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By Gerald Sherman

MINE fires are always dangerous and are frequently accompanied by loss of life during the period of confusion which is apt to follow their discovery. In metal mines, fires may result from the accidental ignition of some combustible material, or they may be of spontaneous origin resulting from the rapid decomposition of sulphide. They occur frequently enough to demonstrate the need of some special equipment and a general plan for controlling them, in addition to measures for their prevention. Nearly all the large copper mines have had fires, and the methods of fighting them have been well described in recent papers, notably-by C. L. Berrien,1 R.. E. Tally,2 and others in discussion. The Copper Queen mine has many shafts, which are connected by underground workings. The deeper parts are ventilated by fans, but in the shallower portions natural ventilation through numerous connecting raises or shafts is sufficient. There have been several fires in the past few years, of which all are extinguished but one, and that is under control. These notes cover particularly the preparations that are being made-to check a new fire, partially or temporarily, and to facilitate the escape of workmen in the interval before a definite plan of fighting it can be established. A fire may occur in a shaft or station, or in some working father away. In case of fire in an upcast shaft, the gas would escape to the surface and the immediate result would probably not be dangerous to workmen, as they could escape by other shafts. In a downcast shaft, however, the gas would be distributed very rapidly throughout a large portion of the mine, and men might easily be trapped and asphyxiated. A fire in mine workings, drift, raise, or stope, would be dangerous in a degree, but the draft there is likely to be less strong than in a shaft and probably a smaller area would be invaded by the fire gases.

Jan 1, 1918

By Richard Downs, Stephen P. Burke

The oxidation of pyritic sulphur associated with coal is important for the following reasons: 1. It is the predominant cause for the formation of acid mine drainage issuing from bituminous coal seams. This drainage pollutes the streams in coal regions, making the water unfit for industrial purposes and causing the disappearance of fish and wild life. As a consequence the movement of industry and especially recreational industry, so important to the population of these areas as an alternative means of livelihood, is impeded. 2. It is instrumental in effecting a series of reactions, which bring about the degradation and ultimately result in the spontaneous combustion of bituminous coal in storage. 3. It contributes to the breaking and shattering of the coal structure and is one of the causes of roof falls and floor lifts in mines. 4. It is the primary cause of the many nuisances associated with gob piles and the refuse of cleaning plants. 5. Its proper control and application might possibly be employed in the preparation of coal to reduce the inorganic sulphur content. The research that has been conducted on the oxidation of the various forms of iron sulphides is extensive and the literature voluminous. Nevertheless, few efforts have been made to determine experimentally and with some precision the mechanism of the reaction under the conditions of its occurrence in bituminous coal mines. It has generally been accepted that the oxidation of pyritic sulphur proceeds according to the reaction 2FeS2 + 2H2O + 7O2 ? 2FeSO4 + 2H2SO4 [1] The ferrous sulphate in the drainage streams is further oxidized by air to produce ferric hydroxide and more acid. Yet the literature fails to disclose any pertinent investigation on the mechanism of this reaction. The present contribution is concerned primarily with this aspect of the problem and while the work to be reported is not yet complete, it has yielded results from which interesting and significant conclusions may be drawn.

Jan 1, 1938

By Richard Downs, Stephen P. Burke

The oxidation of pyritic sulphur associated with coal is important for the following reasons: 1. It is the predominant cause for the formation of acid mine drainage issuing from bituminous coal seams. This drainage pollutes the streams in coal regions, making the water unfit for industrial purposes and causing the disappearance of fish and wild life. As a consequence the movement of industry and especially recreational industry, so important to the population of these areas as an alternative means of livelihood, is impeded. 2. It is instrumental in effecting a series of reactions, which bring about the degradation and ultimately result in the spontaneous combustion of bituminous coal in storage. 3. It contributes to the breaking and shattering of the coal structure and is one of the causes of roof falls and floor lifts in mines. 4. It is the primary cause of the many nuisances associated with gob piles and the refuse of cleaning plants. 5. Its proper control and application might possibly be employed in the preparation of coal to reduce the inorganic sulphur content. The research that has been conducted on the oxidation of the various forms of iron sulphides is extensive and the literature voluminous. Nevertheless, few efforts have been made to determine experimentally and with some precision the mechanism of the reaction under the conditions of its occurrence in bituminous coal mines. It has generally been accepted that the oxidation of pyritic sulphur proceeds according to the reaction 2FeS2 + 2H2O + 7O2 ? 2FeSO4 + 2H2SO4 [1] The ferrous sulphate in the drainage streams is further oxidized by air to produce ferric hydroxide and more acid. Yet the literature fails to disclose any pertinent investigation on the mechanism of this reaction. The present contribution is concerned primarily with this aspect of the problem and while the work to be reported is not yet complete, it has yielded results from which interesting and significant conclusions may be drawn.

