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In presenting this paper the author has in view a twofold purpose. In the first place, it is intended to set forth for the benefit of practical mining men some account of the principles underlying the design of tungsten carbide rock bits in the hope that a knowledge of these principles will assist in the successful application of the bits to mine and quarry drilling. Secondly, by evoking discussion, and even controversy, it is hoped that manufacturers and users of these rock bits will be encouraged to contribute to the general knowledge of a subject which to date has been rather lacking in published literature. The author well appreciates that many of the conclusions drawn in this paper are highly controversial, but if the only result is to call forth contradictory testimony from other workers in this field, at least one of his purposes will have been achieved.An attempt will first be made to set forth and examine those characteristics which go to make up an efficient rock bit, and the main factors affecting these characteristics will be discussed. In a paper of this general nature, consideration of any particular aspect must necessarily be rather brief, and there is little opportunity for detailed presentation of test procedures and data. Thus, it has been found advisable' to confine discussion of some minor points to a simple statement of fact. However, an attempt has been made to support all major conclusions by reference...

Jan 1, 1950

By E. V. Potter

For a nurnber of years, it has been known that manganese made by electro-deposition under certain conditions is ductile while under other conditions it is very brittle. The ductile metal is gamma manganese normally stable only between 1100 and 1138°C1; the brittle metal is alpha manganese, stable up to 727OC. The ductile metal is not stable, but gradually changes to the brittle form; the time required to complete the transfornlation is about 20 days at room temperature. Other observations have indicated that the transformation is completed in 10 to 15 min. at about 125°C, while at — 10°C, no appreciable change occurs in 9 months. The properties of gainma and alpha Illanganese in the pure state are ordinarilj difficult to determine because the gamma structure cannot be retained by normal quenching procedures and alpha manganese is so brittle, it is difficult to obtain specimens free from flaws. In a recent investigation2 some properties of gamma and alpha manganese were determined by studying the ductile electrolytic metal and determining the changes in its properties as it transformed to the brittle alpha form. These investigations provided an excellent opportunity for following the progress of the transition and studying its mechanism. The results of a series of such investigations are reported in this paper. Procedure Various properties of manganese were determined starting with the metal in the original ductile gamma form and following the subsequent changes in its properties as the metal transformed to the brittle alpha form. These observations were made at various temperatures, the data providing information regartling the mechanism of the transformation as well as the effect of temperature 011 the transition rate. Structure and resistivity values gave the most significant results, so this paper is concerned primarily with them. The structure was studied microscopically as well as by X ray diffraction. The resistivity was determined on strips of the metal by measuring the potential drop across a given length of the specimen. Current was passed through the specimen by wires soldered to its ends, and the potential connections were made by wires looped around the specimen near its center. The current was determined by the potential drop across a standard resistor connected in series with the specimen, the potential drop being measured on a potentiometer. In the temperature range from room temperature to 100°C an ordinary drying oven was used to heat the specimen. This was entirely satisfactory except at 100°C, where the time required to heat the specimen was long compared to the transition time, making the initial section of the resistivity curve unsatisfactory. To overcome this limitation, at 100°C and higher a thermostatically controlled oil bath was used to heat the specimens. The block on which the specimen was mountetl was plunged into the hot oil at the start of each test. The heating time was thereby reduced from 5 min. to about 6 sec, and dependable resistivity values could be obtained through 160°C. At this point the whole transition, including the warm-up time for the specimen, required only about 20 sec and it was not considered worth while trying to extend the temperature range further. Aside from the heating problem, the problem of making a sufficient number of accurate resistivity determinations became more and more difficult as the temperature was raised. Using the manually operated potentiometer, 100°C was about as far as it was possible to go. At this temperature and above, a self-balancing photoelectric recording potentiometer was used. Its response was quite rapid, and it proved to be entirely satisfactory all the way through 160°C, where the tests were stopped because of the specimen heating problem rather than any limitation of the potentiometer recorder. The metal used in these tests was prepared at the Salt Lake City laboratory of the Bureau of Mines. The method of preparation is discussed in a paper by Schlain and Prater.3 The sheets were about 2 3/8 by 5 3/16 in. and varied from 10 to 16 mils in thickness. They could be cut readily into pieces suitable for the various tests. X ray and microstructure determinations were made on pieces about 1/8 to 1/4 in. wide and about 1 in. long, while resistivity measurements were made on strips as long as possible and about 55 in. wide. The thickness of each sheet was not uniform over all its surface. This had no bearing on the X ray and microstructure determinations, but sections as nearly uniform and free from flaws as possible were chosen for the resistivity determinations. The gamma manganese was electro-deposited at 30°C, the time of deposition ranging from 5 to 12 hr for each sheet. Whenever possible, the tests were started directly after the metal was stripped from the cathode; otherwise the sheet was placed immediately

Jan 1, 1950

By Harold L. Gibbs

The purpose of this paper is to encourage petrographers to prepare thin sections, as the demand for microscopic examination of rocks, ores, and metallurgical products is becoming more general in .the mining metallurgical-fields. This paper has been written with the knowledge that the successful thin-section methods and the several new techniques herein described will be helpful to all engaged in this field of work. Detailed instructions for making thin sections by methods developed at the Salt Lake-City laboratory of the Bureau of' Mines are given in this paper. Several new techniques are described for making thin sections of loose grains and fine sands and for making good thin sections of highly altered and friable rocks. The methods described insure superior bonding of the rock to the glass slide, thus virtually eliminating any failures due to parting of the section from the glass slide, and permit fast mounting, grinding, and finishing, thereby materially lessening the time required to prepare thin sections.

