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By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines or phosphate ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pct. When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900" to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines Or phosphate Ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pet' When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900° to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By T. R. A. Davey

The fundamental principles by which bismuth may be removed from lead by pyrometallurgical processes are enumerated. Qualitative discussion of the phase diagrams concerned is followed by presentation of quantitative diagrams. Brief mention is made of the practical aspects. Data presented show how chemical lead (<0.005 pct Bi) may be produced by the Jollivet, Dittmer, and Kroll-Betterton processes. BISMUTH is an element whose properties are very similar to those of lead, and its minerals are associated to a greater or lesser extent with nearly all the major lead ore deposits of the world. In consequence, the lead produced contains bismuth unless a special debismuthizing process is practiced. There does not appear to be any published comprehensive review of the effect of bismuth as an impurity on the properties of lead, and there is some uncertainty as to the beneficial or detrimental effects of small amounts of bismuth in lead used for particular applications. Standard specifications of most countries stipulate a maximum bismuth content of 0.005 pct for chemical lead. With one exception, this specification is met only by lead producers whose ores are relatively bismuth-free, or whose refineries are electrolytic. In lead for certain applications a higher bismuth content is preferred by some fabricators.&apos; In contrast to the refining for silver, copper, and antimony, the refining for bismuth does not usually result in a large overalil economic gain from the sale of byproduct. Thus, debismuthizing is customarily practiced only if market conditions demand a lead with a bismuth content lower than that produced by classical refining; i.e., softening, desilverizing, and zinc refining. Until this century, debismuthizing was carried out only a:; an incidental result of desilverizing, as later described. This century has seen the development of the Betts electrolytic process—the only successful one of many electrolytic processes proposed and patented —and a number of precipitation processes based on the pioneering work of Kroll,&apos; although only one of these, the Kroll-Better ton process,&apos; is in wide commercial use today. It is proposed in this paper to deal only with the pyrometallurgical processes, excluding those based on aqueous electrolysis. Historical Background Prior to 1833 silver (and incidentally bismuth) could not be removed from lead. To recover silver from rich bullion, the only course available to lead refineries was to cupel the bullion, forming a slag of litharge. The first slag tappings or skimmings, on reduction, yielded soft lead of very low silver and bismuth contents. Intermediate skimmings yielded a lead of somewhat higher silver and bismuth contents, which could if desired be recupelled. The last skimmings yielded a litharge very high in bismuth, which could by successive oxidation and reduction cycles be separated into high purity bismuth, and lead returns. This oxidation reduction process, with modern operating techniques, can still be used today to produce a portion of a lead refinery&apos;s output of very low bismuth content, as discussed later. In 1833 Pattinsone discovered that lead bullion could be purified of silver and bismuth by fractional crystallization, the primary lead crystals formed by partial freezing being purer than the melt. Repeated processing in pans produced a lead containing as little as 0.02 pct Bi. The simple explanation of this