Jan 1, 1938

By H. F. A. Hergt, K. L. Sutherland, J. Rogers

In 1938 Ralston4 reviewed the many attempts to find a satisfactory collector for the separation of cassiterite from its ores and in 1944 Dean and Ambrose2 summarized some further attempts. Generally, soaps and amines are effective but nonselective collectors. We have discovered that reagents containing either the polar group — SO4- or — SO3- are excellent collectors for cassiterite and promise to be selective. The response to these reagents of cassiterite and of minerals commonly associated with it was studied, using the captive-bubble technique. Using these fundamental data, the practicability of separation and concentration of tin was investigated by flotation tests in a 2000-gram Denver cell. Experimental Methods Reagents All inorganic reagents were of Analar grade. The preparation and purification of sodium cetyl sulphate, cetyl trimethyl-ammonium bromide, laurylamine and cetylamine have been previously described.5 The other collectors were commercial surface-active agents, which were used without purification. Tests Captive-bubble Tests and Cylinder Flotation Tests have been previously de- scribed.5,10 The full lines shown in the figures relating solution composition to contact (flotation) distinguish solutions in which contact (flotation) is possible from those in which it is not. The curve is derived by successive approximation and the accuracy is ±0.2 pH units and ± 5 mg. per liter of collector or depressant. Solution compositions for which it was shown experimentally that contact or flotation is possible are marked by a circle, those for which it is not possible, by a cross, and those for which it is doubtful by a cross within a circle. It is assumed that at all other solution compositions within (without) the curve, contact or flotation is possible (not possible). After testing in the collector solution, the mineral was removed, washed, and tested by the captive-bubble method in water, or floated in a cylinder with frother only. In this way the composition of solutions in which a mineral was conditioned but in which properties of the collector solution prevented contact or flotation can be defined. The composition of solutions in which conditioning was possible lie between the full and broken curve, and the experimental results are indicated by a square. The curves show the limits of composition of solutions in which flotation in a cell would be possible but many factors other than those considered in these the determine recovery and grade of a given mineral in practice. Flotation Tests in 500 and 2000-gram Denver SUbmachines,—Melboume tap water, containing 12 parts calcium, 6 parts

Jan 1, 1947

By J. Rogers, H. F. A. Hergt, K. L. Sutherland

In 1938 Ralston4 reviewed the many attempts to find a satisfactory collector for the separation of cassiterite from its ores and in 1944 Dean and Ambrose2 summarized some further attempts. Generally, soaps and amines are effective but nonselective collectors. We have discovered that reagents containing either the polar group — SO4- or — SO3- are excellent collectors for cassiterite and promise to be selective. The response to these reagents of cassiterite and of minerals commonly associated with it was studied, using the captive-bubble technique. Using these fundamental data, the practicability of separation and concentration of tin was investigated by flotation tests in a 2000-gram Denver cell. Experimental Methods Reagents All inorganic reagents were of Analar grade. The preparation and purification of sodium cetyl sulphate, cetyl trimethyl-ammonium bromide, laurylamine and cetylamine have been previously described.5 The other collectors were commercial surface-active agents, which were used without purification. Tests Captive-bubble Tests and Cylinder Flotation Tests have been previously de- scribed.5,10 The full lines shown in the figures relating solution composition to contact (flotation) distinguish solutions in which contact (flotation) is possible from those in which it is not. The curve is derived by successive approximation and the accuracy is ±0.2 pH units and ± 5 mg. per liter of collector or depressant. Solution compositions for which it was shown experimentally that contact or flotation is possible are marked by a circle, those for which it is not possible, by a cross, and those for which it is doubtful by a cross within a circle. It is assumed that at all other solution compositions within (without) the curve, contact or flotation is possible (not possible). After testing in the collector solution, the mineral was removed, washed, and tested by the captive-bubble method in water, or floated in a cylinder with frother only. In this way the composition of solutions in which a mineral was conditioned but in which properties of the collector solution prevented contact or flotation can be defined. The composition of solutions in which conditioning was possible lie between the full and broken curve, and the experimental results are indicated by a square. The curves show the limits of composition of solutions in which flotation in a cell would be possible but many factors other than those considered in these the determine recovery and grade of a given mineral in practice. Flotation Tests in 500 and 2000-gram Denver SUbmachines,—Melboume tap water, containing 12 parts calcium, 6 parts

Jan 1, 1947

By Fred C. Bond

DISCUSSION Milton C. Shaw and Donald R. Walker (Massachusetts Institute of Technology)—For many years the two empirical rules of comminution bearing the names of Rittinger and Kick have been applied to grinding processes to account for the energy required to crush brittle materials. The basis for Rittinger's rule was the thought that the energy to grind a substance went largely into supplying the additional energy associated with the new free surface, while Kick's law was prompted by the concept of dimensional similarity and is a special statement of the concept of constancy of strain energy to cause rupture. According to Rittinger the energy to grind a given volume varies as the change in the square of the mean particle size, v - A (D2), while according to Kick this energy varies

