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By J. A. Coll, R. W. Cahn, A. Lawley

WHILE the creep properties of pure face-centered-cubic and close-packed-hexagonal metals have been thoroughly investigated and are well established, body-centered-cubic metals have been studied less extensively. Moreover, very few fundamental studies on the creep of solid solutions, irrespective of crystal structure, have been reported. The present study is concerned with the creep of a series of body-centered-cubic solid solutions. The present position concerning creep of pure metals is, briefly, as follows.1"3 Creep at first takes place at a steadily decreasing rate; this is the stage termed primary or transient creep. Except at the lowest temperatures this is succeeded by a stage of secondary or steady-state creep. At high temperatures and stresses, this may be succeeded by an accelerating stage, termed tertiary creep, with which we shall not here be concerned. There is no well-defined physical model at present for the transient stage; in general terms, transient creep is best regarded simply as a manifestation of work-harden ing. Steady-stage creep can certainly take place by several different mechanisms: the choice of dominant mechanism depends primarily on temperature. We shall here be concerned only with high-temperature steady-state creep, a term usually reserved for creep at absolute temperatures higher than 0.5 Tm, where T, is the melting point. In this range, the activation energy for creep is, for many metals, equal to the activation energy for self-diffusion, and this is generally interpreted in terms of a "climb mechanism.1-4 The creep rate is determined by the speed at which dislocations, impeded by obstacles the nature of which is disputed but which are probably established during transient creep, can climb by means of a diffusion process, until they are able to by-pass the obstacle. In solid solutions, the intrinsic resistance to the slip motion of dislocations may be much larger than in the solvent, to the extent that the motion of dislocations in the glide plane, rather than their escape by climb out of this plane, may become the rate-controlling factor. weertman5 has considered this possibility from a theoretical point of view, and concluded that some form of "viscous slip" is likely to be rate-controlling at comparatively low stresses. The resistance to slip may arise from "atmospheres" of impurities forming around dislocations; a high Peierls force in materials of high cohesion; or some structural peculiarity such as clustering or ordering of solute atoms.= We shall be concerned here with the case of ordering. The only published investigations concerned explicitly with the effect of order on creep refer to creep in ß-brass by Herman and Brown,7 and in Ni-Fe alloys, by Kornilov and panasyuk8 and by Suzuki and Yamamoto.9 Recently, Herman and Brown's paper has been supplemented by a determination of the tensile yield point of ß-brass as a function of temperature.10 Both studies showed a sharp drop in resistance to deformation of ß-brass over a range of a few degrees just above the critical temperature Tc at which order finally disappears. These observations are especially noteworthy, because in ß brass the degree of order diminishes steadzly to zero as the temperature approaches Tc. It is, therefore, the disappearance of the last traces of long-range order which has the largest effect on the resistance of the alloy to plastic deformation. In the Ni/Fe alloys of various compositions, resistance to creep at a given temperature and stress is maximum at the stoichiometric composition, both below Tc, (long-range order), and above T,. (short-range order).' Near Tc, the creep resistance of an ordered alloy is much higher than that of the same alloy in the disordered condition.9 The aim of the present investigation was to study the creep behavior in the neighborhood of Tc, of another system of ordering alloys. The iron-aluminum alloys were considered the most suitable. because: i) The order again diminishes steadily to zero as the temperature approaches Tc; there is no sudden drop in order at Tc, and it therefore is

Jan 1, 1961

By D. L. Douglas

The ß decomposition of Zr-U-O alloys was studied during an interrupted quench from the a + ß region. Decomposition was more rapid than in binary Zr-U alloys of the same uranium content or qf equivalent ß composition. The rapid transformation has been ascribed to the presence of the primary a phase which provided nucleation sites and which permitted preferential partitioning of oxygen to the a phase and of uranium to the COMMERCLAL heat treatments are designed to produce certain strength and toughness character- istics by controlling the microstructure of an alloy. A desired distribution of microconstituents is achieved in practice by treating an alloy under the correct time-temperature conditions. A knowledge of the phase equilibria in any alloy system is essential in outlining heat treatments, but in addi-

