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By Charles Clayton

CAST-STEEL runners, while not interesting from a commercial standpoint, furnish valuable material for microscopic study. Foley1 found not only the usual ingot structure, but zones of Widmannstattian structure, which he explains as due to differential crystallization. The writer, in examining nickel-steel runners, found a type of pro-eutectoid ferrite that is unusual and most probably new. The runner in question was 8.75 cm. (3.5 in.) in diameter and analyzed at its center 2.69 per cent, nickel and 0.350 per cent. carbon and at its edge 2.62 per cent. nickel and 0.359 per cent. carbon. The macrostructure, which is the usual type, is shown in Figs. 1 and 2, which are a cross-section and the longitudinal surface, respectively. Slight segregation of ferrite can be seen near the center of the segment and at other points near the edges of polyhedral grains. Stead's reagent brings out the dendritic structure within the grains as shown in Figs. 3 and 4. This texture is what would be expected of such a steel. Figs. 5 to 10 show the peculiar swirls of pro-eutectoid ferrite, the finding of which prompted this paper. These swirls, or eddies, if such a term can be applied, occur in all parts of- the runner and in all parts of the individual grains. I he tipper left-hand figure shows this ferrite at a grain boundary; the middle left-hand figure shows the inner portion of a grain; and the lower left-hand figure shows these swirls at the edge of a grain which is also the outer edge of the runner. Straight-lined ferrite associated with swirls is found in the areas of Figs. 11 and 12. The texture of the carbon-bearing constituent of the runner is shown to- some extent in Fig. 12, the pearlite being very fine or sorbitic. Higher magnification does not bring out any new features either in the peculiar ferrite or in the pearlite. Figs. 13 and 14, at 240 diameters, show the curved and straight-lined ferrite.

Jan 3, 1920

By I. W. L. Finlay, H. R. Ogden, R. Jaffee, D. J. Maykuth

Aluminum has been found to be soluble in a titanium to about 26 pct, and to raise the temperature range of transformation from a to 8. Two intermediate phoses exist in the system, a new face-centered tetragonal phase, designated as 7, which occurs between 34 and 46 pct Al, and TiAI Metallographic and X-ray diffraction data were used to determine the diagram. ALUMINUM is one of the useful alloying additions for titanium-base alloys. It is an important constituent in two commercial alloys released in 1950, one containing 4 pct A1 and 4 pct Mn, and the other containing 3 pct A1 and 5 pct Cr as the metallic alloy content. It is probable that, in the future, other titanium alloys containing aluminum will be released. In the course of alloy development work, some of the features of the constitution of titanium-rich Ti-A1 alloys have been developed. These are described in the present paper. The only available information on the Ti-A1 system concerns the aluminum-rich end of the diagram. In summarizing the early work, Hansen' indicated that there is little solubility of titanium in aluminum, and that a two-phase region exists up to the intermediate compound, TiA1a (37.2 wt pct Ti), which has a congruent melting point at 1355°C. Fink' has reported that a peritectic reaction occurs at 665°C between the melt and TiA1³ to form the terminal solid solution of titanium in aluminum. The solubility of titanium in aluminum has been established at 0.24 pct at 510°C and is estimated to be between 1 and 1.5 pct at the peritectic temperature. No information has been published on the diagram above TiAl,, other than a statement- hat aluminum raises the a to ß transformation temperature of titanium. Materials and Preparation of Alloys Iodide-type titanium and high-purity aluminum were used in this investigation. Typical analyses of these materials are given in Table 1. Alloys were prepared by arc melting on a water-cooled copper hearth, using the techniques described previously for contamination-free melting.' Weight losses during melting were very low, ranging from 0 to 0.02 g in a 10 to 15 g charge, indicating that little or no loss of aluminum occurred during melting. This permitted the use of nominal compositions for these alloys. Alloy ingots containing from 0 to 7.5 pct A1 were rolled in air at 850°C to 0.040 in. sheet. Alloys containing from 7.5 to 16 pct A1 could only be rolled about 25 pct reduction at temperatures up to and including 1200°C. Alloys containing above 16 pct A1