Jan 1, 1957

By R. W. Heins, N. Street

An investigation was made of the Rehbinder-Kuznetsov pendulum as an instrument to aid in the study of rock drillability and of the effect of surface active agents on rock drillability. The "critical weight", i.e., the pendulum weight at greatest sensitivity, was determined for several solids and it was shown that a change in surface environment changed the maximum value of the apparent hardness but not the critical weight, and the critical weight increased with increase of hardness of the solid used. Using silver halides as the solid material, and solutions of either silver nitrate or the appropriate alkali halide, it was shown that there is a hardness maximum at the zero point of charge of the particular silver halide. INTRODUCTION The development of relative hardness measurement by the method of damped oscillations, or what has been called the pendulum sclerometer, is due largely to the Russian workers P. A. Rehbinder and V. D. Kuznetsov and their co-workers.1,2 The apparent simplicity of the method has led to its use in a number of fields, particularly those where relative hardness has to be determined on a brittle solid. As will be shown, the simplicity of the method belies its complexities. THEORY OF THE PENDULUM SCLEROMETER The operation of the pendulum is quite simple. Make a sharp point such as a diamond, hardened steel or tungsten carbide, as the fulcrum of a pendulum, with the center of gravity of the system below the point of support. The attenuation of this pendulum is a function only of air friction, provided no fragmentation, breakage or other work is performed by the point. However, when the solid surface breaks or disintegrates under the point, and the latter penetrates the solid, part of the energy of oscillation is absorbed in the work of creating new surface, and the rate of attenuation increases. In the case of brittle solids, the point of the pendulum produces fragmentary particles, whereas the point produces plastic deformation in a plastic solid; but in both cases damping occurs. This pendulum damping is a function of the hardness or strength of the solid. Whereas a pendulum oscillating on a hard material attenuates primarily by air friction, in the case of a soft material, the points penetrate the surface of the solid, producing new surface, and the pendulum damps much faster. Thus the attenuation rate is a function of the hardness or strength of a material in both plastic and brittle substances. Russian workers initially used a single support for the pendulum, but this created many problems because of oscillation of the pendulum in more than one plane. A revised version used two points, or sometimes a prism as support, successfully confining oscillation to one plane. The version using two points was constructed and used throughout this investigation, making the recording of the attenuation as a function of time relatively easy. That both points should be in contact with the solid during the damping is implied. Rehbinder (c.f. Kuznetsov2) proposed that the hardness H as measured by this method could be expressed by a quantity which is inversely proportional to the relative initial rate of attenuation of amplitude. where H = hardness parameter (arbitrary units) a, = initial amplitude da/dt = rate of attenuation of amplitude with time. In the simplest case of damping with a constant logarithmic decrement, H can be regarded as a constant for the duration of the whole damped oscillation process. That is

Jan 1, 1966

By D. L. H. Forbes

(New York Meeting, February, 1912.) Introduction.. THE application of the cyanide process to complex sulphide ores for the extraction of their bold and silver is no longer a novelty, but, as an example showing modern tendencies in this direction, the following description of the new milling process at the Lucky-Tiger mine is offered: The extraordinary richness of the ore in the upper workings of this mine made any thought of milling in its early hays seen unnecessary. Its location in an almost inaccessible part of the rugged mountains just south of the Arizona boundary-line, and 60 miles from the town of Douglas, made the expense of bringing in machinery and supplies a very serious consideration. But, after the property had been shipping high-grade ore on mule-back for a few years, there remained in the mine a considerable tonnage of vein-matter that was too low in value to ship, but which it was thought could be milled near the mine at a profit. The first milling-experiment consisted in the installation of Nissen stamps in conjunction with concentrating-tables and Frue vanners, but, as the ore from the upper workings contained most of its silver in the form of cerargyrite, this type of mill was not successful. The proximity of the mine to the great copper-concentrating mills of Cananea and Moctezuma naturally made the influence of copper-milling, ideas felt in the company&apos;s next installation. At considerable expense a stage-crushing concentrating-mill, with a capacity of 100 tons per day, was built. This mill was modeled along the lines of the most approved practice in copper-concentrators at that time, the main idea being that crushing by stages in rolls was necessary in order to prevent losses from ?sliming." Largely owing to the fact that, when the mill

Aug 1, 1912

By John Griffen

This paper records over twenty years&apos; experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years&apos; experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By John Griffen

This paper records over twenty years&apos; experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years&apos; experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By AIME AIME

THE Engineering Education Committee of the Institute convened at the Engineering Societies Building on Feb. 18, .1929, with. E. A. Holbrook, dean of the Schools of Mines and Engineering, University of Pittsburgh, in the chair. STATISTICAL&apos; STUDY OF STUDENTS IN MINING The first paper, a statistical study of the students enrolled in courses leading to a mining degree in 40 institutions, was presented by H. D. Keiser,, assistant editor of the Engineering and Mining Journal; who gave a brief outline of an article he was about to publish in ?that journal?. The complete article appears in the issue for March 30; 1929: To avoid repetition the results are summarized as follows: A request was sent to 42 institutions which give courses in mining&apos; for the number of students enrolled as freshmen; sophomores; juniors, seniors and special students in mining for the years 1925-26, 1926-27, 1927-28, 1928-29. Two of these colleges replied that there were no students in mining. Students&apos; enrolled in courses in metallurgy, geology and engineering, other than mining, were excluded from the statistics:. The statistics received were arranged in, four groups, according to the completeness of the information. One complete group of 25 institutions &apos;shows as follows:

