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By Randolph C. Specht

A two year study was made of the waste disposal of the pebble rock phosphate industry. Solid slimes are impounded in large settling areas and the process water is re-used. Clear effluent was not found to be toxic to fish, or animal life in field tests or in controlled experiments.

Jan 1, 1950

By A. Raman, P. P. Singh

The constitution of the La-Rh system was studied by powder X-ray diffraction, metallopaphic, and differential thermal analysis techniques and an equilibrium diagram is presented. Eleven intermediate phases occur in the system and the crystal structural data for nine of them were determined. La3Rh crystallizes in an orthorhombic structure of undetermined type, whose unit cell is obtained by doubling the 'a; and 'c,,' edges of an FesC type unit cell. The other intermediate phases of the system are LarRh-3( undetermined structures also occur in the system. LaRh, undergoes a polymorphic phase transformation at 1240°C. LaRh3 and La2Rh7 also exhibit polymorphisnz. The phases Laah and LazRh7 melt congruently. The latter undergoes a eutectoid transformation into LaRh, and Rh at 1205°C. Laah3 is formed by a peritectoid reaction between Laah and La,Rh,,. The other Phases result from peritectic reactions between the liquid and the adjacent rhodium-rich phases. The intermediate Phases of the La-Rh system are compared with those of the La-Co and La-Ni systems. DURING the course of a detailed investigation to study the occurrence of CrB, FeB, A1B2, and related structures in the rare earth alloys it was found that much information is lacking for the rare earth noble metal systems. Although the structures of several rare earth alloys containing the noble metals at the AB and AB2 stoichiometries have been reported, the occurrence of related structures at other stoichiometries has not been studied. We have initiated a project to study the crystal structural features of selected rare earth-rhodium alloys and to map the equilibrium diagrams of representative systems with conventional methods. The results of our investigation in the La-Rh system are presented in this paper. Two phases were known in the La-Rh system. LaRh has the CrB-type structure.' LaRhz is a MgCu2-type Laves phase.z EXPERIMENTAL PROCEDURE Alloys weighing less than 1 g were prepared from commercially pure lanthanum (99.9 pct +), supplied by Lunex Company, Pleasant Valley, Iowa, and rhodium (99.92 pct +), supplied by Engelhardt Industries, Newark, N.J., in a conventional arc melting furnace under argon atmosphere. The buttons were turned upside down and remelted three times to insure homogeneity in the samples. Since negligible loss of material was encountered during melting, a chemical analysis of the alloy buttons was not undertaken. Powder specimens for X-ray diffraction studies in the as cast state were then prepared. The buttons were wrapped in thin molybdenum foils and homogenized by heating in vacuum at suitable high temperatures for more than 1 week. They were then broken into three or four pieces for annealing experiments. The pieces were wrapped in molybdenum foils and annealed at various temperatures in evacuated quartz capsules. The annealing was carried out for 2 hr at or above 1200°C, 1 day at temperatures close to llOO°C, 2 days at 1000°C, and for 1 week at temperatures below 1000°C. After annealing the alloy pieces were again broken and powder specimens for X-ray diffraction were prepared. The powders of the lanthanum rich alloys with more than 80 at. pct La were prepared by filing. The filings were sealed in molybdenum tubings and stress-relieved at 600°C in vacuum. It was not deemed necessary to stress-relieve the powders of the other alloys, since the alloys were very brittle and were ground easily. POWDER X-RAY DIFFRACTION X-ray diffraction photographs of powders (-325 mesh size) of the alloys in the as cast and annealed states were prepared in a Guinier-de Wolff focussing camera with copper K, X radiations. These patterns were studied to identify the stoichiometries and the crystal structures of the intermediate phases. The lattice parameters of the phases were calculated after minimizing the differences between the observed sin2 6 values, calculated from the diffraction angles 8, and the sin2 8 values, calculated using approximate lattice constants obtained from a few lines. These differences were minimized manually to less than 0.0005. The latLice constants are judged to be accurate to *0.005A for values less thp about 10A and to k0.01~ for values greater than 10A. The relative intensities of the lines were calculated using a computer program written by Jeitschko and Parthk.~ No attempt was made to refine the atomic positional parameters in the phases. METALLOGRAPHY The phase equilibria in the investigated alloys in the as cast and annealed states were also studied by metallographic examination. The polished specimen surfaces were etched with 10 pct picric acid in alcohol (alloys up to 25 pct Rh), concentrated picric acid (from 25 to 37.5 pct Rh), 2 pct nital (40 to 50 pct Rh), 10 pct nital (from 50 to 66.7 pct Rh) and with concentrated 48 pct HF for the other rhodium-rich alloys. Selected microstruture~ were then photographed using a Po-laroid Land camera. THERMAL ANALYSIS Differential thermal analysis of the alloys was carried out in DTA-668 Stone differential thermal ana-

