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By W. J. Keep

Manganese is a nearly white metal, having about the same appearance when fractured as white cast-iron. Its specific gravity is about 8, while that of white cast-iron, reasonably free from impurities, is but a little above 7.5. As produced commercially, it is combined with iron, and with small percentages of silicon, phosphorus, and sulphur. It is generally produced in the blast-furnace. If the silicon is under 0.50 per cent., the product will be white. If the manganese is under 40 per cent., with the remainder mostly iron, and silicon not high enough to make the product gray, the alloy is called spiegeleisen, and the fracture will show flat reflecting surfaces, from which it takes its name. With manganese above 50 per cent., the iron alloy is called ferromanganese. As manganese increases beyond 50 per cent., the mass cracks in cooling, and when it approaches 98 per cent. the mass crumbles or falls in small pieces. When the percentage of manganese is high, the planes of the fracture are not so bright as when it is lower. The ferro-manganese which we have used in the experiments reported in this paper, contains Mn, 81.62; Si, 0.256 ; and combined carbon, 6.153 per cent. We have not been able to make test-bars from this, because the alloy cracks in small pieces in cooling. Manganese combines with iron in almost any proportion, but if an iron containing manganese is remelted, more or less of the manganese will escape by volatilization, and by oxidation with other elements present in the iron. If sulphur be present, some of the manganese will be likely to unite with it and escape, thus reducing the amount of both elements in the casting. Hence, in remelting irons, some of the sulphur which has entered from the fuel with which the iron is melted will very likely unite with some of the manganese in irons which make up the mixture and escape. Remelting, therefore,' reduces the percentage of manganese in cast-iron.

Jan 1, 1892

By W. H. Howard

As the result of a series of analyses and volume determinations of gases discharged from the converters at the Garfield Smelting Co.'s smelter at Garfield, Utah, it was found that a considerable amount of lead fume was escaping from the stacks of the converter plant, and that if this lead could be recovered at a reasonable expense an attempt should be made to collect it. Two methods suggested themselves: (1) filtration through bags; (2) electrical precipitation of the condensed fume by a Cottrell plant. Filtration through Bags Tests were made in filtering converter gases through the ordinary woolen bags of lead-smelter bag houses to determine how long such material could withstand the corrosive action of the gases. While blowing on white metal only, the fabric was usually destroyed in 10 to 14 hr. In tests conducted in filtering gases when blowing on matte or slagging stage exclusively and when relatively considerable lead was being volatilizcd, bags in one instance lasted 60.5 hr. It therefore was evident that the Garfield converter gases did not contain in themselves sufficient neutralizing elements for the purpose of direct bag filtration, even during the slagging stage. Such results were more or less anticipated, but it was thus proved that under all conditions of blowing Garfield mattes a very considerable amount of neutralizing agent, zinc oxide or lime, or both combined, as in the Sprague process, would be required to render bag filtration possible. Again, owing to the constantly changing quantities of SO3 from a given battery of converters, depending on the number blowing at a given time and on the stages of the blow, it would be practically impossible to feed automatically the required amount of neutralizer. In order to safeguard the bags sufficient neutralizer would have to be continuously added to meet the maximum SO3 output, thus wasting neutralizers and giving an undesirably diluted final product. Such a condition is entirely different

