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By Willis T. Burns

I. Introduction.......................................... 703 II. History................................................ 703 III. General Description of Plant......................... 704 IV. AnoDes................................................. 711 V. Cathodes............................................... 714 VI. Electrolyte............................................ 718 VII. Electrolytic Slime..................................... 727 VIII. Power................................................. 728 IX. Operating Details..................................... 732 X. Furnace Refining...................................... 736 XI. Physical Testing oF Products.......................... 739 I. Introduction. These notes are submitted, not as a discussion of the modern practice of electrolytic-copper refining, but as the record of a refinery that was among the pioneers in the field and that is to-day, and has been for 17 years, operating under conditions not to be found in any similar plant in the country. I refer to the current density, which is at least 60 per cent. higher than that regularly employed elsewhere, and to the fact that the anodes treated are the direct product of the converter rather than of the reverberatory furnace. The writer has no knowledge of any other refinery treating anodes of this character. II. History. This refinery, now one of the oldest plants of its kind operating in this country, was built in 1892 by the Boston & Montana Consolidated Copper & Silver Mining CO. to refine the product of its Great Falls reduction works. The estimated capacity of the plant as designed was 65,000 lb. of cathodes per day at a current density of 16 amperes per square foot.

Jan 1, 1914

By J. W. Pugh, J. D. Nisbet

F. B. Foley—The use of data published by Wever and Jellinghaus in 1931 to fix boundaries of the sigma phase in the Fe-Cr system, in the face of the author's own references to the suggestions of Bradley and Goldschmidt, Aborn and Bain, and Hougardy that the phase is much more extensive, and the very much more accurate work, which weighs heavily in favor of these suggestions, of Cook and Jones, published in 1943 and evidently disregarded by the authors, makes their derived ternary diagram, especially fig. 15 for 400°C, quite inaccurate on the Fe-Cr side and affects the extent of phase boundaries in the most controversial and important part of this ternary system. In commercial Fe-Ni-Cr alloys the occurrence of sigma has been observed time and time again at lower chromium contents than that of the 24 pct Cr, 16 pct Ni, 60 pct Fe alloy which is the lowest permitted by fig. 15 of practically carbonless metal. Newel1 reports sigma in 27 pct Cr-Fe even with some carbon present to be one of the greatest detriments to its extensive application, whereas Pugh and Nisbet set a low limit of 36 pct Cr for sigma in the binary Fe-Cr system. J. J. Heger—The diagrams presented by the authors do not agree with observations made on commercial Fe-Cr and Fe-Cr-Ni alloys, nor do they agree with two recent investigations made on the Fe-Cr and the Fe-Cr-Fe systems. I refer first to the investigation made by Cook and Jones1' on the sigma region of the Fe-Cr system. On the basis of their results which were published in 1943, Cook and Jones established new boundary limits for the sigma and the alpha plus sigma regions. These new limits have been accepted and are incorporated in the Fe-Cr diagram that appears in the 1948 edition of the Metals Handbook. This diagram is shown here as fig. 34. As will be noted, the boundary limits of the alpha plus sigma region in this diagram extend to much lower chromium contents than do those in the diagram presented by the authors in fig. 2. These new limits indicate that sigma phase should form in an Fe-Cr alloy containing 26 pct Cr, and experience with commercial alloys confirms this finding. Indeed, recent studies on a commercial Fe-Cr alloy containing 17 pct Cr have shown sigma phase to be stable in this alloy at 550°C. I recognize that the authors' work did not include studies on the sigma region; however, I believe this is a serious omission because sigma phase may profoundly affect the physical properties of these alloys and should be evaluated in any investigation which attempts to relate physical properties to the equilibrium diagram. The second investigation to which I refer is that made on the Fe-Cr-Ni system by Rees, Burns, and Cook,'' who used pure alloys and employed heating 17 W. P. Rees, B. D. Burns, and A. 3. Cook: Journal Iron and Steel Institute. (July 1949) 162, Part 3. p. 325. times that extend to 200 days. These investigators published their results in July 1949 and from these results constructed isothermal sections at 800" and 650°C. The isothermal section at 650°C is shown here as fig. 35. This diagram indicates sigma phase should form in a pure 18 pct Cr-8 pct Ni alloy. Although sigma phase has not been observed after 10,000 hr at 1200°F in commercial 18-8 alloys containing carbon and nitrogen, it has been observed under the same conditions in 18-8 alloys modified with titanium and columbium, both of which serve to reduce the effect of carbon and nitrogen. This section and the one at 800°C suggest that the constant iron sections which are presented by the authors, should be considerably altered, if they are to represent an accurate picture of the system. A few of the suggested alterations are as follows: In fig. 9, the section at 50 pct Fe, the gamma plus sigma region should be widened, not narrowed at 800" and 650°C. In fig. 10, the section at 60 pct Fe, the gamma plus sigma region should be widened at 800° and 650°C. In fig. 11, the section at 70 pct Fe, an alpha plus sigma, an alpha plus gamma plus sigma, and a gamma plus sigma region should be added. In fig. 12, the section at 80 pct Fe, the alpha plus gamma region should be widened at 800°C. In fig. 13, the section at 90 pct Fe, the alpha plus gamma region should be widened at 650° and 800°C. Undoubtedly, the chief reason for the discrepancies between the authors' data and those of Rees, Burns, and Cook is that the authors did not employ long enough heating times to allow for transformation. In this connection, I wish to warn that transformations in these alloys, particularly transformations at temperatures below 800°C, are extremely sluggish and may require a year or more to approach completion. Therefore, unless extremely long heating times are employed or steps are taken to accelerate these transformations by such means as mechanical working, the results will not yield an accurate picture of the alloy system. Certainly, an accurate picture is needed for the development of better high-temperature materials. E. J. Dulis-The purpose of this work is not clear. If an improvement of the ternary equilibrium diagram of the Fe-Cr-Ni system was wanted, it seems logical that testing techniques superior to those previously used would be a prime requirement. To approach equilibrium in this system, long holding times are needed; a fact established long ago but apparently ignored by the present authors, who used the continuous heating and cooling tests in equilibrium studies. A publication on the same system by Bradley and Goldschmidt6 was criticized by Monypenny in a