Jan 1, 1929

By George M. Fowler, J. P. LYDEN

T HE relationship of geological structure to orebodies and to the great masses of chert in the Tri-State mining district is of such significance that it prompts a brief recital of the existing information regarding the various periods of structural deformation. We began to note a relationship between structure and orebodies early in our studies of the field, and further work revealed a like relationship of structure to the great masses of chert so commonly associated with the ore-

Jan 1, 1934

By John Johnston

ONLY two of the Institute&apos;s 26 Local Sections were unrepresented at the delegates&apos; three sessions, held on Monday morning and afternoon and Thursday afternoon of the annual meeting. The Philadelphia and Wyoming Sections alone sent no delegates. In all, seven hours were spent most profitably in reviewing the work done by the Institute at headquarters and in the field, comparing ideas and practices in varied activities, and in establishing personal contacts. Cordiality and the co¬operative spirit marked the meeting and apparently no one had any serious grievances to air. Four motions were introduced and unanimously passed. The first, which was introduced by B. L. Sackett, was a recommendation by the delegates that a national committee be appointed to cooperate with Local Sections in the matter of student activities, to promote Institute welfare. It was agreed that this recommendation be made simultaneously to the Board of Directors and to the Mineral Industry Education Division. A subsequent resolution by A. M. Smith was passed asking that a part of the duties of this committee be to secure the opinions of Local Sections as to the most desirable methods of interesting students in the preparation of technical papers, by giving prizes or otherwise, and to report their findings at the Section Delegates&apos; meeting next year.

Jan 1, 1933

CONTENTS PAGE ANDERSON, ARVID E., and CAMERON, FRANK K.-Recovery of Copper by Leaching, Ohio Copper Co. of Utah. Discussed by Thomas B. Brighton, H. C. Goodrich, Arvid E. Anderson, Thos. P. Billings, Frank K. Cameron 2 BEESON, J. J.-Mining Districts and Their Relation to Structural Geology. Discussed by Charles Keyes, Chas. A. Porter 4 DOBBEL, CHAS. A.-Deep-hole Prospecting at the Chief Consolidated Mines. Discussed by G. W. Crane 10 CRANE, G. W.-Notes on the Geology of East Tintic. Discussed by G. W. Crane, H. E. Havernor, N. O. Lawton 11 SACKETT, B. L., BARDWELL, CARLOS, JACOBSON, SIMON, and JENSEN, N. H. Lead Smelting in Utah. Discussed by D. W. Jessup, B. L. Sackett, E. H. Laws 13 BARDWELL, E. S., and MILLER, Roy H.-Pulverized Coal as Fuel for Copper Refining Furnaces. Discussed by L. D. Anderson, E. S. Bardwell, C. P. Crawford 15 WIGTON, G. H.-The Chief Consolidated Volatilization Process and Mill. Discussed by Thomas Varley, S. G. Olmstead, V. O. Lawton, O. C. Ralston.... 16 JESSUP, D. W.-Evolution in the Preparation of Ores for Lead Blast Furnaces. Discussed by W. Spencer Reid, D. W. Jessup, Carlos Bardwell 19 ALLEN, H. P., and MADGE, WM. C.-Chloridizing Mill of the Standard Reduction Co. Discussed by Thomas Varley, H. R. Hanley, H. P. Allen, George H. Dern 20 TUCKER, E. L., and HEAD, R. E.-Effect of Cyanogen Compounds on Floatability of Pure Sulfide Minerals. Discussed by Blarney Stevens 24 Mix, B. E., and BARKER, L. M.-Method of Unloading Ores and Coarse-crushing Practice at Magna Plant of Utah Copper Co. Discussed by W. C. Madge,B. E. Mix 25 KEIGLEY, C. T.-The New Byproduct Coke Plant of the Columbia Steel Corporation. Discussed by W. M. Grant, C. T. Keigley, A. C. Feilder 25 JACKSON, P. W.-Blast Furnace Plant of the Columbia Steel Corporation. Discussed by W. R. Phibbs, H. M. Henton 26 OLDRIGHT, G. L., KEYES, H. E., and WARTMAN, F. S.-Production of Ferric Sulfate and Sulfuric Acid from Roaster Gas. Discussed by G. D. Van Arsdale William E. Greenawalt, G. L. Oldright 27 MAZANY, M. S.-Braden Copper Company Caletones Smelter. Discussed by A. B. Young, G. L. Oldright 36 HARRINGTON, D.-Safety Methods in Utah Coal Mines. Discussed by Edward K. Judd, Otto Herres, Benedict Shubart, A. C. Watts, D. Harrington 36

