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By D. H. Polonis, M. B. Kasen

AN unusual thermal etching phenomenon has been observed on the surface of a Cu-Si alloy. The observations were made during studies of vacancy condensation pit formation which were reported in previous papers.1,2 Figs. 1 and 2 are photomicrographs of the surface of a Cu-4.8 pct Si alloy after heating in vacuo (3 x X mm Hg) to 525°C and subsequently holding at that temperature for 16 hr in an impure (air-contaminated) argon atmosphere. Under these conditions preferential removal of silicon from the matrix by oxidation results in a surface which is entirely a phase overlying an a plus y structure. The light areas of Fig. 1 are regions in which very little thermal etching has occurred, while the majority of the surface has experienced severe etching. Similar regions in Fig. 2 are seen as plateaus above contours produced by net transfer of material from the specimen surface. Twin boundaries are also visible in Fig. 2. The difference in matrix mi-crostructure between Figs. l and 2 reflects the par- ticular orientation of low-index planes with respect to the surface of observation. Thermal etching contours of this type have been studied in detail by Hondros and Moore Both of the photomicrographs show that one or more dark spots are located approximately at the center of every protected area; where two spots are in close proximity, the protected areas have merged. The microstructures indicate that the anomalous resistance to thermal etching is related to the presence of these spots. Williams and Hayfield have shown that local regions of unusually high oxidation resistance can be caused by a surface monolayer of atoms with a work function higher than that of the base material. The result is an inhibition of the rate of electron removal from the base material and a decrease in the oxidation rate. The source of the protected layer may be either surface diffusion from contaminants on the specimen surface or pipe diffusion of solute atoms followed by migration across the surface. In the latter case the driving force is a difference in chemical potential resulting in a Gibbs adsorption phenomenon. In the present case, chemical polishing of the specimen surface precluded contaminants of a strictly surface nature. Consequently, diffusion of impurity atoms from within the lattice appears to be the most tenable explanation of the observed phenomena. The authors propose that the structures of Figs. 1 and 2 reflect the diffusion of impurity atoms along extended dislocations within the individual a phase grains prior to surface oxidation. The central spots within the protected regions are seen as sites of dislocation emergence which have become locally decorated with oxide in the manner discussed in Ref. 1. Williams and Hayfield observed protected regions adjacent to the grain boundaries of iron-contaminated copper which they attributed to the segregation of iron to the grain boundaries coupled with high dif-fusivity along the complex dislocation array at such boundaries. No evidence of such grain boundary protection was observed in the Cu-Si alloy; this

Jan 1, 1964

By Ronald C. Koo

THE effect of grain size on the mechanical properties of the bcc metals has been a subject of intensive theoretical and experimental study.'-l5 Quantitative relationships have been established experimentally between yield stress and grain size and between the ductile-brittle transition temperature and grain size in a number of metals.6"15 Much significance has therefore been attached theoretically to the grain size dependence. In all these experiments, different grain sizes were always achieved by annealing specimens at various temperatures. The implicit assumption was that any change in the dislocation substructure or in the distribution of impurities which might occur simultaneously with the production of various grain sizes had only a negligible effect on the mechanical properties. The validity of this assumption, however, is questionable. transformation in iron-nickel alloys is actually an extraneous phenomenon due to inadvertent compositional changes in the martensite if the heating rate is not sufficiently rapid. This complication can be avoided by up-quenching for the reversal treatment,5 and then the stabilization does not appear. Clearly, the lattice imperfections and strengthening produced by the reverse transformation are not responsible for the marked stabilization which has been observed by other investigators in these alloys. Data will be presented in this note to show that, at least in recrystallized tungsten of commercial purity, the difference in the ductile-brittle transition temperature of specimens annealed at various temperatures is more a result of a variation in the distribution of impurities rather than of the difference in grain size produced by the anneal. Ductile-brittle transition temperatures in tension were determined on powder-metallurgy tungsten specimens recrystallized at various temperatures. Tensile specimens having a gage length of 1.50 in. and gage diam of 0.18 in. were recrystallized at 1450°, 2000°, 2350°, 2550°, 2750°, or 2800" for 30 min in a vacuum furnace at a pressure of 1 x 10"5 mm of Hg and then tested at a constant extension rate of 8 x 10 per sec in an argon atmosphere from 100" to 800". Typical quantitative analyses of impurities are given in the Table. These results show that, within the precision of the analyses, see Table, annealing did not result in any detectable change in the impurity contents of the specimens, nor was there any detectable difference in purity between specimens having large differences in grain size annealed at the same temperatures . The results of the tensile tests yielded three distinct ductile-brittle transition curves approximately 225' apart. As shown in Fig. 1, data of specimens annealed at 1450°C and those of specimens annealed at 2000°C (O) fell on Curve A. Data of specimens annealed at 2350" and 2550°C fell on Curve B, and those of specimens annealed at 2750" and 2800°C fell on Curve C. The decrease in ductility with increase in annealing temperature became so pronounced that one third of the specimens annealed at 2750" and two thirds of the specimens

