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ANYONE who wishes to make the patterns for statues in order to cast them later in bronze, even though he himself is the artificer who made them, must take many things into consideration: first, whether they are large or small, and whether they are wholly round or only in part; then of what material they are made, for the way that can be followed if they are of wax cannot be used if they are of clay, wood, marble, or even of plaster; likewise he must consider whether or not it has irons inside to support it. There is also a great difference in facility if the support is placed more in. one way than in another. Great consideration should be given to whether it is to be made hollow or solid. If you have not been the workman yourself, you must at least have been an active helper in this and in every other part in order to be able to observe everything without a fault. For if they are large it is necessary to make their moulds in pieces, with marks and measurements so that you can replace them-at will exactly in the same order and in the same places after they have been taken out. But if you are the artificer [as well as sculptor] the process is made easier. Taking the ordinary way and, over an iron, make out of well beaten moulding clay, compounded with cloth clippings a core exactly as the statue is to be or however large you wish the hollow to be. Dry and bake it. On top of this work the wax, thickening it all over as much as you wish the bronze to be and finish it completely, remembering to arrange that the base may come out solid with bronze so as to support the whole work. Then take some rods of hammered bronze one dito thick and one palmo long. Pass these throughs the thickness of the wax and fasten them in several places in the core of clay that you made. [81] Do it so that they, project out of the wax a good three or four dita. These are, put in so that when the wax is removed the core is held in the middle of the mould that forms the outer part, and this outer part is held by the inner part, namely the core, so that the mould may not move in any part from its place even when it is handled. This done, take some of that clay made with well-prepared cloth clippings as I taught you, and mix with it if you wish some wash ashes or

Jan 1, 1942

By E. Rudy, St. Windisch

The only complete phase diagram investigation of the Ti-W system is by Maykuth and coworkers,&apos; Fig. 1. Although previous work by Gonser2 indicated complete miscibility, later investigations3°5 essentially confirmed the diagram depicted in Fig. 1. The present study was undertaken because preliminary results on metal-rich ternary Ti-W-C alloys indicated complete solid solubility between 0 titanium and tungsten at temperatures above 1300°C—in contradiction to the presently accepted diagram. The powdered elements, as well as crystal bar titanium and tungsten sheet, were used in the preparation of the experimental alloys. The iodide titanium, obtained from Foote Mineral Co., Exton, Pa., had a total impurity content of less than 410 ppm, of which 150 ppm were interstitials (H, O, N, C). The overall purity of the titanium powder (Var-Lac-Oid Chemical Co., New York) was better than 99.7 pct. Major contaminants, in ppm, included: 0, 800; N, 10; H, 1500; Fe + Co + Ni + Cu + Zn, 50; C1, 200. The measured lattice parameters of the titanium were a = 2.94A and c = 4.68,A. The tungsten sheet and the tungsten powder (Wah Chang Corp., Albany, Ore.) contained the following impurities (in ppm, values in parentheses refer to the sheet tungsten): Mo, 50 (70); 0, 140 (110); Fe, 20 (60); Ni, 20 (not determined); N, 110 (not determined); sum of other impurities, <60 (<100). Samples from the iodide titanium bar stock and from tungsten sheet were prepared by repeated melting of the chipped materials in a nonconsumable electrode arc furnace under high-purity helium. The melted alloys were heavily cored, and homogeneity of the alloys could only be achieved after prolonged heat treatments at subsolidus temperatures. The majority of the alloys were therefore prepared by cold compacting the well-mixed powders in steel dies and pre-homogenizing the compacts for 6 hr at 1300°C in a tungsten-element furnace. The vacuum was better than 10"6 Torr, but for additional protection the sample container was wrapped in zirconium foil. Both preparation methods led to identical results. The measurements of the solidus temperatures were carried oat in the Pirani-type furnace described in a previous paper.6 Incipient melting in alloys containing between 20 and 90 at. pct W was rather difficult to detect visually, and therefore a photoelectric transducer system7 was employed to indicate the phase change. For the investigation of the solid-state equilibria, the specimens were first homogenized for 2 hr at slightly subsolidus temperatures in the Pirani fur- nace and then reequilibrated at, and rapidly cooled (-100°C per sec) from, the selected equilibrium temperatures. After the runs, the alloys were analyzed microscopically and by X-ray diffraction. With the exception of spot checks for oxygen and nitrogen in processed alloys, no other chemical analyses were performed. However, as the weight changes of the alloys during processing were within 0.05 pct and the lattice parameters of indeqendently prepared alloy series agreed within 0.0015A, it is reasonable to assume that no significant concentration shifts with respect to the nominal compositions had occurred. The measured solidus temperatures, shown together with literature data in Fig. 2, ascend smoothly from