Jan 1, 1915

By Oliver C. Ralston

THE two chemical compounds, natural sodium carbonate and sodium sulphate, in their anhydrous condition are dealt in as "soda ash and "salt cake”-names from an earlier day that are not as precise as the present accepted terminology. Once a large part of the sodium carbonate recovered was in the ash of seaweed, so it was logically called soda ash, and the corresponding potassium carbonate, once mainly recovered from wood ashes, was called potash. The term "salt cake" was applied to the sodium sulphate resulting from interaction of sodium chloride with sulphuric acid or sodium bisulphate, to form hydrogen chloride gas and an anhydrous cake of sodium sulphate. It was called "salt cake" and also "glassmaker's salt cake" because much of it was used in the production of glass. The two salts occur in nature in a variety of forms, but whereas an- hydrous sodium sulphate (thenardite) is fairly common, the anhydrous normal sodium carbonate, Na2CO3, is almost never found in nature. The monohydrate (thermonatrite), Na2CO3.2H2O, is seen occasionally, but it is never a major constituent in natural "soda" deposits. The high solubility of normal sodium carbonate is one reason for this. While both compounds occur widely in brines, the solid salts found in nature are usually trona (Na2CO3.NaHC03.2H2O) and thenardite (Na2SO4). The natural brines containing workable quantities of the two sodium salts are largely found in the arid and semiarid portions of the western United States and other countries. The classical discussion of the resources of these brines are those of Clarke,5 Chatard.4 Packard,19 Gale,11 and Wells.28 From North Dakota to California and from Arizona to the Canadian provinces of Alberta and Manitoba, saline lakes, dried-up lakes, playas, and similar land-locked depressions contain accumulations of saline matter. Sodium chloride is the predominant salt but some of the lakes contain predominantly either sodium carbonate or sodium sulphate. Often sodium carbonate

Jan 1, 1949

By H. J. Oberson, J. H. Brown

Comminution devices such as rod or ball mills are characteristically nonselective in their operation in that all material fed to a mill is ground to some extent. This is unfortunate when the object of comminution is liberation, for not only is energy expended in breaking pure gangue particles wasted, but, since gangue is frequently the predominant phase, the opportunity for such loss is great. Ideally, in this application a comminuting device should develop fracture only in locked particles. In the present investigation an attempt was made to use the large density difference which may exist between minerals to prevent comminution of one while grinding the other. EXPERIMENTAL MATERIALS AND EQUIPMENT The two minerals used in this work were -8-mesh galena and -8+ 20-mesh quartz. All tests were made on 2000-g charges containing 50 pct by weight of each mineral. Heavy media suspensions consisted of -100-mesh ferrosilicon in water. For grinding, a laboratory mill 10 in. diam by 12 in. long was filled to about 30 pct of its volume with either a ball or a rod charge. The balls ranged in size from 3/8 to 1 1/8 in. in diam; the rods ranged from 5/8 to 1 1/4 in. diam by 11 in. long. EXPERIMENTAL PROCEDURE In all tests employing heavy media, the ground pulps were treated as follows: 1) After recovering the ground charge from the mill but before drying, the pulp was wet screened on 200 mesh. 2) After drying the + 200-mesh product from step 1, the ferrosilicon was removed magnetically. 3) The nonmagnetic quartz and galena product from step 2 was sized on 10-mesh through 200-mesh screens. 4) The mineralogical composition of each screen fraction was determined by measuring its specific gravity. The specific gravity was related to the weight of galena in the fraction through the relationship

Jan 1, 1961

By C. S. Barrett, C. T. Haller

Twinning in magnesium is known to occur profusely under certain conditions, and when it occurs in polycrystalline materials it brings about a partial or even a complete change in the preferred orientation. The amount of research on this subject, however, has not been in proportion to its importance as a mechanism of deformation and reorientation. Much of the research has been limited to twinning in single crystals. The present study1 concerns polycrystalline sheet, and is an investigation of the twinning produced by plastic deformation as a function of the following variables: 1. Direction and type of deformation. 2. Amount of deformation. 3. Speed of deformation. 4. Temperature of deformation 5. Composition and temper of the sheet. 6. Twins previously produced. 7. Twins previously produced, then annealed. By proper control of these variables it is shown that, on a laboratory scale, samples can be produced that are free from preferred orientations. The materials used were samples of commercial magnesium-alloy sheets with nominal compositions given in Table I. X-ray diffraction was used in most of this work to detect the twins and to estimate the amount of twinning. Since twinning on (102) planes causes a re-orientation of the crystal lattice by 86", it is very easy to study with X-ray diffraction patterns, particularly if one notices the diffraction ring from the basal plane— the 002 reflection. By properly directing the X-ray beam, it is possible to produce 002 reflections from twinned material at, say, the 6 and 12 o'clock positions on the 002 ring, with the untwinned material reflecting to 3 and 9 o'clock positions; and by judging the relative intensity of the arcs on the diffraction ring at these positions, it is possible to judge the amount of twinned material relative to the untwinned. This X-ray method of estimating the amount of twinning is generally superior to metallo-graphic observation, for it was found

