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By B. T. Brady

Rock mechanics, like all engineering disciplines, must have a theoretical foundation. The subject of this chapter is the formulation of analytical methods that may aid in the rational design and determination of the stability of engineering structures in competent brittle rock materials. By an engineering structure in rock, we mean any (1) excavated sub- surface opening, (2) open surface structure, or (3) system of openings in rock that is virtually self-supporting-that is, not requiring artificial systems to carry structural stresses.

Jan 1, 1970

By T. H. Courtney, G. W. Pearsall, J. Wulff

The mechanism, of the decrease in transition temperature of Nb3Sn was studied. Presintered Nb3Sn powder compacts were heated in a dynamic vacuum at elevated temperatures. Loss of tin resulted from these heat treatments and transition temperature as a function of tin loss was discontinuous at a critical value of tin loss. In general, decreases in transition temperature were correlated with deviations from AS a class, the A15 structure-type (also known as Cr3O or ß W) compounds have the highest superconducting transition temperatures.' In this class are stoichiometry and presence of vacancies in niobium sites. The effect of vacancy distribution on the physical properties of Nh3Sn is considered. It is concluded that the close-packed atomic A chain in the A3B A15 structure must be preserved both in coutinuity and direction if good superconducting properties are to he obtained. tempting to associate these high transition temperatures with the most unique crystallographic feature of the A3B structure: the three close-packed, orthogonal chains of A atoms.'0"' There are at least three principal mechanisms by which the perfection of these "A-atom chains" could be interrupted: changing stoichiometry, inducing disorder, or introducing vacancies. Departures from stoichiometry in either direction in V3Ga lower the transition temperature.7'8 In niobium-

Jan 1, 1965

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO2 oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO2 can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1 ½ ¢ per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past year that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 ½ million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ¾ ton of high-grade pyrite and results in ½ ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15¢ per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1952

By M. G. Corson

The iron-silicon alloy series has always been one of the most puzzling among the binary alloys. Examining the well-known mechanical properties of the iron-rich alloys only we meet the following situation: The ultimate strength and yield point of alloys worked and annealed under the usual conditions represent nearly a straight line function of the silicon content up to 1.5 per cent. silicon (Fig. I). Starting at the ultimate strength of 35,000 lb. and at a yield point of 16,000 lb. for pure electrolytic iron, one reaches a maximurn strength of 93,000 lb. and a maximum yield point of 74,000 lb. at 4.5 per cent. silicon. This corresponds to an increase in ultimate strength by 13,000 lb, per each per cent. of silicon added and nearly the same rate applies to the yield point. The data mentioned are taken from the investigation by T. D. Yensen,1 and while they do not exactly corroborate the results obtained by earlier investigators, they seem to be the most correct. Plastic properties of the alloys are controlled by a much more complicated relationship. Up to 2..5 per cent. Si, the elongation stays close to 50 per cent. (in 2 in.) and the area reduction amounts to 85 per cent., a situation that forms the usual feature of all alloys representing strictly solid solutions. The plasticity of the latter is always nearly equal to that of their base metal. Beyond 2.5 per cent. silicon, the situation becomes much more complicated. Yensen observed an enormous drop in elongation and area reduction at about 2.65 per cent. silicon, where the first hardly reaches 10 per cent. and the second is still lower; he found, however, a considerable recovery in plasticity at about 2.8 per cent. Si and up to 1.5 per cent. Si it is represented by 20 per cent. for the elongation and 30 per cent. for area reduction. Yensen's graphs, as shown in Fig. 1, indicate also a considerable drop in strength and yield point at the same point of 2.65 per cent. Si,