Jan 1, 1952

By William P. Shinn

IT seems proper to present in the Transactions of the Institute, from time to time, formal record of the advances made in the arts and sciences to which our organization is devoted-milestones in the highway of progress-from which our successors may measure what they in their turn have gained. The admirable address of the Hon. Abram S. Hewitt, then president elect of the Institute, on "A Century of Mining and Metallurgy in the United States," read at the opening of the meeting held in this hall in June, 1876, forms a point of departure, and the conclusion of the eighth decade of the nineteenth century and the end of the first five years of our country's second century seem to offer a fitting time for again making up our record. The period since 1875 has been marked by a degree of activity in mining and metallurgical industry never before equalled in our history, while the progress made in the science and art has been, if possible, still more noticeable. Whether we consider the mining of iron, ore, and coal, or of the precious metals,-the advances made in smelting the former and in reducing the latter, the extraordinary increase in the production of Bessemer steel, or the science of Thomas and. Gilchrist, which makes the hitherto rejected phosphorus the corner-stone of success,-the past five years are alike memorable for the progress they have shown and to some of the most marked features of this progress I will now oriefly call your attention. It may be of interest in this connection to note the following figures, showing the growth of the population of the United-States during the last three decades :

Jan 1, 1881

By C. C. Hawes

SILICA is the main impurity in iron ore. It is intimately associated with the iron oxide, sometimes free but more often in the combined state, as a mineral silicate. Its separation and purification sometimes present difficult problems to the analyst. In the absence of any definite knowledge to the contrary, orthosilicic acid. H4SiO4, is supposed to be the only silicic acid made, all other forms, such as metasilicic acid, H2SiO3, representing different degrees of its dehydration. These acids are exceedingly weak but form the most stable compounds found in nature. They are known as the silicate minerals and are formed by the combination of the different metallic oxides

Jan 1, 1936

By H. K. Birnbaum, H. B. Aaron

he authors suggest that a number of the types of yield phenomena which have been observed in high-purity fcc metals are due to dislocation interactions bccurring at the surface of the crystals rather than due to volume effects. Since rather different inter-pretationS2,7,12,14,15,21-23 of these effects have been suggested, it is of some importance to ascertain the role of the surface in these yield phenomena. It should be first noted that the high-purity fcc metals appear to exhibit several distinct types of yield phenomena. The "unloading yield point" which is observed on reloading the specimen after removing the applied load has been found to be independent of the time of holding in the unloaded condition. If however the specimen is aged while under load and re-tested without removing the load, a yield point is observed which does depend on the aging time and tem-perature. Feng and Kramer suggest that the removal of the unloading yield point on removing sufficient material from the surface of the specimen indicates that the effect is associated with dislocation interactions at the specimen surface. Their observations are counter to those of Haasen and Kelly5 and Birnbaum.12 Since the unloading yield points are extremely small (-2 pct of the flow stress) the results of the surface removal will depend on whether the material is removed uniformly over the gage length. The introduction of small etch pits or the removal of more material from one portion of the specimen than from another will result in preferential yielding and obliteration of the yield point. With the normal electropolishing techniques it would be difficult to avoid these effects. The data obtained by Aaron and Birnbaum, Fig. 10, on aluminum polycrystalline specimens indicate that, although the unloading yield point and the strain-aging yield point are not simply additive, a time- and temperature-dependent yield point is observed under conditions where no creep is observed. This does not agree with the results of Feng and Kramer but does agree with the results on copper15 and gold.7~23 The source of the discrepancy may be the procedure used by Feng and Kramer to correct their time-dependent yield points for the effect of creep. Since the strain-aging yield point is not independent of the pre-aging strain rate, Fig. 11, the effects of creep cannot