Jan 1, 1947

By C. S. Barrett, C. T. Haller

Twinning in magnesium is known to occur profusely under certain conditions, and when it occurs in polycrystalline materials it brings about a partial or even a complete change in the preferred orientation. The amount of research on this subject, however, has not been in proportion to its importance as a mechanism of deformation and reorientation. Much of the research has been limited to twinning in single crystals. The present study1 concerns polycrystalline sheet, and is an investigation of the twinning produced by plastic deformation as a function of the following variables: 1. Direction and type of deformation. 2. Amount of deformation. 3. Speed of deformation. 4. Temperature of deformation 5. Composition and temper of the sheet. 6. Twins previously produced. 7. Twins previously produced, then annealed. By proper control of these variables it is shown that, on a laboratory scale, samples can be produced that are free from preferred orientations. The materials used were samples of commercial magnesium-alloy sheets with nominal compositions given in Table I. X-ray diffraction was used in most of this work to detect the twins and to estimate the amount of twinning. Since twinning on (102) planes causes a re-orientation of the crystal lattice by 86", it is very easy to study with X-ray diffraction patterns, particularly if one notices the diffraction ring from the basal plane— the 002 reflection. By properly directing the X-ray beam, it is possible to produce 002 reflections from twinned material at, say, the 6 and 12 o'clock positions on the 002 ring, with the untwinned material reflecting to 3 and 9 o'clock positions; and by judging the relative intensity of the arcs on the diffraction ring at these positions, it is possible to judge the amount of twinned material relative to the untwinned. This X-ray method of estimating the amount of twinning is generally superior to metallo-graphic observation, for it was found

Jan 1, 1947

By R. D. Pehlke, J. F. Elliott

The equilibrium pressure of nitrogen gas over pure silicon metal and silicon nitride has been measured in the temperature range 1400° to 1700°C. From the experimental data, the standard free energies and enthalpies of formation of Si3N4(a) have been calculated as functions of temperature over the above temperature range. Utilizing data in the literature and the results of this investigation, the specific heat, molar enthalpy, molar entropy, standard enthalpy of fornation, and standard free energy of fornation are estimated for the temperature range 298° to 1400°C. THE high-temperature thermodynamic properties of many metallic nitrides are not well established at present. This is a serious deficiency because nitrogen and nitride-forming metals are important alloying additions to steels, and knowledge of the interaction between nitrogen and these elements is essential to a better understanding of the behavior of steels. Metallic nitrides also are used as refractory materials. Consequently, the investigation reported here was undertaken to provide more complete information on the standard entropy and free energy of formation of silicon nitride. The experimental technique provided a direct measurement of the equilibrium pressure of nitrogen gas in the range of 1413 to 1700°C for the reaction: 3 Si(1) + 2N2 (g) = Si3N4 (a) [ 1 ] and below 1413oC for the reaction: 3Si(c) + 2N2(g) = Si3N4(a) [2] APPARATUS The experimental system consisted of a crucible assembly, a reaction chamber, and a gas supply and vacuum system. The Crucible Assembly which contained the equilibration specimen and molybdenum radiation shields with a susceptor for induction heating is shown in Fig. 1. Two covered, high-purity alumina crucibles, one nested within the other, contained the susceptor, shields and specimen. Central openings in the upper shield and crucible covers permitted on optical pyrometer to be sighted into the central cavity of the specimen. The specimen itself was a 5/8 OD by 1-in.