Jan 1, 1970

By Leonard D. Jaffe

IN recent years, the advantages of tempered martensite as a microstructure for steel parts have been well established. For parts that must not fracture brittlely when loaded at high rates, at low temperatures, or under conditions of combined stress, tempered martensite is especially desirable. One of the factors that determine whether a steel part will have a martensitic structure after heat treatment is the hardenability of the steel. Despite the large amount of recent research, there are still many unsettled basic questions concerning hardenability and its relation to chemical composition, and, indeed, considerable disagreement among investigators in the field. Much of the uncertainty that beclouds the existing knowledge of hardenability arises from the fact that hardenability is still a wholly empirical concept, and has not been quantitatively related to the more fundamental knowledge of isothermal decomposition of austenite. Hardenability is, essentially, a measure of the time required for small percentages of austenite to decompose, during continuous cooling, to pearlite, bainite, proeutectoid ferrite, or proeutectoid carbide. In the practical heat-treatment of steel, also, the decomposition of austenite is usually not isothermal but rather during cooling, or, in general, is not at a single constant temperature. Thus, both hardenability and practical heat treatment involve anisothermal decomposition of austenite. In a previous publication1 the author and his collaborators outlined the principles that govern the relations between isothermal and anisothermal decomposition of austenite, and reviewed the experimental information available in 1945. To avoid repetition, familiarity with that article (Paper 1) will be assumed, and attention given primarily to work carried out or published since 1945. The method generally employed was explained in Paper 1: anisothermal transformation is studied through investigation of the effect of holding austenite at one temperature upon its subsequent decomposition at another temperature. Both temperatures are below the stability range of the austenite. This paper is confined to studies in which only a small amount of transformation occurred at the first temperature. This is the only case of importance in practical heat treatment when an essentially martensitic structure is sought. MATERIALS AND PROCEDURE It was necessary to use different steels for the studies involving proeutectoid ferrite and for those involving proeutectoid carbide, because the two constituents have not been found to form in the same steel. Several steels were tried and rejected

Jan 1, 1947

GEORGE F. KUNZ, New York, N. Y. (written discussion*).-We have been informed that the nickeliferous ores of Sudbury could furnish more palladium than the whole of the present world supply, together with a large amount of platinum. As the quantity of ore melted in Sudbury in 1916 totaled 1,521,689 tons, giving 80,010 tons of matte, and as the average of 3 years, up to 1915, are stated to show an average of 0.10 oz. of platinum and 0.15 oz. of palladium per ton of matte, this would indicate the presence of 12,000 oz. of palladium and 8000 oz. of platinum in the ore mined in 1916.1 Interest in the platinum situation is now more intense than it has ever been, and the great importance of this metal for many special uses is being increasingly appreciated. For munition manufacture, for chemical apparatus, in dentistry, and for jewelry, it has been and continues to be in great demand in spite of the phenomenal rise in its value. It is true that this has necessarily led to a search for substitutes, especially for dentistry and chemical uses, but while some of these have been found to answer the needs fairly well, for many 'other purposes platinum is still employed in spite of its great cost. As to the jewelry industry, the most public-spirited efforts have been made to restrict its use, and its employment in the manufacture of watchcases, purses, bracelets, and catches has been abandoned, gold being used in its place. The prime cause of the present shortage is, of course, the elimination of 'the Russian supply which, in normal times, constituted probably more than 90 per cent. of the world's production. This makes an intensive