Jan 1, 1967

By B. R. Hollebone

The chemical action of molten anhydrous pyridinium chloride (pyridine hydrochloride) on oxy salts and ores of some Group IV and V metals are discussed-in particular zirconium, hafnium, niobium (columbium) , and tantalum. Laboratory-scale experiments are described whose results suggest that this reaction might provide a practicable means of converting ores of these metals to the anhydrous metal chlorides. Experimental results are given which provide some insight into the nature of the reactions and some of the compounds which could be present at intermediate stages. The treatment of ores of the more refractory metals is difficult and expensive and often demands the use of gaseous chlorine, hydrogen fluoride, or other reagents which are, if nothing else, expensive and dangerous to use. This situation is, of course, due to the strong metal-oxygen affinity resulting from the high charge and the small size of the metal ion which combine to produce such high lattice energy in the metal oxide as to defeat standard reduction methods. Thus, the characteristic of these metals which leads to one of their most important properties—corrosion resistance—is the main hurdle in obtaining the metal. Many common methods' for treating ores of zirconium and niobium (columbium), for example, proceed as directly as possible to the preparation of anhydrous halides. These halides are purified by now standard processes of solvent extraction, fractional distillation, and so on. This paper reports on preliminary stages in the development of a method of treating these ores—particularly of niobium—so as to prepare the anhydrous, volatile halides by the use of chemical reagents which are much more easily handled than those used at present. CHEMICAL PRINCIPLES The solvent action of aminium halides has been referred to periodically in the literature over several decades. The chemistry of pyridinium chloride has been discussed particularly by Audreith2 and Starke.3 However, since this is not familiar chemistry in its nonaqueous setting, it is perhaps well to point out some general metallurgical principles and applications. One can use, as point of departure, the fact that HC1 dissolved in pyridine is an acid very much the same as aqueous HC1. It thus gives essentially all the reactions of the HC1 with which we are familiar. However, when the ratio of HC1:pyridine reaches 1:l (a ratio far higher than is possible with H2O) the solution becomes solid C5F5NHC1, pyridinium chloride.* Inter- estingly, the acidic properties of HC1 persist in this medium, although they are manifest only in the molten state of the compound. Some typical reactions of divalent metal compounds which we have carried out in our laboratories are the following: 2PyHCl + Zn — (ZnCl2) + H2 + 2Py 2PyHCl + MnO — (MnCl2) + H2O + 2Py 2PyHCl + CuS — (CuCl2) + H2S + 2Py 2PyHCl + CaCO3 — (CaCl2) + H2O + CO2 + 2Py In these equations, the parentheses indicate a solvated compound which in water would be predominantly a hydra ted cation but which in molten PyHCl will more likely be a complex chloro-anion of the type MC14-. Some important physical properties of pyridinium chloride (PyHCl) are given in Table I. One sees from the data in Table I that at the temperature of molten pyridinium chloride both water and pyridine will boil away. Thus the system will be kept anhydrous. For this reason the reaction 2NbCl5 + 5H2O— Nb2O5 + 10HC1 has no counterpart in molten PyHCl. Instead, the chloride is made even more stable by the formation of a hexachloro complex: Nb2O5 + 12PyHC1 — 2PyHNbC16 + 5H2O + 10Py One interesting property of these complexes is their thermal instability which for the case of niobium can be illustrated as PyHNbCl6 — PyHCl + NbCl5 With a view to possible use of this chemistry in ore treatment, we have attempted to dissolve Pyrochlore (essentially FeO - Nb2O5) and recover the NbC15 from it.

Jan 1, 1968

By Albert Sauveur

IT is not without some hesitation, and even misgiving, that I venture into a discussion of the now classical Roberts-Austen Roozeboom diagram, lest I too fail, like so many other writers, in giving a satisfactory and logical explanation of these interesting curves. Roozeboom's original diagram is reproduced in Fig. 1. The generally accepted meaning of the various curves has been so often described before this Institute that it will suffice to recall it briefly before passing to my own interpretation of the changes involved.

Nov 1, 1906

By Charles M. Rolker

The Mariposa estate, a grant made by the Mexican Government to Juan 13. Alvarado, during the time when California was still under the dominion of Mexico, was purchased in 1847 by J. C. Fremont, and the United States patent issued to him in 1856. By sale, or its equivalent, money advanced, the property came into the hands of the Mariposa Company, June 25th, 1863, and through a varied existence of work and temporary inactivity, has as yet been kept as an undivided property, and rests at present with the Mariposa Land and Mining Company of California. I pass over this period of 15 years without giving the history of the ups and downs of the estate, partly because it is familiar to many, and also because I think in such a case this paper would assume too much the character of a narrative. The estate is situated in Mariposa County, California, about 170 miles southeast of San Francisco. It begins in lat. 37' 26' 38'' north, and long. 119' 55' west, and runs 6 miles north; thence 69 miles west; thence 5 miles northwesterly towards the Merced River; thence 84 miles southerly; thence 3 miles easterly; thence 1 mile southerly; thence 2 miles easterly; thence 2 miles southerly; thence 2 miles easterly; thence 1 mile southerly; thence 2 miles easterly; thence 1 mile northerly, and thence 2 miles easterly to the place of the beginning; embracing 44,387 acres (70 sq. miles) more or less. To describe this vast estate in detail as to its full geology, its numerous quartz veins, its many facilities to exploit them and work them by cheap methods, would lead me further than I intend to go at present, and I will content myself to describe the latest workings of the company and its immediate field of operations. As long ago as 1860 the late Dr. Adelberg advised the running of an adit from the northwestern edge of the estate at Benton Mills 011 the Mereed River. His object was to undercut the deepest workings of the then paying mines on the estate—the Pine Tree and Josephine. Following this advice, the tunnel known as "River Tunnel " was commenced. Its entrance is in S. 5, T. 4, s. R. 17 E. It is run in