Jan 1, 1961

By Leo F. Reinartz

This is the fourth and final installment of Leo Reinartz' summation of the steelmaking industry over the first one hundred years. ALTHOUGH basic open-hearth steel has been in the limelight for many years, it has had to share some of its glory with two lesser but nevertheless important rivals, one 50 years old and one a rejuvenated process. Electric steel Electric-steel manufacture began in a humble way in the United States in 1906, when the first heat was tapped from a 3-ton Heroult electric furnace at the Halcomb Steel Co., Syracuse, N. Y. During the intervening years, it has completely eliminated the crucible process. For many years electric melting furnaces were located predominantly in relatively small, nonintegrated plants, making all kinds of high-grade alloy steels. Most of these furnaces, with capacities ranging from 5 to 20 tons, had basic bottoms. During and after World War II, many large basic electric furnaces, some up to 200 tons in capacity, were installed in integrated plants to make commodity steels. There were several reasons for building electric furnaces. Capital investment per annual ton of capacity was low and furnaces could be installed quickly. Since these furnaces used mostly scrap iron in the charge, no extra expensive coke- and blast-furnace capacity had to be added. The perfection of the top-charging design, and the development of large transformers and furnaces, made possible tons per hour as high as from modern 350-ton. open-hearth furnaces. Electric furnaces have high availability and repair costs are low. The process is flexible. Alloy and plain carbon steels of high quality can be made. Where power costs are low, the operating cost can

Jan 1, 1961

Admissions: Joseph C. Gordon, Chairman K. F. Anderson Allen B. Gibson H. J. Gruy R. F. Madera Dan H.-McLendon K. W. Robbins Aurel E. Smith W. F. West Advertising: John. D. Wisenbaker, Chairman Bernard F. Klein Henry N. Lyle R. E. Howard Clark R. Edgecomb, Jr. [ ]

Jan 1, 1961

By P. D. Rao, H. B. Charmbury, D. R. Mitchell

As a result of a changing market pattern for anthracite, there is now an increasing demand for steam size and a decreasing demand for domestic sizes. To help the producer meet this new demand, the authors present the technical aspects of the problem of crushing domestic sizes to steam fineness and size. The market pattern for anthracite coal has changed considerably in recent years. The demand for domestic sizes has been gradually decreasing and the demand for the steam sizes has been increasing. The companies however aim for a maximum production of domestic sizes, since these sizes command premium prices. To meet the market demands, particularly in certain periods of the year, further size reduction of the coarse clean coal is necessary. The present practice of multistage crushing in toothed rolls leads to complicated flow arrangements and is high in cost. Data presented herein indicate that crushers with high reduction ratios could be used with a resulting simplification of crushing flows and an appreciable reduction of capital and operating costs. The different types of crushing equipment investigated included the rotary-type breaker, the jaw crusher, the impactor-type crusher, and the Aerofall mill. CURRENT PRACTICE The standard sizes of anthracite are presented in Table I and market specifications require that any given size be maintained with only a limited amount of oversize and undersize material. The 1958 prices of the various sizes also are presented, so it may be seen that size reduction is only performed when dictated by market demands. At the present time the general practice of crushing anthracite consists of removing the top size and crushing it in a toothed double roll crusher to the next largest size. Consequently, to reduce the domestic egg, stove, and nut sizes to steam sizes of pea, buck, rice, and barley, the original material must be screened to produce the three top sizes and each size must be crushed in three separate rolls to the next largest size. Each crushed product must then be rescreened to remove the oversize for re crush ing. Thus a typical operation would consist of scalping the egg size from a mixture, crushing it to stove size, and rescreening the product. This same cycle would be repeated for the stove and nut sizes requiring three screens and three crushers as well as conveying equipment to handle the crushed products. This results in an expensive series of operations to make a product which sells for a lower price. If the three top sizes are crushed in one operation to the smaller sizes then the yield of steam sizes is low and the production of Nos. 4 and 5 sizes is large. These sizes sell for lower prices than the steam sizes. A typical operation is presented in Fig. 1. From the figure it may be seen that the original cleaned anthracite contains 67.4 pet domestic sizes, 26.4 pct steam sizes, and 6.2 pct fines. After multistage crushing of the domestic to steam sizes the composition of the product is 0.6 pct domestic sizes, 89.6 pct steam sizes, and 9.8 pct fines. From a technical standpoint these results are reasonable but the handling and crushing costs are

Jan 1, 1961

The development and introduction of a preventive maintenance scheme in the coke ovens department, coal preparation section, of The Broken Hill Proprietary Co. Ltd., Newcastle steelworks, is described together with its operation and the results achieved during the first 12 months.'fhe scheme consists essentially of an engineering analysis of the plant to establish maintenance schedules, and to record the details on reference cards.Items are reviewed progressively and a weekly work schedule is prepared, from which work orders are passed on to the maintenance foremen. These foremen in turn originate action orders and issue inspection sheets as required. Results are then reported back to the maintenance planner and used to revise norms, practices, and cycles.INTRODUCTIONWhat is maintenance? Some people consider maintenance to be the work involved in putting things together again after they have fallen apart. The author regards it as the mechanical care and repair of plant machinery and equipment to keep it operating in sound mechanical condition, without undue interference with production processes.Maintenance may be organized or haphazard, and, too often, what has been fondly imagined to be nicely controlled and organized maintenance proves, under pressure, to be more correctly described as haphazard. Haphazard maintenance deals with breakdowns as they occur, and frequently finds itself confronted with more breakdowns than can be handled. Even organized maintenance, if not based on sound planning, can, at times, be found in a chaotic state, because of unexpected and unanticipated breakdown.Properly planned and scheduled maintenance permits an orderly and controlled approach to the problem of maintain...