Jan 1, 1969

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

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

Jan 1, 1961

By 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 M. L. Requa

This paper is abstracted from the report of the Appraisement Committee of the Independent Oil Producers&apos; Agency, of which the writer was Chairman. The other members of the committee were M. V. McQuigg and R. S. Haseltine. To Thomas Cox, engineer in charge of the investigation, and to his efficient staff of assistants, is due the credit for the detail work necessary in carrying out the plan of the Committee. It treats only of the valuation of the lands, wells, and piping, without going into details of equipment and other features. It was recognized, early in the operation, that the value of oil lands was the value of oil to be extracted, and marketed, less the costs attendant thereon. Referring this back to the costs in cents per barrel, it became obvious that surface improvements, absolutely necessary in the production of oil, from wells then drilled; had no separate and independent value aside from the value of the land, but that any excess equipment, not necessary for the production of oil from present wells, such as drilling tools, well casing, etc., should be given a present condition value. In valuing the properties of the Independent Oil Producers&apos; Agency, the Committee has considered the following facts: 1. That from each property there will ultimately be. produced a certain total quantity of oil. 2. That in the production of that oil a certain total quantity of money will be expended. 3. That a certain total amount of money will be received from the oil. 4. That the total net receipts will be the total gross receipts less the cost of development and production. 5. That the present value of the net receipts must be such an amount that when invested by a purchaser it will be returned with 8 per cent. interest additional during the life of the property.

Jan 2, 1918

By Alfred R. Powell

LATE in 1944 a group of petroleum and coal technologists was organized in Wash¬ington under the sponsorship of the Petroleum Administration for War and the U. S. Bureau of Mines. This group, known as the Technical Oil Mission, was placed under the authority of the Technical Industrial Intelligence Committee of the Joint Chiefs of Staff, which was then promoting the organization of other missions to be sent to Germany to obtain technical information regarding various industries in that country. The objective of the Technical Oil Mission was to obtain technical data on German production of synthetic oil from coal and all other commercial practices and research projects connected with the oil industry. The primary purpose was to obtain such data as were immediately essential to assist the war against Germany and Japan, with secondary emphasis on technical information that would assist American industry after the end of the war. Since oil-production facilities in Ger¬many and the occupied countries con¬trolled by Germany were the top targets for strategic bombing, the Technical Oil Mission was the first group under the Technical Industrial Intelligence Com¬mittee to enter Germany. A few members of the group flew to London in February 1945 and the majority of the remainder followed in March. In London they were joined by an equal number of British investigators and the combined, mission was then sent into Germany under the authority of the Combined Intelligence Objectives Subcommittee (CIOS). A sum¬mary of the more important information gained by this combined British and American mission is presented below. GERMAN OIL PRODUCTION The oil available to Germany, both natural and synthetic, and including pro¬duction from occupied countries, reached a peak in late 1943 and early 1944. Pro¬duction at this time was at the rate of 1,200,000 metric tons per month, equiva¬lent to i00,000,000 bbl. per year, or about 7 per cent of the wartime production in the United States. Of this production, 50 per cent was crude oil from Romania and Hungary, 12 per cent was crude oil from Germany itself, including Austria, 9 per cent com¬prised coal-carbonization products, that is, tar and light oil, and only the remaining 29 per cent was synthetic liquid fuel made from coal or tar. Synthetic liquid fuel made in Germany was equivalent therefore to less than 2 per cent of war¬time production of crude oil in the United States. The success of the strategic bombing campaign against German oil-production facilities is demonstrated by the fact that by September 1944 the Ploesti fields of Romania and much of the syn¬thetic oil production had been eliminated, so that the amount of oil available to the German war machine was only 25 per cent of the peak quantity mentioned above. By early 1945, when the combined British and American oil mission was entering