Jan 1, 1960

By Michael R. Cartwright

The Gold Acres outcrop was located by Lee Lakin in 1922, according to Mr. H. W. Treweek. There was no activity, however, until 1935, when prospect drifts and pits penetrated the deposit. It was then drilled by the Willow Creek Mining Company. Beginning in the late 1930s and continuing to the late 1940s more than 150 churn drill holes were drilled in the Gold Acres area. From this drilling came what is now known as the London Extension pit. Only oxidized ore amenable to cyanide vat leaching was mined during this time. Our records indicated that this period of mining removed some 1.4 Mt (1.5 million st) of ore at an average grade of 0.006 kg/ t (0.2 oz per st). Union Pacific and Newmont Mining Company worked in the area prior to acquisition by the Cortez Joint Venture. Union Pacific drilled a number of core holes in the area. The unoxidized nature of the ore, small tonnage, and low grade undoubtedly prevented mining at this time. From these indications of mineralization shown by previous companies, the Cortez Joint Venture developed and mined 2 Mt (2.4 million st) at 0.005 kg/t (0.15 oz per st) in the Gold Acres pit. This pit is located immediately southwest of the older London Extension pit. Cortez Gold Mines has continued to drill to the west and south of these pits along the downdip continuation of the former ore bodies. The downdip extension is unoxidized and more metallurgically difficult than the normal Carlin type ores, but, more than 2.4 Mt (2.7 million st) have been outlined. It depends now upon the gold price and milling costs to turn this mineralized area into an ore body. A possible lesson to be learned from this deposit is that continued mining success depends upon an ongoing adequately funded exploration-development-metallurgical research program. These three disciplines must always be present to locate the mineralization and to turn this mineralization into ore. [ ]

Jan 1, 1985

By Arthur Phillips

The progress made in recent years in the art of dispersion-hardening has naturally led to an intensive study of alloy systems capable of yielding supersaturated solid solutions at ordinary temperatures. In general, the investigations have been confined to rather precise determinations of solubility limits and to experimental approaches designed to furnish more complete information regarding the rate and nature of precipitation under appropriate thermal conditions. Efforts are being made to focus more sharply on the atomic changes that are a direct consequence of the instability of the supersaturated solution. Here precision X-ray methods have come to be of considerable importance because lattice parameter measurements indicate very precisely the concentration of solute atoms in the solvent lattice. In spite of the general acceptance of parameter values as a criterion for the placing of solubility limits, it may be suggested that no thorough appraisal has been made of the several factors that conceivably may affect such measurements. More specifically, little attention has been devoted to the significance of surface energy conditions, specimen size and method of preparation, with particular regard to the genesis of the grains and the surface concentration changes as a consequence of heat treatment and etching. Stenzel and Weertsl recently published the results of a detailed study of precipitation reactions in aluminum-rich copper alloys. They found that, although copper lowers the lattice constant of aluminum, a 5 per cent copper alloy quenched from 150" and 225' C. had parameter values unquestionably higher than that of pure aluminum. Their results seem to substantiate, at least qualitatively, the earlier work of von Göler and Sachs,2 who indicated a parameter value, for the same alloy annealed for 24 hr. at 200° C., some 0.0008Å. larger than that of pure aluminum. Stenzel and Weerts attributed the abnormal lattice dimension to the effect of strain induced by the precipitation of CuA12. Weerts3 later

Jan 1, 1934

By Richard S. McCaffery

A Research on iron blast-furnace slags has been in progress since 1922 in the metallurgical department of the University of Wisconsin, supplemented by observations, tests and critical examination of operating records at 10 furnaces. From the data thus collected information has been put into forms suitable for use by furnace operators and is given in this report and the three accompanying papers.' In the viscosity results reported, it is estimated that temperatures were measured within 5" C. and that the viscosity determination of the particular slag investigated is accurate within ± 4 per cent. In the viscosity diagrams in the various papers, the equal viscosity lines were drawn in such a manner that in all cases they are located correctly with respect to the actual experimental results, but when determining a viscosity for a particular composition from the diagrams, the viscosity may. be slightly less accurate than the experimentally determined viscosity for the same slag, for the reason that the iso-vis lines interpolated between the experimental results may be shifted slightly without violation of the experimental data. The viscosity diagrams represent our judgment of their position under the above conditions, and we believe the relative accuracy of viscosity data taken from the diagrams is more than sufficient for any use to which the data may be put. Hundreds of samples of slags synthetically produced in the laboratory and of actual furnace slags have been studied in various ways. The ranges of compositions of slags used for viscosity measurements are: silica 25 to 55 per cent.; alumina 0 to 35 per cent.; lime 25 to 55 per cent.; magnesia 0 to 20 per cent. These measurements were made at temperatures of 1250°, 1300°, 1350°, 1400°, 1450°, 1500°, 1550°, 1600" and 1650" Centigrade. Descriptions of apparatus, methods used and statements of the problems investigated have been published in the papers listed in the bibliography beginning on page 120.