By G. B. Walker, S. A. Falconer, S. J. Swainson, H. H. Vogel

During the past few years much interest and attention has been focused on a relatively new method of ore concentration, which utilizes the principles of sink and float and employs as the heavy medium a suspension in water of a finely ground solid, such as galena or ferrosilicon. Although a relatively recent development, much progress has already been made in applying heavy-media separation processes to a considerable number of ores. Eight plants in this country and abroad are now treating approximately 10 million tons of ore annually by heavy-media separation (sink-float) processes, including lead-zinc, lead, iron, garnet, tin and tungsten. At the annual meeting of the A.I.M.E. in 1941, three interesting papers (unpublished) dealing with heavy-media separation (sink-float) processes of ore concentration as practiced in four commercial units in the United States were presented by Victor Rakowsky, Grover Holt, Elmer Isern and Robert Ammon. These papers gave an excellent exposition of the principles of sink and float, as well as the operating practice in two iron-ore-beneficiation plants, the Harrison and the Merritt, of Butler Brothers, in Minnesota; the Central mill of the Eagle Picher Mining and Smelting Co.; and the Mascot mill, of the American Zinc, Lead and Smelting Co. Augmented by pictorial descriptions and a working model demonstrating the processes, these papers evoked a great deal of interest and discussion. In fact, in reporting the high lights of the meeting in the April 1941 issue of Mining and Metallurgy, Professor Taggart made the statement that " preconcentration by heavy suspension was Miss A.1.M.E 1941 so far as milling is concerned." Although to date the greatest application of heavy-media separation from the tonnage point of view has been for the treatment of lead, zinc and iron ores, these processes have a much wider field of usefulness. In this regard, it is to be noted that recently new heavy-media separation plants for treating cassiterite ore and garnet ore have been placed in successful operation. Moreover, numerous other types of ores have been found to respond well to these processes when tested on a semi-commercial scale, and it is expected that in the near future a considerable number of new plants will be erected. In view of the fact that heavy-media separation processes offer such attractive possibilities, particularly now, when the mining industry is being called on for an even larger production than ever before has been attempted, despite shortages of labor and supplies, it was felt that a description of some of the successful separations that have been worked out during the past year in the Ore Dressing Laboratory of the American Cyanamid Co. on a variety of ores of strategic importance would be of interest. The purpose of this paper, therefore, is to present the essential data relating to investigations of heavy-

Jan 1, 1943

By G. B. Walker, S. J. Swainson, S. A. Falconer, H. H. Vogel

During the past few years much interest and attention has been focused on a relatively new method of ore concentration, which utilizes the principles of sink and float and employs as the heavy medium a suspension in water of a finely ground solid, such as galena or ferrosilicon. Although a relatively recent development, much progress has already been made in applying heavy-media separation processes to a considerable number of ores. Eight plants in this country and abroad are now treating approximately 10 million tons of ore annually by heavy-media separation (sink-float) processes, including lead-zinc, lead, iron, garnet, tin and tungsten. At the annual meeting of the A.I.M.E. in 1941, three interesting papers (unpublished) dealing with heavy-media separation (sink-float) processes of ore concentration as practiced in four commercial units in the United States were presented by Victor Rakowsky, Grover Holt, Elmer Isern and Robert Ammon. These papers gave an excellent exposition of the principles of sink and float, as well as the operating practice in two iron-ore-beneficiation plants, the Harrison and the Merritt, of Butler Brothers, in Minnesota; the Central mill of the Eagle Picher Mining and Smelting Co.; and the Mascot mill, of the American Zinc, Lead and Smelting Co. Augmented by pictorial descriptions and a working model demonstrating the processes, these papers evoked a great deal of interest and discussion. In fact, in reporting the high lights of the meeting in the April 1941 issue of Mining and Metallurgy, Professor Taggart made the statement that " preconcentration by heavy suspension was Miss A.1.M.E 1941 so far as milling is concerned." Although to date the greatest application of heavy-media separation from the tonnage point of view has been for the treatment of lead, zinc and iron ores, these processes have a much wider field of usefulness. In this regard, it is to be noted that recently new heavy-media separation plants for treating cassiterite ore and garnet ore have been placed in successful operation. Moreover, numerous other types of ores have been found to respond well to these processes when tested on a semi-commercial scale, and it is expected that in the near future a considerable number of new plants will be erected. In view of the fact that heavy-media separation processes offer such attractive possibilities, particularly now, when the mining industry is being called on for an even larger production than ever before has been attempted, despite shortages of labor and supplies, it was felt that a description of some of the successful separations that have been worked out during the past year in the Ore Dressing Laboratory of the American Cyanamid Co. on a variety of ores of strategic importance would be of interest. The purpose of this paper, therefore, is to present the essential data relating to investigations of heavy-

