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By F. M. Lewis, J. F. Myers

The crushing of ore, as defined by Taggart, is "usually a stage process, utilizing . . . machines especially suitable for the reduction of particular sizes. . . . down to a final stage, employed for the reduction of the finest sizes. ... In contradistinction, comminution to 20-mesh or finer is called grinding. There is a twilight zone in which the product is 6, 10, or 14-mesh limiting size which is either crushing or grinding. ..." Since the data in this paper pertain to a rod-mill operation, comminuting a particular ore size of ¾ in. to a final stage, coarser than 20-mesh, the rod mill is designated as a "fine crusher machine" in this paper. The published literature discloses a wealth of data regarding a rod mill as a fine-grinding machine but relatively little as a fine-crushing operation. Typical of the latter operation is: (I) a description by F. J. Tuck2 in which a fine-crushing rod mill replaced crushing rolls and (2) a description by the present authors3 in which a fine-crushing rod mill was installed instead of a collventional fine-crushing machine. The experience of the authors is that a definite distinction must he made between the rod mill functioning as a fine crusher and as a fine-grinding machine. As a fine crushing machine it prepares relatively large amounts of ore for a subsequent fine grinding operation; as a grinding machine, it grinds relatively small amounts of ore to some predetermined finished product. Pulp Movement and Action in Mill In considering the work performed by a fine-crushing rod mill, it seems important to understand just what takes place within the mill. We were inspired to study this matter by a paper by E. W. Davis,4 which indicated that definite water classification of ore particles took place in a ball mill. However, in thinking of the rod-mill action from this point of view, we were confronted with the fact that our rod-mill discharge product showed no evidence of action by water classification. The Tennessee Copper Co. ore is relatively high in sulphides, and the gangue rock and the sulphides are of about the same hardness. In the presence of any action by classification, the coarser sizes are definitely very lean in sulphide minerals.

Jan 1, 1947

By J. F. Myers, F. M. Lewis

The crushing of ore, as defined by Taggart, is "usually a stage process, utilizing . . . machines especially suitable for the reduction of particular sizes. . . . down to a final stage, employed for the reduction of the finest sizes. ... In contradistinction, comminution to 20-mesh or finer is called grinding. There is a twilight zone in which the product is 6, 10, or 14-mesh limiting size which is either crushing or grinding. ..." Since the data in this paper pertain to a rod-mill operation, comminuting a particular ore size of ¾ in. to a final stage, coarser than 20-mesh, the rod mill is designated as a "fine crusher machine" in this paper. The published literature discloses a wealth of data regarding a rod mill as a fine-grinding machine but relatively little as a fine-crushing operation. Typical of the latter operation is: (I) a description by F. J. Tuck2 in which a fine-crushing rod mill replaced crushing rolls and (2) a description by the present authors3 in which a fine-crushing rod mill was installed instead of a collventional fine-crushing machine. The experience of the authors is that a definite distinction must he made between the rod mill functioning as a fine crusher and as a fine-grinding machine. As a fine crushing machine it prepares relatively large amounts of ore for a subsequent fine grinding operation; as a grinding machine, it grinds relatively small amounts of ore to some predetermined finished product. Pulp Movement and Action in Mill In considering the work performed by a fine-crushing rod mill, it seems important to understand just what takes place within the mill. We were inspired to study this matter by a paper by E. W. Davis,4 which indicated that definite water classification of ore particles took place in a ball mill. However, in thinking of the rod-mill action from this point of view, we were confronted with the fact that our rod-mill discharge product showed no evidence of action by water classification. The Tennessee Copper Co. ore is relatively high in sulphides, and the gangue rock and the sulphides are of about the same hardness. In the presence of any action by classification, the coarser sizes are definitely very lean in sulphide minerals.