Jan 1, 1946

By Edward L. Fidler, Thomas E. Finch

Thin seam coal mining in the western US nominally equates to recovery of stray seams associated with thicker, major seams. The thin or stray seams encountered are generally recovered, but not in all cases. The alternative to extraction is, of course, spoiling. The majority of thin seams are encountered somewhere in the overburden above the major seam or in the parting between thick seams. Although some are true strays and appear anywhere in the geologic sequence, others might be classified as thin "rider" seams immediately above or below a thick seam and separated by some equally thin parting. For the purpose of this discussion, stray or thin seams are defined as those which are less than 1.2 m thick or represent less than 10% of the expected extractable coal. [Figure 1] is a graphic representation of typical thin, stray, or rider seams which actually appear in a western operation. Operationally, stray seams create irritating scheduling problems and require special handling techniques which do not always lead to efficient equipment utilization. Additionally, coal recovery from a thin seam is considerably less than the 95% usually recovered from thick seams. Higher recovery in thin seams leads to dilution problems, which depreciates Btu quality, and most western coals are not in a position to lose heating value. Thin western seams also undulate and pinch or swell more than thick seams, and this leads to more difficult recovery. Often they are associated with equally dark carbonaceous shales and visual selectivity is difficult at best during daylight hours. Resource Recovery Regulations Thin seam recovery in western mines is encouraged by both federal and state regulations and by the mining industry&apos;s concern for total resource recovery. [Table 1] lists the western coal states and their recovery regulations. The federal regulations are found in the Surface Mining Control and Reclamation Act of 1977. They state, "General performance standards shall . . . require the operation as a minimum to … conduct the surface coal mining operations so as to maximize the utilization and conservation of the solid fuel resource being recovered." Economic stimulation for thin seam recovery appears to be marginal or at the least controversial. Scheduling and operational problems often outweigh the nebulous profits created by engineering analysis. The only alternative to extracting thin seams occurring above major production seams is to strip directly through them and lose the resource to the spoil pile. Although this is not palatable to environmental emotions or resource promoters, it is the most simple expedient to operations people.

Jan 1, 1981

By A. Gatti

Studies were made on the system iron plus alumina. Various methods of dispersing and various amounts of alumina were used. Powder metallurgy techniques were used to produce the final product. Microstructures and some mechanical properties are presented. Alloys containing up to 16 pct alumina by weight are ductile at all temperatures tested. The ivon-alumina materials have a higher yield stress and improved creep resistance over that of pure iron. SINCE the development of SAP&apos; an A1- Al2O3 powder metallurgy product, great interest has been shown in the development of a SAP-like structure in metals and alloys other than aluminum. The outstanding properties of this class of materials are: 1) Good creep resistance at high temperature. 2) High yield strengths and high hardness at high temperatures. 3) High recrystallization temperature. However, the lack of ductility and formability has thus far limited the applications of SAP-like materials. The following report deals with methods of dispersing A12O3 in Fe, processing of powders into rod, structures produced after processing, and mechanical properties of the iron-alumina material produced. Methods of dispersing the alumina and also the volume fraction of alumina added were varied. MATERIAL Powder Processing— Four methods of obtaining dispersions of alumina in iron were studied. 1) Oxidation-reduction of a Fe + 8 pct Al alloy. .2) Coprecipitation of Fe(OH)3 + Al(OH)3 from aqueous solution. 3) Colloidal mixing of Fe,O3 and Al2O3. 4) Colloidal mixing of Fe powder and A1,O3. Fig. 1 is a processing flowsheet which outlines the processing steps taken to produce a final extruded bar 1/4 in. in diameter and about 30 in. long. The - 100 mesh iron-aluminum alloy was oxidized by placing a thin layer of powder on an alumina plaque. The reduction step was modified to include 1/2 hr at 1000°C since the powders tended to become pyro-phoric unless sintered slightly after low-temperature reduction. Compacting was done in a 1-1/8-in.-diam pressing die. The slugs produced averaged 1 in. in length. The billets were encased in low-carbon steel cans shown schematically in Fig. 2. The blank is included for structure and property comparison purposes. TESTING PROCEDURE All tests were made on as-extruded material.