Jan 1, 1963

By H. S. Cannon, F. N. Rhines

The application of a static load to a copper powder compact during sintering at an elevated temperature accelerates the rate of sintering in such a way that a given load induces the same proportional increase in rate for all times of sintering. It is shown that sintering under a load is like creep under a fixed load in that the stress required to accomplish a given degree of densification is proportional to the logarithm of the sintering time. VIRTUALLY all of the theories of sintering that have been put forward within recent years have contained the assumption that the chief driving force of the process is the surface tension resident in the exposed surfaces and internal pores of the compact, or powder mass. An externally applied load might be expected to provide a somewhat equivalent driving force for sintering. It is known that the application of a compressive load, during sintering, hastens densification, but it is by no means clear whether sintering under the influence of surface tension alone and sintering under an applied load are fundamentally similar processes. The present study was undertaken in an effort to obtain an answer to this question and with the hope that the answer may contribute to an understanding of the mechanism of sintering. It has been found that there is a close relationship between sintering and creep processes. Copper powder compacts were sintered in hydrogen at 1000 °C under dead loads of 0 to 165 psi and the progress of sintering was observed by means of density measurements. Using a commercial reduced oxide copper powder, see Table I, cylindrical compacts Yz in. x x in. high were made at a pressure of 12,500 psi. For sintering, these were placed in a cylindrical graphite container into which they fitted snugly. A compressive axial load was applied to the compact through a graphite rod of similar diameter, which rested upon the upper end face of the compact; iron weights, in suitable amount, were placed upon the graphite rod to provide the load. This assembly was encased in a vertical silica tube within a resistance-type furnace capable of maintaining a substantially constant temperature within the working zone. After various predetermined times at temperature, the samples were removed from the furnace and their density was measured by conventional means. The heating-up and cooling-down times have been included in the computed sintering time, using the assumption that the sintering rate doubles for each increase of 10 °C in temperature. The experimental results are presented in Table 11. Upon metallographic examination of the specimens, it was found that there was no apparent change in the shape of the pores as a result of loading, i.e, no flattening, Fig. 1. Substantial contraction away from the container walls was observed in all samples sintered with loads of 10.5 psi and less.

Jan 1, 1952

By R. I. Jaffee, J. W. Roberts, D. N. Williams

DISPERSION hardening as an alloying process has aroused increasing interest in the past few years. This alloying procedure, in which an insoluble phase is dispersed randomly through a metal or alloy, is believed to be especially suited for the attainment of creep strength at elevated temperatures since the dispersed phase is metallurgically stable. Another advantage is that the physical properties of the matrix phase, such as electrical conductivity, remain relatively unaffected by alloying. These two characteristics resulted in interest in the applicability of dispersion hardening to copper and copper alloys. A study of copper alloys prepared by powder metallurgy methods has shown some strengthening from oxide dispersions.' However, powder metallurgy processing would not be economical for large scale production of dispersion hardened copper-base alloys. Therefore, the research described in this article was initiated to examine methods of forming dispersion alloys which would not require powder metallurgy processing. If powder metallurgy processing is to be avoided, a method must be found by which the dispersion material can be introduced into and dispersed through the melt before casting. When the dispersed phase is an oxide or similar nonmetallic, it is extremely difficult to form a satisfactory liquid: solid dispersion since successful mixing requires that the melt wet the dispersed phase. It is almost impossible to force a nonwetted particle beneath the melt surface. This difficulty can be avoided, however, if the oxide particles are produced in molten copper by reaction within the melt. The problem would then reduce to maintaining randomness until solidification. Four procedures were considered for producing an oxide particle within molten copper by reaction. In all cases, the intent was to react oxygen (0) with a reactive metal (X) to form an oxide (XO) within the copper melt. These reactions were as follows: 1) (Cu-X), + (O)g (Cu)1 + (XO)S [1] 2) (Cu-X)1 + (Cu2O)s (Cu)x + (XO)S [2] 3) (Cu-OK)1 + (Cu-X)1 (Cu)1+(XO)9 [3] 4) (Cu-OK)1 +(X)s (Cu)1 +(XO)s [4] The reactions were carried out in an 8-lb melt of 99.98 pct Cu maintained in the vicinity of 2200oF by induction heating. Where possible the amount of the reactants was controlled to give 1 to 5 vol pct oxide. Mechanical stirring during the reaction period was used to minimize agglomeration of the reaction product. The presence of a reaction product was determined metallographically by study of an ingot cast from the melt at the end of the reaction period. Reaction [I] was studied by dissolving aluminum in the melt followed by forcing oxygen through the melt. The melts became quite sluggish during the reaction period suggesting that A12O3 was formed. However, the reaction product rose out of the melt rapidly, possibly due to the flushing action of the oxygen stream, and little dispersed phase was formed. Reaction [2] was examined by dissolving aluminum, silicon, or nickel i; the melt and then adding Cu20 (melting point 2250F). This procedure was unsuccessful. Although Cu2O was easily stirred into cop-

Jan 1, 1961

By G. D. Van Arsdale

THE paper of Mr. Posnjak presents a laboratory confirmation of the well known facts that air ox-idation of pure solutions of ferrous sulfate is slow and difficult and that, like many other similar reactions, the speed, may be greatly accelerated by the presence of other substances, often in very small amounts, the effect of the presence of copper being determined ex-perimentally in this case. However, it should not be. overlooked that copper is always present in heap leach-ing solutions in amounts equal to or greater than those in the experiments described; therefore the paper may be considered as an explanation of present results rather than a possible improve-ment to be added to present practice. It is not practicable and it does not pay to precipitate copper by ce-mentation below a point about 5 lb. per 1000 gal., equivalent to about 0.0005 mol. Cu. per liter, compar-ing with the approxi-mately 0.0009 mol. per liter of curve 2 in the diagrams given of course, also, in the heaps local concentrations are usually very much higher than this, which is also true of the effluent li-quors, so that I feel that it is safe to say that copper is already present in amount sufficient to realize, as large an accelerating effect as is practicable. Too much im-provement or refinement cannot be expected in a method of this kind, due to commercial limitations. It should be remembered that chemically and physically there is no sharp dividing line between leaching in place, heap leaching, and the usual tank leaching practice, practi-cally the same reagents being used in all three. From the viewpoint of extractions, etc., it is not common sense to apply leaching in place to an orebody carrying enough copper to pay for mining; and the same is true of heap leaching, for an ore carrying enough to pay for the crushing, handling, etc., of ordinary leaching. That is to say, leaching in place should properly be consid-ered only for an ore too low-grade to pay for mining, and heap leaching for the next higher grade, still how-ever too low-grade to pay for crushing and other re-finements. There is no advantage in leaching without mining or in heap leaching except lower cost of in-stallation and handling; on the other hand there are very distinct disadvantages which are inherent and which should prohibit their use except where necessity requires. The object of metallurgy is to make money out of the handling of ore, and it is purely a question of money as to how far we can go in "improving" or