Jan 1, 1879

[Secretary's Note.—The following official tables, prepared by the government of the State of Nuevo Leon, and presented to the Institute at the fifth seseion of its Mexican meeting, held at Monterrey, November 26, 1901, present in compact form a general summary of the mineral and metallurgical industry of the State for the preceding calendar year, together with a statement of the railway-shipments of ores from January 1 to October 31, 1901, which, in the absence of direct official mining and metallurgical statistics (not yet available for that year), clearly indicates an increased activity and productiveness of these industries. In other words, the mines are reported to have yielded in 1900 a total product of 83,960,169 kilos of ore; whereas, the railway is reported to have carried, in 10 months of 1901, a total of 108,318,647 kilos of ore. At this rate, the railway-shipments for 1901 mould amount to nearly 130,000,000 kilos. It is, of course, possible that the railwaystatistics include other freight than ores, or that they include duplications, such as would be involved in the record, first of certain ores transported, and afterwards of the " base bullion " produced therefrom. An accurate comparison is therefore impossible ; yet, after making all reasonable allowances for such items as have been named, it seems safe to infer that the mining and metallurgical industry of this State accomplished in 1901 a substantial progress.]

Jan 1, 1902

By Andrew Hyslop

IN our never ending search for new and better ways of underground mining, we find that transportation has had its share of new ideas in the past few years. The old and still effective method of track haulage has had some competition from rubber tired cars and belt conveyors, each of which has its own safety hazards which demand close scrutiny in any safety study. The use of shuttle cars, in particular, has introduced clearance hazards caused by the absence of the more or less positive guides formed by the rails of a track sys¬tem; likewise the requirement of greater clearance tends to increase roof hazards which we all know to be great enough with the best of methods. However, the subject of transportation hazards is too broad to be covered entirely in one paper, so I will confine myself to track methods which carry the preponderance of underground tonnage. I am quite sure that to this group it is an evident fact that haulage accidents are of major importance. However, to stress the point, it may be well to look at a few figures. Safety statistics can be quite boring and often are subject to question as to mathematical accuracy because of different

Jan 1, 1948

By R. G. Grieves, G. Thodos

A method is presented for the accurate calculation of the cricondentherm and cricondenbar pressures of multicomponent hydrocarbon mixtures of known composition. The mixtures may contain six and quite possibly any number of components including paraffins, isoparaffins, olefins, acetylenes, naphthenes and aromatics. The approach is similur to that used for calculating critical pressures and cricondentherm and cricondenbar temperatures. The critical pressure, the normal boiling point and the approximate vapor pressure behavior of each component are all that art required. A stepwise culculation procedure is necessary for mixtures containing more than two components. From an analysis of 22 binary systems and 118 mixtures, the average deviation of calculated cricondentherm pressures from reported values is 2.4 per cent. For nine multicomponent mixtures the average deviation is 1 per cent. Considering 19 binary systems and 108 mixtures, the average deviation of calculated cricondenbar pressures from reported values is 1.7 per cent. For 15 multi-component mixtures, the average deviation is 2.2 per cent. INTRODUCTION A knowledge of the phase behavior in the critical region of multicomponent hydrocarbon mixtures is of value both in industrial processing operations and for the optimum operation of gas condensate reservoirs. Accurate methods of calculating the critical temperatures and critical pressures4,6,7,8,9,23 and cricondentherm and cricondenbar temperatures6p10 of multicomponent hydrocarbon mixtures are available in the literature. If the cricondentherm and cricondenbar pressures could be calculated with equal accuracy, the entire phase diagram of a multicomponent hydrocarbon mixture could be well approximated. The work of Etter and Kay6 is limited to systems containing the normal paraffins and has not been tested on systems containing a heavier component than heptane. In addition, the development of their multicomponent equations is based upon a limited number of mixtures containing from three to six components. Silverman and Thodos have considered systems containing both paraffinic and non-paraffinic hydrocarbons, but their correlation is limited to binary systems and is highly inaccurate for methane systems. Eilerts has done extensive work on the cricondenbar pressure; he has produced an excellent correlation for binary systems. However, his procedure for multicomponent mixtures is chiefly useful for highly complex mixtures requiring a knowledge of the vapor-liquid equilibrium behavior of the mixtures; he has not considered the cricondentherm pressure. The objective of this study was the development of an accurate and rapid method for the calculation of the cricondentherm and cricondenbar pressures of multicomponent mixtures containing all types of hydrocarbons and having a wide volatility range. The approach that was adopted is similar to that used by Grieves and Thodos for cricondentherm and cricondenbar temperaturesl0 and for critical pressures.9 However, methane systems had to be considered separately and a modified stepwise calculation procedure was utilized for the cricondertherm pressure. The correlations were developed in a manner similar to those for critical pressures and cricondentherm and cricondenbar temperatures. Based upon binary data reported in the literature it was observed that the ratios of cricondentherm and cricondenbar pressure to the pseudocritical pressure (molar average), pt/ppc and pp/ppc, respectively, in two-component systems depended upon the mole fraction of the low-boiling component and upon the diversity in properties of the two components. A dimensionless boiling-point parameter Tb,/Tb was chosen to represent the diversity in properties of the components. For a binary system, Ti,'is the molar average of the normal boiling points of the two components involved. For a multicomponent system it is either the molar average of the normal boiling points of all of the components involved or the molar average of the normal boiling point of the pure low-boiling component and of the atmospheric boiling point of the low-boiling-free mixture of