Jan 4, 1918

By Jr. Ensign C. O., J. W. Trammell

The two principal methods of room-and-pillar mining practiced at White Pine make it important to predict variations in the thickness and rock types of a stratum called the upper sandstone. In full column mining, the nearly barren upper sandstone occurs between two ore horizons (called upper shale and parting shale, respectively), and is mined as a part of the ore column. In general, full column mining is not practiced where excessive upper sandstone thickness causes full column ore grade to be substantially lower than parting shale ore. In parting shale mining, the upper sandstone forms the mine roof. Difficulty in predicting upper sandstone character arose because the extent and directions of its variations were not apparent in the property's diamond drilling, in which holes are spaced on 1000-ft centers. A study of sedimentary features, undertaken to improve the predictability of trends in the upper sandstone, led to an interpretation of the sedimentary environment in which the sandstone was deposited. Ripple marks, mud cracks, and cross-bedding, as well as other, less well known features, such as channel casts, flute casts, and current crescent casts, were mapped or recorded. These data, coupled with the knowledge of regional facies changes gained from studying drill core, show that the upper sandstone was deposited by a series of streams flowing northeastward over the underlying parting shale. Deposition of the lowermost bed of the upper sandstone, near the ancient shoreline, was locally preceded by erosion of the parting shale, and the greatest thicknesses of sandstone are found in channels scoured out of the parting shale. Awareness of the rather strong linear trends in the upper sandstone makes it possible to project continuous areas of thick upper sandstone through apparently isolated "highs" in the upper sandstone thickness contour map, which is based on drill-hole information. Since local exceptions to general stratigraphic trends exist, a method was also needed for esti- mating, in detail, the thickness and degree of shali-ness of upper sandstone forming the roof in active parting shale mining areas. Because locally the parting shale was thinned by erosion during the deposition of upper sandstone, parting shale thickness is inversely proportional to upper sandstone thickness. Utilizing information taken from short drill holes into the roof, curves were constructed for the correlation of parting shale thickness (measurable in the mine) with the upper sandstone total thickness, as well as the thickness of its basal member. The White Pine mine is situated in Ontonagon County, approximately 6 miles south of Lake Superior, in the Upper Peninsula of Michigan (Fig. 1). The orebody mined by White Pine Copper Co. occurs in the lowermost 20 to 25 ft of the Nonesuch formation, a series of middle- to upper-Keweenawan shales, siltstones, and sandstones. This paper discusses the solution of problems in ore grade control and mine planning arising from the presence of an essentially barren sandstone stratum within the orebody. THE OREBODY AND MINING TYPES A brief description of the orebody and mining methods is necessary to show how the upper sandstone affects mine planning and grade control. In the mine, the cupriferous zone of the Nonesuch shale is divided into three major portions, as shown in Fig. 2. Lying conformably on top of the Copper Harbor formation (or "lower sandstone") is the lower part of the orebody, the parting shale. The parting shale is overlain by the upper sandstone. On top of the upper sandstone, the basal portion (8 ft) of the upper shale is quite similar to the parting shale. Both the upper shale and parting shale are divided into a number of smaller stratigraphic units, each having a characteristic copper content, as shown by the histogram of Fig. 2. Other authors1'3 have dealt in detail with the stratigraphy and mineralogy of the orebody. However, in this article, we are concerned with the major subdivisions only, and it will suffice to note the low copper content of the upper sandstone, shown in Fig. 2. At present, two types of room-and-pillar openings are created in the orebody. In one type, called parting

Jan 1, 1964

By N. J. Grant, I. S. Servi, A. Chaudhuri

IN a recent paper by Servi and Grant,' it was illustrated that the slip band spacing and stress based on creep test data were related by the equation: d =1/ where d is the spacing in millimeters and v is the stress in psi, based on a more complete equation proposed by Orowan. It was shown further,' that for a. given grain size, the slip band spacing became larger with decreasing stress (and was independent of temperature), until such time that the slip band spacing becomes greater than about one-half the grain diam- eter. When this occurred, the mechanism of deformation in creep of 2s aluminum at high temperature changed from slip to grain boundary deformation (sliding). Additional metallographic evidence is now available which confirms this in a most interesting fashion.

Jan 1, 1954

By David A. Summers, John Corwine, Li-King Chen

The results from fracturing plexiglas beams are combined in eight existing equations to determine the surface energy of the material. The reliability of each equation is tested, and the equation in use to determine fracture toughness is found to be the most exact. This equation is then used in determination of the surface energy of Barre Granite.

Jan 1, 1971

By H. R. Banks

The Sullivan concentrator has completed 20 years of operation. During this period a considerable amount of data has been accumulated concerning the characteristics of several types of flotation machines in their application to the problem of the differential separation of the lead sulphide, zinc sulphide and, in more recent times, iron sulphide, in the ore from the Sullivan mine. It is the author's intention to sketch briefly experiences encountered when the several types of flotation machines were introduced into the flow. Original Flotation Equipment and Ore The original flotation equipment consisted of four 18-cell lead-roughing machines, whose concentrate was cleaned in an 8-cell and recleaned in a similar 8-cell machine. Zinc concentration was carried out in four 18-cell machines. One 8-cell zinc re-treatment machine received the reground zinc middling. The 168 cells handled some 3000 tons of original daily feed. These cells were of the standard Minerals Separation type, equipped with 24-in. impellers. Spindles were driven through geared line shaft at a speed of 255 r.P.m. The impellers were the 45' type, except for the last three cells of the zinc-rougher machines, which were of sub-aeration type, with full bottom shroud and half shroud for the top side of the impellers. Four-pound air was introduced. This recognition of the benefit of intensified flotation operation for "gleaning" was an important phase in the early operation and will be discussed further on in this paper. Sullivan ore during this period is described as a finely disseminated mixture of four sulphides, totaling go per cent made up of 13.6 per cent galena, 14.4 per cent mar-matite, I per cent pyrite and 61 per cent pyrrhotite. The specific gravity of the ore was approximately 4.4. There has been a gradual change in the ore content. The 10 per cent of nonsulphide constituents has increased to between IS and 20 per cent and all three sulphides have dropped somewhat. Added equipment and changing technique have made possible the treatment of 8000 tons per day of mill feed instead of the 3000 originally handled. Special Features of Flotation at Sullivan Sullivan flotation differs greatly from customary practice. The very heavy return of lead rougher middling and lead cleaner tailing, totaling at times 125 per cent of the new feed tonnage, has been the subject of much investigation. The practice is still employed. This operating condition is mentioned because it is the author's belief that the in-