Jan 1, 1951

By E. R. Morgan

A series of manganese-rich ferromagnetic alloys has been studied. The alloys are based upon an interstitial solution of carbon in manganese. In each of the systems investigated, the ferromagnetic phase was found to be face-centered-cubic in structure. This magnetic phase is based upon the composition (Mn, X)4C; where X is a metallic element which has both a positive size factor with respect to manganese and a high positive valence. Typical elements of this type are aluminum, indium, and tin. Preliminary magnetic measurements indicate that the effective magnetic moment of manganese in the alloys is at least 1.0 Bohr magneton per atom. It is emphasized that the critical atomic separation beyond which ferromagnetism can occur will be governed by the electronic configuration of the manganese atoms. Lattice parameter measurements show that the ferromagnetic condition in the present alloys is associated with an increased atomic separation of only 3 to 6 pct compared with a hypothetical face-centered-cubic lattice of pure manganese. MANGANESE which is paramagnetic and not ferromagnetic in the pure state can become ferromagnetic if suitably alloyed. Under favorable conditions manganese can have an effective ferromagnetic moment of the order of 3.41,2 or even 4.0 compared with a moment of 2.2 Bohr magnetons for iron. The incidence of ferromagnetism is believed to be controlled by the degree of atomic separation.4,5 This is consistent with the fact that the atomic separation, when considered in terms of the ratio of the atomic diameter R to the diameter r of the 3d-shell, is larger in the ferromagnetic metals. The values of R/r calculated by Slater,' for metals of the iron group are given in Table I, and Slater has suggested that the critical ratio of R/r is 1.5. Consideration of the criterion of atomic separation alone would suggest that there are three classes of possible ferromagnetic manganese alloys: 1— Ordered structures, of either solution or compound types, such that manganese atoms are not near neighbors; 2—Structures in which the separation of near-neighbor manganese atoms is sufficiently increased by the presence of suitable solute atoms; and 3—Dilute solutions of manganese. Typical ferro- magnetic alloys of types 1 and 2 are the Heusler alloys and Mn4N, respectively. There are no recognized ferromagnetic alloys of type 3. In view of the proximity of the ratio R/r for manganese to the critical ratio of 1.5, the present work was concentrated on manganese-rich alloys rather than on ferromagnetic alloys of low manganese content, several examples of which are already known. Calculation of the distance of closest approach of manganese atoms in each of its allotropic forms, Table 11, indicates that the high temperature ? phase provides the most suitable basis for ferromagnetic alloys of high manganese content. However the high temperature forms of pure manganese cannot be retained on quenching, although a face-centered-tetragonal form apparently can be produced by electrodeposition. Since the critical separation which must be exceeded in order to obtain ferromagnetism is dependent upon the number of electrons in the d-shell, no exact critical value of R/r can be assumed. Not only will this separation vary from one pure metal to another, but it will also vary with the valency of the given type of atom when it is in an alloy. Raynor8 has shown that metals of the iron group probably

Jan 1, 1955

By C. G. Dunn, W. W. Martin

Various X ray diffraction techniques have been developed for determining orientations of crystals. Transmission and back-reflection Laue methods1,2,3 in particular have been found to be very useful. In many applications, however, a need still exists for a simplification of the details of analysis so that orientations can be determined more rapidly, and operators with little or no training in crystallography can be easily instructed in the necessary procedures. Filling this need in part is a system of analysis that has been in use for several years in the Metals Section of the Laboratory, General Electric Co., Pittsfield, Mass. This method is an extension of the one of Majima and Togino,4 which uses fifty-five indexed standard transmission Lauemams to cover the entire range of crystal orientations. Greater rapidity in determining the orientation of a crystal is achieved by using a slightly larger set of standard Lauegrams and by systematizing the steps followed. This system of analysis in its present form constitutes the subject matter of the present paper. Although the method uses transmission X ray Laue patterns, many of the operations are applicable to the analysis of back-reflection Lauegrams. Very briefly the method is based on the fact that Bragg and azimuthal angles for two or more identified Laue diffraction spots completely determine the orientation of a crystal, and standard Laue photographs of the Majima and Togino type provide a ready means for identifying Laue spots. After a preliminary analysis, which consists of identifying the pattern in terms of one of the standard patterns, a correctly marked Lauegram is selected to carry out the final analysis. From measured Bragg and azimuthal angles, the normals (poles) of the reflecting planes for the selected Laue spots are located on a stereographic projection and the orient.at.ion of the crystal determined (in terms of the positions of cube poles for example) by a few additional operations on a \I1ulff net. The actual procedure followed will be illustrated in detail after a brief treatment or X ray problems connected with orientation determinations and a discussion of useful principles and techniques. X Ray and Plotting Problems RELATIVE ADVANTAGES OF BACK-REFLECTION AND TRANSMISSION LAUE DIFFRACTION METHODS In the back-reflection Laue method it is easy to make a direct measurement of the tilt of a zone of Laue spots. Specimen thickness also is generally no problem. In the case of thin specimens, however, the time of exposure is less for the transmission Laue method. For example, with silicon ferrite this advantage holds up to a thickness of 0.6 mm if tungsten radiation at 40 kv peak voltage is used. Furthermore, the transmission photograph represents the entire thickness of the crystal whereas the back-reflection picture represents only a surface layer (a thickness of less than 0.05 mm for silicon ferrite), the deeper layers supplying a negligible amount to the diffraction pattern. If a crystal is distorted, the transmission Laue pattern gives more information particularly with regard to the orientation of the crystal and the overall amount of spread in orientation. Finally a 3 1/4 X 4 in. film suffices for transmission photograms; a 4 X 5 in. or 5 X 7 in. film with a central hole is needed for back-reflection patterns Not only is the manipulation of film easier because of the size of film, but