Jan 1, 1950

By D. Turnbull, J. H. Hollomon, J. C. Fisher

Application of the concepts of nu-cleation and growth to the analysis of experimental transformation data has led to valuable descriptions of phase transformations, an outstanding example being the transformation austenite —* pearlite which has been examined with particular care by Mehl and co-workers.&apos;-5 In addition to the pearlite transformation, the proeutectoid fer-rite and proeutectoid carbide transformations are known to proceed by nucleation and growth. Martensite, on the contrary, until recently was thought to form by a mechanism involving neither nucleation nor growth; however, extension of standard nucleation theory6 suggests that martensite, bain-ite, and the other products of austenite decomposition all grow from nuclei in the parent phase. The theory that martensite forms by nucleation and growth is strongly supported by recent experimental work of Kurdjumov and Maksimova.7 The concepts of nucleatioli and growth have been fruitful also in providing a sound basis for quantitative theoretical treatments of the kinetics of phase transformations. For example, Volmer and Weber8 and Becker and Döring9 developed the theory of nucleation from fundamental considerations to a point where excellent agreement was obtained with the results of experiments on the condensation of supercooled vapors. As a result of their analysis, the kinetics of vapor-liquid transformations now can be predicted. It seems probable that application of the theories of nucleation and growth to a quantitative study of austenite decomposition similarly will clarify the nature of the individual transfor: mations involved, and will enable the calculation of austenite transformation kinetics. In the present paper, the theories of nucleation and growth are applied to the austenite ? martensite transformation in steels. The analysis begins with a discussion of nucleation in single component systems. Martensite appears to be coherent with the parent austenite, hence the nucleation theory is modified to include the effects of elastic distortion. Nucleation in the two component iron-carbon system then is discussed, for most steels are primarily alloys of these two elements. Finally, M. temperatures and martensite transformation curves are calculated for each of several alloy steels of varying carbon and chromium content, and are compared with those determined experimentally by Lyman and Troiano10 and Harris and Cohen.11 Nucleation Theory NUCLEATION IN SINGLE COMPONENT SYSTEMS6,12-14 The work required for reversible formation of a region of phase within the parent a phase is expressed conveniently as the sum of two terms: W1 = Aa, the product of the area of the interface and the interfacial free energy, and W2 = VAf, the product of the volume of the region and the free energy increase per unit volume associated with the transformation. The total work is therefore W = Aa + VAf. When a is more stable than ß, Af is positive and W increases without limit as the volume increases. The transformation a ?ß does not occur. It is nevertheless true that small regions of phase ß enjoy temporary existence here and there in the a. The equilibrium number of ß regions of given size is proportional to exp(— W/kT) per unit volume of a, assuring that larger (ß regions occur with diminishing probability. When a is less stable than ß, Af is negative and W passes through a maximum as V increases. Assuming for simplicity that regions of ß are spherical, as is true when the interfacial tension is isotropic and there are no elastic strains, W = 4r2a + (4/3)*r3Af The maximum value of W is W* = 16iro3/3Af2 [1] for regions with radius r* = -2o/Af. [2] For single component condensed systems it has been shown14 that the steady rate of nucleation of 0 per unit volume of untransformed a is nearly proportional to exp[- (W* + Q)/kT] where Q is the activation energy for the unit processes of adding or removing one atom from an embryo or nucleus. If To is the temperature at which a and ß are in equilibrium, the rate of nucleation is a maximum at a temperature 0 < T < To where (W* + Q)/kT is a minimum. P regions smaller than critical size are called embryos; they tend to grow smaller and disappear, only exceptionally growing larger. Regions equal to or larger than critical size are called nuclei. A critical size nucleus may grow indefinitely large or may shrink back to a, either process decreasing the free energy of the region.

Jan 1, 1950

By Hugh P. Godard, K. Winslow

"AbstractA number of panels of various aluminum alloys in different combinations and secured by steel and three types of aluminum rivets were exposed underground at two locations in a Northern Ontario gold mine for two years and then examined to determine the extent of corrosion. The results indicate Alcan 65S-T to possess considerable advantages over Alcan 178-T, which has been employed in the construction of light skips and cages with considerable success for almost twenty years. Information on the behaviour of these alloys when used with steel rivets and in mixed combinations with themselves was obtained.IntroductionBy the complete substitution of aluminum alloys for steel in the fabrication of mining skips and cages, their weight can be reduced as much as 50 per cent. Lesser savings are possible though its savings are possible through its plates, doors, reinforcing members, etc. Reduced weight leads to economy through increased hoisting efficiency.For a new installation, the basis of design is the rate of production, which thus fixes the live load to be hoisted per trip. A lighter skip permits the use of a lighter rope and a smaller hoist. In deep shafts with high-speed hoists operated in balance, the power required for accelerating is considerable. The hoist size is determined by the weight of skip, load, and rope plus the force required to accelerate or decelerate these three elements A light skip may result in a considerably lighter rope. This is important because, even though the skip can be completely balanced this is not true of the rope. Partial balancing is possible, but conical hoist drums or other devices for this purpose are costly; thus a smaller rope results in a smaller hoist, with consequent general reduction in cost of the whole installation. A reduction in power consumption is also obtained as a result of keeping to a minimum the amount of unbalanced load and the weight of all moving parts which must be accelerated."