Jan 1, 1947

By H. R. Banks

The Sullivan concentrator has completed 20 years of operation. During this period a considerable amount of data has been accumulated concerning the characteristics of several types of flotation machines in their application to the problem of the differential separation of the lead sulphide, zinc sulphide and, in more recent times, iron sulphide, in the ore from the Sullivan mine. It is the author's intention to sketch briefly experiences encountered when the several types of flotation machines were introduced into the flow. Original Flotation Equipment and Ore The original flotation equipment consisted of four 18-cell lead-roughing machines, whose concentrate was cleaned in an 8-cell and recleaned in a similar 8-cell machine. Zinc concentration was carried out in four 18-cell machines. One 8-cell zinc re-treatment machine received the reground zinc middling. The 168 cells handled some 3000 tons of original daily feed. These cells were of the standard Minerals Separation type, equipped with 24-in. impellers. Spindles were driven through geared line shaft at a speed of 255 r.P.m. The impellers were the 45' type, except for the last three cells of the zinc-rougher machines, which were of sub-aeration type, with full bottom shroud and half shroud for the top side of the impellers. Four-pound air was introduced. This recognition of the benefit of intensified flotation operation for "gleaning" was an important phase in the early operation and will be discussed further on in this paper. Sullivan ore during this period is described as a finely disseminated mixture of four sulphides, totaling go per cent made up of 13.6 per cent galena, 14.4 per cent mar-matite, I per cent pyrite and 61 per cent pyrrhotite. The specific gravity of the ore was approximately 4.4. There has been a gradual change in the ore content. The 10 per cent of nonsulphide constituents has increased to between IS and 20 per cent and all three sulphides have dropped somewhat. Added equipment and changing technique have made possible the treatment of 8000 tons per day of mill feed instead of the 3000 originally handled. Special Features of Flotation at Sullivan Sullivan flotation differs greatly from customary practice. The very heavy return of lead rougher middling and lead cleaner tailing, totaling at times 125 per cent of the new feed tonnage, has been the subject of much investigation. The practice is still employed. This operating condition is mentioned because it is the author's belief that the in-

Jan 1, 1947

By N. D. Gerald

The mathematical analysis, which is an essential part of this paper, has revealed the interesting fact that contrary to the opinions of many, determinate and measurable economic factors are largely responsible for the price of crude oil. It should, however, be well understood that the results of this analysis are applicable, without qualification, only to the particular period and market studied. There are certain peculiarities of the petroleum industry which make the consideration of its economics very difficult, and at times practically impossible. Noneconomic factors often become highly important and have serious economic effects, which make unqualified predictions based on economic data extremely hazardous and ill advised. However, in spite of the attendant difficulties there are many problems, especially in price analysis, to which approximate solutions must be made for the proper guidance of public and private policy, but their limitations should be kept clearly in mind. An understanding of the nature of prices is of fundamental importance for a number of purposes. Measurement of the motivating factors giving rise to trends and other price movements is essential to forecasting for the intelligent engineering and financial administration of the industry. The fixing of prices, maintenance of an intelligent industrial dictatorship, or the making of other industrial adjustments, demand, for success, an accurate understanding of their probable consequences. The establishment of a system of equitable relations within an industry, or between industries, by methods other than the natural competitive ones, requires a thorough knowledge of the price structures and the nature and trends of the normal and basic prices of their products. This research has shown that in this case, even though the statistical records are only fair, quantitative economics offers a direct and satisfactory approach to the problems of economic consequences dependent on the reactions of mass intelligence. A consideration of commodity prices naturally leads to costs. The true cost of petroleum is so complex and elusive that its accurate deter-