Jan 1, 1947

By Leslie Burris, J. C. Hesson

The diffusivity of uranium in liquid cadmium has been measured as a function of temperature by the capillary -bath technique. The diffusivity measurements were made at 4509 500°, 575º, and 650°C, over which temperature range the diffusivity of uranium varied from 1.3 to 2.3 x 10-5 sq cm per sec. The diffusivity of uranium in cadmium was less than that predicted by the Einstein-Stokes equation. VERY little information is available on the diffusivity of uranium in liquid metals. Bonillal has reported a single measurement of the diffusivity of uranium in bismuth. Measurement of the diffusivity of uranium in cadmium was undertaken, not only to supply data in an area in which a dearth of information exists, but also because of a need for this physical property measurement in connection with mass transfer studies now underway in the U-Cd system at Argonne National Laboratory. The capillary-bath method used in this work has been successfully used by other workers2"4 to obtain diffusivities in liquid metal systems. Although considerable experimental technique is required to obtain reliable diffusivity measurements, the basic information required is only that of temperature, capillary tube length, the initial and final compositions of the alloy in the capillary tube, and the composition of the bath into which the solute element is diffusing. From these measurements, diffusivities may be calculated using Fick's law. The diffusivity equations are presented in the Appendix. Equations which are often used to predict diffusivities in dilute solutions5 are the Einstein-Stokes equation: D= (kT)/(6pµr) and the Eyring equation: D = (?,kT/?y?zp), where D= diffusivity (sq cm per sec), k = Boltzman's constant, T= temperature ("K), p = viscosity (poises), r = radius of diffusing molecule or species (cm), and A,, Ay, and A, are the diameters of the diffusing molecule in the three axial directions. For round molecules, the Eyring equation gives values which are 37r times larger than the Einstein-Stokes equation. In this work, the experimental values were compared to the Einstein-Stokes equation. Experimental values of diffusivities in liquid metal systems usually lie between those predicted by the Einstein-Stokes and Eyring equations. The viscosity values for cadmium were obtained from the Liquid Metals Handbook.6 EXPERIMENTAL APPARATUS AND PROCEDURE The general procedure employed consisted of rapidly immersing in a molten cadmium bath containing a negligible amount of uranium a capillary

Jan 1, 1963

By R. E. Grace

WHEN an Fe-C alloy, partially reacted to pearl-ite, is quenched rapidly enough to suddenly stop the growth process, it may be expected that any carbon concentration gradients will freeze in situ in the newly formed martensite. A short tempering treatment should precipitate finely dispersed iron carbide particles in approximately the same distribution in which carbon existed during the growth process. Examination with the electron microscope of areas of tempered martensite adjacent to the pearlite-martensite interface should reveal the existence of preferred areas of precipitation. Fig. 1 is an example of an electron micrograph used in this study. Small sections of a commercial steel (0.9 pct C, 1.6 pct Mn, 0.25 pct Si) were austenitized at 945°C in air for 45 min, furnace-cooled to 697°C, iso-thermally reacted to pearlite for 30 min, and quenched into iced brine. Two samples were tempered 5 and 20 min, respectively, in air at 498°C and were subsequently quenched in water. The samples were mechanically polished and etched in 1 pct picral. Collodion replicas were made by the conventional technique; these were shadowed with chromium at about 650' from the horizontal, and electron micrographs were taken at a linear magnification of 14,200X. Prints were made at a magnification of 44,100X. Lineal analysis of the prints was made as a function of the interlamellar spacing and volume percent of iron carbide. Straight lines were drawn into tempered martensite between adjacent lamellae at a distance S/4 or S/6 apart, where S is defined as the interlamellar spacing measured at the pearlite-martensite interface. The volume percent of iron carbide was measured along each of these traces with a "gage length" of S/10. The results of analyses of two samples are shown in Table I. The area included between the S/2 and S distance from the interface could be considered homogeneous within the sensitivity of the measuring technique. This value was used as a proportionality factor to evaluate an arbitrary carbon content of homogeneous austenite, and also as a parameter by which error in measurements could be evaluated. Results of the lineal analysis are shown in Figs. 2 and 3. Fig. 4 shows the statistical nature of the interlamellar spacing for all lamellae measured in this

Jan 1, 1954

By A. Goldberg, T. E. Tietz, J. E. Dorn

Introduction The concept that the flow stress for plastic deformation of metals in the work hardening range is a function of the instantaneous values of the strain, strain rate and test temperature was promulgated by Ludwick' many years ago. This thought has been more or less a guiding principle throughout the immediate period of development and growth of mechanical metallurgy. It has been tacitly assumed or knowingly applied to many investigations on the plastic behavior of metals. Spurred by the theoretical interest and practical importance of the plastic behavior of metals, many investigators have attempted to determine empirical, semi-empirical, and theoretical equations2-l8 relating the flow stress with the strain, strain rate, and temperature. Although the applications of such equations have been extensive, the experimental verification of the existence of a mechanical equation of state is weak. Furthermore, some of the available evidence on plastic flow appears to suggest that no simple functional relation between stress, strain-rate, strain and temperature can exist. The mechanical equation of state demands that Where a = true stress = loadlinstantaneous area e = true strain _ Instantaneous gauge length ~ initial gauge length i = true strain rate T = temperature This formulation implies that da is an exact differential. Therefore the flow stress (a) is a function of the instantaneous values of the strain (c) the strain rate (i) and the temperature (T), independent of the previous mechanical or thermal history. Having reached a certain strain, strain rate, and temperature, by any path whatsoever, a definite and fixed value of the flow stress should be obtained. One check on the validity of the mechanical equation of state is easily obtained. Consider the following three tests on the stress-strain curve at constant temperature: A. Strain at rate :I B. Strain at rate it C. Strain at rate 11 to el and continue test at &. A graphical representation of possible results of such tests are shown in Fig I. If, for strains greater than el, test (C) data coincide exactly with the data from test (B), as shown in Fig ra, a limited verification of the mechanical equation of state is obtained for the range of strain rates from il to iz. But if the data from test (C) differ from those of (g) beyond strain t\ where the strain rate is changed from il to k2, the mechanical equation of state is not