Jan 1, 1960

By C. E. Berry

Natural stone or manufactured porcelain pebbles are used as the grinding elements in pebble mills and the mills are lined with stone or porcelain blocks. Steel balls usually form the grinding medium in ball mills, and the mills may be lined with steel. Pebble mills are used to avoid contamination by metal of the material being ground, to avoid rapid wear due to corrosion, and for other specialized reasons. Belgian silex (silica) blocks formerly were used widely for lining pebble mills, because of their good resistance to wear, light color, and low cost, but the importation of these blocks was suspended early in the war. The stocks of Belgian silex diminished so rapidly that it was found desirable to compare the properties of a number of domestic substitutes for use in critical operations. Accordingly, a study was made to determine the best substitutes for fine grinding of white pigment and for certain heavy-chemical process work such as the disintegration of phosphate rock in weak phosphoric acid solution. A comprehensive survey of substitute lining materials was not contemplated. It was stipulated at the beginning of the investigation that only materials on the market, available in large quantities, would be tested and that these should be reasonable substitutes from a wear-resistance viewpoint as determined from known physical characteristics. Marble and limestone, for example, could be eliminated immediately. Quartzite, granite and porcelain were selected as likely substitutes for Belgian silex and were subjected to quality tests. Materials Tested 1. Belgian Si1ex.—Sample blocks of a standard grade were taken from crates in which they had been imported. The stone was light gray in color and contained inclusions of small clusters of quartz crystals as well as light colored portions that might be calcite. This was a siliceous rock which appeared to be of igneous origin. 2. Quartzite.—Samples of quartzite were obtained from a quarry in Minnesota. The blocks exhibited a crystalline fracture. The stone was pink in color and composed of fine quartz crystals bound together with a siliceous cement. 3. Granite A was a fine-grained granite quarried in South Carolina. 4. Granite B was coarse-grained granite quarried in South Carolina. 5. Granite C was a coarse-grained granite quarried in North Carolina. 6. Porcelain A.—Sample blocks of the pebble-mill lining material were obtained direct from the manufacturer. It was white in color and of lower specific gravity than the natural materials. It did not have a crystalline structure. 7. Porcelain B.—Sample pebble-mill lining blocks of the porcelain were received direct from the manufacturer. It was white

Jan 1, 1947

By Norberto Dominguez

In the year 1820 a commission, assisted by Sr. C. Fernando de Ainada, was appointed by Sr. Jose Ramon Mila de la Roca to report on the condition of the mining region of Parral, in the State of Chihuahua, Mexico, where in recent years the mines of Batopilas have reached an important development under American management. The object of the inquiry was to provide Sr. De la Roca with facts to support a project for reopening mines that two centuries before had produced a large amount of silver. Although this plan failed, many facts of interest in the history of the mines were presented.. Much of the report dealt with questions that had only local application or were restricted to the conditions of labor and equipment in the opening of the nineteenth century. The historical part is presented in the following paragraphs: Abstract of the Report. Notwithstanding the difficulty of forming a just estimate of the system of mines in this mineral group in the valley of Senor Sail Jose del Parral without an examination of each of them, that which we shall say in a general way will suffice to convey an idea of their value and of the urgent necessity of reopening them. If this is undertaken, we believe the results will be of the highest consequence to the Republic, as well as to this immediate locality. The group of the Villa del Parral covers the mineral territory in which the village of that name was located. Its moantains contain large metallic veins, from which, as from the trunk of a tree, depend others much more slender, locally called ,fibras, crossing and interlacing each other in many directions. The first and oldest of these very rich deposits of gold

Jan 1, 1902

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine ( — 325-mesh) sand assaying 95.5 per cent SiO2, and a very fine (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotalion Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine frac- . tion was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1948