Jan 1, 1932

By M. W. Fabjanczyk, E. G. Alexander

INTRODUCTION The purpose of this paper is to summarise development of design processes used in the 1100 oreoody as a result of experience encountered. Design methods have been developed as a result of the history of ground behaviour, free-standing qualities of cemented back fill and mass blasting techniques. This paper contains an historical review of design methods previously described by Matthews ( 1972 ) , Hornsby and Sullivan (1977 and 19781, Kirkby, Blair and Hutton (19781, and Stewart (1980). The reasoning behind design philosphy changes are discussed and placed into context with methods now adopted. The 1100 orebody is Mount Isa's main producer of copper ore (5 million tonnes per year ) . Proven reserves are 120 million tonnes at 3.2% Cu. Location of the 1100 orebody in relation to overall mine geometry was well described by Dayton (1976). Geology The orebody is located in a series of massive brecciated shale lenses, in either a siliceous or dolomitic matrix, surrounded and separated by a uniform shale sequence. The whole formation is truncated at depth by a faulted contact with a greenschist facies metamorphic sequence below (Figure 1 and Figure 2) Mathias and Clark (1976). The breccia is known locally as silica dolomite and contains nearly all the economic mineralisation. Interfingering of shale and breccia often results in irregular boundaries to the silica dolomite and the ore limits. Geometry The geometry of the orebodies is shown by Figures 1, 2 and 3. Figure 3 also shows the location of the copper orebodies in relation to the lead orebodies. Structure Although the orebody is contained within the massive breccia (u.c.s. 150-250 MPa), it is cut by shear zones sub-parallel to the regional bedding (north-south striking, dipping approximately 60o to the west), and diagonally by major transverse faults (Figure A). These major faults have continuity in excess of 300 metres, with varying amounts of displacement up to 200 metres, and are zones of potential movement.

Jan 1, 1981

By John B. Hastings

The origin of the watery vapors of vulcanism has always been an object of interest and speculation to the seismologist, and as theories of the genetic origin of ore-deposits have of late years been pretty well narrowed down to the expiring forces of plutonic action, the same question has attracted mining engineers and geologists, as is well shown by the discussions of the subject in our Transactions. The important part taken by volcanic emanations in the origin of pegmatites and quartz-veins, described in my paper, Origin of Pegmatite,' and their latent power to concentrate into useful deposits such scattered gold as occurs in the Hartsel granite, make a discussion of their derivation but a natural third and final step. It is coilceded that enormous amounts of vapor accompany vulcanism, though perhaps we are apt to forget that steam has 1,700 times the volume of water, besides which the column seen over Vesuvius and other volcanoes is greatly mixed with air; the immensity of such volumes compared with solids and liquids is shown by the experiment of Gautier, who, by heating dessicated granite to 100' C., evolved from it gases 20 times, and steam 90 times, its own volume. Dana appraised the average amount of water left in ordinary rocks as 2.5 per cent. T. M. Read estimated that the Mississippi river carries to the sea annually 150,000,000 tons of rock material. If these Mississippi sediments, as deposited, contained 20 per cent. of water, it would be 600,000,000 cu. ft., or 4,500,000,000 American gallons, annually. Allowing the vapor suspended over a volcanic cone to be mixed with 80 per cent. of air, this amount of water converted into steam would replenish anew, every 9 min., a column 15,000 ft. high, 2,000 ft. in diameter at the base, and • 10,000 ft. at the top. Read's estimate is only half of the amount given in a more recent and probably correct one.2