Jan 1, 1943

By Charles H. Moore, H. Sigurdson

When lime and magnesia are used as fluxes in the smelting of titaniferous ores fluid, digestible slags low in iron oxide and high in titanium dioxide are produced. The mineral phases present in such slags are (Fe,Mg)O.ZTiOp, CaTiO3, and minor amounts of silicate glass (or its devitrification products). The flux additions are added to the ore in relative amounts to crystallize a ratio of (Fe,Mg)O.2TiO2 to CaTiO, of near the eutectic proportions of 60 pct (Fe,Mg)O.2TiO2 and 40 pct CaTiOl. FeO.TiOz (ilmenite) remains present in the slag until the FeO content is reduced to approximately 8 pct. Optimum slag composition is obtained by reducing the FeO content to between 2 and 6 pct. Continued reduction shifts the (Fe,Mg)O.2TiO2 to oxygen deficient structures classified as MgTi306 and MgTi2O.. Further reduction yields titanium oxycarbides and freezes the melts. The(Fe,Mg)0.2TiO2 structure shows a complete isomorphous replacement of iron by magnesium. It also forms a solid solution compound with as high as 12 mols of TiO2 under reducing conditions without the occurrence of rutile or reduced titanium oxides in the slag.

Jan 1, 1950

By E. F. Poncelet

Forty years ago A. A. Griffith developed a theory explaining why brittle materials displayed such low tensile strengths.' He based his views on two points. First, he found himself compelled to assume that all brittle materials are replete with flaws, cracks, and other defects that act, although quite invisible, as large stress raisers. Second, he applied the "theorem of minimum potential energy," which says that the total potential energy of a system must pass from the unbroken to the broken condition by a process involving a continuous decrease in potential energy. By this means he satisfactorily accounted for the noted low strength of such solids and also for the wide spread obtained in experimental measurement of these strengths. So successful has the theory been that it is favored by some to this day. Unfortunately this theory is of limited use beyond the explanation of these two noted phenomena and it is keenly felt that a better theoretical insight into the physics of the fracturing process is needed as the volume of experimental evidence accumulates. The author proposes in the following to build on the fundamentally sound concepts of Griffith and, with the help of increased theoretical knowledge over that available to Griffith, develop a mechanism for frac-ture which will provide far greater understanding of the experimental evidence accumulated to date than is possible from the original Griffith idea. THE GRlFFlTH THEORY Very little progress indeed can be made without accepting the first postulate of Griffith which supposes all brittle solids to be full of microcracks. It would be difficult indeed to find a better mechanism for the small strength of such brittle materials, in conjunction with the fact that the energy that must be expended for comminution is by no means small. The postulate of the existence of the microcracks permits the breakup of the various bonds a few at a time by concentrating the stress at the tip of the progressing crack, while the total energy expended is the same as if they all had been ruptured simultaneously. The only flaw in the argument is that no reasonable explanation has been proposed to account for the genesis of such cracks. Indeed their very presence is in violation of the Griffith second postulate, the potential energy theorem. This theorem is straightforward for isothermal processes, and, in spite of Griffith, there is some doubt that treating the problem isothermally is legitimate. The surface energy of bodies is a free energy, not a potential energy as stated by Griffith, and the production of new surface free energies is not necessarily an isothermal process. There is ample evidence to the contrary. Generally speaking, if heating a body increases its surface area, then, by virtue of Le Chatelier's principle, any increase of that area by other means will tend to lower its surface temperature. Lord Kelvin calculated the actual cooling that resulted in drawing out a film of liquid.2 R. A. Houston calculated the surface cooling that resulted in stretching a metal wire.3 These calculations were made by applying the Carnot cycle to the process and evaluating the thermodynamics thereof. IRREVERSIBILITY OF THE FRACTURING PROCESS While Griffith was very careful not to say so, the impression gained from studying his papers is that he considered the fracturing process as reversible, that is, a succession of quasi-equilibrium states. There is ample evidence that it is not. The indication that the new surfaces produced by the propagation of a crack are cooler than the original body points to an irreversible heat flow from the interior to the new surfaces to equalize the temperatures. If the process be reversible, any crack accidentally formed should immediately close up as, in the absence of any strain energy, the potential energy would thereby be lowered. The fact that they do not, constitutes a paradox. Such paradoxes are nothing new where certain phenomena that propagate from minuscule nuclei are assumed to be reversible. Such is, for instance, the condensation of a pure saturated vapor that is suddenly chilled by adiabatic expansion. At the beginning the tiny droplets that are formed should be only a few angstroms in size, but the vapor pressure at such droplets is so high that they should evaporate at once. A similar situation arises if a saturated pure solution becomes super-saturated upon cooling; the first tiny crystal nuclei should dissolve as fast as