Jan 1, 1950

By H. M. Thorne

The oil-shale deposits of the Puertollano basin in the Spanish province of Ciudad Real appear to have been give little consideration in discussions of oil-shale resources throughout the world. A small, integrated oil-shale operation has been conducted at Puertollano since about 1922 by a French company, Sociedad Mimora y Metalurgica de Penarroya, hereinafter referred to as "Penarroya", but only within recent years have the potentialities of the Spanish oil-shale deposits been recognized. Empresa Nacional "Calvo Sotelo" do Combustibles y Lubricantes, hereinafter designated as "Calvo Sotelo," which was created in 1942 by the National Industrial Institute of Spain to produce liquid fuels from oil shales, has made marked progress in the design and construction of a complete oil-shale plant at Puertollano. Penarroya is mainly a coal-mining company, and the oil-shale operations were on a small scale of approximately 220 tons a day in October 1947. It is an integrated operation comprising oil-shale mining and retorting and shale-oil refining. Motor gasoline, Diesel fuel oil, light burner fuel oil, lubricants, paraffin wax, cresols, and ammonium sulfate are manufactured.

Jan 1, 1950

By Robert T. MacMillan

An essentially chemical technique has been developed on a small scale for the efficient utilization of certain domestic titaniferous iron ores. Three main steps are involved: (1) Sintering the ore at 1,050° C. with carbon and soda ash to reduce the iron to metal powder, retaining the titania in the slag, (2) separating the powdered iron from the slag, after wet grinding, by magnetic or gravity methods, and (3) sulfuric-acid decomposition of the titania fraction to produce pigment and certain byproducts. The technique was first applied to the tailings from the Bureau of Mines bauxite-concentrating plant at Bauxite, Ark. A good grade of powdered iron was produced from this material. Subsequently, the work was diverted to the study of the titaniferous magnetite from the MacIntyre Mine at Tahawus, N.Y., because of the potentially higher recoveries of iron, titanium and vanadium. Ninety percent of the iron in this ore has been recovered as an iron powder analyzing 90 percent metallic iron and containing 2 to 3 percent TiO2. The remaining titania was found in the slag and was recovered as pigment-grade TiO2 by application of a suitable technique. The vanadium was distributed between the TiO2 fraction and the alkaline liquor resulting from the wet-grinding and magnetic-separation operations, with only a minor proportion retained by the metallic iron.

Jan 1, 1950

By J. D. Davis

The Bureau of Mines Coal Carbonization Laboratory has been interested for the past 20 years in the quality of coke in connection with the selection of coals for the manufacture of coke. The principal experimental. method of carbonization, whereby 100- to 200-pound charges are carbonized in sheet-steel cylindrical retorts with recovery and determination of the properties of coke, gas, and products, was developed in cooperation with the American Gas Association,4 5 and it is known as the BM-AGA method. Results of the investigations of 90 coals included in the BM-AGA Survey of the Carbonizing Prop¬erties of American Coals have been published; 6 and, in addition to these, about 200 coals have been carbonized over limited temperature ranges. Because coke from cylindrical retorts differs in shape from that made in rectangular ovens, the BM-AGA method has been criticized, although those who developed it have always believed that it served well as a means of measuring the relative strengths of cokes from different coals. It is recognized that no laboratory-method carbonization, even on a large twit, can yield results that exactly duplicate those obtained in commercial ovens and retorts. However, with the completion and satisfactory performance of an experimental vertical slot oven of 500-pound capacity at the University of Illinois Geological Survey Laboratory,&apos; the opportunity to directly compare BM-AGA coke with coke made by two-sided heating became available and the necessary test apparatus was installed. The Bureau of Mines slot oven, which is similar to that designed by the University of Illinois Geological Survey, has been used to carbonize 11 coals (or blends) tested by the BM-AGA method. This report compares the results obtained by the two methods.

Jan 1, 1950

By A. B. Needham

A number of the pegmatites in the southern Black Hills of western South Dakota contain sheet mica; and as the demand increased during World War II the area became an important producer of strategic qualities of mica. During this period of maximum activity, the U. S.. Geological Survey made detailed studies of the geological structural control, the mineralogical associations, and the production Of many pegmatites in the Black Hills. On the basis of results of this study, and at the suggestion of the Geological Survey, several of the pegmatites were examined by the Bureau of Mines, and a few were selected for further investigation. Accordingly, the White Bear, Silver Dollar, Buster Dike, and Hot Shot mines were diamond-core-drilled in 1947. The Hot Shot mine was investigated further by deepening the shaft and crosscutting the pegmatite in 1948. The mica-bearing deposits occur in lenticular intrusions of pegmatite in schists of various mineral composition. Eleven diamond-drill holes, totaling 2,111.4 feet in length, were completed on the four deposits. The shaft on the Hot Shot deposit was deepened 44 feet to the 112-foot level, and crosscuts were driven from the bottom of the shaft to the hanging Wall and foot wall of the pegmatite. This publication briefly describes the location, ownership, and history of the deposits and gives details of the work done by the Bureau of Mines and the Geological Survey.