By E. H. Rose

THE present symposium on flotation machines provides a unique opportunity for group reappraisal of standard or habitual viewpoints. A symposium is by definition a trading post for observations that might prove complementary when put together, thus leading to useful thoughts that might not have been evident from the original observations individually. The present paper, pertaining to power consumption in flotation, is deliberately an experiment of that kind. Data were solicited from the operating superintendents of several of the major flotation plants on this continent. The information received, covering a total of 14 separate operations in seven plants, has been collated in terms of power intensity; that is, the total power input per cubic foot of cell volume. This departs somewhat from the conventional expression of power per ton of ore treated, though the latter figure is also given for the plants concerned. Because of wartime restrictions, particularly in Canada, identifying names of plants and machines are withheld. Not as a firm conclusion, but as a tender for symposium discussion, the following points are made: I. That flotation is a power operation in much the same sense as ball milling. 2. That contact time is the complement of power intensity and may be decreased as the latter is increased. 3. That a decrease in flotation power per ton of ore can result from an increase in power intensity. ANALOGY WITH GRINDING In grinding, it has become axiomatic that the more the power put into a given operation, the greater the work performance. The analogous view is now proposed that the effectiveness of any flotation operation will be related to the mechanical energy usefully applied to that operation. Under this view, each increment of properly directed input energy should give either an increase in unit tonnage or an improvement in the metallurgical separation (concentrate grade or tailing grade), until either the theoretical limit imposed by mineral constitution or surface condition is reached, or until further increments of power become uneconomic under the law of diminishing returns. Thus, as in grinding, there are two forms of output that greater power input or improved utilization of power can produce. As for the law of diminishing returns, lower unit cost of power through improved generating efficiency in the past decade has pushed the limit back. Over and above this, however, there is strong indication in the operationg data herein presented that higher power intensity results in an actual economy of over-all power consumption. Like crushing and grinding, flotation accomplishes a rupture of the forces of molecular attraction, by taking particles from the common environment and placing them in the two separate environments of

Jan 1, 1944

By A. A. Collins

OF widespread interest to all lead metallurgists is a dross smelting process that will consistently give mattes and speiss of low lead and high copper contents. It is a problem that has intrigued operators through the years. In the past few years, with increased amounts of copper and overcrowded metallurgical furnaces, it has been almost imperative to evolve some method. Experience at El Paso has shown that high copper-to-lead ratios can be consistently attained by use of the new soda process. † Older Methods Three older methods, i.e., 1. Dross plus siliceous ore and limestone. 2. Dross plus one fourth blast-furnace speiss and siliceous ore, 3. Dross plus sulphides and siliceous ore and scrap iron, do not give consistent results and the products are higher in lead and lower in copper than those produced with the soda process. Any radical changes in the blast-furnace drosses will throw any of these older methods into a tailspin. For almost ten years, El Paso used the third method, with varying results. It was noted that as the arsenic and anti- mony contents of the dross increased, the speiss and matte steadily increased in lead and decreased in copper. A dross containing four parts of copper to one part of arsenic was considered to be ideal. At times this balance of copper to arsenic was almost impossible to obtain, especially when handling copper Cottrell dusts, which vary so greatly in copper and arsenic. Custom smelters like El Paso, handling a great variety of dusts, concentrates and ores, are apt to have anywhere from I to 15 pct copper on charge, with a variable in arsenic of from 0.5 to 2 pct. With such conditions, and drosses varying from 12 to 40 pct copper, an operator using the older dross reverberatory practices would find it just about impossible to control the reverberatory operations and to produce a good product low in lead and high in copper. These conditions prevailed at El Paso from the year 1940 to the present. In May 1942 it was decided to experiment with the soda process. The Soda Process After considerable work the soda method was finally evolved to the present state. It has never failed to produce a good product even under the most adverse conditions. It might be well to remember that many men of thought in the smelting industry have at one time or other sugge~ted the removal of all copper, arsenic and antimony from the lead blast-furnace charge as an aid to better metallurgy.