Jan 1, 1964

By Willard C. Lacy

Rather than attempting to present a summary of the many and highly varied papers that have been presented at this symposium on sampling and grade control, I will attempt to extract the general philosophy of analysis and approach, and attempt to identify the trend of future developments. First, the term "sampling" is used with its broadest connotations. A sample consists of a representative portion of a larger mass, and must represent the mass not only in the grade of contained metals or minerals, but also in all other respects in terms of mineralogy and mineral quality (1, 5), deleterious materials, recoverability of economic components, physical behavior, geophysical response (I), and even archaeological and environmental aspects (7, 11). The sample must be taken from a locality and in such a manner and quantity that it is representative of the larger rock mass. This calls for complete and accurate geological control and an understanding of the nature and distribution of the contained chemical and physical elements and a record of the effectiveness of the different sampling methods. Second, value of a given mass of ore material is based upon its profitability - the difference between recoverable value and costs to achieve recovery, beneficiation and sale. There is a strong movement in mining geology control toward more complete analysis in determining cutoff grades and in grade control, as illustrated by the kriging of metallurgical recovery factors as well as grade at the Mercur Mine (8). To achieve a "profit- ability factor" as a guide for economic mining practice requires further integration of: 1) the value of contained metal or mineral, 2) percentage recovery of values, 3) dilution of ore with waste rock, 4) addition to, or loss of value as a consequence of by-product materials or deleterious components, 5) cost of producing a saleable product plus mini- mum profit to justify the effort (cutoff), and 6) cost of land restoration (7, 11). All these parameters vary with the rock type, rock structure, mineralogy, depth, geometry, mining and metallurgical methods, but they must be sampled and analyzed if sampling and grade control are to reflect profitability. A wide variety of deposits has been presented at this symposium; each deposit with its own problems and special solutions. Deposits containing high unit-value components, e.g. precious metals and diamonds, present special problems in the obtaining of accurate samples and generally require statistical analysis control methods or may disregard or modify occasional high or occasional low values, based upon experience (12 ) Grade control may be accurate for the long term but may vary for the short term. Bulk sampling is always essential. Deposits containing metals or minerals with low unit value are very sensitive to transport costs, and they are often very sensitive to small amounts of deleterious components or differences in physical or chemical behavior. Problems of sampling and grade control change with the genetic type of deposit, with the stage of deposit development and with the size of the information base. Precious metal epithermal deposits (2, 6, 8), because of rapid vertical zonation and erratic lateral distribution of values, have always been difficult to evaluate and maintain grade control and ore reserves. On the other hand, evaluation and grade control are relatively easy in bulk-low- grade deposits (4, 13). However, these deposits generally have a low margin of profit and are sensitive to mining and beneficiaton costs, price fluctuations and political costs. Industrial mineral deposits (5) often must be evaluated on the basis of their behavior, rather than by chemical analysis. Environmental impact generally increases with the scale of the operation, but certain elements or minerals have especially high impact effects (7, 11). In the exploration phase there is no production control of sampling procedures and careful geological observations are particularly essential. The greatest number of problems is related to the oxidized outcrop where the chemical environment of the ore body has changed and the contained values may have been enriched, depleted or values left unchanged (2, 6). Present evidence suggests that gold values may be very mobile under certain conditions (2, 6) and stable under others. Everything must be sampled in detail. Principal values and by-product or deleterious elements may vary dependent upon their position within the soil profile. Such factors as geomorphic position, erosion rate, vegetation, climate, etc., may affect the interpretation (1, 3). During the development phase it is equally easy to overtest, to have "paralysis by analysis," as to undertest (3, 6). Bulk samplng and testing are