Jan 1, 1961

By L. W. Reynolds, R. F. Campbell, K. G. Carroll, S. H. Ballard

Based on metallographic and X-ray data probable equilibrium conditions from 1340" to 2400°F are presented for the composition range investigated. These are correlated with investigations of Takei and Kuo. Provision is made for existence of Mo, C and (Fe, Mo)2,C, carbides heretofore omitted in diagrams in this composition range. These carbides and (Fe, Mo),C, in equilibrium with ferrite and austenite, restrict the austenite and austenite + ferrite fields to lower carbon contents and higher temperatures than in the Fe-C system. OK groups of steel of different molybdenum content, but with approximately the same carbon variation in each group, Table I, were studied. The alloys were melted in an acid-lined 30-lb high-frequency induction furnace, and the ingots forged to l-in. rounds, which were then ground to 0.75-in. diam. To minimize chemical segregations a short length of each bar was given a high-temperature heat treatment as listed in Table I. Specimens 0.062 to 0.125-in. thick were thereafter cut from the homogenized bars for heat treatment to determine phase changes in the alloy. Heat treatments were performed in either a lead bath or muffle-type electric furnace, depending upon the temperature required. Maximum temperature variation was * 2" F in the lead bath and * 5"F up to 1900°F and +8"F above 1900°F in the electric furnace. Adequate protection of the specimen from the oxidizing atmosphere in the electric furnace was obtained by sealing the specimen in either a pyrex or a silica capsule which was evacuated or not as circumstances required. For austenitizing followed by quenching to an isothermal temperature, the specimen was placed in a closed-end refractory tube the mouth of which, extending beyond the furnace, was connected to a vacuum pump. The specimen could be withdrawn from this tube and quenched as desired. IDENTIFICATION OF PHASES The principal method of identification of phases was metallographic examination conducted on surfaces well removed from exposure to lead or atmosphere effects during heat treatment. Hardness measurements (Dph/lOkg) were also made on these surfaces. The etching reagents used and their effects are listed in Table II. Martensite and ferrite were readily recognized by their customary appearance after etching with saturated picral or 1 pct nital. The presence of liquid at high temperature was indicated by curved interfaces characteristic of boundaries between solid and melt, by intergranular pools of dendritic areas in a carbide matrix, and by a black network coincident with austenite grain boundaries. Examples of these structures are shown in Figs. l(a) and (0). The types of carbide which were attacked by the electrolytic chromic acid etch and the alkaline KMnO, etch, Table 11, were identified by X-ray analyses of extracted carbides of five alloys. Carbide residue was obtained from a metallographic specimen by etching the specimen for 24 hr in hydrochloric-picric acid in alcohol (grain-size reagent), after which the carbides, standing in relief, were scraped from the surface of the specimen and mounted on a glass slide. Diffraction patterns were obtained with a Philips spectrometer using iron-filtered cobalt radiation. Figs. 2, 3, 4, and 5 illustrate typical etching behaviors with chromic acid and alkaline KMnO,. Chromic acid attacked Mo,C as shown in Figs. 2(a) and (b)-but not (Fe, Mo),C, Fig. 3(b). Alkaline KMnO, attacked (Fe, Mo),C, Fig. 3(c), and also attacked M0,C. In accordance with these behaviors, a typical differentiation of coexisting Mo,C and (Fe, Mo),~ is shown in Figs. 4(a) and (b). The carbide (Fe, Mo),,C, could not be obtained free of other coexisting carbides for etching experiments. The behavior. however. of coexisting carbides identified by x-ray diffraction as Mo,C ind (Fe, Mo),,C, indicate that (Fe, Mo)23C6 is not etched by chromic