Jan 1, 1909

By W. F. Carew, F. A. Crossley

Tensile, creep-rupture, and stability tests were employed to evaluate the effect of a-strengthening and ß-complexing additions on the base composition Ti-6Al-3Mo. The ultimate objective was to produce an alloy of greater strength than the Ti-7A1-3Mo alloy. Strengthening of a was accomplished by additions of 2 and 4 pct Ge, 4 pct Sn, 4 pct Zr, and 2 pct Sn-2 pct Zr. Making the ß more complex while keeping the total of ß-stabilizing additions constant at 3 pct resulted in higher tensile strengths but lower creep and rwpture strengths, especially at 1000°F. None of the more complex versions of the base alloy offers greater promise than the Ti-7A1-3Mo alloy. However, some improvement in the strength of the latter may follow from additions of zirconium perhaps up to 4 pct. PREVIOUS work directed at the development of titanium-base alloys with: 1) minimum room-temperature tensile elongation of 10 pct in 1 in., 2) high creep resistance at elevated temperature, and 3) the ability to retain original room temperature tensile ductility following exposure to creep conditions at elevated temperatures, demonstrated Ti-Al-Mo ternary compositions in the range 5 to 7 pct Al-3 to 6 pct Mo to be the most promising. Of this group the Ti-7Al-3Mo alloy was the strongest.'-= The objective of the present work was to determine if alloys of this type could be further strengthened by making additions which dissolve preferentially in the a-phase and by making the ß-phase more complex. For this study the base composition Ti-6Al-3Mo was selected rather than Ti-7Al-3Mo in order to minimize the possibility that in a more complex alloy with a-strengthening additions, low ductility might obscure the strength contribution. The metals

Jan 1, 1959

By M. D. Cooper

For the purpose of obtaining some first-hand data in regard to the shooting down of coal in bituminous mines, it was the writer's good fortune to be employed as a shot-firer for almost one year. In all, 6020 shots were fired by him during this period. All of the work was done in what is known as the South Main section of the No. 2 Mine of the Ellsworth Collieries Co., Ellsworth, Pa., a subsidiary of the Lackawanna Steel Go. Position of Shot-firing in the Cycle of Operations The established practice of the Ellsworth Collieries Co. calls first for the undercutting of the coal by electric chain machines across the face of the entry or room. This is followed by replacing any posts removed by the machine-runners, or the setting of new posts or timber, as may be required. After this has been done the machine cuttings are loaded into cars. Then the coal is spragged, and a shot hole drilled to the depth of the undercut. Next in order, the shot-firer loads the hole and fires the charge. The working place is then cleaned up, timbered if necessary, the sides and face squared up, and loose material taken down. The face is again ready to be undercut. Materials Used Permissible explosives were used exclusively in blasting coal. These were Carbonite No. 2, a nitroglyccrine explosive; and three different makes of ammonium nitrate explosives, Tunnelite, Red H, and Mine-ite 5-D. The last-named was used during the greater part of the period. It is interesting to note that even at the beginning of this work, the gases resulting from the explosions of the charges caused no especial distress on. the shot-firer's part and this continued to be the case except on the occasions when a change was made from one class of explosive to another.

Jan 1, 1918

By Henry G. Granger

It still rains in the Choco! I well remember one evening a dozen years ago, while traveling toward the town of Negua in a small champa or dug-out canoe, the rain fell so heavily that I had to work my hardest to bail out the water with a large totuma or gourd in order to keep the canoe afloat, the native boatman meanwhile plying his paddle in the stern. If any one thinks bailing is easy work, let him scoop out a quart or more at a stroke, 40 times a minute for half an hour, and I venture that he will change his opinion. These small canoes required only one paddler to make the 3 miles an hour which is the standard rate of aquatic progress on the Choco. Another time, on the Atrato river, about six o'clock in the afternoon, the rain fell so heavily that an empty gourd, 9 in. in diameter and 4.5 deep, was filled to the brim in 90 sec., timed by a watch. Such downpours are typical, but unless caught in one it is not specially noticed. These downpours seldom last more than an hour, and frequently cover only a limited area, occasionally falling on one bank of a river and not on the other. During the night-time, the rain, if gentle, will usually last until dawn, and sometimes an intermittent drizzle will keep up all day as well, although, as a rule, it rarely rains during the day. The rise of the rivers, due to the rain-fall, varies from a few inches to several feet. One night in August, 1899, the spuds of the dredge Margaret, operating on the Atrato river, showed a rise of 16 ft., and for 3 days thereafter occurred the greatest flood known in that district for 32 years. The dredge had all three spuds down, and the 2-yd. shovel dragging 011 the bottom. A 0.75-in. cable fast to a stout tree, 700 ft. away, stretched tight