Jan 1, 1950

By D. F. Kelsall

The hydraulic cyclone, in simple or modified form, is finding increasing application in metallurgical processing. In this article, the author considers several aspects of conventional applications, less conventional applications and some potential uses of modified cyclones. Information concerning the mechanism of separation within a cyclone and the effect of the major variables is summarized. During the past twenty years, and especially in the last decade, hydraulic cyclones have found increasing application in metallurgical processing as classifiers and, under special circumstances, as thickeners. To a more limited extent cyclones have been used to extend heavy, or dense, medium separation into the finer size ranges, and several attempts have been made to separate dispersions of immiscible liquids with a view to use in the field of liquid-liquid extraction. A large proportion of published work reports incomplete data, applicable only to particular systems. The more useful papers, covering data of relatively wide significance, have been reviewed recently in considerable detail1,2 and no attempt will be made to cover the same material in the present paper. However, before considering several aspects of conventional application, less conventional applications and some potential uses of modified cyclones, it is considered pertinent to summarize information relating to the mechanism of separation within a cyclone and the effect of the major variables. LIQUID FLOW PATTERNS At any point within a simple cyclone the liquid velocity may be resolved into three components, i.e., the tangential or swirling velocity (v), the vertical velocity (w), and the radial velocity (u). The relative values of these components at selected positions within a typical cyclone have been established for water. 3 The extent to which similar flow patterns exist in systems of considerably higher viscosity, or apparent viscosity, operating under conditions of reasonable pressure drop, is not known with precision, but indirect evidence indicates that up to approximately five centipoise there is no marked change. Beneath the bottom of the vortex finder, envelopes of constant tangential velocity are cylinders coaxial with the cyclone, and, at any horizontal level, starting from near the conical wall, this component of velocity increases with decrease in radius (r) according to the relationship vrn = constant (where 0 < n < 1) This holds to regions near the center of the cyclone where v reaches a maximum value at a radius somewhat smaller than that of the inside wall of the vortex finder. At this point, the centrifugal acceleration, v2/r reaches a maximum value. As the radius is further decreased the tangential velocity decreases so that v = r. Above the bottom of the vortex finder there is a tendency for the first relationship to persist to near the outside wall of the vortex finder, until v reaches a maximum and then decreases rapidly as the wall is approached. The two most significant features of the vertical velocity distribution are an approximately conical Table I. List of Symbols v = liquid tangential velocity at any point in the cyclone w = liquid vertical velocity at any point in the cyclone u = liquid radial velocity at any point in the cyclone r = radius from cyclone axis to any point ? = half angle of conical portion of the cyclone ?= fluid viscosity d = particle diameter s = liquid specific gravity p = particle specific gravity d50 = the particle diameter which distributes 5090 to overflow, 50% to underflow e = overflow diameter, in. b = feed diameter, in. G = gpm of feed slurry

Jan 1, 1963

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

The Goodman Mfg. Co., Chicago, builders of underground electric mining machinery, celebrated its 50th Anniversary on April 23rd. The company has a complete line of cutting machines, loaders, conveyors, locomotives, and shuttle cars for service in both coal and metal mines. The Joy Mfg. Co. has received, under the provisions of the Economic Cooperation Administration, an initial order in excess of $1,000,000 for equipment to be used in the mechanization of iron ore mines in France. While the equipment will be used in various privately owned mines, the order was placed through the French Society for the Development of Mechanical Mining. Included in the equipment are the company&apos;s recently developed hard rock loaders and shuttle cars. American Brake Shoe Co. has consolidated all operations of its Canadian subsidiaries into one corporation, the Dominion Brake Shoe Co. Ltd., Montreal, Canada. The products of Dominion Brake Shoe include brake shoes, track equipment, manganese steel castings, automotive brake lining, air compressors, and paint spray equipment. Exhibits of more than 2000 electrical products, processes, and techniques, ranging from precise aircraft instruments to complex working models of steel mill, textile and other industry equipment, will be displayed in the "More Power To America Special" when it makes its nationwide tour this spring. C. H. Lang, Vice-President in charge of sales for General Electric&apos;s apparatus department, said that the train will visit the country&apos;s key industrial centers for inspection by utility and industrial executives and municipal leaders. The train will be hauled by an Alco¬G-E two-unit, 4000-hp., Diesel-electric locomotive which in itself will be part of the unusual series of displays. Sizing of Excavators and Hauling Equipment. The Power Crane and Shovel Association has released a 24-page bulletin dealing with the proper sizing of excavators and hauling equipment, Technical Bulletin No. 3. Rubber Conveyor Belting. "Installation, Care and Maintenance of Rubber Conveyor Belting" is the title of a booklet just published by the Carlyle Rubber Co., New York. Power Transformers. The manufacturing procedure of Allis-Chalmers small power transformers is described in a new 28-page bulletin, 61B6094A, covering the core and coil unit, no-load tap changer, tank construction, cooling methods, oil preserving systems, bushings, terminal compartments, indicators and testing of transformers. Safety Equipment. General Scientific Equipment Co., Philadelphia, has released a 46-page, fully illustrated catalog on personal protective equipment and industrial safety devices. This new catalog covers respiratory devices, eye protection, hats, gloves, carboy pumps, drum pumps, and miscellaneous safety equipment for use in mines, industries, etc. Osmosalts. The Osmose Wood Preserving Co., Buffalo, N. Y., has issued a bulletin on Osmosalts, the wood preservative which, say the manufacturers, makes timber last three to five times longer, slashes maintenace costs, stops rot, and resists wood-destroying insects. Jigs. The Dorr Co. has acquired the Pan American jigs and they are now known as the Dorrco-Pan American jigs. Originally developed as a gold-saving device for dredges, the simplicity, ease of operation. efficiency, and negligible maintenance of these jigs have resulted in wide adoption on all types of placer operations, in the closed grinding circuits of gold mills, and on many applications where a coarse bed concentrate is separated from the ore gangue.