Jan 1, 1949

By John R. Gould

A mathematical treatment of multiple-pass floating-zone refining of semiconductors, including the evaporation of solutes from the melt at reduced pressure, is presented. Vapor pressures and evaporation rate constants for phosphorus, arsenic, and antimony in germanium have been determined from floating-zone experiments conducted at 10 mm Hg. Rate sensi-tivity of arsenic in germanium has been examined over the range of 1 to 10 in. per hr growth rate. The equilibrium value of the segregation coefficient of arsenic in germanium, as determined by an extrapolation of these data, was found to be 0.09. PFANN has treated the segregation of nonvolatile solutes in the zone-melting process, Bradshaw and Mlavsky2 considered the evaporation of solutes in a vacuum during the growth of crystals by pulling from the melt, and van den Boomgaard3 has shown that it is possible to bring a volatile impurity element homogeneously into an ingot by means of zone-melting under a constant pressure of that element. This paper is concerned with the volatilization of solutes during zone - melting in a vacuum. The mathematical analysis by Lord4 of multiple-pass zone melting is extended to include evaporation of solutes. A method is presented in which evaluations of vapor pressures, evaporation rate constants, and distribution coefficients may be derived from the variation of solute concentration in the floating-zone refining of germanium at 105 mm Hg. EXPERIMENTAL Ingots of zone refined germanium, approximately 1/4 in. in diam and 10 to 12 in. in length, were subjected to several floating-zone passes at 105 mm Hg in order to provide intrinsic material for these studies. These ingots were then seeded in a [Ill] direction and solid-doped to a constant impurity concentration of approximately 1015 atoms per cmS in a forming gas ambient (10 pct H - 90 pct A) at 1/2 atm gage. The resulting crystals were subjected to repeated floating-zone passes at 105 mm Hg and a concentration profile along the length made after termination of each pass. Concentrations were determined from four-point probe resistivity measurements, and proper precautions were taken to prevent contamination during handling. A constant zone length, crystal cross section, and growth rate were maintained during zoning. A stationary coil and moving crystal system were employed for floating-zone growth with neither the section of crystal above nor below the melt rotated.

Jan 1, 1962

By C. W. Lawr

The ores mined by the Santa Gertrudis Co. at Pachuca, Mexico, arc mainly silver-bearing; they also yield some gold and carry a little copper. Strong cyanide solutions are used to dissolve the silver and gold, and the mechanical loss of cyanide at the tailings filter has been considerable. Prior to the year 1923, the pulp, after agitation, was given a 35 to 37-min. wash with barren solution in a Butters-type leaf filter; the discharged slime cake was repulped and pumped into eight Merrill self-sluicing slime presses in which the slime was given a further wash with barren solution for 25 to 35 min., followed by a 10-min. water wash. At that time the barren solution titrated around 0.350 per cent. free KCN and 0.500 per cent. total KCN. (Aero brand cyanide is used.) It can readily be seen that a 10-min. water wash would leave considerable cyanide in the filter cake, which of course was a total loss. This averaged 0.522 kg. (1.15 Ib.) KCN per ton of solution. As the daily discharge to the dam was 1600 tons' of slime of a 2:1 ratio of liquid to solid, the value of this mechanical loss of cyanide was an important item and it's recovery held alluring possibilities. Principle of Cyanide-recovery Process As it had acquired the patent rights to theLayng (H. R.) and Halvor-sen (A. L.) processes, also those of Mills (L. D.) and Crowe (T. B.) after considerable experimental work in its own laboratory and at Tonopah, Nev., The Merrill Company of San Francisco early in 1923 notified the Mexican Corporation and compania de Santa Gertrudis of its process for the recovery of cyanide from weak wash solutions, or for the transference of cyanide from a foul mill solution to a clean alkaline solution. This method is known as the Mills-Crowe process, the principle of which is as follows: The solution from which the cyanidc is to be removed is made acid by bringing it into contact with sulfur dioxide (SO2). The acidified solution is the; transferred to a closed tank in which air and solution are brought into intimate contact. The air leaving the tank charged with HCN is then passed to another tank in which it is mixed with an alkaline solution,