Jan 1, 1961

By H. J. Modi, D. W. Fuerstenau

pH effectively regulates the flotation of corundum through its control of the surface charge. Since collector ions function as counter ions in the double layer, the collector must be anionic when corundum is positively charged and cationic when it is negatively charged. Even a long-chained xanthate will float positively charged corundum. Association of the hydrocarbon chains of the adsorbed collector ions gives rise to increased adsorption and increased rate of flotation. Because of this phenomenon, neutral organic molecules aid flotation although the unionized collector cannot promote flotation independently. Na2SO4 was found to depress the flotation of corundum with animic collectors through ionic competition whereas the same salt activates corundum for flotation with cationic collectors. A close connection between ion association, electrokinetic behavior, and floatability is suggested. As part of the increasing interest in the physical chemistry of flotation, strong experimental evidence has been obtained which indicates that organic ions are adsorbed as counter ions by non-metallic oxide minerals mainly because of the electrical forces which exist at interfaces(1-3). Thus flotation of such minerals without activation should then be appreciable only when the mineral and the collector are oppositely charged. Since only limited work has been reported on the effect of surface charge upon flotation behavior(3) this general hypothesis on the flotation of insoluble oxide minerals has not been tested thoroughly. Investigating the flotation of a mineral for which extensive electrokinetic data exist yields a way to test this hypothesis more extensively. The electrokinetic behavior of corundum in solutions containing both organic and inorganic electrolytes has been studied quite thoroughly' Electro-kinetic measurements have shown that in the case of corundum, simply by controlling pH the solid surface can be charged either positively or negatively. For this reason corundum was chosen for study. The object of this paper is to present a study of the flotation of corundum and to explain the results in the light of our knowledge of adsorption mechanisms gained from streaming potential investigations. This paper is concerned mainly with the effect of surface charge on flotation with anionic and cationic collectors, but in addition such flotation phenomena as collector hydrolysis, Van der Waals attraction between hydrocarbon chains, ionic competition, and activation are discussed. While the present paper is confined to a reporting of the data obtained and some of the more immediate conclusions, it will be seen that the results, in combination with those on other systems, do promise a

Jan 1, 1961

By J. M. Lommel

DURING an investigation into the effect of heat-treatment on the creep properties of the magnesium alloy ZW1, (1 pct Zn, 0.6 pct Zr), the previously published methods of final polishing were found to be unsatisfactory. A new diamond-polishing technique has been developed to facilitate the study of fine precipitates in this and other magnesium alloys, and is described in this note. The sectioning and coarse grinding techniques are those of Fox and Hall,' and Hess and Cieorge.' Pre- liminary mechanical polishing is done with 6-µ, and 1-µ( grades of diamond compound on 'Microcloth' using a lubricant of 'Hyprez' oil. When the conventional technique is used with a 1/4- µ diamond wheel for the final polishing of magnesium alloys, scratching is likely to result from the abrasive action of the 'Microcloth' nap on the soft metal surface. This problem has been overcome by the use of a cream on the wheel in order to separate the specimen surface from the polishing cloth. To prepare this cream, a solution of 6-ml tri-ethanolamine in 75-ml water is stirred into 12.5 g stearic acid at a constant temperature of 80o to 90°C. The cream is allowed to cool, and is then 'whipped' to a smooth consistency with 100-ml 'Hyprez' oil. Before polishing, the diamond loaded 'Microcloth' is moistened with a few drops of 'Hyprez' oil, then a small amount of cream is worked into the nap. For the best results a wheel speed of 500 rpm should be used for polishing, and the specimen should be held by hand with a light pressure while being rotated in the opposite direction to the wheel. After polishing it is necessary to swab the specimen thoroughly in methylated spirits to remove residual cream and oil from the surface in order to avoid uneven etching and drying stains. Fig. 1 shows the surface finish obtained on a specimen by polishing on a 1/4- µ diamond wheel using the cream, and Figs. 2 and 3 the precipitate observed in ZW1 after polishing, and etching for two seconds in nitric acid 15 pct in ethylene glycol. This new diamond-polishing technique has made a valuable contribution to the programme of metal-lographic research on ZW1 and has become a routine laboratory practice for other low alloy content magnesium alloys.

Jan 1, 1961

By Kenneth M. Haley

THE development of Reserve Mining Company's pelletizing practices began in 1949 in a pilot plant located at Ashland, Kentucky. Three years later the participating companies -Armco Steel Corporation, Cleveland Cliffs Iron Mining Company and Republic Steel Corporation concluded the work and the small plant was shut down. Oglebay Norton and Company was the operating manager and Armco Steel Corporation furnished labor and services. Arthur G. McKee and Company were the designers and constructors. The equipment flow sheet that was employed in that pilot operation is shown in Figure 1. This was Reserve's first step from the laboratory work per-formed by the Mines Experiment Station of the University of Minnesota in developing a commercial operation. The plant used New York state concentrate, purchased from Jones and Laughlin Steel Corporation's Benson Mine. This concentrate, minus 14 mesh, 64.3 per cent Fe, 5.7 per cent Si02, was processed at Ashland to 80 per cent minus 325 mesh in conventional ball mill, classifier, thickener, filtering