Jan 1, 1934

By J. E. Hackford

In some research work carried out by the writer, certain results have been obtained which bear on the fundamental nature and origin of coal and the relationship between coal and petroleum. Without entering into a discussion of the details of the experiments, which were conducted on petroleum and derived bitumens, there are given here, by way of definition, some of the relations that the writer has established between certain classes of bitumens of petroliferous origin. Bitumen.—-A natural organic substance, gaseous, liquid, or solid, consisting of hydrocarbons and the oxy-or thionic derivatives of the same, or of a mixture of all three. Diaspha1tenes.—-Those portions of bitumens that are soluble in ether or carbon disulfide, but are insoluble in a mixture of equal parts of ether and alcohol. Diasphaltenes are produced by the oxidation or thionization of petroleum oils; they have, as the name indicates, twice the molecular weight of asphaltenes, into which they are converted when subjected to moderate temperature. For example, an artificially produced diasphaltene, which was readily soluble in pentane and ether, was quite insoluble in either of these solvents after heating for three weeks at a temperature of 100" C., and was converted into an insoluble asphaltene. Aspha1tenes.—Those portions of bitumen that are insoluble in ether or ether alcohol but are soluble in carbon disulfide. Aspha1tites.—Those solid or semisolid natural bitumens that are composed, for the most part, of asphaltenes or diasphaltenes. A pure asphaltitel would be composed wholly of asphaltenes and diasphaltenes, but most asphaltites contain small percentages of oil and wax, which have not yet been converted into asphaltenes; they may also contain A small percent.age of kerotenes, which represent the next stage of the metamorphosis of asphaltenes. Among the naturally occurring oxyasphaltites may be mentioned grahamite; and among the thioasphaltites, gilsonite.

Jan 1, 1921

By F. L. Wolf

THE tests here described were made to obtain information regarding costs, efficiency, etc. of baking cores in an oil-fired oven, and two electric ovens, which were installed, early in 1920, in the core room of The Ohio Brass Co. and were used almost continuously during the rest of that year. Operating costs were rather high so a complete test was undertaken early in 1921. The ovens will be referred to as Nos. 1, 2, and 3. Nos. 1 and 2 are electric ovens of different makes while No. 3 is an oil-fired oven made by the firm that made No. 2. The three are equipped with similar ventilating devices. They have full-height doors at front and rear. The cores are placed on iron racks, which are lifted by Cowan trucks and rolled into the ovens through the front doors. When baked, they are removed through the rear doors. Fig. 1 shows the general plan of the electric ovens while Fig. 2 shows the plan of the oil-fired oven. Oven No. 1 is 91 in. (231 cm.) wide, 83 in. (210 cm.) high, and 79 in. (200 cm.) deep, inside measurements. The walls are constructed in, panels about 18 in. (46 cm.) wide and 4 3/16 in. (10.6 cm.) thick, packed with blocks of insulating material and covered with sheet iron. The concrete floor on which the ovens set is 13 in. (33 cm:) thick. A recess about 5 in. (13 cm.) deep was cut in the floor to provide room for the heating units. Channels and plates supported. on I beams serve as' supports for the trucks and racks. The ventilating system consists of a Sirroco No. 1 ½ fan, connected so as to draw air from the top of the-oven or from the room, through damper C, and to deliver it either to the bottom of the oven or to the floor; the delivery is controlled by the vane A. The fan and pipes are covered with ¼ n. (6.3 mm.) asbestos cement lagging.

Jan 4, 1922

By A. R. Anderson

UNTIL recently all discussions directed at justify-ing the use of mechanical-loading equipment and conveyors have referred chiefly to tons per man and cost per ton. But there is another consideration that ? will speed the adoption of mechanical loading in mines even though tons per man on present day stand-ards were not increased nor the cost per ton decreased. With restricted immigration, improved facilities for education in our mining communities, and increased opportunities for recreation, the type of man who can or will throw 20 tons of coal per day over the side of a mine car is disappearing. Whether we will it or not provision must be made to replace muscular effort by mechanical means. Fortunately, indications are that the process will be accompanied by increased output per man and lower production cost, and though the outlook may now be a bit clouded in spots, all should profit by the change. The operator should enjoy a more stable business with more assured profit, the miners should get a better livelihood with less arduous effort and under safer and pleasanter working conditions, and the consumer should gain the assurance of a supply of coal always available at reasonable prices. That sounds like the millennium in the coal business perhaps. Of course, it is not, but there is just enough of that millennium stuff in it to encourage those of us who are working with today's problems of applying loading and conveying equipment to underground opera-tions. If we discern clearly that there can be no back-sliding, and that all will share in whatever advantages accrue, we will carry on more vigorously, and may right-fully expect to have the necessary cooperation of all the elements within the industry. As nature does not seem to have followed a definite plan in the formation of coal deposits it is not likely that any one standard method or type of equipment can be evolved for mining the coal. The method of mining that maybe most advantageously employed with load-ing machines and conveyors depends on the character of the strata overlying the coal as much as anything else.