Jan 5, 1950

[A session of the Institute of Metals held during the February, 1926, meeting of the Institute, was devoted to a symposium on &apos;&apos;gas" in copper. S. Skowronski, Perth Amboy, N. J., was chairman. On the program were papers by T. S. Fuller; W. H. Bassett and J. C. Bradley; Susan B. Leiter, and W. H. Bassett and H. A. Redworth. These papers follow the chairman&apos;s introductory discussion.] S. Skowronski, Perth Amboy, N. J.—There are today two groups of metallurgists, one group who believe in oxygen-free copper and another group who believe that oxygen is absolutely essential in copper; not that the two groups disagree, but that a certain grade of work calls for oxygen-free copper, while another grade of work does not. I can sum up the problem by reading the abstract of an article by Frederick Johnson of Great Britain in The Metal Industry of Sept. 4, 1925. Few, who have watched the trend of things during the last two decades, can have failed to notice the steady growth of opposition to the prevalent use of copper in the "tough-pitch" condition, which for many ages enjoyed an unchallenged supremacy. Broadly speaking, there are two schools of thought in the metallurgical world, one holding the view that copper containing a small proportion of oxygen is an eminently suitable material for users, the other expressing its preference, in no hesitating or uncertain manner, for copper from which oxygen has been removed by means of a deoxidizer. In the old days, before the advent of the converter and the electrolytic refining process, the copper of commerce was so impure that the presence of oxygen was imperative in order to neutralize the influence of impurities which militated against the hot-working properties of the metal. For electrical purposes, oxygen must be present, because sound bars and billets cannot be produced without it, except at the expense of electrical conductivity. For most engineering purposes, however, electro-deposited copper may be used either by itself or with the addition of alloying elements, e. g., arsenic and nickel, which experience has proved to be beneficial, such copper being freed from oxygen by the aid of a deoxidizer, e. g., phosphorus. Such material has positive advantages over "tough-pitch" copper as at present produced from refinery furnaces. It has greater toughness and malleability, and is immune from the action of reducing gases at high temperatures, the destructive action of which, in the case of "tough-pitch" copper, is well known. The reasons why there has been no wholesale replacement of "tough-pitch" copper by deoxidized copper are commercial in nature; copper can be more economically melted in large-capacity furnaces which lend themselves to the poling process and are not so suitable for control of uniformity of action of deoxidizers as the smaller electric or crucible furnaces, whilst they are also more convenient for the production of very large castings. Moreover, the equipment of most large copper manufacturing plants is designed for casting "tough-pitch " copper in open moulds, e. g., wire bars and cakes, and these could not be used for deoxidized copper, which must be cast in vertical moulds as in brass-shop practice. Further, there are still certain impure brands and

Jan 1, 1926

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

Discussion of the paper of DORSET A. LYON and ROBERT M. KEENEY, presented at the Salt Lake meeting, August, 1914, and printed in Bulletin. No. 92, August, 1914, pp. 1791 to 1800. LAWRENCE ADDICKS, Chrome, N. J.-I read this paper for the first time this afternoon, and it is really worthy of a much more careful reply than I can make with such a short time to consider it, but I just want to say that I differ with Dr. Lyon in his premises and conception of what happens in a refining furnace. He cites steel as an instance where the production of special grades justifies the high cost of power in an electric furnace, and goes on to say the same will be true in the case of copper, for the reason that you will get a better quality of product. I don&apos;t think you will get a higher grade of copper,. or a better conductivity, as I will explain in a moment. The cost of fuel in the Eastern plants is somewhere between 30c. and 40c. per ton of copper refined, and, if my memory serves me correctly, in an open-hearth steel furnace it is something like $1.25, so it is evident that steel could be handled to better advantage than copper on the basis of fuel. Now, taking the Great Falls plant to compare with is not quite fair, because that plant operates under peculiar conditions. The furnaces are relatively small; the coal is very bad and is very high in sulphur. No one would think of using a coal with 3.6 per cent. sulphur were anything better available. Due to these conditions, the amount of slag is inordinate, it being quoted as 3.5 per cent., and it is necessary to use lime, as there is a tendency to have troubles from arsenic owing to the very high current density used. I think he has taken about as hard a case as he could pick out. Now as to the conductivity of copper, my own idea about the reason that melted copper does not give as high a conductivity as pieces drawn directly from a cathode, or native copper, is that in the latter case the impurities are present as a mechanical mixture, while after melting they are chemically combined with the copper, which is a very different condition. Further, it is necessary to take the sample from a relatively smooth and consequently pure part of the cathode, as otherwise it would not draw. The same with the mass of copper; you pick out a good solid part of the copper, which is not representative. In fact, the refiners have all found they could not take fair samples of cathodes for a chemical analysis. It is always customary to use wire-bar copper -for the purpose of sampling, and the impurities are several times as great as shown by samples taken from .the cathodes. I think if we examine all of the examples he has given of cases where copper is higher in conductivity, it will fit with that explanation. For example, the Elmore process is a direct electrolytic deposit. Incidentally, at the present time there are a number of ex-