Jan 1, 1909

By Ian Wark

SODIUM sulfide is used extensively to increase the recovery of oxidized copper and lead minerals by flotation, particularly when using xanthates as collectors. It is generally assumed that the sodium sulfide converts a film of the oxidized mineral to the sulfide, and that the xanthate, while not a very effective collector for the oxidized minerals, is a good collector for the sulfide. On the other hand, Taggart and his collaborators' claim that unless sulfide minerals are oxidized they cannot adsorb xanthates and do not float. Gaudin2 states that amyl xanthate, used in conjunction with sodium sulfide, gives better results than ethyl xanthate. Rabone3 and Mayer-Schranz4 agree that excess of sulfide functions as a depressant. Owing to the greater solubilities of the oxidized minerals than of the sulfides, the consumption of sodium sulfide is greater than that of most other flotation reagents, up to 10 lb. per ton of ore being used in some plants. Luyken and Bierbrauer5 recommend a water washing to remove soluble heavy metal salts; by this means sulfide consumption may be reduced. It was desirable to determine the limits of the sulfide concentrations between which contact is possible at surfaces of anglesite and cerussite in xanthate solutions. Similar determinations were desirable for the sulfide minerals with which the oxidized minerals are usually associated, since conditions should be chosen, if possible, so that the two types of mineral will float together. Because of the formation, especially in acid solutions, of hydrogen sulfide and of the hydrosulfide ion (HS-), the permissible limits of sulfide addition are obviously dependent upon the pH value. Sphalerite responds to ethyl xanthate only in the presence of an activator; for example, copper sulfate. It was necessary, therefore, to determine whether sodium sulfide interferes with the activation of sphalerite by copper sulfate. The combined effect of sulfide and copper sulfate was determined also for galena, chalcopyrite and pyrite.

Jan 1, 1936

By C.P. Proctor, J. T. Crawford, J.A. Younkins

Coal-cleaning plants using the launder process generally wash the fine coal (minus ? or minus 5/16-in.) separately in a plant consisting of washing launders or troughs placed one below another and sometimes supplemented by additional washing equipment such as tables and jigs. The coal and its associated impurities enter the feed end of the top launder, with a current of push water. As the material moves along the impurities, being heavier than coal, settle toward the bottom of the launder and are drawn off, with some coal, through "free discharge" boxes to the launder below. These boxes were given their name because their discharge falls freely; that is, it does not discharge into a water seal. Thus the impurities are progressively concentrated in the lower launders until box products are sufficiently free of coal to be discharged to refuse. The overflows of the upper launders go to clean coal while the overflow of the lowest launder, and perhaps that of the one immediately above, is usually recirculated to the plant feed. Thus the bulk of the clean coal is removed rapidly and the impurities are concentrated more slowly. The chief purpose of this paper is to present the cleaning data of such a plant under different operating conditions. OpeRating Equipment The washing units considered here are part of a complete plant cleaning minus 4-in. coal. The primary cleaned product is a high-quality gas and metallurgical coal and another portion of the output may be shipped as a steam coal. The units consist of two parallel Rheolaveur free-discharge units each containing five launders, together with three 7 by 14-ft. Plato tables. The launders are designated by the letters A-B-C-D-E (Fig. I), launder A being the top and longest; E being the bottom and shortest. Launder details are given in Table 1. The launders are equipped with a series of free-discharge Rheolaveur boxes, which are integral parts of the launder (Figs. 2 and 3). Each box with-

Jan 1, 1940

By C. P. Proctor, J. A. Younkins, J. T. Crawford

Coal-cleaning plants using the launder process generally wash the fine coal (minus ? or minus 5/16-in.) separately in a plant consisting of washing launders or troughs placed one below another and sometimes supplemented by additional washing equipment such as tables and jigs. The coal and its associated impurities enter the feed end of the top launder, with a current of push water. As the material moves along the impurities, being heavier than coal, settle toward the bottom of the launder and are drawn off, with some coal, through "free discharge" boxes to the launder below. These boxes were given their name because their discharge falls freely; that is, it does not discharge into a water seal. Thus the impurities are progressively concentrated in the lower launders until box products are sufficiently free of coal to be discharged to refuse. The overflows of the upper launders go to clean coal while the overflow of the lowest launder, and perhaps that of the one immediately above, is usually recirculated to the plant feed. Thus the bulk of the clean coal is removed rapidly and the impurities are concentrated more slowly. The chief purpose of this paper is to present the cleaning data of such a plant under different operating conditions. OpeRating Equipment The washing units considered here are part of a complete plant cleaning minus 4-in. coal. The primary cleaned product is a high-quality gas and metallurgical coal and another portion of the output may be shipped as a steam coal. The units consist of two parallel Rheolaveur free-discharge units each containing five launders, together with three 7 by 14-ft. Plato tables. The launders are designated by the letters A-B-C-D-E (Fig. I), launder A being the top and longest; E being the bottom and shortest. Launder details are given in Table 1. The launders are equipped with a series of free-discharge Rheolaveur boxes, which are integral parts of the launder (Figs. 2 and 3). Each box with-