Jan 12, 1925

By AIME AIME

ALL technical men who are planning to visit the Century of Progress Exposition in Chicago this summer-and all reports indicate that it will be worth visiting-should try to be there during Engineers&apos; Week, beginning June 26. At that time most of the national engineering societies of the country will hold important meetings in that city. Engineers will thus have an opportunity not only to see the Fair at a time when special programs have been arranged for their benefit, but will have the opportunity to meet their professional brothers and to attend technical sessions of direct interest to them. The A.I.M.E. headquarters will be at the Stevens Hotel, and it will be to the interest of every Institute member who expects to attend the Fair at that time to secure his accommodations as promptly as possible. A request to the Chairman of the Chicago Section, A.I.M.E., or to the Secretary&apos;s office in New York, will bring a copy of the program as soon as it is available, together with any in- formation desired about hotel accommodations or other matters. The Chairman of the Chicago Section is J. R. Van Pelt, and his address is Museum of Science and Industry, 1525 East 53rd St., Chicago.

Jan 1, 1933

By AIME AIME

WILBER JUDSON is one of that fairly large group of mining engineers that graduated at an Eastern college, worked his way up in various jobs in the West and in the Latin-American countries, and finally came to New York as an administrative officer of a large mining company. Since coming to New York he has been a loyal friend and counselor of the A.I.M.E., and is now serving a second term as one of its Directors, his first term covering the period 1935-1938. Mr. Judson was born in the capital city of Michigan on July 26, 1880. Probably it was his roving spirit rather than professorial urging that was responsible for the fact that four colleges were responsible for his education: Northwestern, University of Michigan, Harvard (S.B., 1901), and the Michigan College of Mines (graduate study). In his first ten years of professional work he might have been found serving as mine foreman at Catamarca in the Argentine, as superintendent of the Inter Ocean mine at Sunshine, Colo.. as foreman for the Cerro Prieto, Magdalena, Sonora, as superintendent of the Hudson Iron Co. in New York, or as manager of the Dolores mine in Chihuahua for the Mines Co. of America, or in Spain, or in Panama.

Jan 1, 1942

By G. R. Speich, L. Zwell, H. A. Wriedt

HORNBOGEN1 has recently shown that it is possible to measure the degree of segregation accompanying cellular precipitation in ferritic Fe-Zn alloys by measuring the lattice parameter of the ferrite phase within the cell. In extending this work to study a wider range of composition and to evaluate the free-energy change accompanying the growth of the cells, the lattice parameter and the boundary of the a! phase in Fe-Zn alloys were redetermined to check the earlier work of Schramm.2 This appeared to be necessary since the Fe-Zn diagram has been drastically revised since the work of Schramm. Stadelmaier and Bridgers3 have recently shown that, contrary to prior belief: the y iron field is closed and the a phase boundary extends to an a + T + L peritectic at 782°C where peritectic a contains 42 at, pet Zn. The phase T in equilibrium with the a solid solution has a composition near Fe3Zn10 and a cubic 081 type structure.&apos; The alloys were prepared by two different methods. In the first method carbonyl iron powder (400 mesh, 99.8 pet pure) and high-purity zinc powder (200 mesh, 99.99 pet pure) were mixed in a ball mill and then pressed in a die at 40,000 psi pressure to make compacts with a diameter of 5/8 in. and a thickness of l/4 in. These were sealed in evacuated quartz capsules, gradually heated to 1100°C over a period of 1 week (3 days at 600°C) 3 days at 800°C, 1 day at (1100°C), and then quenched in iced brine. For the alloys containing 23.4 and 30.1 at. pet Zn, the final homogenization anneals were performed at 1000" and 850°C) respectively, to prevent melting. In the second method, alloys were prepared by saturating high-purity iron millings (0.001 in. thick, 99.9 pet pure) in zinc vapor. For each alloy the impregnation was conducted at a fixed vapor pressure of zinc and a fixed temperature between 750&apos; and 800°C. When saturated, the millings were quenched in water. The compositions of the alloys prepared by both methods are shown in Table I. The lattice parameters of the ferritic solid solutions were determined by comparing the 26 values of their 310, peaks (using Co Ka) or of their 211 , peaks (using Cr Ka) with those of corresponding peaks for high-purity a iron. The 20 values were measured at room temperature with an X-ray dif-fractometer (G.E.-XRD5). By using a fiducial value of 2.8664Å for the lattice parameter of pure a iron at 25oC,5 the lattice parameters of the zinc-bearing