Jan 7, 1927

By Sherman L. Pease

An inexpensive and relatively rapid method that can be used by field crews is described. Fluid movement is determined by releasing a tracer (dye) in the well at a predetermined level and, after an interval of time, sampling the well fluid at another level. The testing equipment consists of a combined crusher and sampler unit that is run into the well on a wire line. The rate of movement can be calculated from time-volume relationships. Description of Method It is often desirable to determine whether there is occurring any transfer of fluid from point to point in a shut-in well, and, if so, to measure the rate of movement. When it is desired to examine a large number of wells for possible "thief" sands, the time and expense will be considerable unless the method used is inexpensive and rapid. After experimentation with several possible solutions to the problem, the Production Department of the Shell Oil Co., Inc., developed the apparatus described herein, which has proved satisfactory. Routine operations with it can be carried out by field production crews with a minimum of technical supervision and with negligible investment in equipment. The method is based upon releasing a suitable tracer in the well fluid at a known level, and recovering a sample of well fluid at another level after a measured interval of time. Repetition of this process with different time intervals between release and sampling gives the necessary data. For purposes of illustration, the setup for measuring migration of fluid in an upward direction is described in the following paragraphs. The arrangement of equipment is shown diagrammatically on Fig. I. The elements of the combination unit are the tailpiece, the crushing device, the spacer and the sampling device. The lengths of the tailpiece and spacer are varied to place the crushing and sampling devices in the desired positions in relation to the sands being tested. The crushing device breaks the glass bottle containing the tracer when the tailpiece strikes the bottom of the well and the tension on the sand line is slackened. Because of the open construction of the crushing device, the tracer is quickly and completely disseminated in the well fluid. The sampling device is a suction bailer, which obtains a sample of the fluid surrounding the perforated nipple just before the unit is picked up off bottom. The sample is retained in the barrel by the ball valve for examination at the surface. Normally, on the first run into a well, the sample is obtained near the releasing point of the tracer, and thereafter samples are taken near the base of the upper sand if migration is indicated by the dilution or absence of the tracer in the initial sample. The length of the time interval between the releasing of the tracer and the sampling of the well fluid is varied for successive runs, attempting first to establish the rate within limits and then to define it more accurately by additional runs if necessary.

Jan 1, 1946

By Jon J. Connor, Richard J. Ebens

The trace element geochemistry of the cherty clay-rich carbonate residuum cover in much of southern Missouri and adjacent parts of Arkansas was investigated during the earl 1970's as part of a statewide geochemical survey of Missouri.1-8 The study was environmentally oriented and focused on the character of the "natural" geochemical environment of the state. A great deal of effort was made to collect samples that would represent the geochemical variability of the near-surface environment in an unbiased way. In general, this was accomplished by use of randomization procedures in both collection and analysis. Most of the sampling sites were selected randomly to prevent personal bias from entering into sample selection. In the laboratory, samples were analyzed in a randomized sequence to circumvent the effects of any systematic bias that may have entered into the laboratory procedures. Several residuum samples contained 1000 parts per million (ppm) or more of lead, zinc, and barium. All these samples were collected in Washington County in or near the barite district. The frequency distributions shown in [Fig. 1] indicate just how unusual the residuum in Washington County is with regard to these three metals and copper. The shaded parts of the histograms represent the samples from Washington County. The histograms are based on analyses of 141 samples from 38 localities scattered over the subcrop of the Cambrian and Ordovician formations of southern Missouri and northern Arkansas. Eight localities were selected randomly from over each of four bedrock units-Bonneterre dolomite; Potosi and Erninence dolomites; Roubidoux formation; and the Jefferson City, Powell, and Cotter dolomites. Six localities were selected randomly from over the Gasconade dolomite. The geochemical "background" for these elements in residuum is defined by samples from outside of Washington County (unshaded part of the histogram); lognormal distribution curves have been fitted to the corresponding parts of each histogram. Clearly, these lognormal models (curves) describe well the background distributions of zinc, barium, and copper. The background histogram of lead concentrations is notably skewed to the right compared to the model. Lognormal distribution curves were also fitted to the data on lead, copper, and zinc in the residuum of Washington County. No model was fitted for barium in Washington County because of the distinct bimodality in the data. All samples containing more than 1000 ppm were collected in the barite mining district and obviously reflect the elevated levels of barium in the district. The frequency distributions for Washington County are clearly anomalous with respect to the background data, those for lead, zinc, and barium being exceptionally so. The models for anomalous lead, zinc, and copper seem to fit reasonably well, although the scatters in the lead and zinc anomalies are considerably larger than the scatters in the corresponding background data. The copper anomaly, although probably real, is greatly subdued compared to that of lead and zinc. Also, it has a scatter similar to that of the copper background, rather than being enhanced like the scatters of the lead and zinc anomalies. On the basis of techniques described by Miesch,4 the chemical abundance of copper, lead, and zinc in residuum in