Jan 1, 1961

By B. Paul

1. INTRODUCTION It is known that the elastic modulus of a metal may be considerably increased by dispersing

Jan 1, 1961

By W. H. Porter, M. E. Wadsworth, J. R. Lewis, K. L. Leiter

The oxidation of Cu2S in oxygen and the sulfating of Cu2O in oxygen-sulfur dioxide atmospheres was carried out under a variety of conditions. The oxidation of Cu2S was found to be retarded by entrapment of SO, and O2, which stabilized internal sulfates for long periods of time. The course of the reaction was followed by measuring weight changes and also by SO, evolution. Sulfating of Cu2O was a maximum at ratios of SO, to O2 approximating maximum SO, production. At elevated temperatures SO, was found to increase the rate of oxidation of Cu2 O to CuO even though sulfates did not form. All sulfating reactions followed the parabolic rate law indicating diffusion. MANY studies of the roasting of copper sulfides have been reported in recent years. Diev et al.1 investigated the roasting of chalcocite (Cu2S) in air, and oxygen enriched air. Lewis et al.2 also studied the oxidation of natural and synthetic chalcocite in air and oxygen atmospheres and their studies indicated that the maximum formation of water soluble sulfates occurred at approximately 450oC. Ashcroft3 reported that oxide production during the roasting of chalcocite resulted only from secondary decomposition of sulfates which were formed as primary products. peretti4 refuted this claim by showing that a layer of Cu2O appeared directly adjacent to the Cu2S during roasting of cylindrical briquettes of cupric sulfide, CuS. The linear advance of the Cu2S-Cu2O interface was used as a measure of the kinetics of the roasting reaction. The reactions proposed were: 2 CuS—Cu2S + 1/2 S2 [4-1] 1/2 S2 + O24 SO2 [4-2] cuzs +3/2 O2 4 Cu2O + SO2 [4-3] cu2o + 1/2 O2—2 cuo [4-4] At temperatures above 663oC, CuO was the only final solid phase reported. Below 663" C increasing amounts of sulfate were found mixed with the CuO. McCabe and Morgan5 investigated the roasting of discs of synthetic chalcocite and reported the following sequence of products beginning at the sulfide surface: Cu2O, a mixture of Cu2O and CuSO4, Cum,, CuO . CuSO4, and CuO. The principal reactions were reported to be: Cu2S + 3/2 O2-Cu2O + SO2 [5-1] CU2O + 2 SO2 + 3/2 O2—2 CUSO4 [5-21 2 CUSO4— CUO . cum, + SO3 [ 5-31 cuo . cuso4—-2 cuo + SO, [ 5-41 Eq.15-11 supports the claim of Peretti, Eq. [4-31, that CuzO is formed directly from Cu2S rather than as a secondary product from a sulfate as suggested by Ashcroft. On the other hand CuO was found to form as a secondary product from the decomposition of copper sulfate and basic copper sulfate, Eqs. [5-31 and [5-41. The formation of sulfates was explained by McCabe and morgan5 to be a direct reaction of Cu2O with 0, and SO, or SO, at distinct regions in which the partial pressures of each were such as to form the sulfate. Thornhill and pidgeon6 roasted both natural and synthetic chalcocite grains in air at temperatures between 420" and 550° C. They found a dense primary oxidation layer in contact with the sulfide. A secondary layer of porous oxidation products was found to expand with roasting time. The oxide products were leached away and the remaining core was studied by X-ray diffraction. The X-ray patterns showed an increased conversion of chalcocite to digenite with time. Digenite,7 a defect structure of cuprous sulfide, occurs naturally as Cu,-,S where x = 0.12 to 0.45, with an average analysis of Cu, ,S. The mechanism of digenite formation was proposed as: Cu2S + oxygen—Cu1-8S + 0.1 Cu2O [6-1] Cuj.eS + oxygen—0.9 Cu2S + SO2 [6-2] It is apparent from the above studies that the oxidation of Cu2S, ultimately ending in CuO, may be divided into ihree general stages (all of which may occur simultaneously): 1) primary oxidation to Cu2O; 2) secondary sulfate formation; and 3) sulfate decomposition. Consequently reactions of O2 and SO, with Cu20 constitute important aspects of the roasting of chalcocite. Virtually no studies have been made regarding sulfating reactions involving Cu,O. Mills and Evans8 noted the effect of sulfur dioxide on the oxidation of copper at low temperatures and low SO, partial pressures. They reported a measurable increase in the oxidation rate of copper when SO2, was present. Interest in the Cu2O-CuO-0, system has been limited predominantly to misciblllty studies and determinations of heats of formation by