Jan 1, 1953

By Fred C. Bond

DISCUSSION Milton C. Shaw and Donald R. Walker (Massachusetts Institute of Technology)—For many years the two empirical rules of comminution bearing the names of Rittinger and Kick have been applied to grinding processes to account for the energy required to crush brittle materials. The basis for Rittinger's rule was the thought that the energy to grind a substance went largely into supplying the additional energy associated with the new free surface, while Kick's law was prompted by the concept of dimensional similarity and is a special statement of the concept of constancy of strain energy to cause rupture. According to Rittinger the energy to grind a given volume varies as the change in the square of the mean particle size, v - A (D2), while according to Kick this energy varies

Jan 1, 1953

By A. H. Larson, R. J. McClincy

Laboratory-scale decopperizing experiments with multiple sulfur addifions were conducted at 330°C on ternary Pb-Cu alloys containing, as the third elenlent, Sn, Ag, As, Sb, Bi, Zn, and Au, common impurities in lead blast-furnace bullion. For silver and tin, an increased rate and extent of 'cofifier removal was obsert3ed. The elements As, Sb, Zn, Au, and Bi had no effect or less effect as compared to sulfur additions with no i)npurily additions. THE production of primary lead in the blast furnace yields an impure lead frequently containing such impurities as copper. antimony. arsenic. tin, gold, silver iron, oxygen. and sulfur. By cooling this lead to a temperature near its melting point. most of the iron, sulfur, and oxygen and part of the other impurities are removed in the form of a dross. With incipient solidification of the lead, the copper concentration wil have been reduced to 0.02 to 0.05 pct. depending upon the concentration of the other impurities. according to Davey.' Since copper interferes with the treatment of silver after the desilverizing process, it is desirable to decrease the copper content of the lead still fur-ther before the lead is desilvered. The decopperizing of the lead is accomplished by stirring a small quantity. approximately 0.1 pct. of elemental sulfur into the lead at a temperature near its melting point, 330" to 360°C. The copper is removed as a copper sulfide which constitutes a small fraction of a voluminous dross consisting mostly of lead sulfide and entrained metallic lead. The residual copper concentration following the decopperizing operation is frequently as low as 0.001 to 0.005 pct. Thi fact has aroused considerable interest because the equilibrium copper concentration of lead in contact with solid PbS and solid Cu2S is at least an order of magnitude greater, 0.05 pct Cu at 330C. 1, 2 Most investigators have suggested that various impurities in the lead bullion are responsible for the very low copper concentrations frequently encountered in practice. There is little agreement, however? as to which of the impurities are helpful and which are not.3"11 Also. few investigators have sought to explain the mechanisms responsible for the removal of copper to very low concentrations. Willis and Blanks9 have proposed that a nonstoichiometric copper-deficient cuprous sulfide forms in place of the supposed Cu2S. Being copper-deficient, this sulfide phase would possess a low copper activity, and the diffusion of copper dissolved in the liquid lead into this phase would be greatly facilitated. Pin and wagner2 have investigated the removal of copper from liquid lead by studying the effect of impurity-doped lead sulfide on the decopperizing of pure Pb-Cu alloys. Samples of the doped PbS were held in contact with copper-saturated lead for 1 week at 33'7°C. They reported a beneficial effect on decopperizing with bismuth and antimony and no effect with tin or silver. which is directly opposite to the results observed in practice and those reported by Davey 3 and this studv. The purpose of this paper is to describe the effects of certain additive elements on the extent to which copper can be removed fro111 liquid lead by successive additions of sulfur. The impurity elements were added individually to prepared Pb-Cu alloys. The resulting ternary alloys as well as a binary Pb-Cu alloy were then decopperized with repeated additions of sulfur. EXPERIMENTAL Materials. Granulated test lead with a purity of 99.999 pct and the additive elements Cu. Ag. Sb. Bi. Zn. Sn. and Au with purities of 99.99 pct were American Smelting and Refining Co. research-grade materials. The major impurities in the lead were 1 ppm each of iron and copper. all others being less than 1 ppm. The arsenic used was a technical-grade arsenic of 98+ pct purity. Reagent-grade flowers of sulfur were melted under argon to provide small pieces free of fines. Apparatus. The decopperizing experiments were carried out in a 25-mm-OD by 375-mm-long Pyrex tube sealed at one end. The tube was mounted vertically in a resistance-heated. hinge-type tube furnace controlled to within ±lcC. Temperature measurement was accomplished by means of a standardized chromel-alumel thermocouple sealed into the base of a silica. paddle-type stirring rod. All decopperizing experiments were carried out under an argon atmosphere. Procedure. A Pb-Cu starting alloy containing 0.05 pet Cu was prepared under carbon and poured into cold tap water to produce shot. The ternary alloys were prepared by melting together 100 g of the starting alloy and a sufficient amount of the impurity element to yield the desired concentration. The resulting alloy was then homogenized in a Pyrex tube at 450C with continuous stirring. The furnace temperature was then lowered to the operating temperature of 330°C. When thermal equilibrium had been obtained at the operating temperature, individual additions of 0.2 pct (0.2 g) of solid sulfur were added to the melt and stirred in. Stirring was continued for a period of 3 min. discontinued for 5 min. and resumed for the remaining 2 min of a 10-min cycle. This cycle was repeated for as many sulfur additions as desired. When the decopperizing experiment had been completed the lead bullion was quenched and samples of the bullion and dross phases were taken for analysis. Results. The results obtained in the decopperizing