Jan 1, 1965

By John Jermain Porter

(San Francisco Meeting, October, 1911.) THE value of any particular ore, coke, or limestone, for iron-making, depends upon its effect, first, upon the quality or value of the resultant product; and second, upon the cost of smelting. The factors under the first head are the percentages of phosphorus, manganese, and other elements, which remain in the pig-iron ; while those under the second head are the percentages of iron, carbon, or lime, the slag-forming constituents, which must be fluxed, and any other ingredients which give rise to variations in cost of fuel and flux. The first of these criteria of value is perhaps not capable of being reduced to a general formula embracing all cases, since it must vary to a great extent with market-fluctuations. For example, Bessemer pig is worth sometimes more and sometimes less than foundry-pig ; and the premium paid for high manganese varies within wide limits. The second, or smelting-value, however, can be expressed numerically with a fair degree of accuracy; and it is with this phase of valuation that the present paper is concerned. The utility of a means of comparing numerically the relative smelting-values of different ores, cokes, and limestones, is self-evident, although it may not have importance in Northern practice, where materials are largely standardized. On the other hand, there are, especially in the South, many furnace-managers who must use a great variety of materials of widely varying grade; and it is with special reference to this district, and in connection with the efficiency-methods which I have elsewhere proposed,1 that the methods herein described were devised.

Mar 1, 1912

By F. N. Rhines, A. S. Nemy

The mode of high-temperature tertiary creep of 523-0 aluminum alloy was found to be strongly stress dependent. The occurrence of necking and/or fissures during tertiary creep exhibited a sequence with varying stress, wherein fissures with no necking occur at a minimum stress and necking with no fissures occur at a maximum stress. At an intermediate stress, tertiary creep begins without necking but with fissures in a region just beneath the external surface, followed by necking and the development of fissures along the axis of the specimen. The results are interpreted in terms of grain boundary gliding and the hydrostatic stresses produced by restraints to grain boundary gliding. THERE appear to be two possible reasons for the existence of a period of accelerating elongation prior to rupture during high-temperature creep in metals; namely, 1) a change in the structure of the metal itself, leading to a change in its response to loading, and 2) a reduction in the cross section of the metal resulting in a higher unit loading. Howe,' by annealing and retesting broken low-temperature creep specimens, found that irreparable damage is done once a metal is subjected to creep. Andrade2 first pointed out the association between the reduction in cross section and the increase in creep rate. The present research was designed to investigate both of these factors by retesting specimens machined from material which had been subjected to high-temperature creep testing short of rupture. EXPERIMENTAL METHOD An aluminum alloy, 52s-0, was selected for these studies because it is constitutionally simple, is strong enough to be relatively insensitive to mechanical damage during machining,3 and undergoes accelerating creep in a convenient temperature range. This alloy is hardened by 2.6 wt pct Mg in solid solution and is not subject to overaging effects; it has a very small quantity of an insoluble impurity phase, Mg2Si, finely dispersed throughout its micro-structure. Creep specimens, similar in design to ASTM 0.252-in. tensile specimens, were machined from 3/4-in. round stock and the surfaces were electro-polished. A 24-hr anneal at 500°C was then applied to stabilize the grain size to an average diameier of 0.27 mm. Subsequent heating for as long as 700 hr at the intended test temperatures produced no change in the tensile properties, indicating a well-stabilized state of the material. Corresponding to each of the typical testing conditions listed in Table I, one creep specimen was tested to destruction and this test was employed as a guide in the selection of times for subsequent tests. Duplicate creep tests were then run to various points along the creep curve to produce specimens which were used to determine changes in mechanical properties and microstructures. The latter were examined at the midsections of the specimens. Other specimens were tested to various stages of tertiary creep, after which they were remachined to new cylindrical test specimens, electropolished and re-tested in creep. RESULTS The major findings of this investigation may be illustrated by a comparison of the results of three sets of creep tests. These were run at 400°C at three stress levels, namely 750, 1000, and 1500 psi. In presenting these results in Figs. 1, 5, and 8, the creep curves are given in pairs: That curve of each pair which is drawn through filled circles corresponds to the initial test, and the curve drawn through open circles (and designated by - R) represents the result of retesting the same specimen after machining to a new cylindrical surface. The load was readjusted to the original unit stress at the beginning of each retest.