Jan 1, 1950

By C. S. Smith, K. K. Ikeuye

In an earlier paper1 one of the authors called attention to the significance of the relative free energies of grain boundaries and interphase boundaries in alloys in determining the shape and distribution of micro-constituents. The ratio of the interphase and inter-crystalline boundary energies (that is, —?12, where ?12 is the energy of the interface between a crystal of phase 1 and a crystal of phase 2, and ?11 is the energy of the boundary between two crystals of phase 1) will be denoted by F. It is this ratio that determines, by the simple triangle of forces relation, the dihedral angle, ?, of phase 2 where it adjoins a boundary between two grains of phase 1. On a metallographic sample various angles are actually obtained because of the random orientation of the plane of observation, but the statistically most frequent angle is the true one. These basic principles were discussed in detail in the earlier paper. The experiments described below show the effect of both composition and temperature on ? and T for some alloys containing a liquid phase, and provide data on the rate of approach to the equilibrium angles in both solid and liquid phases. All the alloys were melted in a graphite crucible and cast into ingots 3/4 by 11/8 in. cross-section. Some of the aluminum alloys consisted of two immiscible liquids from which uniform castings could not be obtained. The amount of liquid phase varied locally to a considerable extent and the analyses were meaningless. Nevertheless, the composition of the liquid would not differ from that in an alloy of the intended gross composition, and the synthetic composition of the alloys was therefore used in considering the surface energy results. The ingots were given initial deformation by cold-rolling to 50 pct reduction, annealed, and given a second 50 pct reduction prior to cutting into small samples for heat treatment at the times and temperatures subsequently reported. The resulting micro-samples were carefully prepared metallographically and examined at magnifications of 650, 1280, or 2500. The objectives used were 21X, 0.40 NA (dry); 41X, 0.65 NA (dry); and 80X, 1.40 NA (oil emulsion). The appropriate angles were measured by rotation of the stage of a Bausch and Lomb metallograph to align the appropriate boundaries successively with a cross-hair in a micrometer eyepiece. About 250 angles were measured on each phase. Though the angles were read to 1°, the accuracy is probably not above 3°, because of uncertainty in setting the cross-hair tangent to the sometimes curved boundary. The corners of very small particles, particularly in early stages of adjustment after cold-rolling, were difficult to measure and the results were slightly dependent on the magnification used (compare Fig 1-C and 1-D). With particles of larger size and more nearly equilibrium shape, magnification made no difference (vide Fig 1-F and 1-G). The measured angles were grouped in 10° intervals (5° if ? was less than 20°) and plotted in bar

Jan 1, 1950

By George L. Gates

It is generally recognized that if the lose and contamination of the fluid content of a core specimen from an oil- and gas-bearing formation in a well were eliminated, the value of core analysis would be increased greatly. In an effort to obtain uncontaminated cores, extensive field tests were carried out in the Mercy field, San Jacinto County, Tex., with the Carter pressure core barrel and an oil-base drilling fluid having very low rates of filtration as measured by the standard A.P.I. drilling-fluid filter teat. A tracer was added to the drilling fluid filter test. A tracer was added to the drilling fluid to provide a means for quantitatively measuring the volume of oil filtrate in the cores. Unfortunately, as stated in the report on the Mercy tests, very little oil nand was penetrated in coring the well; and because the performance of the pressure core barrel was not entirely satisfactory, the tests did not provide a complete answer to the question: Does a filtrate from an oil-base drilling fluid enter cores., and how does it affect the results of the core analyses?