Jan 1, 1961

By Lionel A. York

THE first application of the cementation process took place in the year 1864, according to H. Standish Ball, when a break in the brickwork lining at a depth of 270', in one of the Rhine Preussen pits near Homburg, was sealed by running in milk of cement from the surface by means of a small hand pump. At about the same time Monsieur Portier adopted the method in France for repairing breaks in shaft lining. Figure 1 is descriptive of the Portier system and is of interest when compared with modern practice. During the next thirty years numerous attempts were made to improve the process; the general method used in sinking being to put clown bore holes outside the perimeter of the shaft for the purposes of injection. The cementation process was first recognized as an important as-set to mining in 1896 when Monsieur A. Francois developed his method of boring and injecting inside the shaft below the permanent water level. The Francois cementation process rapidly attained fame on the Continent and many shafts were sunk by its use in France, Belgium and

Jan 1, 1961

By J. Nutting, D. H. Kirkwood

IT is clearly of great metallurgical interest to know the conditions of temperature and gas composition under which different products will form in sulphide roasting. A method of graphically presenting the relevant thermodynamic data will be demonstrated, such that these conditions are readily obtainable. The products of sulphide roasting may be oxide, sulphate, or oxysulphates, and the equilibria between these may be represented in the following way: 2 MSO4 = MOMSO4 + SO3 MO-MSO4 = 2 MO + SO3

Jan 1, 1961

By A. Gatti

A dispersion of AI2O3 was produced within chromium by the hydrogen reduction of a solid solution of Cr2U3-A1203. The yesulting powder. was compacted, extruded, and the material tested in tension at various temperatures and in creep at 1800°F. The material is ductile above 1200°F and is at least twice as strong as electrolytic chvomium at all temperatures. Its creep resistance, rupture strength, and ductility at 1800°F warrant further work to improve its Low -temperatutre properties. BECAUSE of its high melting point and excellent corrosion resistance, chromium has long been considered as a base for high-temperature alloys. Most chromium base alloys, however, are extremely brittle at low temperatures. Elements such as nickel and iron lower the melting point of chromium and thereby limit its high-temperature usefulness. Chromium-alumina materials made by othersl, 2 have shown good high-temperature creep resistance but poor ductility and thermal shock resistance. Pure ductile chromium3 is weak at high temperatures and reverts to a brittle state if exposed to nitrogenous atmospheres. In sintered aluminum powder (SAP)4 and internally oxidized alloys of silver, copper, and nickel, fine dispersions of oxide phases can control hardness, recrystallization, and yield stress. The possibility of formation of a ductile, high-strength material by precipitating dispersions of alumina in a chromium matrix appears worthy of investigation. MATERIAL Powder Production—Chromium oxide and alumina in amounts to produce a final alloy of Cr + 16 pct AI2O3 by weight were colloid milled, filtered, dried and high fired for 16 hr at 1600°C in air. The resulting solid solution of oxides was ball-milled to a fine powder and the chromium oxide reduced to chromium by heating in pure dry hydrogen for 64 hr at 1400°C. Final Materials Produced by Powder Metallurgy Techniques—-Chromium powder, containing a disper-sion of alumina, was pressed into 1 1/8-in. diam slugs at 25 tons per sq in. and sintered 16 hr at 1400° C in pure dry hydrogen. The sintered slugs were encased in a stainless steel can, 2 in. in diam, and extruded at 1200° C through a 1/2-in. die. The 16 to 1 extrusion ratio produced a final chromium-alumina rod 1/4 in. in diam about 30 in. long. Samples of the rod were plunge-ground to small rupture bars of 0.82-in. gage length with a 0.01-sq-in. cross section. RESULTS Microstructure Studies—Fig. 1 is the microstruc-ture of the material in the as-sintered condition while Fig. 2 is the microstructure of the material as-extruded. Comparison of the as-sintered and as-extruded micrographs show the redistribution and realignment of the alumina inclusions by the extrusion