Jan 1, 1960

By P. R. Sperry, P. A. Beck, J. Towers

Dahl and Pawlek1 found that electrolytic copper develops extremely coarse grains at 1000°C after about 90 pct reduction by rolling. This coarsening occurs only under conditions of penultimate grain size, deformation, and alloying which lead to the "cube" recrystallization texture.l,2,3,5 The peculiar angular shapes and straight grain boundaries of the coarse grains were noted by several investigator.1,4,5 On the other hand, coarsening in Fe-containing aluminum or in AI-Mn alloys8 does not depend on a "cube" (or any well developed) recrystallization texture. It is true that increasing deformation by rolling, and, therefore, an increasingly well developed re-crystallization texture, are associated with decreasing incubation periods of coarsening.6-7-8 Nevertheless, coarsening readily develops in aluminum even after only 30 pct reduction by rolling, where the recrystallization texture is very weak.6,8 Also, coarsening was observed by Jeffries9 many years ago in sintered thoriated tungsten, which presumably has no preferred orientation. In all these cases coarsening is associated with grain growth inhibition by a dispersed second phase.8,9 The annealing temperature has to be suficiently high to overcome the inhibition at a few locations. But if it is too high, growth starts at many points, and the resulting grain size becomes much smaller.9 Normally, the coarse grains are more or less equiaxed, and the boundaries have a typical ragged appearance.6.8 Cook and Macquarie4 demonstrated that, in addition to the texture-dependent coarsening previously found at 1000°C,l electrolytic tough pitch copper may also coarsen at 800°C after 50 pct reduction by cross rolling. The coarse grains formed under such conditions have rounded shapes and ragged boundaries, like those in aluminum. When the annealing temperature is higher, the tendency for their formation decreases. All these observations suggest that the coarsening at 800°C is associated with inhibition by a second phase. Actually, coarsening at 800°C after 50 pct reduction by cross rolling was observed only in tough pitch copper,4 which contains Cu2O particles. On the other hand, the texture-dependent 1000°C coarsening occurs in both tough pitch and oxygen-free copper;4 it does not appear to depend on the presence of a dispersed second phase. However, the interpretation of the 800°C coarsening in Cu after 50 pct rolling as an inhibition-dependent process, similar to the coarsening in A1-Mn alloys, is somewhat weakened by the fact that this coarsening was reported4 to occur only after cross rolling, and not after straight rolling. It was, therefore, decided to re-examine this question. A 1 in. diam electrolytic tough pitch copper rod, No. 2 hard drawn, was annealed for 20 min at 700°C, rolled to 0.5 in., annealed 10 min at 700°C, and straight rolled to 0.064 in. It was then given a penultimate anneal of 20 min at 500°C and it was cut into four sections, which were given final reductions by straight rolling as follows: A 30 pct reduction of area B 50 pct reduction of area C 70 pct reduction of area D 90 pct reduction of area Specimens cut from the four sections were finally annealed at 800°C in an oxidizing atmosphere. Strip A remained fine grained up to 10 hr, but the specimen annealed 12 hr consisted of only 2 large grains. Strip B had a few scattered large (1/2 to 3/4 mm) grains after 1 min, although the balance of the specimen consisted of fine grains of about 0.02 mm. After 5 min there were several 10 to 15 mm grains present, and after 1 hr strip B was completely coarsened. The coarse grains had the same characteristics (see Fig 1) as those obtained by Cook and Macquarie at 800°C after cross rolling. Strip C had several grains of 0.05 to 1 mm after 1 min, but it was still largely fine grained after 12 hr. After 48 hr it consisted entirely of grains of about 0.5 to 4 mm, with an extraordinarily large number of twin bands. Strip D remained com- pletely fine grained after 4 hr at 800°C. These results indicate that, in the deformation range of 30 to 70 pct reduction, the incubation period for coarsening as well as the rate of growth and the final size of the coarse grains decreases with increasing deformation. Similar