Jan 1, 1950

By Hugh H. Waesche

Of all the military services, the Signal Corps is the most concerned with piezoelectric minerals because of its function as a supply service to the strategic and tactical military forces. Consequently this paper is written from the point of view of one associated with that organization. The Signal Corps is responsible for the research, development, and supply of communications, radar, and components to the using services of the Department of the Army and to some extent the Other branches of the National Defense Department. Their work therefore includes the study of the sources* characteristics, and application of quartz and other piezoelectric materials. These materials have become a vital consideration in strategic planning and are essential for efficient tactical operation by all the Armed Forces. The Signal Corps at the beginning of world War 11 Was respon-sible for both Army Ground and Air Force electronic equipment. Since that time this Army service organization has probably done more in the development of frequency control devices using piezoelectric materials than any other group. The U.S. Department of the Interior, Bureau of Mines, Minerals yearbook of 1945, shows that during the four war years, 1942 through 1945, 9,598,-410 Ib of quartz crystal were imported for all uses and of this total, 5,168,000 lb were consumed to produce 78,320,-000 crystal units for electronic application. Other government records confirm these data which conclusively show that approximately 53 pct of the crystalline quartz imported was consumed in the production of electronically applied quartz crystal units. It may be assumed that some effort was made to maintain a stockpile over demands for all purposes. and this would mean that the actual percentage of quartz used electronically was considerably over the 53 pct figure. These data only emphasize that electronic application of crystalline quartz was the greatest requirement, and per- haps the actual value in this application to national defense is many times greater in importance than is apparent on first inspection. Current electronic research and development programs of the Armed Forces are planned around the fundamental use of piezoelectric minerals for frequency control and this at present, at least, means quartz. Definition and Early Development The word piezoelectricity is formed from combination of the Greek word "piezein". meaning "to press," and "electricity." It is that property shown by numerous crystalline substances whereby electrical charges of equal and opposite value are produced on certain surfaces when the crystal is subjected to mechanical stress. It appears to be intimately associated with the better known property, pyro-electricity and in fact, the two may be manifestations of the same phenomeuon. This property was discovered by Pierre and Jacques Curie in quartz, tourmaline, and other minerals in 1880 while studying the symmetry of crystals. The converse effect, that is, mechanical strain in the crystal when placed in an electrical field, was predicted by the French physicist, G. Lippman, in 1881, and verified by the Curies almost immediately. As has been the case with many discoveries of similar character in the basic sciences, not much attention was paid to this property for man)- years except as an entertaining curiosity. Between 1890 and 1892 a series of papers was published by W. voigt in which the theoretical physical properties were put into mathematical form. The first practical application of piezoelectricity occurred during World War I when professor P. Langevin of France used quartz mosaics to produce underwater sound waves. The same mosaics were used to pick up the sound reflections from submerged objects which were in turn, amplified by electronic means and used to determine the distances to such objects. This device was intended for use as a submarine detector but development was not completed in time for war service although it was used later for determining ocean depths. About the same time, A. M. Nicholson, of Bell Telephone Laboratories, developed microphones and phonograph pickups using Rochelle salt crystals. A major step in the application of piezoelectric quartz came in 1921, when professor W. G. Cady, of wesleyan university, showed that a radio oscillator could be controlled by a quartz crystal; from that date, this use of quartz has increased steadily, reaching its peak in world war 11 as is shown by the figures previously given. Essentially all American electronic equipment for communication, navigation, and radar, utilized quartz crystals for the exacting frequency control required. Crystalline Minerals with piezoelectric Properties QUARTZ Hundreds of piezoelectric crystalline materials are known, most of which are water soluble. Of these, quartz appears to be without a peer for electronic frequency control. Unfortunately, the quartz must be of superior quality. It must be free of mechanical flaws, essen-tially optically clear, free of both Brazil and Dauphiné twinning and must be, for average uses, over 100 g in weight. Because of these stringent requirements, raw quartz of the quality desired is of rare occurrence. In addition to quartz, several other naturally occurring crystalline materials are known to have the piezoelectric property and could perhaps be substituted for quartz in some applications. These

Jan 1, 1950

By W. J. Kroll

This investigation was undertaken with the idea of studying a number of unconventional methods for making chromium to determine whether production cost could be lowered and to discuss some of the more recently described processes. These comprise carbon reduction of the oxide in vacuum, nitride dissociation, fusion electrolysis of the chloride, amalgam method, iodide dissociation, calcium reduction of the oxide, hydrogen reduction of the chloride, and reduction of this compound with zinc and with magnesium. Preference is given this latter method, which is shown in a flow sheet. Special attention is given purification of the chloride and separation of the vacuum at about 900° C. A reduction method for the elimination of traces of oxide contained in pure chromium by pure hydrogen obtained from the dissociation of zirconium hydride is described. The various methods are compared in a table in the appendix. Aluminothermic and silicothermic methods for the reduction of chromium oxide and the aqueous electrolysis have been disregarded in this investigation.

Jan 1, 1950

By Bruce F. Grant

This report discusses research being done by the Petroleum and Natural Gas Branch und the Applied Physics Branch of the Bureau of Mines to deter-mine the most effective uses of explosives in shooting oil and gas wells, as measured in terms of rates of production from the wells. An experimental investigation of the effects of detonating explosives in shallow holes drilled in rocks, chosen to be similar to those forming oil and gas reservoirs, forms the basis of this study. Records obtained from strain gages placed in the rock around the experimental shots permit determination of the generation and propagation characteristics of the strain wave. Water-injection tests made on the test holes before and after shooting allow the increase in fluid conductivity resulting from the shot to be determined. By rotating these data to the physical properties of the rock, and by conducting shooting experiments over the range of rock types common to oil and gas reservoirs, it is expected that relationships between the fluid conductivity and the properties of the explosive and rock can be developed and used for designing efficient shots for oil and gas wells. Four series of experiment.al tests were completed during the summer of 1949 and are described herein. The results of these tests confirmed the feasibility of the experimental approach and permitted outlining the plans for future research discussed in this report. This report is presented to acquaint industry with this research program and the progress to date, so that the Bureau may have the benefit of suggestions from interested readers.