Jan 1, 1985

By A. W. Furbank

In reviewing the aerial cameras produced in different countries, it becomes apparent that in nearly all of them an attempt has been made to secure the greatest possible angle of view. This angle, of course, is determined by the objective. The objectives most commonly in use in aerial cameras have a maximum angle of 65". The objective developed for this new Zeiss camera combines freedom from distortion and brilliant definition, with an angle of 95". Aerial mapping, it is claimed, has been placed on a different economic basis, and new possibilities presented, particularly in connection with automatic stereoscopic plotting instruments, to which some short references are made. One way of enlarging the angle of view is to couple several instruments with their axes convergent, thereby producing a multilens camera. Typical of cameras of this sort in the United States is the Bagley four-lens camera, which later was developed into the five-lens camera (T3-A) of the U. S. Air Corps. More recently, in order to get away from the unfavorable Maltese cross image field of the five-lens camera, and to obtain a more practical form, two five-lens cameras were combined. By using them in tandem and rotating them by 45" relatively to each other, a ten-lens camera was produced, covering an octagonal ground area, with a total angle of view of 140". Similar designs have been evolved in other countries. In all of them, however, a number of difficulties have arisen, quite apart from their large weight and high cost. Because of their complex mechanism, numerous shutters and magazines, and the high precision with which these parts must be interadjusted and synchronized, these cameras are far from dependable. Moreover, a multilens camera by itself does not yet constitute a complete outfit for producing pictures suitable for photogrammetric survey. Because all the component pictures, except the central one, are obliques, it is always necessary to employ rectifying apparatus which transforms the marginal views to the perspective of the central one. In multilens cameras with many individual magazines, it is not convenient to

Jan 1, 1941

By A. W. Furbank

In reviewing the aerial cameras produced in different countries, it becomes apparent that in nearly all of them an attempt has been made to secure the greatest possible angle of view. This angle, of course, is determined by the objective. The objectives most commonly in use in aerial cameras have a maximum angle of 65". The objective developed for this new Zeiss camera combines freedom from distortion and brilliant definition, with an angle of 95". Aerial mapping, it is claimed, has been placed on a different economic basis, and new possibilities presented, particularly in connection with automatic stereoscopic plotting instruments, to which some short references are made. One way of enlarging the angle of view is to couple several instruments with their axes convergent, thereby producing a multilens camera. Typical of cameras of this sort in the United States is the Bagley four-lens camera, which later was developed into the five-lens camera (T3-A) of the U. S. Air Corps. More recently, in order to get away from the unfavorable Maltese cross image field of the five-lens camera, and to obtain a more practical form, two five-lens cameras were combined. By using them in tandem and rotating them by 45" relatively to each other, a ten-lens camera was produced, covering an octagonal ground area, with a total angle of view of 140". Similar designs have been evolved in other countries. In all of them, however, a number of difficulties have arisen, quite apart from their large weight and high cost. Because of their complex mechanism, numerous shutters and magazines, and the high precision with which these parts must be interadjusted and synchronized, these cameras are far from dependable. Moreover, a multilens camera by itself does not yet constitute a complete outfit for producing pictures suitable for photogrammetric survey. Because all the component pictures, except the central one, are obliques, it is always necessary to employ rectifying apparatus which transforms the marginal views to the perspective of the central one. In multilens cameras with many individual magazines, it is not convenient to

Jan 1, 1941

NEW MEMBERS The following list comprises the names of those persons who became members during the period Apr. 10, 1919, to May 10, 1919. BEAVER, JAMES J., Supt., Mech. Ore Loaders, Tennessee Coal, Iron & R.R. Co., Box 62, Bessemer, Ala. BLACKWELDER, ELIOT Geol. & Prof. of Geol., Univ. of Ill., Urbana, Ill. BRAECKE, GUSTAVE E., -Min. Engr., Mgr., Societe Curas et Soldado, La Carolina, Spain. BRINK, CYRIL GORDON, Reduction Officer, Charge of Reduction Works. Fairview Devonian Montrose Gold Mines, Ltd., P. O. Sheba, Transvaal, So. Africa. COMERFORD, JOSEPH S., Chief Chem. & Met., American Manganese Steel Co., New Castle, Delaware. DROSDOWSKI, N. C. V., Asst. Supt.,, The Spassky Copper Mine, Ltd., Karagandy, Akmolinsk, Oblast, Siberia. DUNGLINSON, GEORGE, JR., Asst. to Gen'l Mgr., Norfolk & Western Railroad, Bluefield, W. Va. FARAGHER, PAUL VANCE, Research Fellow on Light Metals, Mellon Institute, Pittsburgh, Pa.

Jan 6, 1919

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By F. R. Morral

HOT-DIPPED heat-treated zinc coatings on sheet steel were examined with X-rays. The phases identified were compared with those listed in the modern literature of the zinc-iron system. The zeta (FeZn13) phase is shown to correspond to the phase found by Schueler (1925) in heat-treated zinc coatings. An investigation was undertaken to determine the phases present in hot-dipped FIG. I.-SYSTEM IRON-ZINC ACCORDING TO J. SCHRAMM.11 heat-treated zinc coatings on steel sheets and wire. This kind of coating on wire was studied some years ago by Schueler1 by microscopic and chemical methods. He reported a new phase, FeZn10. Later, Clark and Stillwell,2 using X-ray crystal analysis, identified a new phase in ordinary galva-nized wire, which was assumed to be the FeZn10 phase.

