleated graphite crystallization. According to the early work summarized by anssen,' nickel carbide may form at temperatures far above the melting point and may be recovered by very rapid quenching. More recently, Giardini and Tydings, working on diamond synthesis, found evidence for "free nickel carbide'' in high-pressure reactions carried to temperatures far in excess of the limit for diamond growth, followed by rapid thermal quench. They stated that nickel carbide was not found in association with diamond growth. In the present work, observations were made on the melting and solidification temperatures of Ni-C alloys as a function of composition, temperature, and pressure. The nickel-nickel carbide eutectic was found at pressures above 10 kbar and its variation with pressure followed up to 60 kbar. Nickel-nickel carbide microstructures were recovered by quenching under pressure from temperatures of about 1400°C and higher. The formation of nickel carbide structures under ordinary diamond-growing conditions was usually observed but the role of carbide formation as an intermediate step in diamond growth needs more study. EXPERIMENTAL METHODS The melting and solidification temperatures of various nickel-graphite compositions were observed by latent-heat arrests in the sample holder illustrated in Fig. 1. The samples were composed of mixtures of carbonyl nickel powder and National Carbon Co.'s SP-1 graphite packed to densities of about 6.4 g per cu cm. They were contained within

Jan 1, 1965

By James T. Waber, Allen C. Larson, Karl Gschneidner, Margaret Y. Prince

The original graphical method proposed by Darken and Gurry for predicting which elements are soluble in a given element was slightly modified and was used to predict terminal solubilities in 1455 known binary combinations. it was found that the modified Darken and Gurry method had a reliability of 61.7 pct in predicting extensive solid solubility (greater than 5 at. pct) and of 84.8pct in predicting limited solid solubility (less than 5 at. pct), an over-all reliability percentage of 76.6. Using the same experimental data, it was found that using only the simple Hume-Rothery size-effect criterion was 50.1 pct reliable in predicting extensive solid solubility, only slightly better than a purely random choice. and 90.3 pct reliable for predicting limited solid solubility, an over-all reliability percentage of 67.6. THE first rational rules for determining a priori whether two metals are not freely soluble in the solid state were announced by Hume-Rothery and coworkers.1,2 These well-known rules have been accorded a special place in metallurgical literature. Since the "electronegative" and "relative valence effect" rules of Hume-Rothery were only qualitative in nature it was difficult to use them in a quantitative analysis. The next major step was made by Darken and gurry3 and apparently went unexplored for a number of years. They prepared graphs illustrating the influence of both electronegativity and size difference on the solubility of elements in silver, magnesium, and aluminum, and they drew elliptical figures, whose maximum width was ±15 pct of the solvent radius and the maximum height was ±0.4 electronegativity units. The purpose of the present paper is to assess and compare the effectiveness of using the simple size-factor and the Darken-Gurry criteria to predict which elements will be soluble and insoluble in a given solute. In 1957, waber4 prepared Darken-Gurry plots for determining the solutes that were likely to be soluble in 6 and plutonium and introduced an inner ellipse as an aid to determining which elements should be soluble to a considerable extent. As illustrated in Fig. 1, the major axis of the outer ellipse corresponded to ±0.4 electronegativity units (OD) while the minor axis corresponded to ±15 pct of the radius of the solvent phase (OB). This simultaneous "action" of the two Hume-Rothery criteria was used to indicate which elements would have little or no solubility in the host lattice, namely those elements whose points lay outside this ellipse. An inner ellipse, "centered" over the same point of the solvent was drawn with ±0.2 electronegativity units (OC) and ±8 pct (OA) as the major and minor axes. Solutes with points lying within the inner ellipse were expected to be capable of forming extensive solid solutions with the host element. Use of this method of applying the earlier Hume-Rothery criteria was moderately successful for plutonium. For 6-phase plutonium, only one element out of the eight predicted to be extensively soluble was known to be almost insoluble, and of the sixteen elements predicted to have more restricted solubility, five were known to be essentially insoluble. The success of discerning solubility in the E phase by this method was comparable. One prediction made on the basis of this analysis was that the heavier rare earth metals, such as holmium, erbium, and thulium, should be soluble in the 6 phase, whereas the lighter rare earth elements, such as lanthanum or