Jan 1, 1950

By S. S. Cole, D. L. Armant

The smelting of titaniferous ores for the past hundred years has not been successful because of thickening of the slags in the furnaces. The interest in the utilization of these iron bearing materials has been due to the extensive deposits which have been found in various parts of the world.1 The literature contains references2 to the use of titaniferous ores in blast furnaces. None of these attempts has met with success due to the lack of sufficient information as to the cause of failures and to economic considerations. Evaluation of the published information, as to the probable cause of failure in various attempts, has indicated that the presence of lower valence titanium oxide and formation of titanium carbide contributed to the failure. It was apparent that severe reducing conditions which are present in a blast furnace had been one of the factors for such failures. The previous smelting operations had as a primary objective the recovery of iron from these ores. Under these conditions the highest recovery of iron could only occur with severe reducing conditions. Application of Melting Point Data The application of the melting point data, from the system CaO-MgO-Al2O3-TiO2 reported in another paper,3 to the smelting of titaniferous ores required an extrapolation of these melting points due to the presence of other impurities from the ores. Likewise, a study of the effect of varying FeO and reduced titanium content of the slag was necessary before the data could be fully applied. The melting point data under oxidizing conditions had clearly indicated that a fluid slag with a melting point in the temperature range of 1370°C could be obtained. The smelting of a titaniferous ore had as a primary objective the production of a high titanium slag which was fluid and substantially free of iron and low in trivalent titanium. Since production of a high titanium slag was of primary importance, rather than the recovery of iron from these ores, the smelting concept could be altered to a marked degree. The ferrous oxide could be maintained at higher levels than is the normal practice in iron ores slags without being detrimental to the slags, although a loss in iron recovery resulted. During the reduction of the molten mixtures of ferrous iron and titanium, carbon acts as the reductant for changing the valence of the titanium and iron. The reaction can go further than in the case of the reduction which occurs in iron-titanium solutions. The formation of metallic iron removes iron from the system, so that the titanium oxide can be reduced to trivalent and divalent titanium oxides and also form titanium carbide which removes the titanium from the slags as a soluble constituent. The problem, therefore, resolves itself into maintaining the proper amount of ferrous iron in the slag thereby avoiding the formation of appreciable amounts of trivalent titanium and no divalent titanium oxide or titanium carbide. The latter two have been found in the viscous

Jan 1, 1950

By A. A. Collins

When we speak of high copper on a lead blast furnace we think in terms of 4 to 5 pct, or. any lead charge carrying over 1 pct. Any copper on charge will produce its corresponding troubles such as lead well, extra slag losses, drossing problems, and the working up of the dross. This is indeed a very interesting subject and one that used to give the old-time lead metallurgists such as Eiler, Hahn and lles many worries, not so much in the actual operation of the hlast furnace but in the working up of the copper. When the American nletallurgists commenced with the American rectangular-shaped lead blast furnace in the 1870&apos;s and got away from the reverberatories such as were in use in Germany and other parts of the world, they went to greater tonnages, as 80 to 100 tons per day in comparison to the 20 to 30 tons per day in the other processes. With the greater tonnages along with insuficient settling capacity, the silver losses in some cases were increased. Hence the lead-fall was low, for there were no leady concentrates in those days to assist the metallurgist to gain lead or an absorber for the precious metals; and in some cases copper sulphides were added intentionally to the charge to produce a copper matte to lessen the silver losses through the dump slag. The operators in those days thought that where some copper was always present in the lead ores the copper should not enter into the reduced lead and alloy with it. This, by the way, is just the reverse of our present-day practice, when we try to put all of the copper into the blast furnace lead and to remove the same through the drossing kettles. Therefore the furnace was operated to produce a certain amount of matte or artificial sulphides, since, due to the great affinity of copper for sulphur, any copper present would enter the matte almost completely. Thus, the lead bullion produced was practically free from copper. The products of the furnace were metallic lead or lead bullion, containing 05 to 95 pct of the lead and about 96 pct of the silver which were in the ore—a lead-copper-iron matte which contained nearly all the copper in the ore and the slag, the waste product. In the United States, up through the year 1092, we find the small furnace 100 X 32 1/2 in. with 12 tuyeres, some 6 on each side, plagued with a small amount of poorly roasted sulphides— either from heap or hand roasters that produced matte. This matte was roasted and if poor in copper was returned for the ore smelting. Otherwise it was smelted either alone or with additions of rich slags or argentiferous copper ores, the products being lead and a highly cupriferous matte, the latter being subsequently worked up for its copper. The lead metallurgists kept trying and improving on furnace and roasting equipment designs until we find blalvin W. Iles constructing at the old Globe Plant at Denver what came to be the modern furnace. That is, in 1900 he built a furnace of 42 in. width by 140 in. at the tuyeres with a 10 in. bosh and a 16-ft ore column. This type has been more or less standard to the present time, though modified in width and length to meet the demand for large tonnages and improvements in structural details. In 1905 at Cananea, Mexico, Dwight and Lloyd developed the present down-draft sinter machine that has meant so much in producing a well-processed material for the lead blast furnace. In 1912 Guy C. Riddell came forth with double roasting at the East Helena Plant of the American Smelting and Refining Co., which removed the "zinc mush plague." Incidentally, with the introduction of double roasting, which most lead plants were forced into after 1924, when lead flotation came into its own, less matte or no matte was produced. When this stage arrived, the copper was forced into the dross and the casting of lead at the blast furnace lead-wells was stopped. In plants with a fair copper carry 1 pct or better on the blast furnace charge, the lead wells became inoperative once the production of matte stopped. The copper drosses clogged the lead wells and even with bombing, either water or dynamite, the operators could not keep them open. Thus, the lead wells were abandoned in some plants, such as at the El Paso and Chihuahua smelters of the American Smelting and Refinillg Co., and all lead taken out through the first settlers. The elimination of sulphur, espccially sulphide sulphur, from the blast furnace charge and the nonproductiori of matte resulted in a great saving of tinie, energy and equipment in the recirculation of the copper, With the copper content in the dross and dross-fall ranging in quantities from a few percent up to 60 pct, such as at El Paso, a drossing problem was created. As the old-time operators hated dross and buried the same in the shipping bullion, the modern metallurgists from 1925 on decided that with increasing dross-falls they would have to adopt the lead refiner&apos;s ideas of drossing kettles with subsequent treatment of the lead with a sulphur addition to have the shipping lead of 0.01