Jan 12, 1914

By H. D. Kessler, M. Hansen, R. J. Van Thyne

The phase diagram of the titanium-rich portion of the system Ti-Cr-Fe to 70 pct Ti was established by means of isothermal sections at 900°, 800°, 750°, 700°, 650°, 600°, and 550°C, using arc-cast alloys. An isotherm at 800°C was determined for the section bounded by Ti, TiFe2, and TiCr2. Micrographic analysis was employed as the principal method of investigation, supplemented by X-ray diffraction and incipient melting studies. A ternary eutectoid, ß ? a+ TiCr2 + TiFe, occurs at approximately 8 wt pct Cr, 13 wt pct Fe, and about 540°C. AMONG the first titanium-base alloys in commercial production were the Ti-Cr-Fe alloys. For this reason, the Materials Laboratory, Wright Air Development Center, sponsored an investigation of the system. No information on the constitution of the titanium-rich corner was available. Vogel and Wenderott studied titanium-poor ternary alloys.&apos; As part of this program, the binary systems Ti-Cr and Ti-Fe have been determined previously by the authors.&apos; Other investigators have also studied the binary systems Ti-Cr,3-7&apos; Ti-Fe.³,8-10 Both of the binary systems are characterized by the eutectoid decomposition of the ß phase. Most of the effort of the present work has been concentrated on the titanium-rich corner with compositions of less than 30 pct total chromium and iron content. The determination of isothermal sections was the experimental approach used to obtain the ternary phase diagram. Vertical sections were used to check the consistency of the isothermal sections. While this investigation was in progress, a study of the Ti-Mo-Cr system at Armour Research Foundation disclosed the existence of a high temperature modification of TiCr2 above 1300 °C.11 This allotrope has a hexagonal structure of the MgZn2 type, iso-morphous with TiFe2, whereas the low temperature modification of TiCr2 is cubic," of the MgCu2 type. The existence of allotropy in TiCr2 of course means a more complicated set of phase relationships in this general region than if the phase were monomorphic, as was first assumed. Additional studies were made late in the investigation in order to clarify the part that the hexagonal TiCr2 phase plays in the ternary phase relationships. These will be discussed in a separate section. Because they are of greatest practical significance, the phase relationships in the titanium-rich corner will be presented first, omitting the equilibria involving the hexagonal modification of TiCr2 in an effort to simplify the presentation somewhat. Experimental Procedure Materials: The titanium used in the preparation of the alloys was iodide crystal bar (99.9 pct pure) produced by the New Jersey Zinc Co. High purity chromium and iron were obtained from the National Research Corp. Table I gives the analyses of these materials. Melting Practice: Over 100 alloy ingots weighing 10 to 20 g were melted in a nonconsumable electrode arc-melting furnace. As the techniques are identical to those reported previously,12,13 they will not be detailed here. The ingots were melted in the cavity of a copper melting block insert under a slight positive pressure of helium. No measurable hardness

Jan 1, 1954

By P. S. Roller

P. S. ROLLER,*College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author&apos;s quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should he defined as the distance between an F line and a y line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and y lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. x between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling. It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz." This would require many data to prove, but no proof is given; any consideration would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.," Cambridge, Mass. Contrary to Mr. Bond&apos;s stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids., Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystalline solids. Andreasen and Berg&apos; have ground

Jan 1, 1941

By D. H. Thompson, A. W. Tracy

E. A. ANDERSON*—At the outset, I note that you are using a humid atmosphere containing ammonia but that you make no reference to the variable of carbon dioxide content. Edmunds in his work in this laboratory found very marked effects on the rate of failure due to variations in carbon dioxide in the environment. This point is, of course, strongly related to the overall corrosion rate problem since the role of the carbon dioxide seems to be to permit the formation of hygrogroscopic salts which give a controlled moisture film on the specimen surface, which in turn seems to support the intergranular penetration more consistently than would otherwise occur. (2) Your results with various copper binary systems are very striking. It is curious, however, that the time-to-failure minima all occur at percentages of the alloying element which are considerably below the solubility limits as given in the conventional equilibrium diagrams. If your theory regarding the presence of these impurities in some concentration in the grain boundaries is correct, then either the existing diagrams are not accurate in the room temperature region, or there is a difference in solubility in the grain boundaries as compared -with the grains. A study along these lines might be very revealing. (3) The undesirable effect of silicon in copper calls to mind our own observations on the influence of silicon on the stress corrosion cracking resistance of alpha brass. In this case, silicon was found to have a very marked beneficial effect, particularly when a heat treatment capable of producing a grain boundary constituent was used. I should appreciate it if in the discussion of your paper this point should arise, you would attempt to clarify the difference between the copper-silicon binaries and the copper-zinc-sili-con ternary alloys. D. H. THOMPSON and A. W. TRACY (authors&apos; reply)—(1) We have no data on the constancy of the carbon dioxide content of the air used for the test atmosphere but have believed that the normal carbon dioxide content of the air was as constant as could be obtained by using Edmunds&apos; method of removing carbon dioxide from the air and then adding a measured quantity. In any event, the test atmosphere was the same for all specimens in a test run. We agree with Mr. Anderson that the carbon dioxide content is important in the formation of a hygroscopic film on specimens. (2) We can imagine that there could be a concentration of solute atoms at the grain boundaries which would be so small that present methods would not detect it even if it exceeded the solubility limit of the alloy. There could be somewhat greater solubility at the grain boundaries for an interstitial solute. There is a possibility that phosphorus dissolves in this way but we believe that none of the other solutes studied does. In our present studies we hope to find out something about conditions at the grain boundaries of some of the systems mentioned in the paper. (3) Mr. Anderson points out that silicon brass, which has had a high temperature heat treatment followed by a quench. is highly resistant to stress-corrosion cracking. Although the test conditions used by Anderson* et al were different from our test conditions, it is noted that the silicon brasses which had not had the special heat treatment were less resistant to stress-corrosion cracking than copper containing approximately equal amounts of silicon. G. R. GOHN†—Nonferrous metallurgists and particularly, materials engineers, have been accustomed to thinking