Jan 1, 1940

By R. J. Brison, O. F. Tangel

The development of a new method of mineral separation was sponsored by the International Salt Company, which requested Battelle Institute to investigate means for improving the quality and appearance of rock salt from the Company's Detroit mine. Although developed specifically for removing impurities from rock salt, the general method may be applicable to other separation problems. The principal impurities in rock salt from the Detroit mine are dolomite and anhydrite which represent 2 to 5 pct of the weight of the mined salt. In the size range from 1/4 to M in. (the range of primary interest in this project) the impurities are only partially liberated from the halite in normal production. Further size reduction to improve the liberation of impurities is not practicable in view of the market requirements for the coarse grades of rock salt. Laboratory separations in heavy liquids showed that, to improve the quality and appearance of the rock salt substantially, it would be necessary to remove not only free gangue particles but also a large proportion of the locked-in particles. Because rock salt is an inexpensive commodity, a low-cost process was required. Gravity methods were, of course, considered. The heavy-liquid separations indicated that a split at an effective specific gravity of 2.2 to 2.3 would be required. (The specific gravity of pure halite is 2.16.) Heavy-media separation was investigated but had the disadvantages that it was necessary both to operate with saturated brine and to dry the cleaned salt, and that the cleaned salt was darkened by the magnetite medium. Air tabling was tried but did not give the desired separation. It soon became apparent that established methods would not provide a satisfactory solution and work was undertaken on the development of a new process to solve the problem. PROCESS DEVELOPMENT Preliminary Experiments: At the start of the investigation, an analysis of the problem indicated that the diathermacy of rock salt—that is, its ability to transmit radiant heat—might form the basis for an efficient separation process. Under this theory, the impurities might be selectively heated by radiant heat. The particles could then be fed over a belt coated with a heat-sensitive substance so that the warm impure particles would adhere preferentially to the coating. After the initial experiments, made by heating the rock salt with an infrared lamp and separating the product on small sheets of resin-coated rubber, proved encouraging, a small continuous separation unit was set up. This comprised 1) a simple heating unit consisting of a vibrating feeder covered with aluminum foil and an infrared lamp mounted above the feeder and 2) a separation belt 6 in. wide and 36 in. long. A sketch of the device is shown in Fig. 1. Results with this apparatus confirmed the fact that a good separation was possible. It was apparent, however, that a considerable amount of experimental work would be needed to develop the scheme to a practical and economical process. The Process: Basically, the process consists of two main steps: 1) selective heating by radiation and 2) separation of the heated particles on a heat-sensitive surface. Because neither of these steps had previously been utilized commercially in mineral processing, it was necessary to do basic research on both aspects. Factors studied in the investigation included type of heat source, design of heating unit, design of separation belt, selection of heat-sensitive coating, removal of heated particles from the belt, contact between particles and coating, and maintenance of the heat-sensitive surface. Part of the experimental work was carried out on a small-scale unit consisting of the 36x6 in. belt and auxiliary apparatus, and part on a larger unit. For simplicity, discussion of work on both of these units is grouped together. SELECTIVE HEATING Radiant-Heat Source: The essential requirements for a radiant-heat source were 1) that the radiant heat be in a wave length range which is effectively absorbed by the impurities but not absorbed appreciably by the rock salt and 2) that it be dependable, practical, and economical. Selection of a heat source of suitable wave length range was one of the first considerations. It is well known that pure halite is highly transparent to radiant energy in wave lengths from 0.3 to 13 microns. However, the available data on infrared transmission by dolomite and anhydrite, particularly in the range below two microns, were not complete enough to serve as a reliable basis for selection of a heat source. Although it may have been possible to obtain sufficient data on infrared transmission and absorption to enable one to select the best heat source, a more direct procedure was used. This consisted simply of exposing the crude rock salt to each of several types of radiant-heat source on the small continuous separation device. The heat sources investigated, approximate source temperature used, and calculated wave length of maximum radiation are tabulated in Table I. Of the two types of tungsten-filament lamps investigated, both the short wave length photoflood lamps and the longer wave length infrared lamps were satisfactory from the standpoint of selectivity