Jan 1, 1940

By Walter E. Duncan, Oscar F. Tangel, A. M. Gaudin, Franklin Dewey, R. A. Johnson

Xanthates have been used as collectors of sulfide minerals in the concentration of ores by flotation for over a decade, yet much remains to be done to elucidate the underlying reasons for the remarkable effectiveness of these compounds in this connection. Briefly stated, the known facts concerning the effect of these reagents on the floatability of minerals are as follows: 1. Xanthates are without effect on silicates, silica, and generally on minerals currently regarded as "gangues." 2. Xanthates in very small amount increase enormously the floatability of sulfides of copper, lead, silver, iron and mercury. 3. If zinc sulfides are first '(activated," as by treatment with a copper salt, they are readily floated by xanthates; otherwise not. 4. Elemental copper, silver and gold have their floatabilities increased by xanthates. 5. Sulfides of elements that are not typically metallic—e.g., stibnite, realgar, molybdenit—are not made to float markedly better by treatment with xanthates. 6. Oxidized minerals of lead and copper are floated if treated with large amounts of xanthates, but not so if treated with amounts of the same order of magnitude as required for sulfide flotation. Oxidized zinc minerals are not floated unless xanthates having unusually long hydrocarbon chains (e.g., eight carbon atoms or more) are used. Iron oxides and carbonate are not floated by xanthates. From a flotation standpoint the most striking feature about xanthates is the very small amount required to accomplish startling changes in the behavior of some minerals. The amounts used in practice to effect excellent concentration are frequently of the order of 0.05 lb, of agent per ton of ore; that is, of the order of one part of agent per forty thousand parts of ore. Changes in amount of agent of the order of 0.01 lb. per ton of ore (or 0.002 lb. per ton of pulp, on a 4:l basis) have marked effects on the metallurgical results obtained. Yet these changes represent a variation of but one part per million parts of pulp.

Jan 1, 1935

By C. S. Barrett

It has been observed that deformation in iron takes place by slip on (110) + {112) + (123) planesl, 2, but in silicon ferrite with low deformation temperatures or high silicon contents (exceeding 4 per cent) the authors have found that slip is confined wholly to (110) planes, at least with small amounts of deformation. The purpose of the present investigation was to determine the effect of this alteration of slip mechanism on preferred orientations in cold-rolled and recrystallized strips. Incidental observations on the effect of the recrystallization temperature were also made. The numerous determinations of the texture of cold-rolled iron and steel have been in satisfactory agreement as to the principal features of the texture (with the exception of a few conclusions that apparently have been based on insufficient data). The texture is chiefly one in which 11101 directions of the grains lie along the direction of rolling, with a deviation of a few degrees, and in which the (100) planes lie in the plane of the rolled sheet, with a deviation from this position chiefly about the rolling direction as an axis. This deviation about the rolling direction as an axis has been measured by different observers and lies between about 45" and 50" or 60" for iron5, 6, 9 and is usually in the neighborhood of 50" for mild stee16, 7, 8. The range of this deviation is a function of the percentage of total reduction8, 10. Post8 reported that it decreases with increasing percentage reduction, but apparently he studied only the surface material, which Gensamer and Meh17 found to be somewhat differently oriented from the material in the inside of the sheet. McLachlan and Davey10 found that the deviation also decreases with increasing percentage of total reduction for the material in the interior of the sheet and is independent of the percentage of reduction per pass. The deviation about the cross direction as an axis (the direction in the

Jan 1, 1937