Jan 4, 1927

By J. P. Morris

PREVIOUS investigations1,2 have shown that alloying elements in liquid iron influence the thermodynamic activity of sulphur and thereby affect the partition of sulphur between metal and slag in the desulphurization process. For example, the greater efficiency of desulphurization in the blast furnace as compared to the open hearth can be attributed in part to a higher level of sulphur activity in blast-furnace metal due to the higher concentration of carbon and silicon. In the present investigation, a short study was made of the influence of manganese on the activity of sulphur in liquid iron and iron-carbon alloys. In contrast to carbon and silicon, manganese was found to decrease the activity coefficient of sulphur; and in iron-carbon alloys it counteracts to some extent the influence of carbon. However, at manganese concentrations normally present in the blast furnace or open hearth, the effect of manganese is small. Since manganese sulphide has a limited solubility in iron, manganese can act, under certain conditions as a desulphurizing agent. Considerable data on the manganese-sulphur product in carbon-saturated melts were obtained in the investigation and have been included in this report. The experimental procedure was the same as that used in the earlier investigations on the effect of silicon' and carbon' on sulphur activity. Briefly, the method was as follows: The molten alloy, contained in a graphite or sintered alumina crucible, was brought to equilibrium at a constant temperature with a mixture of hydrogen and hydrogen sulphide of constant composition by bubbling the gas through the metal. Samples of' the melt were taken for analysis at regular intervals by suction through a 2 to 3 mm bore silica tube dipped into the metal. The experiments were run in a graphite spiral resistance furnace with melts weighing 50 to 60 g. The gas bubbling tubes were made of sintered alumina and were 5/16 in. OD, 1/16 in. ID, and 24 in. long. Equilibrium was assumed to have been attained when the sulphur content of the liquid metal reached a constant value. During an experiment there was a rapid loss of manganese from the melt by volatilization. To offset this loss, small additions of manganese were made periodically. The rate of manganese addition needed to maintain a constant manganese concentration was determined in preliminary tests. In all of the experiments, deposits of manganese sulphide formed above the melts in a cooler region of the furnace. Apparently, these deposits resulted from a reaction between manganese vapor and hydrogen sulphide in the gas. To prove that manganese sulphide did not volatilize from the melts to a measurable extent, an experiment was run in which helium was bubbled through liquid iron containing both manganese and sulphur. Although manganese volatilized rapidly in this test, there was no appreciable loss of sulphur. Volatilization of manganese sulphide from a melt would have led to an apparent equilibrium condition in which the sulphur content of the metal was lower than the true equilibrium value. The experimental results are shown in the first seven columns of Table I. The data in the last two columns were obtained from the previous work on the effect of carbon' and show what the results would have been in the absence of manganese but with temperature, gas composition, and carbon content of the metal remaining the same. Comparison of the last four columns show that, in the presence of manganese, the sulphur content of the metal increased at equilibrium and the activity coefficient of sulphur decreased. However, the results show that, for manganese concentrations below 3 pct, the effect of manganese is small. The values for activity coefficient of sulphur given in Table I were calculated from the following relations: S (in liquid metal) + H2 (gas) = H2S (gas) [l] K ph2s/?s X %S X phg = 0.00251 [2] where K is the equilibrium constant for the reaction, PH2S and ph2 are the partial pressures of hydrogen sulphide and hydrogen, respectively, and ?s is the activity coefficient of sulphur. The standard state for sulphur was taken to be a 1 pct solution of sulphur in pure iron. The numerical value for K at 1600°C was determined in the earlier work. For the purpose of showing graphically the results of the tests run at 1600°C, the activity coefficients of sulphur were recalculated so as to correspond to a manganese concentration in the metal of 2 pct in each case. In the calculation it was assumed that the increase in sulphur content of the metal at equilibrium caused by the presence of manganese

Jan 1, 1953

By R. M. Treco

The effect of small additions of oxygen on the hardness, density, and lattice parameters of high purity zirconium has been investigated. Precision parameter values and density of oxygen-free zirconium are given as: a,, - 3.2258 kX units, C = 5.1373 kX units, O23 = 6.507 g per cc. THE effect of oxygen on the hexagonal-phase lattice constants of pure zirconium is of interest for several reasons. The extensive hardening effect of oxygen noted by previous investigators' must be due in part to the lattice distortion accompanying its introduction and possibly to the preferred position of the oxygen atoms. De Boer and Fast' have shown that oxygen is taken into solid solution in large amounts, appearing to occupy an interstitial position. These authors also found that expansion of the lattice was not uniform since the a axis attained a constant value after the addition of 10 atomic pct while the c axis continued to expand. It was thought in the present work that this expansion of the zirconium lattice with increasing amounts of oxygen