Jan 1, 1950

By Ming-Kao Yen

A considerable amount of work has been reported in the literature in regard to the texture and earing behavior of copper strip. The rolling texture of copper has been confirmed as (110) [112] and (112) [111], which yields ears of a drawn cup at the position 45" from the rolling direction.1-3 The recrystallization texture has been established as the cubic or (100) [001] texture, where the earing positions are at 0" and 90" to the rolling direction.4-8 It has also been reported that in the development of cubically aligned grains of copper strips, the percentage of this cubic texture increased with an increase of final reduction and final annealing temperature.8,9 A comprehensive study on H.C. copper (British commercial copper of high-conductivity quality = Cu 99.95 pct, O2 0.03 pct, Ag 0.003 pct, Fe 0.005 pct and Pb < 0.001 pct) was made by Cook and Richards.6 They concluded that the recrystallization textures could be described as one or more of the following textures: (1) a single texture (100) 10011, (2) a twin texture (110) [112] and (3) a random orientation, depending upon the previous history of the specimen concerned. The effect of various alloying additions in copper was reported by Dahl and Pawlek.10 They found that certain alloying additions, such as 5 pct Zn, 1 pct Sn, 4 pct Al, 0.5 pct Be, 0.5 pct Cd, or 0.05 pct P suppressed the formation of cubic texture. Brick, Martin and Angierll reported that the cold rolled textures due to various additions fitted a rather simple pattern. However, the recrystallization textures were subject to very considerable variations. In the discussion of this paper, Baldwin stated that deoxidized copper containing 0.02 pct P gave a complicated recrystallization texture at lower temperature. When this copper was annealed at high temperature, a single texture appeared which was described as (110) [ill] but. according to a pri- vate communication from Baldwin, this orientation reported was in error and should have been reported as (110)[112]. He also reported that the earing positions of drawn cups were at 60" to the rolling direction.12 Recently, Howald, in his discussion on the paper by Hibbard and Yen,13 reported that the rolling texture of phosphorus deoxidized copper, containing from 0.006 to 0.020 pct phosphorus, was of the pure copper type. When these coppers were annealed at lower temperatures, they exhibited a random orientation, and when they were annealed at higher temperatures they had a mixed (111)[110] and (100)[001] texture, depending on the severity of the final reduction and annealing temperature. However, the specific influence of phosphorus and other impurities on the recrystallization textures and the deep drawing properties of copper strip has not been thoroughly reported. Therefore, an attempt has been made in the present work to determine the rolling and recrystallization textures and also the earing behavior of five types of commercial copper and thereby to evaluate the effect of phosphorus and some other significant impurities on the development of texture for cold reductions of about 87 to 89 pct. Materials Used The five types of copper employed in the present investigation were two phosphorus deoxidized coppers of different phosphorus content (0.007 and 0.013 pct P), an oxygen-free copper (OFHC), an electrolytic tough-pitch copper, and a fire-refined tough-pitch copper. These materials were subjected to a thorough spectroscopic and chemical analysis. The designations and the chemical compositions were as shown in Table 1. The coppers, FA1, FA2 and FA3. were hot-forged from 3-in. billets into a ½ X 6-in. plate and cold rolled to the ready-to-finish gauge indicated below. FA4 and FA5 were hot rolled and scalped to ready-to-finish gauge. The grain size of all the materials in the ready-to-finish condition was about 0.030 to 0.045 mm. Table 2 shows the last stage of the production schedule for each copper strip used. Experimental Procedure ANNEALING, GRAIN SIZE AND HARDNESS DETERMINATIONS Specimens of each type of copper were finally annealed in air for periods of one hour at temperatures ranging from 300 to 1600°F and were subsequently cooled in air. The average grain diameter of the annealed specimen was estimated by comparing with a standard grain size chart. Hardness was determined on the Rockwell 15 T scale. CUPPING TESTS Cups were made in a blanking and drawing set, in which blanks of 2-in. diam were drawn to a cup of 1.25-in. diam with an average depth of about 0.75 in. The clearance between the punch and die was about 0.032 in. The ears of the cup were measured with a special fixture which read the height of ears to one-thousandth of an inch on every ten-degree interval along the circumference of the cup. POLE FIGURES The usual transmission diffraction method with unfiltered copper radiation was employed to determine the pole-figures of the specimens cold-rolled or annealed at 900°F. All the pole-figures were derived from the positions of intensity maxima on 111 diffraction rings of the X ray photo-grams taken at 10 rotation of a