Jan 1, 1961

By S. S. Shen, M. J. Pool, P. J. Spencer, J. R. Guadagno

VERY few thermodynamic data are available for rare-earth alloys, partly because of the difficulties of obtaining the metals in sufficiently pure form and also because of their comparative scarcity up to the present time. The rare earths form a very interesting series in the periodic table in view of the progressive filling of their 4f subvalence electron shell and would be expected to provide useful information with regard to alloying characteristics of transition elements. It has been found in recent calorimetric investigations that the heats of solution of a number of the rare-earth metals in liquid tin are highly exothermic.1-4 Exceptionally strong bonding between rare-earth and tin atoms is clearly indicated in the liquid solutions. Continued thermodynamic studies of the behavior of rare earths in solution could thus contribute a great deal to present theoretical knowledge of alloying behavior and would yield an insight into bonding of a strength hitherto unencountered in liquid alloys. The determination of the heats of solution of the pure components in liquid tin is an essential preliminary to calorimetric work of this nature. Heats of solution of praseodymium, neodymium, and samarium in liquid tin at 750" ' and the heats of mixing of Pr-Nd solid solutions' have already been determined in this laboratory. The present investigations with lanthanum continue the study of rare-earth elements. Details of the construction and operation of the liquid metal solution calorimeter used in this work have been published previously by Pool and Peluso and no further description will be given here. The lanthanum used in heat of solution determinations was supplied by Lunex who quoted the purity as 99.9 pct. Their detailed analysis of the material is given in Table I. Samples of 0.0003 to 0.0005 g-atom were used for each run and great care was taken to remove all surface oxide prior to weighing. The samples were stored at all other times in a beaker of oil to prevent their otherwise rapid oxidation. The compounds LaSn (Cudu, L1P type structure) and La&, (structure not classified) were prepared from the appropriate quantities of the two pure components by arc melting under a purified argon atmosphere and once again great precautions were taken to avoid any subsequent oxidation of the specimens. Although no structural examination of the compounds was made, the consistency of thermodynarnic measurements carried out on twenty or so individual samples of each indicates that the compounds were in fact homogeneous, single phase alloys. Samples of the compounds consisting of 0.0003 to 0.002 moles were used in the calorimetric measurements. A solvent bath consisting of approximately 80 g of 99.9 pct pure Sn was used for each series of runs and its heat capacity was determined at regular intervals throughout the experiments using pure tin or tungsten calibration samples. Measurements were made of the heat of solution of lanthanum in liquid tin at 725", 750C, and 775°K and at compositions up to 2.21, 4.76, and 5.25 at. pct La at the three temperatures respectively. In every case, the heat of solution of lanthanum remained constant at its limiting value, within experimental limits, up to the highest lanthanum concentration investigated. Since no change in was observed, it was concluded that the liquid/liquid + LaSn, phase boundary had not been

Jan 1, 1969

By Stephen H. Stow

A relationship exists between the color and the chemical composition of the pebbles from the Bone Valley Phosphate Formation. Darker-colored pebbles are relatively high in iron oxide and relatively low in calcium phosphate when compared to lighter-colored samples; the brown color of the phosphate pebbles can be correlated with their iron content and the black and gray colors with their organic carbon content.1,3 It was anticipated that the darkness of a phosphate sample might be quantitatively measured by a simple reflectivity test and that the amount of reflected light could be correlated with the chemical composition of the rock, especially with the elements which cause the color and darkness of the Florida phosphates. Twenty-one samples of production grade phosphate rock from Hillsboro County, Fla., were studied. All samples were ground to -200 mesh for reflectivity tests and analyses. All analyses are accurate to well within ±2% of the reported value, except for aluminum and organic carbon, which are accurate to only ±15% of the reported amount. Reflectivity measurements were taken with a model 610 Photovoltmeter for tri-stimulus amber, blue and green, and white lights. Experimental procedures and data are available upon request. Results Table 1 contains linear correlation coefficients (LCC) for the element-reflectance relationships for the samples. The top portion of the table refers to the LCC between the percent of light reflected off the samples and the percent of a given element (or its oxide) as analyzed. In the lower portion of the table the LCC refers to the relationship between the same reflectance data and the chemical analyses which have been corrected on the basis of 0% acid insoluble material (chiefly quartz) in the phosphate rock. Each LCC presented in Table I is computed for the 14 coarser samples exclusive of the seven concentrates since the concentrate samples appear to have anomalous reflectivities upon comparison with the coarser grades. Only the most significant LCC for each element is presented in graphical form. Discussion of Data Iron: [Fig. 1] is a graph of the relationship between iron and reflectivity (LCC = - 0.98). A good correlation with Fe2O3, is expected due to the relationship between the tan (or brown) color and the abundance of iron oxide, the principal mode of occurrence of iron in Florida phosphates.8 Note the location of the seven triangles which repre-