Jan 1, 1961

By F. D. Hoyt, H. L. Hartman

The development of new processes and methods by The Pennsylvania State University to improve slate quarrying technology has centered in recent years on cutting and sawing stone in the quarry to eliminate a second cutting process in the mill. Two machines exhibit promise for this work: 1) a circular saw mounting diamonds or hard inserts to produce smaller sizes of stone and 2) a chain saw with insert cutting teeth to produce stone in the larger dimensions. Prototype machines have been constructed and tested in several Pennsylvania slate quarries, and one commercial installation has been operated for several months with a circular diamond saw. Other kinds of dimension stones may be cut by these saws. Research at Penn State has begun to study the fundamental cutting action of rotary tools or saws in slate and other dimension stones. A laboratory drill press is being instrumented to permit thrust-torque-rotational speed us penetration rate studies of single tooth cutting surfaces on stone. Machinability studies of slate conducted with tungsten-carbide inserts have been performed. The dimension stone industry generally accepts the rather basic premise that the larger the block removed from the quarry, the more practical and economical the operation. Thus, the concept of cutting to size any dimension stone while it remains in place in the parent bed would receive little consideration from the majority of members of the industry. However, the slate industry, which is usually considered a separate member of the dimension stone family, is pioneering in the development of an in-place sawing method. Before any final decision can be reached concerning a proposed new system, it is essential to take a long, hard look at the present method of operation in order to determine if the new system is indeed an improvement or even desirable. In the following section is a brief description of present quarry practice in the slate quarries of eastern Pennsylvania. PRESENT METHOD OF QUARRYING SLATE In the numerous slate quarries of Lehigh and Northampton Counties of Pennsylvania, the grain and cleavage of the slate are most often at right angles to each other; if a third surface is broken at right angles to these two natural planes of weakness, blocks of more or less rectangular shape can be separated.' In conventional quarrying a large calyx core drill prepares holes of 36-in. diam in which wire-saw standards are positioned. By wire sawing between strategically located core-drill holes, large sinks or benches of virgin slate are opened up. The sides freed by wire sawing will vary from quarry to quarry but generally are rectangular in shape with dimensions averaging 20 ft in length and about 15 ft in depth. Some quarries are fortunate in having a joint or natural parting to work to, which of course diminishes the amount of core drilling and wire sawing required. Once the various benches have been developed either through wire sawing alone or through a combination of wire sawing and natural jointing planes, the block removal proceeds in the following manner. A plug hole is drilled in the block with a compressed-air hammer and a feathering chisel is inserted in the hole to cause a fracture of the rock either with or against the grain as indicated by the positioning of the so-called feathers. This operation is referred to as sculping. After a block has been freed with and against the grain by means of wedging from the cross fracture with long bars or levers, the block is further freed along the cleavage plane by shimming it up with small wooden pieces in an operation known as styling. A large steel loading chain is wrapped around a block thus freed and the chain is attached to the wheel of an overhead cable. The block is then hoisted vertically to the cableway and moved along this cableway laterally to the lip or edge of the quarry. From here it is unloaded onto a rail car or truck for transportation to the processing mill. Except for occasional blasting to free stubborn blocks,

Jan 1, 1961

By J. S. Browning

The U.S. Bureau of Mines has been experimenting with flotation processes to separate the spodumene-beryl ores mined at Kings Mountain, N.C. The success to date as well as the present status of the process is discussed. Further experimentation is underway and studies will be made on the milling costs involved in the flotation process. The pegmatites of the Kings Mountain-Lincoln ton, N.C., area constitute the largest known domestic reserve of beryl and spodumene. The reserve is estimated to contain 90 million tons of pegmatic material with 1,280,000 tons of recoverable Li,O(lithia) as spodurnene.' The pegmatites also contain 0.4 to 0.5 pct beryl disseminated throughout the orebodies. The pegmatites may contain a potential reserve of 240,-000 tons beryl, equivalent to approximately 34,000 tons of BeO. Different flotation methods for separating beryl from feldspar or quartz, or both, have been developed by various investigators.2-7 On the other hand, published information is limited on concentration of spodumene-beryl ores. As the response of beryl and spodumene to flotation is essentially the same, a successful separation of the two minerals depends upon use of a selective depressant or a selective collector for one of the minerals. Several years ago, the U.S. Bureau of Mines, recognizing the need for developing a large domestic supply of strategic beryl, undertook studies to develop improved methods of recovering this mineral. Because of the low beryl content of the pegmatites, economic production of the beryl entails its recovery as a byproduct of spodumene flotation. Thus, the development of procedures that would provide for maximum recoveries of the spodumene, as well as the beryl, was a primary objective of the studies. Tests were made on run-of-mine ore and spodumene flotation tailing from the Foote Mineral Co.'s spodumene concentrator at Kings Mountain (Fig. 1). Pet-rographic analyses of the ore and flotation tailing are given in Table I. Detailed analyses and examinations of the ore and tailings revealed that the beryllium was present in the form of a low-alkali beryl. The crystals were small, rarely larger than 14-mesh diam, and were clear and colorless. About 10 pct of the Li20 content of the ore was present in the mica and feldspar components of the pegmatite. In some weathered ores the associated clays also contained small amounts of lithia. This is typical of pegmatites in the Kings Mountain area. LABORATORY TESTS In the past, a vigorous chemical treatment of the spodumene surfaces with acid or caustic soda has been considered essential for satisfactory selective flotation of the spodumene from the other minerals in pegmatites. Many tests were made during this investigation to determine if chemical treatment of the spodumene could be eliminated in favor of a reagent combination that would depress the other minerals while selectively floating the spodumene. A simple reagent combination and process was developed that gave better spodumene recovery and grade of concentrate than the more elaborate acid or caustic treatment methods previously used. The finely ground ore pulps were conditioned with either an ammonium, alkali, or alkaline earth lignin sulfonate and sodium fluoride, and the spodumene was floated and cleaned using oleic acid as the collector. The spodumene tailing was deslimed, conditioned with sulfuric acid and coco amine acetate, and floated to reject mica. The mica tailing was thickened and conditioned with hydrofluoric acid, then washed to remove the acid. The washed pulp was conditioned with sodium hydroxide for pH control, and the beryl was floated and cleaned with oleic acid as the collector. Table II summarizes the results of the laboratory test work. About 77 pct of the spodumene and 75 pct of the beryl were recovered from the pegmatitic material. The spodumene concentrate assayed 6.1 pct Li20 and 0.01 pct BeO, and the beryl product assayed 3.0 pct Li20 and 1.57 pct BeO. Similar results were ob-