Jan 1, 1950

By Charles Austin

IN an earlier series of studies on 18-8 stainless steel at 600°, 700°, and 800°C.1 one pretreatment was used throughout; viz., 15 min. in hydrogen at 950°C., followed by an air quench. Several manufacturers of this alloy since have expressed doubt as to the suitability of the temperature employed, recommending instead the more customary Strauss treatment as promising superior results. In this subsequent investigation on the effects of the temperature of pretreatment several interesting points have arisen, which are described in this paper. Previous work on our equipment had shown a marked difference between alloys of "similar chemical composition" that had been manu-factured by different companies. Accordingly, one mill-bar length was selected from which to prepare all specimens. It could then be assumed that any marked differences in creep characteristics would be due to effects resulting from variations in the pretreatment of the alloy. The material was furnished by the Allegheny-Ludlum Steel Co. in the form of 5/16-in, dia. rod, finished by air cooling from a mill anneal at 1950°F. (1066°C.). It was analyzed only for the elements shown: carbon, 0,07 per cent; chromium, 18.73; nickel, 9.46; manganese, 0.40. Three pretreating temperatures were chosen, the specimens being held for 15 min. in hydrogen at either: (1) 1150°C., to conform to the recom-mended Strauss treatment; (2) 950°C., to compare with the results of previous tests; or (3) 750°C., to initiate precipitation of carbides in the material before testing. As the test cross sections were only 1/8 -in. dia., air cooling followed the annealing treatment. These treatments per-mitted two direct comparisons: (1) between the 1150° and 950°C. treat-ments to indicate the effect of grain size when all carbides were probably in solution at the start of the test; and (2) between the 950° and 750°C. treatments to indicate the effect of carbide "nucleation" for similar grain sizes.

Jan 1, 1940

By W. F. Boericke

AT Kellogg, Idaho, J. Fred Johnson is driving the 5000.-ft Hooper Tunnel under contract for the Bunker Hill & Sullivan Mining Co. This was visited by a group of engineers during the recent meeting of the A. I. M. E. at Spokane, who were impressed with several features of the work. An excellent record of advance is being made by the contractor, over 47 ft. being driven in 3 days during October, and 223 ft. advanced in 29 shifts. The size of the tunnel is 8 by 9 ft. The ground is Wallace quartzite and, save for the first hundred feet, stands well and does not require timbering. The grade is held uniformly at 0.5 per cent. Two shifts are employed, with two sets of machine men to the shift. Ingersoll-Rand S-70 drills are used. each machine being set up independently in the heading; 26-30 holes are drilled to the round; using 1 1/4: in. round hollow drill steel. A 7-hole cut is required, and these holes are blasted simultaneously, using 60 per cent gelatine, 12 sticks to the hole. Instantaneous and fire delays of electric detonators are used for the complete round. A feature of the drilling is the use of long carriage drills, necessitating fewer changes of steel, and more time spent on drilling, with less time consumed taking steel back and forth from the face. Of course, this means better blacksmithing so that the drill steel shall stand up for the added footage.

Jan 1, 1930

By B. E. Mix, L. M. Barker

THe present methods of unloading ore and coarse-crushing at the Magna plant of the Utah Copper Co. are the developments of the pas five years. Hand dumping and breaking have given way to the rotary car dump and large gyratory; and the bucket elevators and stationary screens have been replaced by conveyors and shaking screens. The coarse-crushing plant may be broadly divided into two units the primary crushing and the secondary crushing, or primary roll, unit The former was erected in 1923; with the exception of the excavation the construction was carried on by forces of the company. Excavation was started in the middle of March, the first concrete was poured July 16, erection of the steel building started Oct. 30, and the unit was com pleted and operating Dec. 30. The unit has a capacity of over 24,001 tons of ore per day, and cost more than a million dollars. Remodeling of the primary-roll unit extended over a period of years preceding this some of the work having been done as early as 1918. Delivery of Ore from Mine, Yard Storage, and Handling Ori Cars to and from Primary Crushing Plant The ore is transported, in steel cars of 80 tons capacity, from the min' to the mill, where a reserve tonnage of loaded cars is maintained. Malle locomotives, handling trains of 45 cars, or approximately 3600 tons o ore per train, deliver the trains to a storage yard which contains thre tracks—a passing track and two storage tracks. All of this trackage i equipped with an overhead trolley system, so that electric locomotive can be used in yard switching. The storage tracks have a total length of 4500 ft. and provide storag space for 140 cars. The passing track, besides permitting switchin room for locomotives, is used for storage purposes between train arrivals thus giving intermittently a total storage capacity for the yard of approxi mately 200 cars. The amount of ore stored at any one time, howeve