Jan 1, 1930

By Lewis A. Riley

I desire to submit, for the consideration and information of the members of the Institute, the following data, drawings, and tables, showing what I believe will be interesting information with regard to the diamond drill, and its uses in developing mineral lands. I would first say that the results given have been obtained during the past two years by means of two drilling machines belonging to the Lehigh Valley Coal Company, operating on their coal lands in the Mahanoy, Lehigh, and Wyoming regions. 1st. A tabulated statement, giving location of borings, date, progress, total cost, itemized cost per foot, etc. 2d. Section showing strata bored in Nos. 5 and 10 holes at Mt. Carmel, see Plate VI. 3d. Section showing strata bored in No. 19 hole at Hazleton, see Plate VI. 4th. Photograph and description of drill No. 2. From the first it will be seen that the total number of test-holes bored up to May 1st of this year is 24; 9 of these are in the Mahanoy region, 6 in the Lehigh region, and 10 in the Wyoming region, in all of which cores have been taken for the whole distance. The total distance bored is 9902 feet. The majority of the holes have been put down for the purpose of proving the lower veins of the coal series, and have had to encounter the harder rocks of the coal formation. Much of the distance has been through the lower conglomerates, going in some cases through the coarse egg conglomerate, the foundation of the coal deposit, and to the green sandstone and red shale which underlie it. These lower sandrocks and conglomerates are among the hardest known to geologists, and present a formidable resistance to the exploring drill. This fact must be taken into consideratiotl in comparisons of cost of drilling, and when considered will still further improve the very favorable results I am able to show in my statement of costs. In illustration, I will cite the work of one bit when boring in the upper or softer rocks of the Wyoming region, where 280 feet were obtained without resetting, while in the Lehigh and Mahanoy regions, in the lower and harder rocks, bits set with the same quality of diamonds would only bore from 3 to 20 feet before being worn out. In this connection it may

By H. E. Howe, A. A. Smith

The creep properties of lead alloys have been studied by a number of investigators but most of the work has been done on extruded material and the amount of data available on rolled alloys is surprisingly limited. Greenwood1 and his collaborators investigated various rolled commercial leads containing relatively small amounts of impurities, Phillips2 likewise studied rolled commercial leads, although he did include a few alloys— such as tellurium lead and one per cent antimonial lead. But as far as the authors are aware no data on the creep properties of so-called "hard leads"-—i.e., containing from 4 to 8 per cent antimony—have been published. In this paper some data are presented on a few antimonial lead alloys and compared with a copper-bearing lead. Preparation OF Specimens All the alloys used in this investigation were cast at a temperature of 400°C. into an open cast-iron mold preheated to Ioo°C., the resultant casting being % by I by 10 in. The copper-bearing alloy was cold-rolled by 10 per cent passes to a thickness of 0.100 in. Creep specimens 16 in. long were prepared by clamping the strips between wooden blocks and machining the edges of the strips parallel by means of a plane. Careful manipulation gave parallel specimens with no reduced section and with a minimum of distortion. The antimonial lead castings were hot-rolled in 10 per cent passes at a temperature of I50°C. to a thickness of 0.400 in., the heating being carried out in an electric oven after each pass through the rolls. After the castings had been broken down to 0.400 in., they were allowed to cool to room temperature and then cold-rolled to 0.100 in. with the usual Io per cent passes. Sufficient time was allowed after each pass for the strips to cool to room temperature. The apparatus and method of measuring strain have been adequately described in a previous paper3 and essentially the same practice was used in obtaining the present data. Time-extension curves for only two alloys are given, since these are typical of all. Figs. I and 2 show the time-extension Curves for the copper-bearing alloy A at 30°C. and loo°C., and Figs. 3 and 4 show

Jan 1, 1945

By Louis A. Riley

I DESIRE to submit, for the consideration and information of the members of the Institute, the following data, drawings, and tables, showing what I believe will be interesting information with regard to the diamond drill, and its uses in developing mineral lands. I would first say that the results given have been obtained during the past two years by means of two drilling machines belonging to the Lehigh Valley Coal Company, operating on their coal lands in the Mahanoy, Lehigh, and Wyoming regions. 1st. A tabulated statement, giving location of borings, date, progress, total cost, itemized cost per foot, etc. 2d. Section showing strata bored in Nos. 5 and 10 holes at Mt. Carmel, see Plate VI. 3d. Section showing strata bored in No. 19 hole at Hazleton, see Plate VI. 4th. Photograph and description of drill No. 2. From the first it will be seen that the total number of test-holes bored up to May 1st of this year is 24; 9 of these are in the Mahanoy region, 6 in the Lehigh region, and 10 in the Wyoming region, in all of which cores have been taken for the whole distance. The total distance bored is 9902 feet. The majority of the holes have been put down for the purpose of proving the lower veins of the coal series, and have had to encounter the harder rocks of the coal formation. Much of the distance has been through the lower conglomerates, going in some cases through the coarse egg conglomerate, the foundation of the coal deposit, and to the green sandstone and red shale which underlie it. These lower sandrocks and conglomer¬ates are among the hardest known to geologists, and present a formidable resistance to the exploring drill. This fact must be taken into consideration in comparisons of cost of drilling, and when considered will still further improve the very favorable results I am able to show in my statement of costs. In illustration, I will cite the work of one bit when boring in the upper or softer rocks of the Wyoming region, where 280 feet were obtained without resetting, while in the Lehigh and Mahanoy regions, in the lower and harder rocks, bits set with the same quality of diamonds would only bore from 3 to 20 feet before being worn out. In this connection it may