Jan 1, 1950

By G. C. Weaver, G. T. Harley

Beds or lenses of potash and magnesium salts are found in a thick salt section (Salado) overlain by Rustler Red Beds in several members of which water is present and from one of which the rehery process water used by International is obtained. Most of the holes drilled are less than 1000 ft. deep, but some wildcat holes have exceeded 3- ft. cable or rotary tools are used to drill down to the salt. when samples of water are desired the cable rig is used, but when this is not necessary, and where caving material is present, the rotary rig has an advantage. The rotary rig has a further advantage in that only one machine is required to drill and core the entire hole, whereas the cable rig must be moved off to make way for the core drill. Casing is placed in the upper section of the hole as required to shut off water, and a 5 3/16 in. string is set in the top of the salt before coring starts. With the rotary rig drilling is often carried to the salt with no casing, as drilling mud seals off caving ground and water-bearing strata, and a string of 4-in. line pipe or 4-in. flush-joint casing is landed on top of salt. Several types of gasoline-powered core rigs have been used in the field and all recover a 2 1/2-in. core. More recently the Sullivan Model 37 rotary rig has been tried, and a Sullivan Model N has been adapted to use a rotary bit in the upper section. Core is recovered in 20 ft. sections with a double-tube core barrel, and the bit used is a bottom-discharge type set with diamonds. The drilling fluid must be a saturated brine to prevent solution and etching of the minerals present. Blowouts are apt to Occur from nitrogen gas trapped under the several polybalite beds, and some have been strong enough to blow the tools out of the hole. Core recovery is practically loo per cent. All holes are plugged after drilling, with cement through the salt section and opposite each water horizon* and mud through the remainder of the hole. Cores are carefully logged and preserved, and in sections containing potash and magnesium minerals they are quartered and analyzed. The correlation of Ore beds between drill holes is briefly described. Introduction Beds Or lenses of potash and magnesium salts are found at varying depths in a thick deposit of salt (Salado formation) in what is known as the Permian Basin, in southeast New Mexico and West Texas. The Salado formation is Composed chiefly of common salt (halite) with numerous beds of anhydrite, polyhalite, and the soluble potash minerals carnallite, sylvite and langbeinite. Many other minerals are present in small amounts but their Presence has no bearing On the coring technique. Before starting to core the salt section it is necessary to penetrate 300 to 600 ft. of unconformably overlying Permian sediments (Rustler formation) which are composed of red and gray clay, sand, anhydrite, dolomitic limestone and occasional lenses of salt. The upper part of the formation is predominantly anhydrite, and the lower part predominantly clastic. A thin bed of dolomitic limestone, known

Jan 1, 1948

By B. Ramaswami, Y. Nakada, R. Schaefer

EARLY investigations1 of the tensile properties of fcc single crystals did not reveal the easy glide region of crystals with axial orientations inside the standard stereographic triangle. With the production of high-purity metal, a region of easy glide became known1 Consequently, modern work of fcc single crystals ordinarily uses high-purity metal. In order to determine whether further purification of aluminum beyond 99.992 pct purity would result in significant alteration of the tensile properties of single-crystal specimens, metal was zone-refined and crystals grown from the purified material were compared to crystals grown from the original material. Some similar studies were made of gold and silver crystals. Aluminum. The material to be zone-refined was cast into bars 1/4 by 1/2 by 8 in. A zone, approximately 1-1/2 in. in length, was passed over each bar eighteen times at a rate of 6 in. per hr. Melting and stirring were carried out by an induction heater under a vacuum of approximately 10"5 mm Hg, the metal being held in a graphite container. he last 3 in. of each processed bar were then cut off and discarded. No quantitative determination of purity was made, but the effectiveness of the zone refining was tested by a method first suggested by Walton, Tiller, Rutter, and winegard3 depending on the relation between cell structure and impurity content. Rutter and Chalmers4 have shown that a cellular substructure is associated with the presence of impurities during solidificatibn, and have shown how solute rejection at the freezing interface can lead to such an association. Thus the observation of cells gives evidence of impurities being present. To look for cells, the last quarter of a zone-refined bar was cut off, remelted, and grown rapidly into a single crystal. It was then heavily electropolished and thermally etched.~ Cells were revealed by the presence of many parallel rows of etch pits. The same process was carried out with material from the first half of the bar, but no cells were detected. These observations give evidence that a significant amount of impurity material was moved to the "dirty" end of the zone-refined bars. Further indication of purification was obtained from recrystallization experiments. It was found that, after cold rolling to a 97 pct reduction in thickness, the material from the pure end of the bar re-crystallized at room temperature, but the material from the "dirty" end did not. Other investigators6 have previously reported that zone-refined aluminum recrystallized at room temperature. The purified material was cast into 1/4 by 1/4 in. bars, and seeded crystals were grown in a (100) orientation and in a single slip orientation where the Schmid factor is m =2. The crystals were then cut to the desired length, annealed for 24 hr at 610°C, furnace-cooled, and etched in Tucker&apos;s etch. The crystals were tested in a floor-model Instron machine, using a gage length of 4 in. and a strain rate of 1/2 pct per min. The resulting curves are shown in Fig. 1. It is seen that the differences between the curves for zone-refined and original-purity materials are small. These differences are within the scatter observed for crystals of the original purity. In particular, the zone refining does not significantly affect the onset of the different "stages" of the curves.