Jan 1, 1964

Prospectors were digging silver in the Ray mine area in 1873, and by 1880 high-grade copper ore was feeding a 30-ton copper furnace. In 1910, D. C. Jackling and his associates organized the Ray Consolidated Copper Company, known in mining circles as Ray Con, to exploit the Ray orebody. For 50 years, all mining was underground and the Ray mine was known the world over for innovations in mining methods. It was the first to mine important tonnage (8,000 tpd) by block caving methods. The mill at Hayden, 26 miles (41.6 km) from the mine, kept pace with the new mining methods developed at Ray and in the early 1920&apos;s flotation replaced gravity concentration by the vanner machines. First mining by open pit methods began in 1950 and in January 1955 all underground mining was discontinued. The new primary crushing plant, installed in early 1967, featured design simplicity with a low capital cost. The 54 in. gyratory crusher was installed within a total height of 6 1 ft (18.65 m) from the truck dump road bed to the floor beneath the feeder-conveyor under the surge pocket. The crusher is fed by 120-ton trucks dumping from one side only into a 200-ton hopper equipped with a horizontal splitter bar to protect the crusher spider. The crushed ore pocket, of 120 -ton live capacity underneath the crusher, discharges to a 0 to 200 fpm (61.2 mpm), variable-speed, 72 in. by 80 ft long (1.83 m by 24.4 m) sloped belt feeder, probably a first in the industry. Wood-picking stations are provided on either side of the belt so that old mine timbers and tramp iron can be removed from the crushed ore and the wood can be dropped into Dempster containers located below the belt. Counter-weighted finger gates control the bed depth on the belt at 14 in. (356 mm) and a low- level radiation unit automatically maintains a minimum of a 3 ft bed (900 mm) of material on the belt directly under the surge pocket. The conveyor is skirted for its full length and is approximately 60 in. (1.52 m) wide inside its metal skirt liners. Ten impact idlers at 12 in. (305 mm) spacing are provided under the loading pocket and remaining idlers are spaced at 18 in. (457 mm). The belt is 1 ¼ in. (31. 8 mm) thick; and, while it has occasionally been damaged by tramp material, belt life averages 11 million tons. When the belt is changed, the coiled new belt is lowered into position onto one of two belt stands mounted over the feeder, the old belt is cut, and one end is attached temporarily to the new belt. The other end of