Jan 1, 1877

By W. G. Woolf, E. R. Crutcher

The electrolytic zinc plant of the Sullivan Mining Co. is in Government Gulch, at Silver King, Shoshone County, Idaho, in the Coeur d'Alene mining district, about one mile from the Bunker Hill smelter and about three miles from Kellogg. The Sullivan Mining Co., which owns both the zinc plant and the Star mine, is itself owned jointly by the Bunker Hill & Sullivan Mining & Concentrating Co, and the Hecla Mining Company. The Tainton high-density, high-acid process is used at this plant, therefore some parts of its flowsheet differ markedly from those of electrolytic plants using the low-density process. While it is a custom plant and has treated a variety of zinc concentrates originating from various sources in a number of states, it has obtained its main supply of zinc concentrates chiefly from the zinc-bearing ores of the Coeur d'A1ene mining district, of which the principal important sources have been the Bunker Hill and Star mines. Thus the main feed has been zinc concentrates obtained from complex lead-zinc-silver ores. For the year of 1935 the concentrates had the following composition: water, 9.5 per cent; Zn, 50.6; Pb, 3.3; Fe, 9.4; S, 31.0; Cu, 0.2; insoluble, 2.9. Flowsheet The concentrates are delivered to the plant in railroad cars, which discharge into storage bins directly under the track. Bottom-dump cars are dropped directly into the bins while a slip scraper, pulled by an air tugger hoist, is used to empty box cars. There are five storage bins for concentrates, each of approximately 300 tons capacity. At the bottom of each bin there are individual belt feeders that discharge onto a common conveyor belt, which in turn discharges to another belt that carries the concentrate onto a weightometer belt for weighing. Immediately after the weighing, the crushed oversize return from the roasters is added to the load on the belt and then the ground-up sweepings, lumps and roaster cleanings. An electromagnet at the end of this conveyor removes tramp iron from the feed. The belt discharges to a set of 30 by 12-in. rolls, which break up lumps or frozen chunks of concentrate, etc. Beneath the rolls there is a screw conveyor, which conveys the roll discharge to a

Jan 1, 1937

By Warren L. Howes

LITERALLY hundreds of designs of mill liners are in use in current grinding operations, varying in contour from smooth to the roughest of surfaces, and in materials from scrap rail to alloy steels. A certain amount of variation in materials can be expected, owing to differences in local market conditions and the character of ores. However, when one attempts to rationalize the designs in current operations of a parallel nature, there is little that can be called standard practice. Designs and opinions of designs are as divergent as day and night. Somewhere in this maze there must be clues to better practice. The purpose of this paper is to examine the factors for consideration in liner design, and to propose a vigorous campaign of exploration in this phase of grinding operations. The paramount considerations in the choice of mill liners for a grinding operation are: (I) durability, in view of the cost, and (2) developed grinding capacity. Because of the strategic nature of grinding circuits in controlling the production of most mining operations,1 grinding capacity usually is the greater consideration. GRINDING CAPACITY Mill liners affect grinding capacity in two ways: (I) by establishing dead space within a mill shell, which could be usefully occupied by live grinding media, and (2) by controlling the grinding action of the media. The extent to which liner thickness affects the potential volume of live grinding media, and thereby the grinding capacity, is shown in Table I. In a ball mill 7 ft. in diameter, for example, an inch of liner thickness reduces ball volume and corresponding grinding capacity by 5 per cent. With grinding capacity at a premium, this makes consideration something more than first cost of liner materials. A low-cost, voluminous material, made extra thick to secure adequate life, can be very expensive in terms of grinding capacity. On the other hand, in increased grinding capacity, a long-wearing material in a thinner pattern can save many times its value. [ ] From the standpoint of mechanics, the function of mill liners is to provide the means of driving or tumbling the load of grinding media. In this respect, an internal-spur gear-and-pinion assembly resembles a rotating mill shell with its contained load of grinding media, the mill-shell liners