Jan 1, 1980

By Walter Rose

It is the purpose of this note to call attention to the circumstance complicating the attainment of uniform gas-liquid distributions in multi-phase flow systems, and especially in those of the so-called Hassler type.' Although the complication arises directly as a consequence of gas compressibility and the dependence of fluid distributions on interfacial curvature phenomena, little reference to the problem has yet appeared in the literature.' In fact, it is only the paper of Geffen et al.3 which gives any experimental evidence that a problem exists. On the other hand, previous authorsl,4,5 have made the tacit, albeit fallaceous assumption that uniform gas-liquid distributions automatically are established in linear flow systems by the expedient of maintaining the pressure gradients equal in the flowing gas and liquid phases. Such a device, it can be shown, will succeed only if the immiscible fluids are both equally compressible or both essentially non-compressible. Another necessary condition required before uniform fluid-fluid distributions can be achieved is that the porous matrix is isotropic. Uniform fluid saturation distribution conditions obtain in isotropic media only when the curvature of each interface of contact between wetting and non-wetting fluids is everywhere the same throughout the interspaces. This is a necessary condition which follows from the consideration that variation in interfacial curvature gives rise to finite capillary pressure gradients and therefore to variation in the saturation distribution. That this is not a sufficient condition follows from the consideration that hysteric possibilities allow for different saturations even though the capillary pressure gradient is zero. In any event, it will- be recalled that it is the intent in the Hassler scheme of relative permeability determination' to obtain initially (by the capillary pressure drainage or imhibition process) uniform conditions of fluid distribution in the core sample, and then to maintain this uniformity during mixture flow so that the resultant fluid mobilities which are calculated will refer to steady-state transfer under fixed conditions of uniform saturation. If the flowing fluids are incompressible, it is recognized that employment of the same value of the pressure gradient in each fluid will result in maintenance of the initially obtained condition of zero capillary pressure gradient. On the other hand, the consequence of gas compressibility is that the gradient in the gas pressure varies from point to point in the flowing stream. and since the pressure gradient in an incompressible liquid is constant it is thus impossible to maintain zero capillary pressure gradient during the simultaneous flow of gas-liquid mixtures. Therefore, one would expect to observe a variation in the gas-liquid saturation in the directions of the capillary pressure gradients, depending in magnitude on the particular way saturation changed with capillary pressure. In order to formulate the manner gas compressibility effects prevent the attainment of uniform gas-liquid distributions during mixture flow, consider a linear flow system of unit length. As noted, it is necessary to assume an isotropic medium (viz. one where permeability is independent of position and direction). One then can arbitrarily set the pressure

Jan 1, 1951

By H. Foster Bain

THIS subject is one that has attracted my attention for a good many years. All of us have had occa-sion to think of the waste that comes from the poor organization of our methods of finding mines and of looking after prospects. Probably all of us, in, examination work, have made a long and difficult journey to see some prospect, in a far country, or often in the mountains, have found the prospect one that never could be developed, and have learned for the first time that it had been previously examined by from two to six of our professional associates, who, each in turn, condemned the property as worthless. I know properties, as all of you do, that have been repeatedly examined by first-class engineers. Perhaps it would not be safe or wise to have an invariable rule that, each prospect should be entitled to only one examination by one engineer; and yet, it seems to me that somewhere there is a middle ground to take; it is hardly necessary that each little prospect should be examined so many times that the walls get greasy from the climbing of mining engineers, or the stopes get filled up with cigarette stumps. 'This mining engineering profession ought to have brains enough to find some way of making this unnecessary, and of making the benefit of an examination, in some cases at least, and in a much larger number of cases than at present, avail-able to the industry as a whole. If we would save the amount of money that is spent unnecessarily on reexamination of properties, and let that money flow into the actual development of )properties, I believe that, in the long run, we would be better off. Just how to do that is another problem. In 1916, while in London, Edgar Rickard and I attempted to do something of that kind. We started to organize a plan whereby mining engineers might, with the consent of their class, turn over to a central organization their reports on properties they had examined, whereby men who had occasion to look up these same properties might, on payment of a fee (which should go back to the original client, if you desire, or might be devoted to some other purpose for the benefit of the industry), have access to these earlier reports. We found, in the short time we considered this matter, that a good many British engineers, and a good many of the companies, were willing to consider this. In other words, they were willing to pool their information, as a matter of practical experience. We all know the enormous value of pooling information. I think it is a problem that may well be considered from the point of view of the industry as a whole, which the Society might attempt' to work out on a businesslike basis. Assuming, however, that it is possible to work out a plan whereby we get the minimum number of examina-tions that are necessary and we make some saving in money, we reach another stage-the problem of the examination itself.

Jan 3, 1922

By Richard Moldenke

Note.—At the request of the American Society for Testing Materials, this paper wau presented at the New York Meeting, February, 1923, in connection with the "A. S. T. M. Tentative Specifications for Foundry Pig Iron" and The Physics of Cast Iron and Its Bearing on All Cast-iron Products and Specifications for Cast Iron," by William R. Webster. The discussion of this paper was not published by the Institute, having been referred to the American Society for Testing Materials, as these papers were again presented for discussion at the meeting of that Society at Atlantic Cit;, June, 1923. The result of these discussions is shown in the following abstract from the Committee A3 report on cast-iron specifications: During the past year, the adaptability of the tentative specifications for pig iron for the steel and the malleable foundry, in addition to the iron foundry proper, have been studied and suitable changes and rearrangements are being made which appear to cover the situation properly. In spite of the experience of foundrymen to the contrary, there is a persistent call from consumers of castinga for defixute compositions for given lines of castings. To respond to this call for information desired by the consumer—information every modern foundryman should have and guide his procedure with—the committee has requested the Sub-committee on General Cast!rigs to take up this matter with a view of constructing tables of analyses rang? for Important lines of foundry product. These analyses ranges however, are to be given as "recommendgd practice" and not intended for actual spkcification. In this way, the consumer of the foundry product will get the information he wants; the furnace-man will be gulded in the production of his pig irons to prepare for analyses ranges which are more closely standardized and have less of "freak" requirements made by Inexperienced metallurgists; and the foundryman will not be bound by chemical requirements he knows will not work out well with his particular reglonal supphes of melting stock. Doctor Moldenke's paper is published in full as it shows clearly the good results of the Institute's work in 1895 and a comparison of the same with the present knowledge of the metallurgy of cast iron—Bradley Stoughton, Chairman, Iron and Steel Committee. In MaRch, 1895, William R. Webster contributed a carefully prepared scheme for the study of the physics of cast iron to the American Institute of Mining Engineers at its Florida meeting. This summary of the interrelation of the chemical and physical characteristics of cast iron with the operations of daily foundry practice caused a wide discussion of the then little understood though universally used material. Mr. Webster has requested the author to bring the subject up to date and, while glad to comply, he must pay that with all the knowledge that has been gained in the quarter century that has passed since the subject was brought to the attention of the Institute—and we now are pretty familiar with cast iron in all its aspects—there is practically noth-

Jan 1, 1923

By AIME

HAPPILY, a large group of the non-ferrous fraternity have long valued a personal association with the new Chairman of the Institute of Metals Division. Many others who will meet him, either officially or unofficially this year, will find themselves sharing the enthusiasm with which he has been welcomed to his new office. Albert A. Smith, Jr., was born at Butler, N. J., on Dec. 28, 1904, and moved to New Haven in time to initiate his preparation for Yale University on the home grounds. After receiving a B.S. degree cum laude in 1926, he stayed on for graduate work at the Hammond Laboratory under C. H. Mathewson, and was awarded an M.S. degree in 1928.