Jan 1, 1978

By R. D. Pehke, J. F. Elliott

The solubility of nitrogen in liquid pure iron has been measured as a function of pressure and temperature. Sieverts&apos; Law is obeyed at all pressures up to 1 atm and the temperature coefficient of solubility is 8 x 106 %/°C. The solubility of nitrogen in liquid iron alloys has been studied by the Sieverts&apos; method and by the sampling method. The solubility is decreased by Al, C, Co Cu, Ni, 0. Si atzd Sn, and it is increased by Cr, Cb, Mn, Mo, Ta, W, and V. Interaction coefficients eN, for the effect of these elements on nitrogen dissolved in liquid iron, are reported. 1 HE effect of nitrogen upon the properties of steel may be desirable or not, depending upon the composition, processing treatment, and the use of the product. To understand the behavior of nitrogen in the various steelmaking processes, it is necessary to have reliable fundamental data pertaining to the solubilities, rates of solution, and chemical reactions of nitrogen in steel, particularly in the liquid state. In an effort to extend our knowledge in these areas, an investigation of the thermodynamics and kinetics of the solubility of nitrogen in liquid iron alloys was undertaken. The thermodynamic aspects of the study are presented in this paper. The solubility of nitrogen in liquid pure iron has received the attention of several investigators whose results are summarized in Table I. Reasonable agreement among the more recent researches1-l8 is found for the solubility of approximately 0.044 wt pct at 1600°C and 1-atm pressure of N,, for the temperature coefficient of solubility being small but positive, and for adherence to Sieverts&apos; law for pressures up to 1 atm. The solubility of nitrogen in a number of binary liquid iron alloys has been studied: Fe-A1,3 Fe-As,16 Fe-C,3, 71 18, 19, 20 Fe-Cr ,5,7, 10, 11, 15 21 Fe-Co,&apos;4, &apos;6 Fe-Cu,&apos;6 Fe-Mn,10, 11, 18 Fe-Mo,13, 16 Fe-Ni,10, 11, 13, 15, 16 Fe-O,16, 17, l9 Fe-p,7 Fe-S,l6 Fe-Sb,16 Fe-Se,lg Fe-Si,4) 97 &apos;77 187 19>2° Fe-Sn,16 and Fe-v.51 &apos;3 A few ternary systems have been studied, most of which were reviewed by ~angenberg" who applied the methods suggested by WagnerPz3 Chipman,24 Darken and G~rry, and Morris and Buehl26 for predicting solubilities in multicomponent alloys. Langenberg22 and Humbert and Elliott15 compared the solubilities computed from simple binary alloy systems with experimental measurements of nitrogen solubilities in liquid iron containing two or more alloying elements. The interaction coefficients for nitrogen in

Jan 1, 1961

By C. S. Barrett, D. F. Clifton

THE transformation that occurs in lithium and its solid solutions containing magnesium1,2 is similar in many respects to other diffusionless transformations of the martensitic type. This general similarity makes it desirable to investigate the transformation curves in detail. The current interest in the phenomenon of stabilization in martensitic transformations has created a need for a careful appraisal of possible stabilization effects in both cooling and heating transformations in lithium alloys. The effect of prior heat treatment on the temperature at which transformation begins and the effect of various temperature cycles in partially transformed alloys also merit attention. Limited data on some of these features were presented previously,&apos; but the present studies have been made with improved apparatus, in which erratic fluctuations were smaller than in the earlier work, and include many independent determinations of certain critical observations. Material and Methods Geiger counter X-ray spectrometers were used: a Norelco instrument, original model, and a General Electric XRD-3. The low-temperature specimen holder has been described elsewhere." Specimens 10 mm in diam and 1.5 mm thick were held in firm contact with a copper block which was attached to the lower end of a stainless steel tube that extended up to a reservoir of liquid nitrogen. A heating coil on the tube permitted a controlled temperature gradient along the tube, so that desired temperatures and rates of heating and cooling could be obtained by controlling the heating current. The specimen was surrounded by an atmosphere of evaporated nitrogen in an insulated chamber having double cellophane windows. The temperature of the specimen was read by a thermocouple imbedded in it close to the irradiated area; the specimen could be held at constant temperature within 1°C for long periods. The body-centered cubic 200 reflection near 8 = 26" was used as the index of the amount of b.c.c. material present; occasional checks were made using the most suitable line of the low-temperature phase, the close-packed hexagonal 110 line near 8 = 29". Copper radiation was used throughout. A continuous record was made of the reflected intensities on a strip chart recorder, both when the peak intensity of a line was being followed and when the lines were being scanned at constant temperature. On the continuous recordings of peak intensity, periodic checks were made on the background intensity between the lines, and on the setting of the spectrometer for maximum intensity. The lithium-base solid solution containing 12.4 atomic pct Mg (33.3 wt pct) that had been used previously&apos; was chosen, as it had a transformation range at relatively high temperatures, so as to permit wide temperature cycling within the range. Just prior to use a pellet was cut from the ingot,