Jan 1, 1945

By Lester Guttman

THE connection between short-range order and thermodynamic activities in binary solid solutions has been pointed out by Birchenall1 who calculated approximate values of the energy of interaction of Zn and Cu and of the degree of short-range order in alpha-brass at one composition and two temperatures. The purpose of this note is to call attention to a somewhat more accurate theory than that used by Birchenall and to extend the calculations to a wider range of temperatures and compositions as a means of testing its applicability. This is an example of the well-known general method of comparing measured thermo-dynamic functions of a system with those calculated from a microscopic model by means of statistical mechanics. In this case, mathematical approximations must be used which make difficult any decision about the validity of the model. Birchenall's statistical method resembles closely that called by Fowler and Guggenheim2 the "quasi-chemical" method which has been shown to lead in certain cases to results identical with those obtained by the method of Bethe.3 The quasi-chemical method has been extended by Takagi4 to give general formulae for the free energy of binary solid solutions based on several lattices at all compositions and temperatures. Specifically, he treats the four systems typified by: I. ß-brass; 2. a- brass; 3. AuCu3; 4. AuCu. When the temperature is so high that no long-range order exists, these formulae reduce to the same form for all lattice types, namely: [ ] Here the symbols have the following meanings: [F-Helmholtz free energy E-energy S-entropy k-the gas constant per molecule z-coordination number, (e.g. 12 in. a-brass). N-total number of atoms c-atomic fraction of A] VAA, VBB, VAB-energy per bond of the corresponding bonds log means log, throughout The superscript 0 denotes the thermo- dynamic function in the standard state, which has been chosen to be pure A or B with the same crystal structure as the solid solution. The quantity x is the ratio of the number of A-A bonds to the total number and for minimum F satisfies the equilibrium condition [ ] The free energy may also be written in terms of µA and µB, the chemical potentials or partial free energies per atom:

Jan 1, 1948

By W. T. Hicks

Visual observation of the oxidation of columbium shows that the protective behavior noted previously in gravimetric work in the early stages of the reaction below 600°C and throughout the reaction at 640°C is associated with an adherent oxide. The fact that the oxide is seen to move outward from the metal at temperatures from 550° to 935°C implies that the reaction takes place at the oxide-metal interface throughout this temperature range. THE work described in this paper further clarifies the complex oxidation kinetics of Cb revealed, for example, by the continuous gravimetric studies of T. L. Kolski.1 Time-lapse motion picture studies reveal details of the reaction not provided by a study of the combustion products. EXPERIMENTAL PROCEDURE Cb specimens were prepared from material with the following typical analysis: 0, 0.03 pct; N, 0.002 pct; C, 0.006 pct; Ta, 0.04 pct; Fe, 0.01 pct; Ni <0.003 pct; and Cr <0.003 pct. Arc-cast ingots were swaged and cold rolled into slabs. These slabs were machined into blocks 7/16 by 7/16 by 7/16 in; 1/16-in. holes were drilled through the large face of the blocks for suspension. The samples were then de-greased and annealed 1 hr at 1200°C in vacuum. The sample was suspended by a platinum and Pt 10 pct Rh thermocouple inside a 1-in. quartz tube in a Kanthal wound furnace (2 in. OD by 4 in. long) equipped with windows 1/2 in. sq. A Cine-Kodak Special II 16 mm motion picture camera was trained on the sample through a Spencer microscope adjusted so that a magnification of X1.5 resulted on the film. Super Anscochrome film was used to record the events in color. Argon was flushed through the furnace while the sample came to temperature. The argon was then replaced by a stream of dry oxygen at 1 atm. Pictures were taken at the rate of eight per second for about 20 sec during the change of atmosphere. Then by use of a Stevens time-lapse device pictures were taken at the rate of four per minute starting within 1 min of the initial oxygen flow. In this manner continuous visual records of the oxidation of pure Cb

Jan 1, 1962

By J. Kittl, D. Arias

THE decomposition on cooling of the high-temperature ß bcc phases in copper- or silver-based binary systems usually takes place by a martensitic. massive, bainitic, or pearlitic reaction depending upon the particular conditions of each system. The AgAl system is one of these systems in which a massive or a martensitic transformation can occur upon cooling from the 3 phase field. Recently Hawbolt and Brown1 observed both massive and martensitic transformation products in a 22.7 at. pct A1 alloy. In the present work the massive and the martensitic transformations were investigated in a number of alloys for the composition range between 23.2 and 27.9 at. pct Al. The alloys were prepared using high-purity silver, 99.999 pct, supplied by ASARCO and aluminum, 99.999 pct, supplied by ALCOA. The metals were melted using a graphite or a sintered alumina crucible which was sealed in quartz tubes under an argon atmosphere. The alloys Were remelted to reduce segregation and homogenized in the 13 phase field for 24 hr. The weight of the alloys after homo-genization compared with the initial weight indicating the alloy composition to be within k0.1 pct of the in- tended composition. Chemical analysis of several samples confirmed the estimated alloy composition. All metallographic observations were performed on samples electropolished with ac current using a 10 pct solution of KCN and a stainless-steel electrode. The samples were cut into -1-mm-thick slices, polished, heat-treated in the ß phase field, and quenched in air, ice water, or iced brine. Some heat treatments consisted of a 2-min anneal carried out in air to allow maximum quenching speeds. Other samples were heat-treated in a vacuum of 10"4 torr and quenched using a helium blast in a quenching apparatus described elsewhere.&apos; In both cases, no appreciable losses of silver or aluminum were noted at the sample surface by sectioning and electropolishing. Table I summarizes the microstructure resulting from the various quenching treatments. The composition of the alloys investigated is also indicated on the equilibrium diagram in Fig. 1. The <, hcp phase, typical of a massive ß — cm, transformation, was observed in all alloys. A minimum temperature of 450°C was estimated for this transformation based on preliminary cooling experiments. the results being included in Fig. 1. In these