Jan 1, 1972

By W. A. Stickney, J. W. Town

Tests demonstrated that aqueous solutions of sodiu~n sulfide would dissolve over 95 pct of the cinnabal- in 5 pct Hg flrotation concentrates and 60 to 90 pct of the cinnabav in low-grade ol-es. Double-leach tests showed that potassiutn sulfide would dissolve 90 pct of the avsenic sulfide in Hg-As satnples and only 1 to 10 pct of the Hg; the sodium sulfide would then dissolve 80 to 90 pct of the rernaining Hg and 30 to 70 pct of the vemaining As. 1 HIS paper describes the extraction, by alkaline sulfide solutions, of mercury from cinnabar-bear ing ores and flotation concentrates. Leaching time and solution temperature had minor effects on mercury extraction when compared to the effects of reagent quantity and ore type and grade. Cinnabar, although readily soluble in sodium sulfide, was only partially soluble in potassium sulfide and essentially insoluble in ammonium sulfide. A double leach process was developed to extract selectively the arsenic from Hg-As products with K2S and to treat the residue with sodium sulfide-hydroxide solutions to extract the mercury.' Mercury was then recovered from the pregnant solutions either by precipitation with metallic aluminum or by electrodeposition. The solubility of HgS in alkaline sulfide solutions was first reported by Volhard in 1878.' The first reported commercial application was made in 1915 when Thornhill applied the method to the treatment of an amalgamation tailing.3 Mullhollen, in 1911. described the chemical reaction of HgS with Na,S and alkaline hydrate to be:4 Most investigators have reported using a 4-pct sodium sulfide and a 1-pct sodium hydroxide soluti~n.3-4 An increase in temperature or excessive dilution of the solution causes the reaction to be reversed. Church Holmes described a similar process used for the extraction of antimony from silver concentrates at the Sunshine Mining Company's mill at Kellogg. Idaho.5 More recently. Erspamer and Wells reported the results of leaching studies on a Hg-Sb ore from Alaska.6 EXPERIMENTAL PROCEDURE Leaching tests were made on samples of mercury ore from eight deposits in Oregon, Washington, and Idaho. In addition. tests were made on flotation concentrates from seven of the samples as well as a flotation concentrate from an operating mill. Since previous investigators reported reagent requirements greatly in excess of that required by the equation HgS + Na2S = HgS . Na2S, initial studies were made on chemically pure HgS. Results showed that the reaction actually required about 1 g of Na,S per g of HgS in an aqueous solution rather than 0.338 g of Na,S as indicated by the equation. To check further the matter of reagent requirements, Na2S and HgS were mixed together in equal mole ratios. Water was then added, a drop at a time, until the components reacted; X-ray diffraction analyses showed no HgS nor Na,S present. The yellow hydrate compound which formed would not redissolve until about three additional mole equivalents of Na,S were added to the solution. This indicated that a stable solution would require at least 4 moles of Na2S for each mole of HgS. While these implications may be valid for pure HgS, the possibility of entirely different results caused by the existence of other elements present in ores and concentrates was recognized. Consequently, experiments of the same order were made on samples of a flotation concentrate from a producing mill. The concentrate, containing about 18 pct

Jan 1, 1962

By Joseph E. Pogue

THE period of new oil discoveries in Illinois began in February 1937, when The Pure Oil Co. found the Clay City field the forerunner of a number of limestone pools. The importance of the area was emphasized greatly in December of the same year with the discovery of excellent sand production by The Carter Oil Co. at Louden and in June, 1938, with the discovery by The Texas Co. of the Salem field producing from two sands and one lime- stone horizon. By September, 1939, new discoveries in Illinois were estimated to have added, under prevailing production practices, some 500,000,000 barrels to the crude oil reserves of the United States and to have produced to that date some 63,000,000 barrels with daily production passing the 300,000-barrel mark by Sept. 1.

Jan 1, 1939