Jan 1, 1948

A SELF-RESPECTING miner doesn't wash the cuttin's off his hard hat until he quits his job but, on the other hand, he keeps his lamp clean and a "spot" focus on the work at hand; the former habit is a matter of pride and the latter serves for efficient operation. During 1949, the mining engineer got all the "wet raises" for the problem of cutting costs to meet the drop of metal prices and keeping research work up to schedule when strikes tied up his work were his headaches. We note, not without admiration, that he hasn't washed the muck off his hat in token of surrender but, on the contrary, his light shines bright on the job to be done as can be seen in the following pages. Few people have any conception of how much ore and rock have to be mined and how much sorting it takes to cull a pound of copper from a 1 % ore body. These people would be even more startled to picture the problems faced by the engineer in mining and extracting, profitably, .2 oz of gold from a ton of ore as is done at some of our lode mines. Strikers have sometimes kept our researchers from their pilot plants and research laboratories, but the end of the year shows progress in making ore out of taconite and ferruginous sandstone, with a pilot plant in operation at Aurora, Minn., and plans for one near Birmingham. Unfortunately, it looks as though the mining engineer will have to make John Doe secure in mineral resources in spite of himself. Coal for steel, power, and to warm our homes is still economically possible, in spite of Mr. Lewis, through development of mechanized tools for mining. During the year, the use of continuous mining machines has extended, transport systems are being adopted to keep the ore flowing from the face to the tune of 40 tons per man-shift. Experiments in the gasification of coal are going forward in several places and may eventually extend our fuel reserves by the addition of coal seams now too lean for mining. Liquid fuel from coal, oil shale, or natural gas will become economic when oil consumption exceeds our supply; but, at less than 30 cents a ton direct cost for mining oil shale, the engineer is almost making it competitive under present market conditions. In this age of specialists, the mining engineer is still holding his own as a broad-gauge man. As a matter of fact, he is emerging from a stage of semispecialization as a coal, nonmetallics, or metal mining man to the realization that, with mechanization, mining is fundamentally materials handling whether it is fluorspar, coal, or copper ore. From the coal business come rubber-tired mechanical loaders to the iron mines of Birmingham. Shuttle cars are also finding application in the lead mines of the Tri-State. Open-pit methods are universal. Belt conveyors, trucks, Diesels, and mobile, jumbomounted drills are finding application in all types of mining. In preparation, such methods as flotation,- heavy density, electrical, and centrifugal are finding application in coal, nonmetallics, and metals. These concepts have been recognized by the mining engineer who has kept himself informed and has been quick to seize a method, machine, or a process from another operation and, with ingenuity, applied it to his problems. This, our Annual Review issue, is written by mining engineers. In gathering these separate articles under one cover for publication, we think of it as a trrbute to the tireless energy and skill of these men. They get little of the publicity accorded physicists, medical men, labor leaders, or company executives, but they supply raw materials without which industry would die.

Jan 1, 1950

By W. W. West

MACHINERY used for size reduction of materials may be divided into three classes-depending upon the manner in which the crushing force is applied. In the first class could be listed the primary crushing units, such as the jaw crusher and gyratory crusher-which crush to coarse sizes by applying pressure intermittently between converging surfaces. In the second class may be listed crushing units, such as the hammer mill, roller mill, and ball mill, which crush principally by grinding or attrition. In the third class is the impact mill which, as the name infers, crushes by impact. The second and third classes generally include machinery used as secondary units for intermediate or fine crushing and grinding. Theory of Crushing by Impact Throw a stone into the air and give it a severe blow while it is momentarily suspended, and it shatters into several pieces. The impact crusher uses this "batting" principle, but carries it one step further. As the material to be reduced enters the crushing chamber, it receives a severe free-air impact blow by hard-hitting beaters suspended from rotor discs. The shattered particles formed are then driven at high velocity against impact blocks, where a further reduction in particle size takes place. The rebounding particles from the impact blocks are again caught by the beaters, and the process repeated until the material is swept out of the open bottom of the crusher. Impact Crushing Versus Attrition Crushing Fig. 1 shows a Hammermill type of attrition crusher. This particular mill has central feed and is reversible. Because of the central feed, an impact section can be designed into it at the feed entrance for preliminary crushing, followed by the cage bar attrition area for final sizing. The material is caught and held in a crushing chamber enclosed by grate bars or perforated plate. It remains in this chamber until it is reduced to particle size small enough to pass through the openings between the bars or in the plate. In this manner the top particle size discharged from the mill is closely controlled. As the swinging hammers pass through the bed of material, the material is broken down by a scrubbing action between hammers and material, material and grate bars, and material against material. This attrition produces a rounded particle shape, accompanied by slivers and a comparatively high percentage of fines. The crusher is especially useful in industries where the material treated is not too abrasive, such as limestone, cement or coal plants. It can be readily understood that from the maintenance standpoint, it is not suitable for hard abrasive materials such as refractories and aluminum oxides. One common type of impactor is shown in Fig. 2. The bottom of this type crusher is entirely open and the passage of the material through the crushing chamber is almost instantaneous. Liberal clearances are maintained between the impact hammers and the anvil type breaker blocks. This eliminates practically all attrition. It is the elimination of attrition that makes the impactor particularly adaptable for reducing hard, abrasive materials. The impactor is also well suited for the reduction of wet or frozen materials, which clog an ordinary Hammermill. Particles broken by impact, when examined under the microscope, show a distinct tendency toward rough, cubical shape, with a minimum of fines. Another important factor in impact crushing is that crushing is accomplished along natural cleavage lines, thus freeing foreign ores or materials.