Jan 1, 1970

By W. Edeburn, H. R. Fife

An extended study of dustproofing Freeport-seam coal has emphasized several fundamental features in the successful treating of bituminous coals. The extent of the exposed surface area and its absorption propcrties for the dust-retaining film were found to be the most critical factors affecting the efficiency of fly-dust prevention. From recently published screen data and specific surface-area measurements on United Kingdom coals, an approximation of the surface exposed by mine-run coal and by 1-in. slack was made. The total surface area exposed by an average tori of this mine-run coal was estimated at 3768 sq. yd., while a ton of 1-in. slack represents a total surface area of 8779 sq. yd. with no additions for porosity. It is important in coal treating to consider that in mine-run coal the fraction above 1/4 in. in size represents 79.5 per cent of the total weight, but only 4 per cent of the total surface, while the fraction below 150 mesh contains 47 per cent of the total surface and only 1.2 per cent of the total weight. Also, one should appreciate that, in a 1-in. slack, the domestic stoker fraction (1 in. to 8 mesh) contains 76.3 per cent of the weight and 8.2 per cent of the surface, while the fraction below 150 mesh represents 3.3 per cent by weight and 54.8 per cent by surface area. Wax extraction data on 1-in. to 0 slack from mine-run treated Freeport-searn coal showed that the coating was distributed uniformly as a function of exposed surface; hence, nonuniformly on a weight basis. This distribution of the coating on mine-run coal therefore limits the extent of treating that is desirable on bituminous coal at this point in the preparation. The treatment of Freeport mine-run coal before it reached the tipple gave a 70 to 80 per cent reduction of fly-dust in the tipple, a satisfactory treated commercial slack coal, and an indicated change in the trend of tipple degradation. The premium-treated domestic sizes required a supplemental treatment for complete dustproofness. The actual cost of the pretipple and supplemental treatments was very low per ton of coal treated, while the net or true cost was sufficiently

Jan 1, 1938

By A. Bar-Or, G. Hirsch, P. Wynblatt

The rates of nucleation and growth at various temperatures in U-0.45, 1.1, and 1.35 at. pet Cr alloys were determined, providing a qliantitative proo-f that the transformation under study is controlled by nucleation and diffusional growth. The linear relationship, log NV US 1/T2, suggests that the nucleation of the alloys is probably controlled by the classical law of nuclei formation. The established linear relationship, log G us 1/T, in the three alloys suggests that growth in each alloy is controlled by bidimensional nuclei formation. It was established that both the rate of growth and the rale of nucleation are retarded by chromium. Using classical formulae, sApN and sApG (the ß/a apparent interphase energies) were calculated from the nucleation and growth results, respectively. It was found that their values are of the same order of magnitude. However, why sApN is unaffected by the addition of chromium, while sApG increases with it, needs further clarification. In the last few years, an appreciable amount of research has been conducted into the mechanism of the ß —a transformation in uranium and uranium alloys. Recently, Burke1 reviewed the published work concerned with this transformation. He concluded that in dilute uranium alloys it is firmly established that, at high temperatures in the a range, ß decomposes by a diffusional process. However, he based his conclusion either on experiments dealing with the over-all transformation,2 or on qualitative results.3 It was felt that a better insight into the mechanism of transformation could be gained from the quantitative measurements of growth and nucleation rates, which we aimed to carry out in the present work. In a previous work,4 we have shown that a rim of a grains is formed during isothermal transformation at the high a range. The thickening of this rim with time was used for measuring the rate of growth in U-0.45 at. pet Cr.5 This method was also used in the present work for two more alloys. For meas- uring the rate of nucleation, a method of quenching and macroetching was developed to allow differentiation between a grains formed during isothermal transformation at the upper a range and those formed after quenching, Fig. 1. EXPERIMENTAL WORK AND RESULTS The Rate of Nucleation. The measurement of the rate of nucleation was effected by counting the num-