Jan 1, 1963

By William J. McCaughey, Richard L. Barrett

FOR sixty years-in fact, ever since the inception of the basic steelmaking process -basic slags have been the subject of study by chemists, metallurgists and petrographers, with the purpose of providing a better understanding of the reactions between the slag and the metal during the refining process. The result of this work and of the researches of mineralogists and physical chemists into the chemistry of oxide systems is a considerable body of literature dealing not only with slags themselves but with equilibrium relationships in systems synthesized from pure components. Since any actual slag is a multicomponent system more complex than any yet worked out in the laboratory, a complete solution to the problem is still far away. The writers have recently published a paper1 which they believe sheds light on the role of the slag in the elimination of phosphorus and it is the purpose of the present paper to discuss the metallurgical applications of that work. Since the pioneer investigations on the constitution of basic slags carried on by Stead and Risdale, Carnot and Richards, Hilgenstock and others in the 1880's, a considerable body of literature on this subject has appeared. Many of the more important of these papers were summarized in the previous paper,' therefore they will not be reviewed here. PETROGRAPHIC STUDY The present writers have had occasion, over a period of years, to examine petrographically a great number of basic steel- making slags having a very wide range of chemical compositions and made under various conditions of operating practice. Although certain of these slags exhibit some unusual features, it is a remarkable fact that the great majority of mature slags show certain typical characteristics, which persist throughout the wide range of chemical compositions. The typical mature slag usually appears under the microscope to consist of two phases. The phase of primary crystallization is light colored and doubly refracting. If separated mechanically and subjected to chemical analysis, it is found to consist essentially of SiO2, CaO, and P2O5. Sometimes, although not usually, two such silicate-phosphate phases are found. The other conspicuous phase is dark in color and optically isotropic, and occurs as the interstitial material between the grains of the light colored phase. This dark colored interstitial phase when analyzed is found to consist essentially of FeO,Fe2O3, MnO, and MgO. In some slags high in iron oxide, particularly those from low-carbon heats, dicalcium ferrite is found. This phase is dark reddish brown in color and anisotropic.

Jan 1, 1944

By Charles J. Potter

There are many sources of energy available for consumption. Some available are solar, tide and wind. However, in terms of relatively large usage, only four principal sources are used for primary energy today. They are coal, oil, natural gas and water power. Such additional items as wood and nuclear energy, although supplying quantities of energy today, do not in the aggregate amount to any real significant quantity, but nuclear energy must be considered for future use. In addition, such a thing as gas from distillation of coal and oil is not considered primary energy. The principal markets for primary energy are the electric power utilities, space heating, industrial uses and transportation. The electric utility use of energy is primarily from coal, followed by natural gas and water power. Space heating is primarily a gas market, followed by oil and then coal. Space heating by electricity is only considered in the generation of electric power. Transportation's principal source of energy is oil. Coal has practically no market left in this field.

Jan 4, 1962

By J. C. Bokros, A. S. Schwartz

The irradiation-induced densification of pyrolytic carbon has been shown to take place by a simple first-order rate process. The values of the densification constant, kd, for isotropic carbons are 1.0 x 10-22, 1.4 X 10-22 and 2.8 X 10-22 (nut)-1 at irradiation temperatures of 520o, 670°, and 1000oC, respectively. The apparent activation energy obtained from the temperature dependence of kd for isotropic carbons is about 0.2 ev. Estimates of the maximum stresses that develop in isotropic carbons owing to shrinkage onto an unyielding support show that stresses in excess of 15,000 psi may be avoided during irradiatian at 600" and 1000°C if the initial coating density is kept above 1.5 and 1.9 g per cu cw, respectively. If a carbon coating on an unyielding support is not completely isotropic, stresses may continue to build up at high neutron exposures after densification is complete because of continued anisotropic distortion. Stresses in excess of 15,000 psi from this source may be avoided during irradiation at 1000°C if' the Bacon anisotropy factor is kept below 1.3. A variety of applications of carbon and graphite in nuclear technology has stimulated numerous studies of the effect of neutron irradiation on these materials One of the most interesting problems associated with the use of carbons and graphites in the high fast neutron fluxes of nuclear reactors arises as a consequence of the neutron-induced dimensional changes that occur during irradiation. Energetic neutrons displace car- bon atoms from their lattice positions; and, depending on the temperature, the interstitials and vacancies produced migrate to form clusters, recombine, or are annihilated at dislocations or crystallite boundaries. The competition of various sinks for the point defects results in structure-sensitive bulk-dimensional changes that vary with temperature in a complicated way. This paper deals with the neutron-induced dimensional changes in turbostratic carbons that are being used as coatings on nuclear fuel particles. Details of the structure of these carbons and the variations of their structure with deposition conditions have been reported.'-3 It has been shown in previous work that the neutron-