Jan 1, 1951

By D. W. McGlashan, K. N. Mcleod

IN the use of free bubbles with precise temperature control and continuous pH measurement, the contact angle apparatus differs from all previous equipment. Experimental procedures differ sharply from the captive bubble method1,2 of introducing the gaseous phase in the three-phase system. When freed from the tube and captured by a solid surface, a bubble more closely conforms to the spherical ideal set up by the usual mathematical analysis of surface free energy or surface tensional relationships of the forces involved. The assembled equipment, Fig. 1, departs from the usual contact angle apparatus in providing: 1) water circulation for fixing temperature, 2) instruments to measure and record pH during testing, and 3) a system for producing and using free bubbles. The apparatus allows testing over a broad range of closely controlled conditions, and the use of free bubbles gives contact angles independent of operator manipulation, which was unavoidable in the older captive bubble method. A bubble released below the specimen rises through the liquid and contacts the mineral. If the mineral is gas-avid the bubble adheres, developing a finite contact angle. Reaction cells (1), 2x33/8x5 3/16 in. ID, are of pyrex glass 5 mm thick. To prevent optical distortion, the sides are optical flats and are planed parallel to within five wave lengths of light over the entire surface. Special cement is required to withstand chemical attack and thermal shock. The cell is placed in a thermal jacket (2) made of welded aluminum. A water seal for the optical ports is made by using O-rings tightened against the flats of the cell by fine-threaded and adjustable ports, Fig. 2. In-flow water is discharged downward from a manifold located near the bottom of the cell jacket. Water overflows the weirs at each end of the cell jacket, discharging from both sides to a collecting manifold (24). A hose (26) leads the overflow water to the thermal-conditioner well (4), which is recessed to accommodate a reaction cell and two beakers (150 ml). Both the thermal jacket and the thermal-conditioner well are insulated to minimize the effects of temperature differentials. Water, discharging from the thermal-conditioner well, returns to a thermostat bath and pump (3), from which it is pumped to the thermal jacket. The thermostat bath permits temperatures to be adjusted from 1" or 2°C to 80°C and will maintain the selected temperature within 0.2ºC. An electrode support (22) groups the four electrodes (8) so that they fit conveniently into the reaction cell, Fig. 2. This assembly holds the glass electrode, the calomel electrode, the temperature-

Jan 1, 1957

By George Comstock

IT seems a common opinion among metallographists that all light-gray inclusions seen with the microscope in polished sections of steel are manganese sulphide. Examples of this belief are continually appearing, as for instance in the paper by Dr. Henry Fay on manganese sulphide as a source of danger in steel rails,1 and in Lieutenant-Commander Cook&apos;s paper on Metallography of Steel for U S. Naval Ordnance.2 Slate FIG. 1-TRANSVERSE VIEW OF THE EDGE OF A SMALL HOT-ROLLED STEEL ROD. ETCHED WITH PICRIC ACID. X 200. colored inclusions are considered to be silicates, and dove-gray inclusions, manganese sulphide. To show the danger in the latter unqualified assumption, it should be sufficient to examine the edge of any piece of steel that is covered with a fairly thick scale, and that has been polished in such a way as to prevent the scale from breaking away entirely below the polished surface. This can be done by protecting the scaled edge of

Jan 12, 1916

By Lewis Robert M.

The importance of phosphate in feeding the people of the world has been recognized by mining companies as they continue their search for new ore deposits and ways of improving phosphate production. An extensive phosphate-recovery research project involving investigation of beneficiation variables by batch testing, flotation pilot-plant processing of the major portion of the ore, gravity separation of the coarse fraction, and recovery of byproduct heavy minerals of a North Carolina phosphate deposit has been completed. Data obtained through batch tests enabled the development of an economical flowsheet which proved successful for the efficient recovery of a good quality phosphate concentrate in a pilot plant.

Jan 1, 1975