Jan 1, 1952

By A. Lee Barrett

Mine power systems are quite different in many respects from those usually found in industrial plants. Wide areas are served, usually by a circuit which is connected continuously throughout the mine. In some mines one circuit may run for several miles. There are usually a small number of branch circuits involved considering the areas covered. In all mines circuits are sectional-ized at approximately 2500 ft by means of switches. All, or at least most branch .circuits, are sectionalized by manual switches. In some mines automatic sectionalizing gear is used to give automatic overload protection at various strategic points in the system. However at the present time this practice is by no means general. Circuit hazards in mine electric systems are much greater than those encountered in industrial installation. Falls of roof or coal form a prominent hazard which may break the wire and may cause ground faults or partial ground faults. Ordinarily electric cables in mines cannot be insulated, therefore a fall can easily cause a short circuit. The humidity in most mines is greater than 95 pct and at many points the floor is wet and water may be dripping from the roof over insulators, increasing the difficulties at that point. All electrical circuits must be contained in the mine entries and space must also be provided for moving equipment and men. Consequently, because of the proximity of the power circuit to the locomotives and other moving equipment a considerable hazard is introduced. Mine haulage track is not always good so that wrecks are not unusual. A wreck frequently does considerable damage to the power circuit. A ground fault frequently results in heavy arcing which may be sustained. If the damage to the circuit is mechanical it may mean that a trip of coal has wrecked and a great deal of coal dust could be thrown into the air as a result of the wreck. The arc resulting from the ground fault could ignite the coal dust and a mine explosion would result. A continued ground fault, even if high in resistance, also may ignite the coal which may be on the bottom of the entry and a mine fire may result. The fire hazard is one of the most serious hazards in any coal mine. Since the mine circuit is completely interconnected, any power interruption resulting. from a fault seriously affects the whole operation. Unless the mine is completely sectionalized the whole mine will be shut down until the fault is located and the proper switches are opened to isolate the fault; even then large areas may still be without power if they are supplied by a single power circuit which may be opened in isolating the fault. Further, mine power is usually supplied from a relatively small number of substations and the fault may be on a stub feed circuit, in which case power cannot be restored until the fault has been cleared. Practice varies widely with respect to placement of substations. Usually they will be spaced from I to 2 miles in 500V opera-tions, and at I mile intervals or less, in 250V operation. Actual mining operations govern substation placements and no general rules can be followed.

Jan 1, 1949

By Jean Howard

ALTHOUGH zone melting has found favor in recent years because of its convenience and its faster rate of production of single crystals, the older technique of strain annealing still has a number of advantages. It has been reported1 that the lattice of a crystal prepared in this manner has fewer defects than one grown from the melt and, of course, the control of the external geometry is easier. Also, the diameter of the crystal which can be grown by zone melting has a theoretical limit determined by the surface tension, thermal conductivity, and specific gravity of the material, while the size of that which can be grown in the solid state is limited only by practical considerations. In a program for the study of the structure-sensitive properties of niobium and its alloys it was desirable, for ease of handling, to have available single crystals of large diameter. This requirement dictated the choice of the strain-anneal technique. Induction heating was chosen in the present work in preference to other methods because of its deep penetration of large-diameter sections. A five-turn coil, powered by a motor generator operating at 10,000 cps, is contained in a vacuum chamber. The work is lowered through the coil for both the recrystallization and the crystal-growth steps. Starting with electron-beam melted material, the bars are cold-swaged, recrystallized, strained in a tensile machine, and then subjected to the growth step. Single crystals have been grown under the following conditions: Reduction in area during swaging— 79 to 99 pct, recrystallization temperature—1500o to 1667°C at the rate of 1.25 in. per hr, strain—1.0 to 2.5 pct, growth temperature—1600o to 2400°C at rates from 0.15 to 1.25 in. per hr. Starting grain sizes achieved after the recrystallization step varied from 0.6 to 1.55 mm. The starting grain size was found to be critical; if it is too fine and has any residual cold work, as shown by the presence of a fiberlike texture, a single crystal cannot be grown. Using this technique niobium single crystals have been grown from 0.250 in. diameter and 12 in. long to 1.00 in. diameter and 5 in. long as shown in Fig. 1. The 1-in. crystal is the largest to be reported for a refractory metal as compared to 0.500 in. for molybdenum grown by zone melting.2 As with zone melting, control of orientation is possible by the use of special procedures. Other investigators, see for example Williamson and Smallman,3 have reported that orientation control may be achieved by a bending technique. In the present work the strained rod is partially lowered through the coil to start the growth of the crystal. Then it is removed and bent at a point ahead of the interface between the single and polycrystalline portions. Finally, it is returned to the chamber and growth is continued "around the corner". The amount of bending which can be imparted to the rod is limited by coil geometry which up to now has been 10 deg. However, by repeating the bending and growing operations it should be possible to attain any desired orientation. The perfection of these single-crystal specimens as a function of process variables is being measured by X-ray and metallographic techniques and the results will be reported in a forthcoming paper. This work was partially supported by the Advanced Research Projects Agency.