Jan 1, 1954

By G. W. Powell

G. W. Powell (Ohio State University)—Appleton and Bever measured the quantity Es which, for a solid torsion specimen, is basically an average quantity. Es is the amount of energy stored in a solid torsion specimen strained to a surface shear strain of divided by the total number of gram-atoms of alloy in the specimen. It would seem more meaningful and rigorous to deal instead with the quantity which is the energy stored per gram-atom of an alloy uniformly strained to the shear strain . The purpose of this discussion is to demonstrate that as a function of y is readily obtained from the torsion data. radius of the torsion specimen, shear strain at the surface of the torsion specimen density of the alloy, atomic weight of the alloy, and total energy stored in unit length of the torsion specimen. The energy stored in unit length of an elemental tubular section of the torsion specimen is, as given by Appleton and Bever, where E is expressed as calories per gram-atom. The energy stored in unit length of the torsion specimen is

Jan 1, 1963

THE Pacific Northwest's joint meeting of the Industrial Minerals and Metals Branch in Spokane, Wash., drew 260 members, 72 students, and 30 ladies. Theme of the meeting was to acquaint men in the industry with students and younger men. The special appeal to the students was coupled with an attempt to attract young technical men from the companies located in the northwestern area. The activities of the meeting were available without charge to students. Contributions were made by the industries in the area to assist student attendance. Following the special student and faculty trip to the Trentwood rolling mills and research laboratory on Friday afternoon everyone enjoyed a complimentary dinner. A. Y. Bethune of the Sullivan Mining Co., spoke to the group on the operations of Kellogg, Idaho electrolytic zinc plant. The Industrial Minerals Div. session was opened on Saturday morning by R. G. Vervaeke, Conference Chairman who introduced A. O. Bartell of the Raw Materials Survey, Portland, Ore. Mr. Bartell in turn introduced C. W. Sweetwood who discussed Open-Pit Phosphate Mining.

Jan 1, 1952

By R. Q. Shotts

Nitric acid oxidation rate analyses of three coals, previously studied microscopically by the Bureau of Mines, revealed three components. Relative quantities agree with those found for the four components given by the Bureau and results are consistent with current ideas of coal constitution. Possible multi-component composition for bright coal and a reactivity-rank relation are suggested. THE physically dissimilar components of bituminous coals often are easily recognized mega-scopically. Under the microscope, reflected light or light transmitted through thin sections reveals the presence of the different components, even when these are intimately mixed. Optical methods for the quantitative estimation of the relative abundance of the various components, both by means of thin sections and by particle count, have been fully described.'. ' It has long been recognized that there are chemical and physical differences between the various petro-graphic components of bituminous coals, although analytical differences usually are small.:'. ' Only in the case of fusain have chemical differences been used for quantitative determination of a component. C. C. Hsiao and associates, at the Mineral Industries Experiment Station of the Pennsylvania State College, have described a method of analysis which is based upon the differences in the rate of nitric acid (8N) oxidation, fusain, and the other components of coal."," The reproducibility of their method and its applicability in checking microscopic determinations of fusain content have been supported by several independent investigations.'. " The writer has proposed the use of differences in oxidizability for the estimation of other components." "' The results of the oxidation of whole coals and of float-and-sink fractions of coals were reported. In most cases the plots of the logarithms of the percent dry, non-fusain, organic residue from oxidation, against time, revealed the presence of at least two distinct components. Both components appeared to oxidize according to a first order law, but the reaction constants for the components were distinctly different. One or more of the dull density fractions were found to contain but one component, and some of the lower rank coals oxidized in such a way as to suggest the presence of three components. A suitable way to check the identity and significance of the components delineated by oxidation would be to analyze a sample of coal both by the nitric acid oxidation procedure and by a microscopic method. The writer was wholly unfamiliar with either of the microscopic techniques commonly used, and to make such a comparison it was necessary to rely upon microscopic analyses made by someone else. It is hoped that some laboratory which is equipped to make both types of analyses will some day make them upon identical samples. During the past 20 years, four Alabama coals have been analyzed petrographically and the results published by the United States Bureau of Mines. They are: 1—Flat Top mine, Mary Lee bed; 2—Empire mine, Black Creek bed; 3—Wylam No. 8 mine, Pratt bed, all in the Warrior field; and 4—Soot Creek mine, Fairview bed, in the Coosa field."-" Of these, only the Flat Top mine is still operating. Because of the closing of these mines, it first appeared necessary to rely upon the indirect and unsatisfactory procedure of sampling the beds in other mines located as near to the closed mines as possible. Upon investigation, however, it was found that the Bureau of Mines still had, in storage, the very same samples which had been used in the published petrographic studies. The Bureau very generously furnished about 2000 g each of the Pratt, Mary Lee, and Fairview bed coals, largely lumps but with some fines. The blocks of coal, when received, still were covered by the paraffin coating which had been placed on the polished surface, in the case of the Mary Lee coal almost twenty years ago. Procedure The procedure for oxidizing the coal sample and removing the alkali-soluble humic acid has been described. In the present study, oxidation periods of 1/6, 1/3, 1/2, 3/4, 1, 2, 3, and 4 hr were used. All oxidations were made in triplicate. After the paraffin had been removed in boiling water and the coal washed carefully with cold benzene, the entire sample of approximately 2000 g, obtained from the Bureau of Mines, was crushed to pass a No. 4 sieve. About 200 g of this material was pulverized to pass