Jan 1, 1926

By R. G. Clay, C. F. Seaman

THE mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that time, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered, the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. ORIGINAL EQUIPMENT AND INSTALLATION A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots and the up position would connect it to the line from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1942

By W. Bradley, R. G. Hall

This paper will attempt to show how a metallurgical problem at one California quicksilver mine was solved, and how the solution was applied successfully at another mine. The pronouns "we" and "our," as used below, refer '0 the Bradley 'lining Co. of San Francisco. The term "concurrent firing" for the purpose of this paper can be defined as feeding and firing a a kiln at the same end—that is, of course, the upper end' This is as opposed to "countercurrent firing," the orthodox method, in which a kiln is fed at the upper and fired at the lower end. The terms "quicksilver" and "mercury" are used interchangeably. Early References The earliest work on nonferrous metallurgy, we believe, was The Pirotechnia of Vannocio Biringuccio,l its first edition having been printed in Italy in 1540. Quoting from its chapter entitled "Concerning Quicksilver and Its Ore": Quicksilver ... is one of the gods and has divine strength in itself, and also, to the annoyance of the alchemists, it is winged. Hence, when it sees that it is in gravest danger, it loosens itself from . . . every bond . . . and fies away into the heavens, escaping from the hands of those who crucify it. Almost laughing, it leaves all its adversaries mocked and scorned, with their phials and filters empty . . . How thorough a job, under certain conditions, it can do of mocking its adversaries and leaving their phials and filters empty, was found out four centuries later at the Sulphur Bank mine, on the east shore of Clear Lake, in Lake county, California, The other mine mentioned in the title is the Reed, in Yolo County, California, 35 miles by road southeast of Sulphur Bank, Reed and Sulphur Bank ranked respectively Nth and seventh in U. S. production in 1943. The earliest direct reference we can find to furnacing troubles at these mines was by Thomas Egleston in 1890.2 Speaking of Sulphur Bank, Egleston says: of there is more sulphur than cinnabar in the ore, which is a detriment to both . . . making the sulphur impure, and rendering the cinnabar difficult to work on account of the soot. In some of the early . . . furnaces the accumulation of soot from the excess of sulphur has been known to penetrate as far as the fan blower at the end of the . . . condensers, and completely prevent its revolution. In order to get rid of the inconvenience of this accumulation . . . , as well as to get a commercial value from the sulphur, it (the sulphur) is now separated by steam and sold. It is claimed that the sulphur alone pays all the expenses of ... the mine. The mine was first worked for sulphur, and the cinnabar it contained was considered to be an inconvenience, and was carefully picked out and thrown away, as its richness and value were not

Jan 1, 1949

By Gerhard Derge, J. Warren Stewart

Extrusion processes are used in the commercial production of a wide variety of products, as indicated by the review presented a few years ago by D. K. Crampton.' Most writers have confined themselves to a general description of the equipment and products and the engineering aspects of the process. The few studies that have dealt with the metallurgical side of extrusion have been concerned primarily with an examination of the mode of flow of the metal during extrusion. This has been accomplished by the use of composite billets, by scribing grids on symmetrical sections of billets and observing the change of dimensions in the gridwork, or by examining microstructures in extruded sections.2"8 Ariotlier limitation of these original studies is that they were confined to the extrusion of very simple shapes, usually rods having a diameter of about one-fourth that of the billet. Tin is particularly well suited to the extrusion of collapsible tubes and similar shapes in which the extruded section is very small in comparison to the diameter of the billet or slug. In modern production, impact extrusion methods are generally employed, therefore the thickness of the slug is small compared to its diameter. Under these conditions the process differs considerably from the processes that have been most fully described in the literature, and it is known that the quality of the product is very sensitive to slight variations in any one of the many variable factors involved, such as coinposition and structure of the alloy, rate of extrusion, temperature, shape of punch and die. An exhaustive examination of all of the above-mentioned problems promises to produce so many data that it seems advisable to initiate the reports at this time, even though all of the factors outlined have not been examined fully, and to confine this paper to a description of the apparatus and method. Only tin and its alloys with small amounts of copper have been considered. These are the alloys most commonly used in this country for the production of the type of articles already described.

Jan 1, 1940