Jan 1, 1964

By Tracy L. Atherton

THE mapping of slant-hole oil fields is complicated by the fact that relationships between wells are subject to variations in three dimensions and are not readily adaptable to representation on a plane surface. Formalized perspectives usually expand into an undecipherable maze of lines and peg models with the necessary supporting dividers serve only to awe the beholder with their perplexity. The usual compromise arrived at is a plotting of the projection of the courses of all of the wells in the field onto a horizontal plane (Fig. I). These complicated plots, when used in conjunction with the individual vertical projections of the wells, heretofore have been the most practical foundation for the study of general field problems. This method has many shortcomings, however, which are not found in the recently developed straight-hole equivalent projection method. The new method supplies a series of maps similar to the type now in common use in straight-hole fields. It is remarkably simple to use and supplies a ready check on clearances between wells. Perhaps the method can best be understood if the steps that led to its development are described. Common practice of oil-well surveyors is to supply the owner with a tabling of the survey and a plan view plotted against the true meridian. If a vertical projection is plotted, it is usually plotted separately and shows vertical depths against the projection of the stations along a section line joining the top and bottom of the hole on the plan view. A more practical method would be to plot the section line of the plan parallel to the abscissa of the vertical projection (Fig. 2). If this is done, any line parallel to the "base line" will pass through the points on each projection that represent a single point in the hole. Another very useful graph, which the well surveyor could provide, would be plotting of vertical depths against measured depths. Fig. 3 illustrates the next step facilitated by this form of plotting. The lines designated "well A" represent the proposed course of a hole to be drilled. Superimposed on the drawing, in its proper relative position, is the plan view and a portion of the vertical projection of a present producer, "well B." From an inspection of the plan view we arrive at the correct conclusion that the only indicated possibility of collision is at the point of crossing of the wells on the plan view. Referring to the vertical projection, note that the vertical depths of stations on B, in the vicinity of the crossing, are plotted on the same plane of projection as A. The clearance is the indicated distance between the two vertical sections measured along the parallel to the "base line," which passes through the point of intersection of the plan view. Repeat this process for the other wells crossing the proposed course of A on the plan view and you have a preliminary check of clearance for the proposed course. A similar comparison of crossings on the vertical projections of all

Jan 1, 1941

By U. F. Kocks

SINGLE slip and a single parallel array of edge dislocations have been obtained in silicon single crystals by homogeneously bending them in a four-point dead-loading device at 1000°C around the [211] axis with [253.] as the tensile axis which encloses an angle of 40° with the glide plane. Other orientations or lower temperatures proved unsuitable. The homogeneity of the bending was checked by the method shown in Fig. 1. The chemically polished specimen was exposed to parallel light from a microscope-stage lamp. The sharpness of the focus of the reflected light is an approximate measure of the circularity of the sample. This method permitted the determination of the actual radius of bent specimens within 10 pct accuracy. The etching solution developed to show up dislocations consisted of the following volume parts: 2 HF, 1 HNO,, 2 CH,COOH, 1 a solution of 5 wt pct of CuNO, in water (Rosi'). The acetic-acid content was adjusted to produce fine well-defined etch-pits in about 2 min at room temperature. The effectiveness of the etch did not depend on the crystal-lographic orientation of the face etched. The number of etch pits produced after a given deformation was checked in the usual way against the Cahn formula' and showed good agreement. As it is to be expected from the orientation of the primary system there are slip steps on the top and bottom faces but not on the side faces (perpendicular to the bending axis), whereas etch pits appear on the side faces but not on the top and bottom faces. This is different only for occasional dislocations of a secondary glide system. All the micrographs included have been taken in places where the most of secondary glide could be seen. From this it is evident that in the homegeneously deformed region the number of secondary dislocations was less than 1 in 1000 of the primary type. Dislocation Distribution—A series of tests with varying amounts of deformation is shown in Figs. 2(a-c). An observation that can readily be made from these pictures is that the linear density of dislocations within slip lines is relatively constant and independent of the amount of deformation. "Slip line" in this context is understood to mean a line that is a straight connection of individual etch pits. The heavily pitted bands which can be seen in each of these three pictures may be resolved fairly definitely, at higher magnifications, into many "slip lines" the distance between them being of the order of 0.5p The distance between dislocations in slip lines is of the order of 2p and it is also fairly constant in going from the neutral zone outward. The higher over-all dislocation density in the outer regions of any sample and the higher total number in a more heavily deformed sample, are effected by a larger number of "slip lines." The relative constancy of the dislocation distance within an individual slip line is surprising. The absence of pile-ups" may imply, if there is no rear- rangement on unloading," that there are no substantial obstacles to slip in this case. It seems that the equilibrium distribution of the dislocations in each active slip plane under the higher external stress on

Jan 1, 1959

By d&apos, K. Mukherjee, C. R. Antonio, R. J. Maciag, G. J. Fischer

Tensile data indicate that over the range of strain rates 10-5 to 10-1 sec-1 and in the temperature range 298° to 743°K the flow stress at a given temperature may be expressed as: C0 = Cem where m changes abruptly at a critical strain rate of 10-2 sec-1. The Zener-Hollomon parameter was tested and found to be invalid except at temperatures exceeding 700° K. Values of AH calculated for temperatures above 700°K ranged from 30 to 47.5 kcal per g-mole. ThE flow characteristics of aluminum and aluminum alloys as a function of temperature and strain rate have been of interest to various experimenters for many years. Trozera et al.1 determined tensile flow curves for pure aluminum and concluded that the combined effect of temperature and strain rate on the flow stress could be related using the Zener-Hollomon parameter. More recently, Hockett2 proposed a relationship between compressive flow stress, temperature, and true strain rate which could be represented by a surface. Holt et al.3 studied strain rate effects on flow stress in a number of aluminum alloys at room temperature. It was determined that there is a critical strain rate above which a strain rate effect is observed. It was also reported that the strain rate sensitivity was significant in aluminum alloys in the "0" condition only. In a recent review paper McQueen4 has considered a number of mathematical expressions which have been used to describe the relationship between stress, strain rate, and temperature. These expressions have been applied to the flow characteristics of a number of different metals and alloys including aluminum. In the present work the strain rate sensitivity of 7075-0 aluminum was determined in the temperature range of 298° to 743°K. The tensile flow stress at 0.2 pct offset was obtained over a nominal strain rate range of 10-5 to 10-1 sec-1. EXPERIMENTAL Threaded tensile specimens were cut from a single heat of 7075-T6 aluminum. The composition of this heat is given in Table I. The test bars had a gage section 1.5 in. long by 0.250 in. diam and conformed to ASTM specification E-8-57T. After machining, the bars were annealed at 413° ± 10°C for 2 hr. They were then cooled to 260°C in 3 hr, and finally air cooled to room temperature. This treatment is commonly used to produce the "0" condition. Load-time curves at constant crosshead speeds were obtained at various temperatures using an Instron testing machine. Crosshead speeds ranged from 0.002 to 40 in. per min. A three-zone resistance tube furnace was used for the elevated-temperature tests. Temperatures were measured at three points along the gage length by means of suitably placed thermocouples. Each tensile bar was held at temperature for 30 min prior to testing and the maximum temperature variation along the gage length was kept below 1°C. In all, 275 samples were tested and each data point plotted is an average of at least two tests. In many cases the point represents up to six tests. RESULTS AND DISCUSSION Values of the flow stress at 0.2 pct offset at six different temperatures are plotted vs strain rate on log-log coordinates in Fig. 1. Over discrete ranges of strain rate, the data for each temperature fit the expression:

Jan 1, 1969

By Robert G. Johnson, Richard J. Prosen

THE present work was done in connection with a program for investigating metal-to-ceramic braze techniques using the well-known reactivity of titanium with certain ceramic oxides. The investigation extended over the range of 0 to 30 pct Ti. The indium was obtained from the Indium Corp. of America and was specified to be 99.97 pct pure. The titanium was supplied by Metals Disintegrating Co. and was stated to be 99.5 pct pure. Each 2 to 4 g quantity of the alloy compositions studied was prepared using a mixture of 300 mesh powders pressed into pellet form. The pellets were heated in thin-walled molybdenum crucibles using a resistance-wound furnace in a vacuum less than 10-4 mm of Hg. Little reaction occurred between the molten alloy and the crucibles at the temperatures and short time intervals used. Only slight spectroscopic traces of molybdenum were found in the melt after a typical period of heating. The points shown on the phase diagram of Fig. 1 represent thermal arrest measurements. The phase boundaries shown are consistent with the thermal arrest and X-ray data obtained. The temperatures were determined using a Pt-Pt 10 pct Rh thermocouple with the junction spot-welded to a short molybdenum rod. The rod was almost completely immersed in the liquid metal, so that good thermal contact with the melt was achieved. The small mass of the crucible and its contents made it possible to obtain many heating and cooling curves in a few minutes time, an advantage which helped to minimize the effects of indium evaporation at the higher temperatures. A Brown strip chart recorder was used to display the thermal arrest curves. To check the technique, the melting point of 99.999 pure Al was measured and found to be within 1 of 660.2c, the accepted melting point. The upper limit of temperature with this technique was about 900°C beyond which the evaporation rate of indium became great enough to alter appreciably the alloy composition during the runs. A peritectic transition was found at 796°+ 5°C from 1.5 to 30 wt pct Ti. It was not found at 0.5 or 1.0 pct. A small super-cooling effect was observed. The melting-point transition to the B + L phase was found up to 20 pct Ti but not at 25 pct. Within the 1 deg accuracy of the measurement, the transition temperature was that of pure indium, 156°C. The liquidus line to the left of the 1.5 titanium point was located only approximately. The two crossed circle points at 1.0 and 0.5 pct Ti place a lower limit on the liquidus line and were determined by observing visually the cessation of vibrations in the melt caused by the damping effect as B-phase crystals were formed during cooling. The B-phase boundary was found at 23.8 pct Ti by X-ray analyses of small single crystals of the B phase isolated from 2 pct Ti melts. The compound was found to be Jetragonal with lattice dimensions of c = 3.052A and a = 10.094W. From the unit cell dimensions and the measured density of 6.42 i 0.05 g per cc, the compound was determined to be In4Ti3 which has 23.8 pct Ti. The calculated density of 6.44 g per cc is obtained assuming 2 mol

Jan 1, 1962

By Mark J. Klein, Robert A. Haggins

The pesence of a small amount of oxygen was found to cause significant grain growth restraint in 99.99 pct Ag. This behavior does not seem to be due to the presence of an oxide dispersion, but instead, to be related to the existence of minute amounts of oxygen m solid solution. In the amounts necessary to cause the observed effects it was found that oxygen migrates with extreme rapidity, probably along extended structural imperfections, such as dislocations and grain boundaries. A number of years ago, hastton' observed that the presence of small amounts of oxygen caused a retardation of grain growth in silver. He attributed this behavior to the formation of a fine dispersion of oxide particles by the internal oxidation of trace impurities. More recently, wood2 also found that oxygen caused grain growth restraint in copper. Experiments using both copper and a dilute solution of aluminum in copper led to the conclusion that the effects observed in copper were due to oxygen in solid solution, rather than an oxide dispersion. During the course of a study of the influence of several solutes on the recrystallization of Silver, the presence of minute amounts of oxygen was found to exert a very significant influence on grain growth. The experiments reported here were undertaken to determine whether the grain growth restraint observed in silver due to the presence of oxygen is related to oxygen in solution or to an oxide dispersion. Because of the difficulty in measurement and control of the very small oxygen concentrations causing the observed effects no attempt was made to treat the data in a quantitative manner. EXPERIMENTAL PROCEDURE AND RESULTS Silver of 99.99 pct purity was melted and cast into slabs in a vacuum of better than 0.1 p. These slabs were reduced 50 pct by cold rolling to a thickness of approximately 2.5 mm. Samples cut from this material were annealed in vacuum for a series of different times at 803°C. Another group of samples was similarly annealed in air. This temperature is well above the recrystallization range for silver of this purity (150" to 200°C). The mean grain diameter of these samples was measured, both in the center and near the edge, using the line-intel-cept method, and is shown as a function of annealing time in Fig. 1. The grain size in the centers of the samples annealed in air was found to be larger than that of the surface regions, while the grain growth behavior of the edge and center regions annealed in a vacuum was approximately the same. he edge measurements were taken at a distance of 0.2 mm from the surface.) A few samples were annealed in nitrogen. Their behavior was identical to that of the vacuum-annealed samples. Samples annealed in oxygen had approximately the same grain growth behavior as those annealed in air, leading to the conclusion that this behaviouer is due to oxygen. The microstructure of a sample annealed in air at 803" for. 30 minutes is shown in Fig. 2. The smaller grain size near the surface is characteristic of the iir-annealed specimens. Possible causes of this grain growth restraint in the air-annealed specimen are the formation of an im-

Jan 1, 1962