Jan 1, 1949

By W. A. West

ONE who attempts to collect and classify equilibrium data from various iron-alloy systems is soon struck with the absence of any quantitative theory that can serve as a general background against which to compare and contrast results from different systems, at different temperatures, and with different concentrations. It is not necessary to wait for a perfect theory; investigation of divergences of theory from observation may prove of great value in acquiring understanding of the subject. For example, one has only to think of Raoult&apos;s law as applied to aqueous solutions. The equations developed by Zener1 represent an effort to provide such a quantitative theoretical background, and it is therefore desirable to subject this theory to such tests as can be applied. The present article is intended to be an application of available tests, and may thus be considered as a supplement to Zener&apos;s paper. It is assumed that this paper is available to the reader. The equations make it possible to calculate the effect on the equilibrium concentration of one alloying element due to the presence of one or more other elements. In particular they are intended to give, for alloy steels, the carbon concentrations in alpha and gamma phases in equilibrium with each other, and in gamma in equilibrium with cementite. They may, however, be modified for use for carbon-free alloys. A summary is given later of the equations used in analyzing the data in this paper, and the symbols are there defined. About 125 papers have been examined for experimental data suitable for this investigation. An effort was made to look up all available reports of ternary systems including iron and carbon, as well as similar systems of higher order. Supplementary material was also obtained from a number of binary systems. A few noncarbon ternary systems were also found to which these methods could be applied. A comparatively small proportion of the papers read provided usable material. In many systems the number of quantitative observations made in the iron-rich region is extremely small, and the smoothed curves to be found in most published phase diagrams are entirely useless for quantitative treatment. This opinion is confirmed by the authors of the series on Alloys of Iron Research, whose remarks show clearly that they have no illusions as to the accuracy of many diagrams that they present. A further difficulty arises from the fact that these equations apply only to rather low percentages of alloying elements. The result is that at the lowest concentrations, where the best agreement might be hoped for, the errors of observation become relatively very large. The experimental difficulties in measurements of this kind of equilibrium are, of course, an old story.

Jan 1, 1946

By Donald W. Scott

GENERALLY speaking, there is nothing very difficult about taking good micrographs of photogenic thin sections or grains with a petrographic micro-scopecamera setup. However, sometimes it is desired to show some detail that cannot be illustrated satisfactorily by the usual micrographs taken with parallel polarized light, with crossed nicols, or even with two micrographs. A simple procedure giving good results has been used at Battelle Memorial Institute for subjects falling into this classification, and it is believed to be of interest, especially to research people who must sell a point to executives through the medium of well-illustrated reports. The following problem illustrates a typical application: It is desired to photograph a thin section of an iron ore to illustrate the character and association of the iron oxides and the individual quartz crystals. A magnification of 75 dia. is acceptable. A micrograph of the field in question (Fig. I[a]) was taken with parallel polarized light. Due precaution was taken to provide critical illumination, correct exposure, a sharp focus, a good darkroom technique, and other essential conditions. It illustrates correctly the iron oxides, but it fails to show the individual quartz crystals. A micrograph of the identical field (Fig. I[b]) was taken under the same precise conditions, but with crossed nicols. It gives a better idea of the character of the individual quartz crystals but is likely to be misinterpreted unless accompanied by a lengthy descriptive note. This micrograph suggests that the centrally located, large black area is composed entirely of iron oxides. That this is an incorrect interpretation is evident by reexamining the micrograph of Fig. I[a]. The correct interpretation is that the dark area contains quartz crystals oriented with their optical directions parallel to, or nearly parallel to, the vibration directions of the polarizer and analyzer of the microscope. Thus, the executive, or anyone who had not done the work personally, would get no more than a fair interpretation, even if two micrographs were included; and in the interests of both brevity and clarity, one micrograph usually is more desirable. Exactly how many quartz crystals occur in this region cannot be told from the two micrographs, and a printing of the two negatives in register would likewise not show the detail desired. Thus, it is evident that some subjects cannot be satisfactorily illustrated by the usual procedures. In photographing Fig. is the light was subdued and, during the entire exposure of 75 sec., the analyzer nicol was rotated uniformly in its mounting from a crossed position to 45° uncrossed position to introduce the desired amount of contrast in the various individual quartz crystals. The true reproduction of the character and association of the iron oxides, and the additional structure and detail in the quartz areas, is apparent. It might be

Jan 1, 1946

[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 R. G. Weavind, R. S. Young, I. Wolf