Jan 1, 1954

By K. E. Beu

IN the measurement of retained austenite concentrations in steels using the integrated intensity method,1 Averbach has pointed out" that the absorption factor A(0) for a flat sample making a glancing angle 4 with the incident X-ray beam can be combined with his constant factor R to obtain our constant factor" G provided that "1—There is no preferred orientation in the sample, and 2—The geometric requirements [for the sample, film, and X-ray beam] have been met precisely."' If A(0) can be combined with R to give G according to the equation G = R . A(0) [14] then the possibility for making non-compensating errors in the austenite determination has been eliminated. This has been described previously in a technical note." Because of the brevity of the previous note,3 it was impossible to emphasize the fact that the two conditions on preferred orientation and geometry can be easily met experimentally, contrary to Aver-bach's statement that: ". . . the necessary conditions must be tested experimentally for each determination, and this is done most easily by observing whether the apparent absorption has the form of Eq. 2 [the theoretical equation for A(0)]."" he way in which these two conditions can be met experimentally will be discussed briefly to help clear up this point. In addition to these two conditions, other factors such as sample shape, homogeneity, and grain size which also affect A will be included in this discussion. A (0) depends on sample shape. The theoretical function for A was derived originally for a flat surface.' In general we have found that a flat sample surface is readily obtainable." If such a surface can * The effect of surfaces which are not flat on the measurement of retained austenite will be discussed later. be obtained, it has the following advantages: 1—it is easily reproducible from sample to sample, 2—it is the form required for metallurgical examination —this type of examination being frequently desirable for this work, and 3—it is an efficient shape for diffraction purposes. Perhaps the most important of these features is that a flat sample surface is easily reproducible; hence, this requirement on the repro-ducibility of A from sample to sample is met for all such samples. Fig. 1—Schematic diagram of the quartz crystal monochromator diffraction unit. The centerline of the main beam is at 6' to the target face. The tangent to the crystal face is at 16.8' to the moin beam. The centerline of the monochromatic beam is at 33.6 to the main beam. The sample surface can be rotated in its own plane. The angle of the sample surface and the monochromatic beam can be adjusted by rotating about the vertical axis, The film holder can also be rotated about B so that the film can be exposed over the desired angular range. For Fe K, X = 1.932A. 1011 planes of quartz have d = 3.35A. A(0) depends on homogeneity and grain size. If the sample is badly segregated or the grain size is large, the effect of micro-absorption and primary extinction1. ' must be considered. It has been shown, however, that for most plain carbon or low alloy hardened steels, neither micro-absorption nor primary extinction effects are present.' A (0) depends on camera geometry. For a given angle 0, A remains theoretically constant from a geometrical viewpoint only if the following factors are kept constant: 1—the angle of inclination .+ of the sample to the X-ray beam, and 2—the centering of the sample with respect to the film cylinder. These are mechanical problems which can be solved readily if the facilities of a good machine shop are available. The arrangement used to insure that the angle 4 remains constant and the centering of the sample is reproducible is indicated schematically in Fig. 1. The sample is clamped against a thin plate with a hole in it by means of a spring-loaded pres-

Jan 1, 1954

By Joseph G. Wargo

An investigation of trends in exploration expenditures for a selected group of mining companies was undertaken for the interval 1968-1971. These trends were compared with financial factors that are associated with various groups of companies and the implications of these trends are indicated. The sample consists of 21 mining companies, of these, 19 are listed on the New York Stock Exchange and 2 are listed on the American Exchange. Twenty are American, one is Canadian. All are fundamentally "mining companies" in that they derive the bulk of their income from the ex- traction, sale, and at times, fabrication of metals from their own mines. This is in contrast to companies which have mineral exploration departments but derive the bulk of their revenues from sources other than mining, such as oil and gas sales. No companies which can be considered strictly as exploration companies are included in the sample. This definition held up in 1968, however, its strict application became more difficult through the years. In 1971, three or four of the companies were changing their character in that they were phasing out of the mining business and deriving a substantial part of their income from activities other than mining.

Jan 5, 1973