Jan 1, 1968

By Bunting S. Crocker

Bunting S. Crocker (Executive Vice President, Kilborn Engineering Ltd., Toronto, Ont., Canada)-This is an interesting report on laboratory-scale testing with pebble charges in 12-in.-diam mills, and the author correctly concludes that useful laboratory testing can be done. In the opening paragraphs he states, "laboratory pebble-milling tests have been considered too difficult to perform and not relevant enough to the actual grinding situation"!! This is definitely not so! The writer published detailed tests done in a 12-in.-diam mill comparing pebble and steel balls in 1952 6 and again in 1954 7. We note that our review paper in 1963 was the only one mentioned in the References. The author states that "a smooth lining is more frequently used in commercial-scale pebble grinding." The writer has designed some 22 plants treating 134,000 tpd in North America and Australia. He has also visited pebble installations in Norway, Sweden, Finland, and South Africa. All these plants correctly use heavier lifters than would be considered standard for ball mills of a similar size. Smooth liners are not recommended in wet ore treating plants. In the 1952 paper6 it was clearly demonstrated in the test work that the individual ore pebbles had to be the same weight-not the same diameter or size-as the steel balls if comparable grinding results were to be obtained. In the tests performed in this study most of the pebbles "passed 50-mm round hole screen openings and remained on a 40-mm round hole sieve. Correspondingly most of the ball mill experiments were performed with 50 or 40-mm steel balls". Thus, the balls and pebbles were essentially the same size or diameter. This means we are comparing apples and oranges. When the feed being ground is fine enough the pebbles will be superior. When the feed is coarse the steel balls will be the better. This is similar to grinding tests in which two sizes of steel balls are compared. In the middle of the article, "Crocker5 has reported that equal weights of pebbles and steel balls achieve the same grinding capacities under similar conditions …", then he proceeds to question this postulate. In the eight reports that I have published, I have never made this statement. What I have said is "for comparable grinding, the individual pebbles in the mill charge should be the same weight as the individual steel balls." This is not the first time that I have been misquoted on this point! It is most unfortunate that this test work has been done with a ball and pebble of the same diameter and not the same weight. It is impossible to properly evaluate the results. I also regret the fact that the screen analyses of the tests were not shown. Many of us who have been 40 years or less in the grinding game can easily read screen analyses at a glance. We can tell whether the top meshes are being ground by the ball charge, etc., etc. The graphs are fine for the summary or to illustrate conclusions but are often misleading in the report itself. However, we do agree that useful pebble grinding tests can be done in a 12-in. mill.

Jan 1, 1973

In our sessions devoted to geophysical prospecting, the greater part of the time heretofore has been given to the presentation of theoretical subjects or highly technical details by specialists directly engaged in the development and perfection of the various methods. This was necessarily so, for we wanted such information available for the benefit of other workers and for our members who were anxious to gain a reliable understanding of the principles involved, to enable them to form an opinion of the adaptability of this new technique to their special problems. The methods have now been actively applied by many reliable organizations for several years and in this period there has been opportunity for operating mining men, engineers and geologists to appraise them. We therefore felt that it would be desirable at this meeting to give some time to interchange of opinion concerning results from actual cases. Consequently, the Committee on Geophysical Methods of Pros pecting sent out letters of inquiry to numerous mining people, asking for frank statements of opinion based on results from geophysical surveys at their properties. We have received many replies, among which 30 or more are of specific interest. On the whole, the point of view is remarkably uniform. Except for one or two who were, perhaps, a bit irritated, the operating men who have replied consider the methods to be capable of yielding valuable information when suitable conditions exist. Their attitude is definitely one of tolerance and of confidence in the integrity and sincerity of the outstanding workers. There is, however, a very general feeling that the chances of attaining positive and trustworthy results are small unless the work is directed by, or at least carried on with the close cooperation of, a geologist or mining man who is thoroughly familiar with the local geologic and mineralogic conditions and with the relations of the orebodies to rocks and structure. From certain comments, I am inclined to think that some operators are not greatly impressed with the cheapness of the work. Probably they are not attaching much weight to cost per acre, but are thinking primarily of total cost in terms of significant results obtained. It is