Jan 1, 1964

By E. D. Glass, R. C. Hessing

The volume of gas in storage reservoirs may be compuled from estimates of hydrocarbon pore volume and gas density. However, both are dificult 10 estimate accurrately. Further, no adequate method has been presented for estimating reservoir performance during operation for Volumetric gas-storage reservoirs. Normally, gas-storage reservoirs exhibit small pressure gradients in the area conr(zilliitg the storage wells, even under conditions of maximum injection or withdrawal. A zone of low permeability usally surrounds, and is in communication with, tile permeable zone which causes the storage reservoir to exhibit a definite pressure lag, as evidenced by lack of pressure stabilization after extended periods of shut-in. An analysis is presented for determining pore volume and gas in place, and for predicting future reservoir behavior of such reservoirs. The mathematical development is presented and a solution is shown for an operating storage reservoir. INTRODUCTION Depleted gas reservoirs arc often used for gas storage. The reservoir usually contains a permeable area in which numerous storage wells arc located. The permeable zone permits high deliverability and injectivity with relatively small pressure loss. However, pressure equilibrium may not be attained even after long periods of shut-in because of the fact that the permeable zone is in communication with an adjacent tight section. Most of these gas-storage reservoirs are sand lenses, such as ancient beaches or offshore bars. The areal geometry of these reservoirs is normally such that flow in the tight zone will be approximately linear, and the solutions presented here should apply. DISCUSSION For purposes of this paper, a gas-storage reservoir is considered to consist of a volume of permeable rock in communication with tight rock, all of which is surrounded by an impermeable barrier. Pressure gradients across the permeable zone are small due to low flow resistance and well density. Hence, the permeable zone is considered to be at a uniform pressure which is only a function of time as gas is injected or withdrawn. The tight zone is considered to be linear with uniform permeability and porosity, and the pressure in the tight zone is a function of both position and time. l&apos;wo solutions for flow in this system are used to define the reservoir and determine its properties. The first solution is for an infinite linear tight zone in communication with the permeable zone. The expression for constant gas-withdrawal rate is Eqs. I and 2 ore superposed to account for varying rates of production or injection and are used to fit gas-storage field reservoir history. The solution for the infinite case, Eq. 1, is first applied to field data for early times following a shut-in period. Constants BDa and c are ohtained by a least-squares fit of pressure and rate data. After the pressure wave from gas injection or withdrawal reaches the external boundary of the tight zone, the infinite solution will no longer apply and pressure behavior should be more nearly represented by Eq. 2 for a gas reservoir in communication with a finite linear tight zone. The values of the constants found from curve-fitting Eq. 1 to the early time data may then be used in Eq. 2 that only the coefficient- L containing the unknown

By W. O. Philbrook, A. E. El-Mehairy

The reduction by hydrogen of individual particles of dense hematite implanted in beds of inert spheres is controlled by single-particle kinetics. No evidence of reagent starvation was found down to low flow rates established by pressure gradients as low as 0.01 in. H2O per- in. of depth in beds of spheres as small as 3/16 in. diam. The optimum size for most rapid reduction of beds of homogeneous ores results from the effect on reducing potential of the gas of competing factors of supply and consumption. A recent publication1 presented the results of a study of the rate of reduction of iron ore in a fixed bed by hydrogen introduced at constant pressure differential. Under these conditions the reagent flow rate varied directly with the bed permeability, which was a function of particle size and shape and bed voidage, and inversely with the temperature-dependent viscosity of the gas. The present note adds data from supplementary experiments designed to reveal whether or not any partial reagent starvation arising from a flow-dependent transport resistance in the gas phase might have been operative. Kinetic studies were made of the reduction of individual spheres of iron oxide inserted into beds of inert spheres, so that the reagent flow rate and aerodynamic conditions were again determined by bed characteristics and temperature. The experimental techniques and apparatus were similar to those previously described.&apos; The synthetic iron ore spheres were wet-rolled by hand from powdered, reagent-grade* Fe2O3 and subsequently dried and then fired in air at 1250°C. The indurated pellets has almost no porosity as indicated by their specific gravity of 5.10. The beds were made of dense alumina spheres, closely sized to the same dimension as the iron oxide spheres, in the following diameters: 3/16, 1/4, 5/16, and 3/8 in. Each bed was hand-packed of particles of a single size to a controlled voidage of 0.32 and depth of 3 in. within a capsule having an internal diameter at least 10 times the particle diameter. A few weighed iron oxide spheres were individually planted in the central plane of the bed of inert spheres at points away from the wall. Reduction was carried out at atmospheric pressure by purified hydrogen at 670°, 800°, and 985°c. The reagent was supplied at precisely controlled pressure differentials of 0.01, 0.02, and 0.03 in. of water per in. of bed depth. This gave a range of flow rates similar to that used in the study of beds of ore and comparable with blast-furnace practice in terms of mols of reducing gas per hr per sq ft of cross-section. Viscous flow prevailed in all beds under the experimental conditions of this study. The reduced particles were checked for loss in weight and then sectioned for metallographic examination. Occasional particles were found to have contained minute internal cracks that interfered with the concentricity of the reduction interfaces, and data for such particles were rejected. An experimental rate constant for the reduction of the imbedded Fe2O3 spheres was calculated on the assumption that the rate was proportional to the metal-oxide interface area by an equation that has been found by McKewan2 to describe the reduction of single spheres