Jan 1, 1950

By A. B. C. Stothart

"IntroductionIn the past four years Ontario Mines have made great progress in the development of reliable mine conveyance safety dogs. History shows that, prior to 1945, a rope failure on a descending conveyance was usually followed by unsatisfactory safety dog operation.It is proposed in this paper to discuss briefly;The history of this development. Progressive .development of one type (E. Long).Free fall testing of safety dogs. Anti-trailing rope device.HistoryOn February 2nd., 1945, a hoisting rope broke at the Paymaster mine. The safety dogs proved ineffective and permitted the cage to crash through to the bottom of the shaft, with disastrous results.Following this accident, the Ontario Government appointed a Technical Committee to investigate The Safety of Hoisting Equipment and Hoisting Practice in Ontario Mines. This Committee consisted of Dr. C. R. Young (Chairman) and Professors T. R. Loudon, E. A. Allcut, V. G. Smith, and Lloyd M. Pidgeon. The Ontario Mining Association (at the request of the Ontario Government) also appointed a Committee consisting of Messrs. R. E. Dye, R. D. Parker, and R. L. Healy."

Jan 1, 1950

By E. B. Spice

"IntroductionExperience at Eldorado has been gained with a variety of mill feeds. That is, the actual mode of occurrence of the pitchblende varies to such an extent that gravity concentration has at times been decidedly unattractive and at others highly successful. Because of this, the type of ore necessary for economic recovery of pitchblende have become well known. Also, those types which do not lend themselves to gravity concentration have become familiar. In general, it may be said that when the pitchblende occurs free, to a large extent, from sulphides or arsenides of high specific gravity, and when the bulk of the pitchblende is localized rather than disseminated, excellent recoveries, high ratios of concentration, and good grades of concentrates , will result from proper gravity concentration. On the other hand, if the ore contains pitchblende intimately or otherwise mixed with large quantities of arsenides or sulphides, or if the pitchblende is finely disseminated, gravity concentration alone will yield high recoveries only at the expense of very low-grade concentrates and low ratios of concentration.According to Taggart, the ease with which gravity separation may be effected is determined by the value of the following ratio: []It is stated that when the ratio is above 1.75, gravity separation can be expected to be easy. Thus with pure U308 (S.G. over 9) and gangue consisting chiefly of siliceous material, the process can he expected to be entirely successful even when carried on at very fine sizes. Unfortunately, the term pitchblende covers a great diversity of material. Although some of the mineral occurs pure enough to have a specific gravity of about 7, the average is much lower than this and is probably well under 5. Some of the pitchblende particles have a specific gravity so low that gravity concentration fails completely. It may also be said that no degree of grinding will produce free pitchblende of sufficient density from such particles. This, then, represents the most important limitation to the application of gravity separation to pitchblende ores."

Jan 1, 1950

By R. W. Stahl, H. E. Sanford, J. B. Benson

"As movement or transportation of coal is second only in importance to actual mining, and as the latter is wholly dependent on the former, the efficiency of the haulage system often determines the efficiency of the entire plant. Safe or accident-free transportation is essential, also, because no attempt is made to divorce safety and efficiency, it being realized that they are inseparable.Most of the Ruhr mines were opened decades ago, and many features that were considered up-to-date then are now obsolete. Progress in developing haulage systems has been slow, chiefly because it would be very costly to change many to accomodate more modern equipment. These systems are being altered or modernized slowly, as capital permits. The only revolutionary changes that have been made in mine haulage are the use of larger locomo¬tives and cars, in the movement of shuttle cars, and in the adoption of skip hoisting. Most Ruhr shafts, however, will not accomodate large cars, the secondary roadways are too narrow for large locomotives, and the coal is too soft to permit skip hoisting because degradation would be too great. Therefore, notable advances have been made only at newer plants, where allowances have been made for modern equipment.The shafts are deep (average 2,300 feet) and of comparatively small diameter. The small bore was designed to reduce costs of sinking and lining. Generally, the shafts are lined throughout with masonry; however, most of the buntons and guides are wooden.The hoisting engines arc installed in masonry structures, and where more than one hoist is used they are placed on opposite sides of the shaft to eliminate as much noise as possible (fig. 1). This arrangement also prevents an engineer from acting on any signals except those transmitted to him. A few hoisting engines are mounted in the headframes (fig. 2). The hoisting engines are equipped with overwind, overspeed, and auxiliary braking devices, which are tested frequently and kept operative at all times. The signal code is displayed in view of the engineer, and a light or dial indicator informs him of the cage's position in the shaft. Many hoists are equipped with re¬cording devices, which show on a continuous chart the speed of hoisting as well as any delays throughout the 24-hour period.A Koepe system of winding is in general use because of the depth of shafts (fig. 3). With this system, a single rope passes over a single-groove, large-diameter sheave attached to the motive power, and the ends of the rope, after passing over the head sheaves, are attached to the cages. A flat tail rope suspended in the shaft and connected to the bottom of the cages serves to balance the system. Both steam and electric engines are used to activate the hoisting system, end the engines are capable of moving at least a 50-percent overload.The ropes are necessarily of large diameter (2-1/2 inches and greater), and flexibility decreases as rope diameter increases. To overcome the unwieldiness .of large-diameter ropes as shafts become deeper, a four-rope system is being tried at one mine (fig.4). In this system the load is equalized between four ropes which permits much smaller ropes to be used. All hoist ropes are inspected daily and discarded after 2 years service unless special permission for longer service is granted by the Mine Inspectorate, and then only after thorough visual and electromagnetic inspection."

Jan 1, 1950