Jan 1, 1970

By W. C. Knopf, I. M. LeBaron

In the Carlsbad area potash is dry-mined and wet-concentrated. Wet concentration involves recircu-lation of saturated brines, with resultant difficulties of brine disposal and inherent losses in recovery through solution of the KCl. A dry beneficiation process for potash minerals would offer many advantages over present methods. International Minerals & Chemical Corp. has been working on development of a dry process for some time. This article is a progress report. Potash ores available in this country are largely mixtures of halite and sylvite contaminated with slimes and often with many sulfate and magnesium minerals. Electrostatic concentration of KCl when the contaminating minerals are in low percentage has been accomplished. Continuous production of better than 60 pct K2O concentrate from high potassium sulfate ores has not been demonstrated at the present time. Laboratory Procedure: Laboratory work on electrostatic beneficiation of potash minerals has been carried out using both pure crystals and mined ore. Electrostatic separation of the two main components of potash ore, sylvite (KCl) and halite (NaCl), is accomplished by first heat-treating the ore to allow for electric charging of the particles. Preferential exchange of' electrical charges between different mineral particles through surface contact is the principal mechanism involved. After proper conditioning of the particle surfaces, the oppositely charged minerals are dropped vertically into a horizontal electric field having a field gradient from about 5000 to 15,000 v per in. The particles are free-falling in the field. A suitable power supply is used to produce the proper total voltage for the electric field. Total required voltage is, of course, a function of electrode spacing. A typical electrode arrangement with power supply is shown in Fig. 1. In general, the total voltage on any system is 75,000 to 90,000 v. Power consumption is negligible, as leakage at insulators and corona discharge are the main losses in the system. Optimum preheat temperatures for pure material components generally are not the same as for mined minerals. Fig. 2 shows clearly that maximum recovery on a 50 pct pure KCl, 50 pct pure NaCl mixture occurs at a preheat temperature of about 600°F. In this test the preheat mixture was allowed to cool to about 225°F before dropping through the electric field. The optimum temperature for sep- aration is practically never the optimum temperature of preheat. Optimum separation for mined Carlsbad potash ores occurs when a preheat temperature closer to 900°F is used, as shown by Fig. 3, which plots percent sylvite (KCl) in the concentrate and tails as a function of preheat temperatures. Here again the ore was allowed to cool to about 225°F before separation. It should be noted that efficiency of separation increases continuously with increasing preheat temperatures until incipient fusion of the ore takes place at about 950°F. The increase in optimum preheat temperature for Carlsbad ores over the pure minerals can be traced to the effect of the clay slimes intimately associated with the mined potash. If the slimes associated with the ore are removed by washing with alcohol or other non-aqueous material, the necessity of preheating to the high temperature is removed. This temperature range has been proved critical within these limits during continuous pilot plant operation. Concentrate and tailings grade and optimum recovery also are functions of the temperature of the potash mineral particles during separation in the electric field. Fig. 4 presents data indicating that feed to the rougher pass should range between 150°F and 275°F for best results. It has been found by subsequent pilot plant work that these temperatures can be maintained in actual operation.

Jan 1, 1959

By H. J. Beattie, G. C. Gould

J. B. Seabrook' recently published properties of a low-alloy Ni-A1 age-hardening steel known commercially as "Nitralloy-N". He mentioned three possible mechanisms of age hardening, viz. order-disorder in the matrix, order -disorder in a precipitate, and a precipitate that goes in and out of solution reversibly without losing coherency. The following results indicate that the last - mentioned of these mechanisms is the cause of age hardening in Nitralloy-N. The reversibility of the age hardening with a sub-critical solution temperature was demonstrated strikingly by Seabrook, whose results are reproduced in Fig. 1. Aging for 22 hr at 975°F (525°C) can reharden the fully-tempered martensite to its original as-quenched hardness of R, 45. A Nitralloy steel was obtained from Seabrook' with the following analysis:

Jan 1, 1962

By Nicholas J. Grant, Yoichi Ishida, Arthur W. Mullendore

An inert particle -marker technique was developed to provide a direct measurement of grain boundary sliding during creep in tile interior of aluminum specimens. Groin boundary sliding in the interior as measured by this technique was found to be nearly the same or slightly lower than that measured on the surface. These data disagree with those obtained by the grain-counting technique developed by W. A. Rachinger. In tests where both techniques were used, the grain-counting technique gave large and variable values of grain boundary sliding. It is shown that the grain-counting technique gines erroneozcs results because of a preferred direction of grain growth during creep. A question which has long plagued those who study grain boundary creep is whether or not surface measurements yield a true representation of the deformation in the interior of a specimen. The first attempt to measure grain boundary sliding in the interior of a creep specimen was made by Rachinger.1 His technique is based on the measurement of average grain diameters, both parallel and perpendicular to the tension axis, followed by a calculation of the grain elongation from these measurements. The difference between total elongation, Et, and the calculated grain elongation, Eg, is considered to be the result of grain boundary sliding, Egb. Surprisingly large values of Egb/ Et (90 pct) came out of creep tests of pure aluminum at temperatures above 250°C (480°F), utilizing the Rachinger technique. This value is very much larger than the values of 8 to 15 pct measured by numerous investigators on the surface of a specimen. Values obtained on the surface by Rachinger, however, were consistent with those measured by the displacement of reference scratches. Chaudhuri and Grant2 repeated the Rachinger method for A1-10 pct Zn, and found the same large values of Egb/Et inside the specimen. However, from purely geometrical considerations, grain boundaries cannot slide without grain deformation; therefore, they questioned the validity of Rachinger's grain-counting method. Several authors2,3 have suggested that grain boundary migration during high-temperature deforma- tion may tend to restore the equiaxed grain shape in order to minimize interfacial energy. Rachinger4 tested this hypothesis and found that it did not hold. Grains elongated by rapid extension retained their elongated shape during a subsequent slow creep test. It appears that some other factor was responsible for the abnormally high values of .Egb/El. McLean and Gifkins5 made a survey of the effect of grain size on the ratio Egb/ Et, and proposed that the high values were the consequence of a small grain size. They failed to explain the low value of Egb/Et which Rachinger observed on the surface of the small grain size specimens, and suggested that the surface effect on the value of Egb/Et does not always occur. Many of the disagreements arising from Rachinger' s work stem from the fact that there was only one method of estimating grain boundary sliding in the interior of a specimen, and that method was of an indirect nature. If a marker line of some sort could be introduced inside of the specimen, one could make measurements just as clearly as on the surface. In the present investigation, the technique utilizes a layer of finely dispersed oxide particles inside the specimen introduced by hot press-bonding of two pieces of aluminum. Rachinger attempted a similar technique but with the oxide distribution he obtained there was so much interference with grain boundary motion that quantitative measurements were not attempted. MATERIALS AND EXPERIMENTAL PROCEDURE A) Materials Preparation. Two pure aluminum rods (impurity content in percent: Si, 0.002; Cu, 0.004; Fe, 0.002; Mg, 0.000: Zn, 0.000; V. 0.001) 1 in. diameter by 2 in. high were jacketed end to end in a 4-in.-diam cylinder of commercial-purity aluminum, 4 in, high, which had a l-in.-diam hole. The l-in.-diam mating surfaces of the two pure aluminum rods had been electropolished initially. This interface provided the oxide-film internal -marker plane after the hot press-bonding process, see Fig. 2. The composite was annealed for 1 hr at 900 F and hot upset more than 50 pct reduction in height in one step. The resulting slab was then rolled to 1/2 in. thickness; the pure aluminum test material was cut out and annealed at 900°F for 1 hr and cold cross-rolled to 1/8 in. thickness. The sheet was machined into specimens with a 1/8-in.-square cross section and a 1/2-in. gage length. The following heat treatment was given the specimens: Specimen 5A series: annealed at 1000°F for 5 min for grain-size control and then cooled to 700°F and held for 15 hr for stabilization.

Jan 1, 1965

By R. C. Buehl, M. B. Royer

High manganese slags of low phosphorus and iron content are produced by air oxidation of high phosphorus spiegeleisen in a basic-lined converter. Control of phosphorus and iron within specification limits for ferromanganese ore feed is obtained by a unique cyclic operating procedure. Various types of slags, or synthetic manganese ore, can be made. AN affiliated paper' describes the production of high phosphorus spiegeleisen containing 14 to 23 pct Mn and up to 4 pct P in an experimental blast furnace from open-hearth slag or manganiferous iron ore. This phosphorus content is too high for use of the spiegel in normal steel-production operations. Consequently, a preferential separation must be effected whereby the manganese is concentrated in a product that is usable in industrial operations. The preferential separation methods being investigated for this phase of the work are confined to pyrometal-lurgical processes for ready incorporation into steel-plant operations. It is fortunate that manganese is more actively oxidizable than iron or phosphorus, and therein lies a preferential separation method for isolating manganese from these two elements. Silicon is more strongly oxidizable than manganese, and its oxidation precedes, or occurs simultaneously, with manganese. Therefore, manganese and silicon are separable into a high manganese oxide slag phase while the phosphorus and iron remain in the molten metallic state. Complete separation of these elements represents an ideal condition which is not generally attainable in actual operations. This pyro-metallurgical method for separating manganese from phosphorus and iron by preferential oxidation was investigated by blowing 500 lb of metal in a basic-lined vessel to obtain preliminary information on the effectiveness of the procedure. Certain conditions of attainment in the separation process are desirable in any method for removing manganese from high phosphorus spiegeleisen, such as: 1—High recovery of manganese in the oxidation product; 2—the product to be of equal or better quality than ferromanganese-grade ore—48 pct Mn minimum, less than 10 pct SiO2, and Mn-P ratio of 300:1; and 3—attainment of the enumerated objectives in a product amenable to industrial handling, such as a slag of high manganese content that will flow from the processing vessel and of sufficient fluidity for ready separation of metallic granules that may have become mixed with the manganese slag phase. Previous Work Recovery of manganese from low grade ores and industrial byproduct slags by smelting to spiegeleisen, with subsequent oxidation to synthetic ferro-manganese-feed ores, has been a subject for periodic investigation during the past four or five decades in the United States and Germany. Joseph' and associates of the Bureau of Mines North Central Experiment Station, Minneapolis, Minn., smelted manganiferous iron ores of the Cuyuna range, Minn., to spiegeleisen and subsequently tried air and ore oxidation procedures for concentrating the manganese into synthetic ferrograde manganese ore. Joseph's results on the operation of a side-blown basic-lined converter for air oxidation of manganese from molten spiegeleisen were so discouraging that the procedure was abandoned after 11 blowing tests. Much slag was thrown from the converter vessel; phosphorus content was four to five times the maximum allowable for ferro ore; and a manganese-iron ratio of only 2 or 3:1 was obtained in the slag instead of the required 8:1. Oxidation of manganese from spiegeleisen with about 0.5 pct P, by adding iron ore to molten spiegel in an experimental open-hearth furnace, proved moderately successful in Joseph's experiments. The main drawback to the operation was the necessity for over-oreing for effective oxidation of the manganese with consequent too high iron oxide residual in the slag. Furthermore, the phosphorus content was proportional to the iron concentration and therefore much too high, so that several hours of reduction by a carbonaceous material were required to adjust the iron and phosphorus to the desired content of high grade ferro feed. A troublesome feature of the slag product of the spiegel oxidation process was its lack of fluidity if the silica content was less than 10 pct, the concentration considered desirable for the synthetic ore. Approximately twice the desired concentration of silica was required to impart enough fluidity for transfer from the vessel by pouring. The acute shortage of manganese in Germany