ONE of the most important fields of investigation at the Diamond Research Laboratory in Johannesburg is concerned with improvements in metallurgical practices for the diamond mining companies, with particular reference to increased recoveries of the smaller industrial stones. Considerable work has been done on the possible application of flotation processes to the diamond industry. From the viewpoint of recovery processes, diamonds in their natural state may be grouped into two classes, the water-repellent diamonds, which adhere to a grease surface, and the water-avid diamonds, which, because of their wettability, will not adhere to a grease surface and must be recovered by means other than the grease tables used for water-repellent diamonds. Experiments in the flotation of small diamonds from both kimberlite and alluvial deposits indicate the following conclusions: Water-repellent diamonds: Satisfactory flotation of —16 mesh diamonds, which are naturally water-repellent, may be accomplished using either Aero-float 25 together with cresylic acid, or Du Pont B23 as frother together with kerosine as auxiliary oily conditioner or froth stiffener. The quantities of reagents required are of the order of 0.15 lb per ton Aerofloat and 0.18 lb per ton cresylic acid. The use of acid circuits is detrimental, and for best results a pH between 7 and 9 should be maintained. R. G. WEAVIND and I. WOLF are with the Diamond Research Laboratory, Johannesburg, South Africa, and R. S. YOUNG, Member AIME, is Director of Research, Diamond Research Laboratory. TN 76B. Manuscript, March 23, 1951. A flotation period of 5 to 10 min is necessary to insure that all floatable diamonds are removed. Diamonds larger than 16 mesh cannot be recovered satisfactorily by flotation. Xanthates do not act as collectors for diamonds, and this is probably associated with their inability to impart a water-repellent surface to this mineral. Water-repellent diamonds smaller than 28 mesh can be floated quite well by the addition of a frother alone, such as pine oil or Du Pont B23. Water-avid diamonds: Partial flotation of the smaller sizes of wettable diamonds can be achieved by preconditioning with fatty acid soaps, cationic amines, and similar reagents, which will impart a water-repellent film to diamonds. Washing to remove excess reagent, followed by flotation, will result in a fair recovery, but the latter is much lower than for water-repellent diamonds. Crushed boart: Recovery of crushed boart from the waste products of industrial uses is best accomplished by first conditioning the diamonds present in the mixture with fatty acids, cationic amines, etc., removing the conditioning agent by washing, and floating in a cell of the pneumatic type. Flotation reagents found to be most suitable are Du Pont B23, and Aerofloat 25 together with cresylic acid. The optimum quantities are 1.2 g Du Pont B23, and 0.55 g Aerofloat plus 1.4 g cresylic acid, per liter of pulp in the cell. A solids-to-water ratio of 1 to 4 is found to be the most suitable, and alkaline circuits are necessary for best results. Diamonds larger than 28 mesh are not recovered satisfactorily by this procedure, but for finer sizes flotation offers an excellent procedure for the reclamation of this strategic industrial material.

Jan 1, 1952

By John Chipman, K. C. McCutcheon

To a very large extent the quality of a mild steel ingot is determined by the manner in which it is poured into the ingot mold and its behavior during the first few minutes after it has been poured. The formation of the solid rim of metal around the periphery of the mold, the effervescent action of the liquid steel within this solidified rim and the upward or downward movement of the metal level are indications of what is going on beneath the surface of the liquid metal, especially at the interface at which solidification occurs. Observation of this "rimming action" enables an experienced observer to predict the thickness of the sound skin of the ingot and the location and extent of cavities or "blowholes" caused by entrapment of gases. The control of rimming action at the hands of a skillful steel pourer is a vital part of the control of quality in mild steel sheet, and the behavior of the liquid steel during the rimming period is a useful criterion of the melting and refining practices employed in the open hearth. The behavior of rimming steel during solidification depends primarily upon the evolution of gases from the molten metal and it is therefore a problem of considerable interest to determine the nature and amount of the gases evolved. A number of attempts to determine either the amount or the analysis of the evolved gases have been reported. Klinger,1 in 1925, analyzed the gases extracted from rimming, semikilled and killed steel ingots and found that the chief constituent of the rimming gas was carbon monoxide. The relatively large amounts of nitrogen reported in his first few samples may have been due to incomplete removal of air from the gas-collecting mechanism. McKune2 reported similar studies of rimming ingots from which surprisingly large proportions of nitrogen and hydrogen were evolved. Ameen and Willners3 determined both the amount and composition of gases extracted from unkilled ingots with the aid of diminished pressure. . Their ingots weighed only 80 kg. (176 lb.) and apparently were solid within two or three minutes, so that their