Jan 1, 1961

[A. B. Cummins, President James C. Gray, President-Elect J. W. Woomer, Past President C. E. Lawall, Eastern Regional Vice-President Donald W. Scott, Central Regional Vice-President H. C. Weed, Western Regional Vice-President Nathaniel Arbiter, Treasurer John Cameron Fox, Secretary AIME 29 W. 39 St. New York 18, N. Y. D. R. Tone, Assistant Secretary]

Jan 1, 1961

By D. W. Fuerstenau, D. A. Sullevan

In the experimental work for this comparison of wet and dry grinding, it was found that the size distributions for wet grinding operations are characterized by a constant value of the distribution modulus, (a; whereas, for dry grinding the distribution modulus decreases slightly with increased size reduction. Wet grindability seems independent of the weight of material in the mill while dry grindability depends on weight. Experimental data and eqmtions for each process are presented for noncleavable isotropic material, for cleavable or material in which there is some form of internal stress, and for energy consumption. The equation ED/EW=BK-? (where ED and Ew are, respectively, energy for wet and dry grinding, k is the size modulus, and B and ? are constants) was found to give the ratio of energy. Dry grinding is generally considered to require more energy than wet grinding. Taggart' states that power consumption in wet grinding ranges from 60 to 90 pct of that required for dry grinding. Bond' states that dry grinding requires one third more power than wet grinding. It is interesting that Rose and Sullivan' note this same difference in power required for wet and dry grinding, but they state that on purely mechanical grounds it is difficult to see any great difference in fundamental principles between wet and dry milling, since dry milling may be regarded as wet milling with a fluid of low viscosity (i.e., air vs water). Yet it must be the viscosity of the fluid that is responsible for the differences between wet and dry grinding. Probably the most comprehensive investigation comparing wet and dry grinding has been that made by Coghill and Devaney. They compared the grind-ability of both chert and dolomite in a 19x36-in. batch ball mill under both wet and dry conditions. In these tests the weight of ore in the mill was varied from 35 to 200 lb and each grinding test was carried out for such period of time that the amount of -200-mesh material in the product was essentially the same for each test with a given material. Cog-hill and DeVaney's data for the grindability of dolomite, which are reproduced in Fig. 1, show that a comparison between wet and dry grinding is com- plex. The comminution of a material by a wet grinding method is characterized by a constant grindability over a wide range of mill loadings. Since grindability is the weight of material comminuted per unit of energy expended, it is equivalent to (w)/ (p)(t), where W is the weight of material being ground, P is the power, and t is the grinding time. In the case of wet grinding, the grindability of dolomite in the Coghill and DeVaney experiments was 0.546 * 0.009 ton per hp-hr when the load of dolomite in the mill ranged from 50 to 200 lb. On the other hand, for dry grinding no such simple relationship was found to exist between power, time, and weight of material in the mill. Dry grinding efficiency appears to depend greatly upon the amount of material within the mill. When the mill contained

Jan 1, 1961

[Alphabetic Section (Page 729): The letter following the member's name indicates grade: M-Member, A-Associate, J-Junior. The numerals following member's grade indicate year of election. The letter following year of election indicates primary Divisional affiliation: A-Mining & Exploration F-Coal B-Minerals Beneficiation G-Petroleum C-Iron and Steel H-Industrial Minerals D-Extractive Metallurgy J-Council o f Education E-Institute o f Metals K-Council o f Economics]

Jan 1, 1961