Jan 1, 1965

By W. Edeburn, H. R. Fife

An extended study of dustproofing Freeport-seam coal has emphasized several fundamental features in the successful treating of bituminous coals. The extent of the exposed surface area and its absorption propcrties for the dust-retaining film were found to be the most critical factors affecting the efficiency of fly-dust prevention. From recently published screen data and specific surface-area measurements on United Kingdom coals, an approximation of the surface exposed by mine-run coal and by 1-in. slack was made. The total surface area exposed by an average tori of this mine-run coal was estimated at 3768 sq. yd., while a ton of 1-in. slack represents a total surface area of 8779 sq. yd. with no additions for porosity. It is important in coal treating to consider that in mine-run coal the fraction above 1/4 in. in size represents 79.5 per cent of the total weight, but only 4 per cent of the total surface, while the fraction below 150 mesh contains 47 per cent of the total surface and only 1.2 per cent of the total weight. Also, one should appreciate that, in a 1-in. slack, the domestic stoker fraction (1 in. to 8 mesh) contains 76.3 per cent of the weight and 8.2 per cent of the surface, while the fraction below 150 mesh represents 3.3 per cent by weight and 54.8 per cent by surface area. Wax extraction data on 1-in. to 0 slack from mine-run treated Freeport-searn coal showed that the coating was distributed uniformly as a function of exposed surface; hence, nonuniformly on a weight basis. This distribution of the coating on mine-run coal therefore limits the extent of treating that is desirable on bituminous coal at this point in the preparation. The treatment of Freeport mine-run coal before it reached the tipple gave a 70 to 80 per cent reduction of fly-dust in the tipple, a satisfactory treated commercial slack coal, and an indicated change in the trend of tipple degradation. The premium-treated domestic sizes required a supplemental treatment for complete dustproofness. The actual cost of the pretipple and supplemental treatments was very low per ton of coal treated, while the net or true cost was sufficiently

Jan 1, 1938

By B. F. Fackenthal

For the past year we have had in use at the Durham furnace a set of hollow pig-patterns made of iron, which have given such satis factory results that I think a description of them would be of interest to the members of the Institute. Before the introduction of these iron patterns, we used the ordi nary wooden patterns, made in the usual way with straps of iron along the sides to protect the corners from wearing off. The size and shape of these wooden patterns are shown in Fig. 1. When the patterns were new, the pigs weighed about 130 pounds each, and were uniform in size, and the moulding was clean. But constant use soon wore the patterns away; and at the end of a few months, the pigs had become, in consequence, so much lighter that the moulding, daily, of several additional beds was necessary to accommodate the same quantity of iron. As the patterns wear away, the moulding becomes ragged and the pigs rough. Moreover, wooden patterns are often destroyed by fire. I have known an entire set to be destroyed by a break-out of iron. The iron pattern which I have designed is free from all the objections to the wooden pattern, being not only more durable, but, in the long run, far cheaper. The iron pattern having the same shape and size as the wooden one, is made of the best flange-iron, of No. 13 gauge. After the sheets have been cut to the proper size, three heats are required for the flanging. At the first heat, each piece is stamped in a cast-iron form, which gives the proper shape to the bottom part of the pat-

Jan 1, 1889

George W. Maynard, New York, N. Y.:—It is well known that there are many magnetite-mines of which the ore is too low-grade for direct use in the blast-furnace. For carrying out the sintering process, the preliminary step, in the case of magnetites, is the separation of the ore from the gangue by magnetic concentration. The resulting fine-grain product is very objectionable in the blast-furnace, and when used must be a very small percentage of the charge. The sintering process has saved the situation by furnishing a product which is absolutely ideal on account of its cellular structure. (In reply to questions by W. J. Tudor, and Joseph W. Richards, the following additional information was given by Dr. Maynard.) The percentage of iron in the product is determined entirely by the degree of concentration of the raw material and is practically not increased by the sintering operation. The magnetic concentrates at Port Henry and at the Benson and one or two other mines contain from 65 to 67 per cent. of metallic iron. The cost of sintering is from 39 to 41 cents per ton of product, including superintendence, depreciation, and repairs. The itemized figures are about as follows : Mixing (including labor), 24; power, 9; ignition, 5; repairs, 3; total, 41 cents. Prof. J. W. Richards, South Bethlehem, Pa.:—From a metallurgical standpoint I question the accuracy of the statement on p. 374: " C.—An increased reducibility of the iron so treated. (This has been conservatively estimated as 5 per cent. better than the best available ore.)" It seems to me that when a material is&apos; sintered in this way, the constituents present, as lime, silica, and alumina, have a chance to form a compound with the iron oxide; that such a compound, if formed before the material goes into the blast-

Jan 1, 1913