Jan 1, 1953

By C. D. Ramsden

DURING the war years of 1941 to 1945, maintenance of mine locomotives and other mine equipment took the form of rebuilding rather than of renewing. Pur¬chase of new parts became increasingly difficult and mine operators found it necessary to maintain equipment by sal¬vaging old parts. Mine-locomotive wheels, although not representing a high cost per pound of material, contain a con¬siderable amount of steel and without rebuilding would have to be scrapped long before even a moderate percentage of the steel in the wheel is actually consumed. Many mine operators have been re¬building locomotive wheels for several years with automatic open arc-welding equipment. In 1927, the Pittsburgh Coal Co. started such a program, using a single Westinghouse automatic welding head. Since that time, many refinements have been added, until in 1944 the setup was operating very efficiently. A desire to increase wheel life and to hold down rising costs led the company in 1944 to investigate the possibilities of the under¬flux arc-welding process for wheel rebuild¬ing, although at that time, little or no work had been done on multiple pass or "build-up" welding with the underflux arc. Work with this process had been confined largely to welding together flat plates in a "down hand" position. Possibilities for savings were intensified by the fact that the company has 300 locomotives in operation, using actively or as spares 1200 sets of wheels. During 1942 and 1943, about 500 sets of wheels were sent each year to the company's maintenance shop at Library, Pa., for rebuilding. Based on having 800 sets in service (many of the company's heavy locomotives have three- sets of trucks), the average life of a set of wheels was approximately 19 months. Anything that could be done to increase wheel life promised sizable annual savings. OPEN ARC WELDING With the open arc process, a set of wheels could be rebuilt for about 40 pct of the cost of new wheels. Underflux arc welding gave promise of greatly increased welding speeds, more homogeneous, dense deposits and the chance for use of a wider range in welding-wire compositions. Against this was ranged the higher cost of materials and the cost of developing the process from the theory to a practical reality. Since 1927, when a single open arc head was placed in operation, fixed welding procedures had been gradually developed, a positioner for handling a set of wheels was built, and a second head was placed in operation in order to weld two wheels simultaneously. The fixture, or posi¬tioner, which was built to position and turn the wheels under the two welding heads, included a motor drive, controlled by a rheostat, which rotated the wheels through a double set of speed reducers and a removable jaw coupling attached to one end of the axle of the wheel truck.