Jan 1, 1962

By J. H. Jackson, P. D. Frost, C. H. Lorig, A. C. Loonam

THIS paper describes (1) the effect of sodium on the tensile ductility of magnesium-lithium base alloys, and (2) the precautions necessary to avoid sodium contamination. Effect of Sodium on Properties The harmful effect of sodium in magnesium-lithium base alloys has been recognized by other investigators.1,2 However, quantitative data on its effect have not hitherto been published. The effect of sodium on the ductility of an 87 pct Mg-9 pct Li-4 pct Zn alloy is shown in fig. 1. The sodium was introduced by adding sodium chloride to the flux, which consisted of 50 pct LiC1-50 pct KC1. Moderately high ductility was obtained at an indicated sodium content of about 0.05 pct. Actually, the value probably was less than 0.05 pct, be-cause analyses of the lithium, which were more accurate than those of the alloy, indicated that the highest ductility was obtained when its sodium content was 0.02 pct or less. In fact, in alloys having somewhat greater strength than this particular composition, the sodium content of the lithium must be under 0.01 pct to avoid embrittlement. The effect of sodium on the ductility of several alloys having fairly low strength levels is shown in fig. 2. These alloys were melted with a flux containing 75 pct LiC1-25 pct LiF, to which various quantities of sodium chloride had been added. Sodium had no significant effect in pure magnesium but markedly decreased the ductility of the magnesium-lithium-zinc alloy, even though the alloy contained only 3 pct Li. As the lithium content of the binary alloys was increased, the ductility of the sodium-free alloys also increased. However, the effect of sodium on ductility became progressively more pronounced. Sodium raised the minimum hot-rolling temperatures, particularly those of the high-strength alloys, but its effect, in this respect, was less critical than its effect on tensile ductility. For optimum cold roll-ability, a low-sodium content was required. No deleterious effect of sodium on the stability of properties at slightly elevated temperatures was found. Control of Sodium in the Alloys Optimum ductility is obtained in the magnesium-lithium base alloys when sodium-free materials are used. Commercial lithium was formerly the chief source of contamination. Now purified lithium is usually available. When lithium of ordinary purity is used, the effects of sodium can be minimized by using a flux consisting of 75 pct LiC1-25 pct LiF. Bubbling nitrogen through the melt also helps remove sodium. The damaging influence of sodium on the ductility of the magnesium-lithium base alloys was observed very early in the alloy development program. The discovery came about as a result of chemical analyses made on the lithium used in the alloys. Ductile alloys had been made from lithium containing about 0.2 pct Na, whereas the lithium used in the brittle alloys contained about 0.9 pct Na. The first magnesium-lithium alloys were melted under commercial fluxes ordinarily used in melting magnesium alloys. These fluxes were unsuited for scavenging the contaminating sodium, and the resulting alloys were not ductile. Next, a flux containing 50 pct KC1-50 pct LiCl was tried. This was somewhat more effective but still was not satisfactory. After a number of different experimental fluxes had been investigated, one containing 75 pct LiC1-25 pct LiF was adopted. This flux made it possible to obtain good ductility in the alloys. The

Jan 1, 1951

By H. W. R. Wardlaw

This paper shows how the cost of the drilling operation per 100-ft drilled can be related directly to bit weight, rotary speed and hydraulic horsepower variations. This enables optimum drilling conditions to be determined in a rapid and concise manner. The paper also shows how data for such calculations cun be obtained quickly and accurately by the "drill-off" method suggested by Lubinski. Examples are given to illustrate the principles and methods used. INTRODUCTION The problem of drilling oil wells at lowest cost and hazard needs little introduction, since it has been with us virtually from the earliest days of our industry. With better understanding of the effect of the various factors of bit weight, rotary speed and hydraulic jet action, however, the problem has recently taken on a more scientific aspect, and engineers or groups of engineers have been assigned to study many of the facets of this problem not immediately apparent to the general observer. Moreover, the availability of better equipment and materials has enabled certain practices to be adopted which were hitherto impractical or uneconomic. This has paved the way for further efforts to reduce footage costs, and so on. The improvements to be made in drilling technology and cost reduction are by no means exhausted, and we still frequently hear of new records being set and new materials and equipment becoming available. Some of the analytical methods used by the author to assist in promoting faster and more economic methods of drilling, particularly in Venezuela, form the basis for this descriptive study. STATEMENT OF THEORY AND DEFINITIONS In a highly competitive industry, it is no longer permissible to regard penetration rate as the sole criterion for drilling efficiency. Every new technique, however spectacular, must nowadays be viewed in the cold light of whether it will eventually lead to lower cost per foot drilled. In the exposition which follows, therefore, actual drilling rates will play only a secondary role, and attention will be focused on the more important aspect of cost per foot drilled. Changes in drilling factors such as weight and rotary speed will be studied in the light of how they directly affect the cost of operations. The advantage of this method is twofold—(1) cost increments are indicated directly; and (2) the relationships are, surprisingly enough, simpler than those relating to drilling rates. To achieve these advantages, it is necessary to introduce a new unit of measurement which, although unfamiliar, is readily understood. The unit in question will be referred to as the "actual drilling cost factor" and will be designated by the symbol h. The unit is defined as the number of rotating hours taken to penetrate 100 ft of formation. The unit can be obtained immediately from the actual drilling rate (r) in feet per hour; thus, This latter transformation is so easily effected that it may be convenient to use h as an international symbol for drilling cost, and to delay conversion to dollars, bolivars, pounds, etc, per foot until actually required (e.g., for fiscal reasons, etc.). In actual operations, h was found to vary in a relatively simple manner with increases in bit weight. The variation in h with rotary speed was found to be somewhat more complex. However, this also could be enormously simplified by introducing a further factor known as L or relative bit life. The factor L is also defined very simply by dividing 1,000 by the rotary speed (N) in revolutions per minute.

By H. R. Stahi

THE ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classifiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9 1/2 -in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3 ½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 5 2 ¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1 3/4in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges are renovated occasionally, broken pieces being removed and the charge brought up

Jan 1, 1945