Jan 1, 1938

By J. C. Sullivan, W. A. Knoll

The sinking of a new shaft at the Newport mine, Ironwood, Mich., was started in May 1931 and completed on Aug. 3, 1932. During this period, 2665 ft. of shaft in granite was completed, at an average advance of 6 ft. per 24 hr. In Novernber of 1930, experimental work was conducted on 65 ft. of shaft through 22 ft. of overburden where ledge was encountered, and this work convinced the operators that all drilling could be done with vertical holes with the blasting relief to a 4½-in. hole drilled vertically in the center of the cut. The reason for not adopting the conventional V-cut was to get away from damaging the shaft steel. The blasting in a V-cut naturally tends to throw the broken material upward, but with the cut as adopted the relief was in a horizontal direction to a large extent. The scheme of mucking employed made it necessary to keep the steel shaft sets within 50 ft. of the bottom at all times. This could not be done without serious damage with a V-cut when blasting in granite. The cut as drilled also broke 10.83 ft. per round of drilling, which was considerably more than the experience with a V-cut in the same material. The extended cut reduced proportionately the delays for charging, blowing smoke and handling drilling equipment in the shaft. Another advantage was the ease in drilling vertical over inclined holes, and the preservation of the proper spacing of the drill holes at the bottom of the cut. Drilling Air was carried down the shaft in an 8-in. flanged pipe line at a pressure ranging from 95 to 100 Ib. This line was extended to within 100 ft. of the bottom, where a valve and a 4 by 4 by 6-ft. tee was connected to it. To each side of the tee was connected a nipple and elbow, the resulting span of the two elbows being about 4 ft. Lengths of short 3-in. hoses

Jan 1, 1940

By J. K. Stanley

Iron-cobalt alloys in the range of 35-50 pct cobalt are of interest in the electrical industry because they possess the highest magnetic saturation of any magnetic material known. l1,2The magnetic saturation induction of the 35 pct alloy is 24,200 gausses, compared to 21,600 for pure iron (Fig I). This difference is significant in electrical apparatus, whcre weight or space is at a premium. Saturation values higher than that for iron can be attained up to about 70 pct CO hut economics restrict the use to as low a cobalt content as possible without sacrifice of saturation. Fig 2 shows the iron-cobalt equilibrium diagram. Up to the present time, two factors have retarded the use of iron-cobalt alloys in the electrical industry for magnetic purposes.? One is the high cost of cobalt ($I.50 per pound) and the other ig the inherent brittleness of the binary iron-cobalt alloys. Straight iron-cobalt alloys up to 25 pct cobalt are ductile and easy to process to thin sheet or strip by conventional methods of hot and cold rolling, but the alloys containing higher percentages of cobalt, including those for which the saturation magnetization is a maximum (35 to 40 pct), are increasingly brittle with increasing cobalt, at least up to 50 pct, and are consequently difficult to process, either hot or cold. Work by White and Rah13 indicated that iron-cobalt alloys could be made ductile by alloying with vanadium, and after quenching the alloys (vanadium up to about 4 pct) were found to he colt1 rollable to very thin gauges. With this as a background, we resumed our work on these alloys with the objective of developing methods for economical commercial production of wide strip. For this purpose we have concentrated on the 35 pct alloy in order to obtain the highest possible saturation value. Experiments were carried out on many alloy additions to the iron-cobalt alloys and on the response of these compositions to heat treatments. This work finally culminated in a commercially producible iron-cobalt alloy which has been named Hiperco.* Materials StudIEd and Procedur&apos; Binary alloys of iron and cobalt in the range of 35 pct cobalt are difficult to hot work (by forging or rolling) without such additions as C, Mn, Si, V, Cr, Mo, W Or perhaps others. Another beneficial effect of such additions is that they increase to a greater or less extent the strangely low electrical resistivity of these alloys about 10 microohms per cm3

Jan 1, 1948