Jan 1, 1946

By Otto H. Henry, Edward L. Badwick

Up to the present time, according to Hansen1 and Haughton,2 the constitution of the bismuth-indium system has not yet been published. The generally accepted value for the melting point of indium, as listed in most scientific journals and textbooks, is 155°C. Recently, however, H. M. Davis3 has determined a new melting point for indium of 257.3°C ± 0.1°C. Since bismuth has been used in preponderance as an alloying element in the majority of low-melting alloys, and indium has been used with considerable success in lowering the melting points of low-melting alloys,4,5 an interest has been developed in the application of these two metals, making a complete knowledge of the constitution of its alloys desirable. In this investigation, the melting point of indium reported by H. M. Davis has been found to be substantially correct. A value of 157.2°C ± 0.5°C was determined. The two eutectic compositions have been placed at 34 and 50 pct bismuth with their respective eutectic temperatures as 72.4'C ± 0.5°C and 90°C ± 0.5°C. The bismuth-indium phase diagram has been determined by thermal and microscopic analysis and is shown in Fig I. THERMAL ANALYSIS Granulated bismuth and indium shot, having a guaranteed purity of 99.99 and 99.95 pct, respectively, were melted in a carbon crucible by means of a laboratory heater. To minimize oxidation of the alloys, a heavy mineral oil (Nujol) having a high flash point was used. A platinum-platinum rhodium thermocouple 0f fine wire encased in a thin-walled Pyrex glass tube, about 1/8 in. in diameter, was connected to a type K potentiometer for temperature measurements. The thermo-couple was calibrated against the melting points of tin, bismuth, and the silver-copper eutectic and the boiling point of water. Since the temperature readings could be reproduced to 0.2°C, and the accuracy of the temperature standards in this range is slightly better than 0.5°C, it seems reasonable to assume that the temperature readings are reliable to approximately 0.5°C. The cooling rate could be controlled by varying the amount of insulation surrounding the crucible. Three or four layers of asbestos paper, 1/16r, in. thick, sufficed to give a cooling rate of from 1° to 2°C per minute. Undercooling was slightly noticeable. This condition was minimized by vigorous stirring, since the advantage of using slower rates was too small to justify the greater expenditure

Jan 1, 1947

By J. S. Johnson, W. C. McCulloch

All phases of plant operation must be synchronized so that the A number of interruptions in the flow of material may be reduced to a minimum. In the majority of plants operating under a labor agreement, time lost during the operating shift cannot be made up with overtime. With the work week restricted to 35 hr., the ratio of productive time to total time justifies extra effort in keeping the plant operating at capacity during the productive shift. This does not apply when central plants with adequate storage are operated by multiple shifts. In such installations the problems become similar to those of any factory or mill where added units, parallel circuits, or excess capacity serve to maintain the average production during intervals of maintenance shut- down. These plants, by special agreement, may run continuously, barring mechanical interruptions. The control of plant operation is in those procedures that are practiced to obtain from the run-of-mine product a high recovery of clean coal products to meet adopted trade standards and consumer specifications at the lowest possible cost. Correlation of all factors for continuous operation requires more than mechanical perfection in the plant. The run-of-mine coal must be delivered to the preparation plant without delays and the clean coal and refuse products taken away. Power supply and water requirements must be adequate. Signal systems must be established for emergency stops and to facilitate resumption of coal flow.

Jan 1, 1943

The annual report f the Secretary of the Navy for the fiscal year, ending June 30, 1917, but including operations and recommendations up to Dec. 1 of that year, contains the following remark about the activities of the Naval Consulting Board. The members of the Institute who are members f this Board are William L. Saunders, Benjamin B. Thayer, Thomas A. Edison, Thomas Robins, Lawrence Addicks, Joseph W. Richards, W. R. Whitney, assisted by 47 other members f the Institute, in the capacity of associate members of the Naval Consulting Board. During the year, the work of this board, organized and approved by Congress in 1915, has increased very materially in importance and volume, its meetings have been frequent and the work of its individual members has been such in some cases as to occupy almost their entire time in the service of the Government. Some time before the active entry of this country into war the board called a special meeting to which were invited some 50 of the leading scientists and industrial managers, whose special study fitted them to advise on the methods of meeting the

Jan 2, 1918

By P. E. Hopkins

Introduction and History IN speaking of the economic geology of southeastern Ontario, W. G. Miller and C. W. Knight' say that "there occurs in southeastern Ontario a variety of minerals and rocks of economic value, probably as great as in any district of like size on the North American continent. Some of these deposits, including marble and trap, are inexhaustible. Others, including talc and iron pyrites, have proved to be of considerable economic importance. From time to time, during the last 50 years, the following minerals and rocks have been mined or quarried with varying success: gold, iron pyrites, zinc blende, copper pyrites, galena, mispickel, magnetite, hematite, talc, actinolite, mica, marble, ophicalcite, feldspar; fluorite, apatite, corundum, graphite and sodalite. All of the economic materials, with the exception of fluorite, appear to be of pre-Cambrian age. The fluorite veins penetrate the Ordovician, Black River, limestone." Accompanying that report was an article by the writer on the Queensboro Pyrite Area which includes one of the two working pyrite properties in southeastern Ontario. In the present paper will be given a brief description' of all the known pyrite deposits in the area which may at some time possess an economic value, with fuller descriptions of the two working mines-The Canadian Sulphur Ore Co.'s mine near Queensboro, and the Nichols Chemical Co.'s property at Sulphide. The earliest mining of iron pyrites in Ontario was done in 1868 on the Billings property near Brockville. The mines were closed down in 1879 under the assumption that they were exhausted. Many other