Jan 1, 1932

By James Norman, O. C. Ralston, John Dasher

MOST mineral products are used in the finely divided state, but some are sold in larger sizes. Coal, gravel, metallurgical fluorspar, phosphate rock, hematite, chromite, and other products are sold in sizes coarser than can be treated by many mineral-dressing methods. Gravity separation of coarser sizes by log washing and jigging is probably as ancient as any process now in use. The performance of jigs on a feed with sufficient difference in density of the minerals is excellent, but many families of minerals are too similar in density to be separated effectively by this method. The use of heavy suspensions of fine solids to separate coarse coal from bone and slate has been in use for some years. This separation by sink-and-float was limited until recently to light minerals. Development of galena and ferrosilicon as suspension media, by Minerals Beneficiation, Inc., at Mascot, Tenn., Pitcher, Okla., and Crosby, Minn., under patents granted during and since 1938, has permitted excellent results to be obtained on minerals of very similar densities up to 3.3. This process has been put in the hands of American Cyanamid Co. for exploitation, and further research now in progress may increase the density of suspension obtainable so that separations at a density of 4.0 may be practicable before this is published. The Huntington-Heberlein dense-medium process, developed in England, is also being commercialized in the United States by Sink and Float Corporation. Du Pont has developed a sink-and-float separation using halogenated hydrocarbons as parting liquids; at this writing, the method is used commercially only on coal but has been successful in the pilot plant on many other minerals using tetrabromethane at densities up to 2.95. While these processes were being developed to extend the range of separation by difference in density to the coarser sizes, the Bureau of Mines was engaged in developing a process that will extend the range of separation by difference in wettability up to 2-in. size. War work has pushed aside its further development indefinitely. The results are published here with the hope that some use may be suggested that will justify post-war development, and so that readers that may be able to make immediate use of the scheme may have the information. EARLIER WORK It has been known since the time of Herodotus (B.C. 484(?)-425) that water-repellent particles will adhere to sticky surfaces. That early historian reports the

Jan 1, 1942

By Carville E. Sparks, Rollin Farmin

Edward Thomas (U.S. Bureau of Mines, Washington, D. C.)—This excellent article on an ingenious and successful installation of wooden rock bolts loses much of its effectiveness through an attempt by the authors to establish the general adaptability of this type of support. Possibly because they have only local experience as a guide, they imply erroneously in the first paragraph that the Bureau of Mines has been opposed to wooden bolting, although later in the article reference is made to a Bureau publication on the subject. In 1947 the Bureau of Mines decided to sponsor experiments in roof bolting in coal mines because the experience of the St. Joseph Lead Co. had indicated that this type of roof support appeared to be a roof-fall accident-prevention measure that might be adopted and also effect economies in operating cost. Justification for this belief has been borne out, particularly in coal mining. This program has been carried out under the following general program: 1—The Roof Control Section of the Bureau of Mines has advocated and initiated underground experiments under controlled conditions with many different ideas related to rock bolting since 1947. In no instance has cooperation in developing an idea been refused, no matter how ridiculous it may have seemed at first glance, if it has been requested by a mining company. 2—The Bureau of Mines has no police powers over the selection of roof-support devices and seeks none. 3—At the request of an operator, the Bureau of Mines will make comparative tests on various types of bolts underground and will submit resultant data for the operator's consideration. 4—Wherever possible, the Bureau publicizes successful installations, so that others may benefit from this experience; however, it is careful not to discriminate between manufacturers of bolts or bolt-installation equipment. Unsuccessful experiments probably should also be publicized, but from a practical standpoint it is usually not feasible. In other words, the role of the Roof Control Section of the Bureau is in line with the original purpose for which the Bureau was first organized: to serve industry through gathering, analyzing, and disseminating factual data. It would seem to be logical that the starting point for any experimentation on roof bolts in coal mines should be the time-proved practice of the

Jan 1, 1955