Jan 8, 1916

By G. W. P. Rengstorff, R. B. Fischer

A study was made of the effect of impurities on the bend ductility of cast molybdenum. High purity molybdenum was repared by re-melting under high vacuum. The ductility of "transverse-grain" specimens indicated that intergranular brittleness decreased with an increase in purity of the metal. THE production of as-cast molybdenum with high ductility at room temperature would make working of the metal easier and increase the usefulness of the as-cast metal. Furthermore, a .knowledge of the factors that influence the ductility of as-cast molybdenum might be applied to the problem of making fusion weldments having satisfactory room-temperature ductility. Individual grains of commercial as-cast molybdenum are slightly ductile at room temperature. However, the presence of carbides at the grain boundaries may cause the metal to fracture in a brittle manner. This behavior has been discussed.' The amount and distribution of the impurities in as-cast molybdenum would be expected, from previous research on the metal, to have considerable influence on its ductility. Some effects of oxygen and carbon on the mechanical behavior of the metal were discussed by Parke and Ham,2 Woodside,3 and Zappfe. Landgraf, and Worden.4 Parke and Ham2 found that, if a sufficient number of molybdenum oxide particles were present at the grain boundaries, the intergranular cohesion was reduced to the extent that arc-melted molybdenum would rupture intergranularly during hot forging. They also found that if the oxygen content was less than 0.003 pct, the cast molybdenum was hot forgeable. It is suspected, however, that forgeable molybdenum may contain a small amount of oxide that could reduce the ductility of the as-cast metal at room temperature. Also, there is the possibility of the unfavorable distribution of other impurities having a detrimental effect on the ductility of cast molybdenum. The results of a study of the influence of purity on the room-temperature ductility of cast molybdenum are presented here. Vacuum Arc-Melting to Produce High Purity Mo One way of determining the effect of impurities on the ductility of molybdenum is to prepare and test metal as free of all impurities as possible. Accordingly, the present investigation was undertaken with this in view. Arc melting in high vacuum (less than 0.1 micron of mercury) was chosen as the method for reducing the impurities below the concentrations occurring in commercial molybdenum. Commercial arc melting of molybdenum is carried out at about 50 microns pressure. During the arc melting, carbon and oxygen in the metal react to form carbon monoxide. By virtue of the vacuum process, the partial pressure of carbon monoxide is maintained at such a low value that the reaction greatly lowers oxygen and carbon contents. An excess of at least 0.03 pct C is used in commercial arc melting to insure that the oxygen content will be below 0.003 pct. The excess carbon remaining after the deoxidation step forms carbides on freezing. Some of the carbides precipitate at the grain boundaries and, as will be shown, have an unfavorable effect on the room-temperature ductility. In the present work, only the small amount of carbon (0.003 pct) contained in the commercial powder-metallurgy molybdenum used for melting stock was present. Partial removal of oxygen from the melt probably occurred as the result of the volatilization of molybdenum oxides, or by distillation of oxygen at the high melting tem-

Jan 1, 1953

By B. F. Bush

THE coal-fields of Missouri, situated hi the northern and western portion of the State, are distributed, in whole or in part, over 57 counties, embracing an area estimated by Mr. Broad-head to be practically 23,000 sq. miles, including small detached beds and outliers in St. Louis, St. Charles, Lincoln and Warren counties in the eastern part of the State. This area includes about 8,000 sq. miles of Upper or Barren Measures, 2,000 sq. miles of exposed Middle, and 12,420 sq. miles of exposed Lower Measures. In a general way, the Missouri coal-field is bounded on the east by an imaginary line entering the northern part of the State just east of the 92c1 Meridian and passing across the State in a southwesterly direction to a point on the Kansas-Missouri line a little below the southern boundary of Barton county. An important part of the field overlaps this line and extends down the northern side of the Missouri river, under¬lying the greater part of Randolph, Howard, Boone and Audrain counties. In all, 34 counties produced coal during the year 1903. Fig. 1 shows the position of the coal-fields in Missouri. GEOLOGICAL FEATURES. The Coal Measure formations, all of the Carboniferous age, rest unconformably upon the rocks of the sub-Carboniferous or Mississippian series, and, in some cases, upon the rocks of the Devonian and Silurian ages. From their outcrop on the eastern and southeastern borders of the field, the Measures dip west to northwest at the rate of 10 ft. per mile, reaching a maximum thickness estimated by

Jan 1, 1905

By J. W. Furness

Before attempting any discussion of the importance of tariff enactments as applied to exhaustible resources, certain basic factors should be emphasized. First, it must be remembered that minerals differ from organic substances in that they are irreplaceable, and when once taken from the earth's crust enter into uses which ultimately dissipate them, so . that they cannot be recovered for future use. A second factor entering into the problem is the unequal distribution of the mineral deposits of the world. No industrial nation has within its political confines a sufficient quantity of all the minerals necessary for the maintenance and advancement of present civilization; hence international exchange of these commodities is inevitable. As minerals are essential to modern industry, the supply available to any given nation constitutes a factor which limits its ambition to advance industrially. A nation which has this ambition endeavors to secure not only a supply of minerals sufficient for its current needs, but also sources which will assure it expansion. Should mineral raw materials be withheld from such a nation by enactment, it seems inevitable that every effort will be made to obtain them, and that, should the nation deem itself sufficiently powerful, its ultimate recourse will be force. Thus the importance of tariff enactments, not only from the standpoint of industrial prosperity but from the standpoint of international peace, can be appreciated.

Jan 1, 1932