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By H. R. Skewes

IN the course of small seale experiments on the electrodeposition, of copper, extremely brittle deposits were obtained from electrolytes prepared by leaching calcined chalcopyrite with aqueous sulfuric acid. It was not possible to strip the metal without fracture after it had been deposited from such electrolytes onto smooth oiled copper sheets, the usual plant procedure for the production of so-called starting sheets. The embrittlement did not occur with electrolytes made either by leaching the calcine with water or by treating an acid leachate to remove iron as hydrous ferric oxide. The electrolytes which produced brittle deposits were found to contain molybdenum in concentrations of 10 to 25 mg per 1, whereas those which produced ductile deposits contained <0.5 mg per 1. Synthetic electrolytes having the composition: Cu", 40 gpl; Fe", 0.5 to 1.0 gpl; Fe3+, 0.5 to 1.0 gpl; and H,SO,, 10 to 50 gpl, were prepared from reagent-grade salts; by adding ammonium molybdate, the embrittling effect of the higher concentrations of molybdenum (VI) was strikingly confirmed. The brittle deposits contained up to 0.03 pct Mo. The importance of oxidation state of the molybdenum was demonstrated when the electrolyte containing it was circulated continuously through a bed of fine copper metal during deposi- tion. Under these reducing conditions, the cathode deposit was ductile and molybdenum could not be detected. Molybdenum has not been recognized as one of the metals causing embrittlement of copper nor is it listed among the usual impurities. In most elec-trowinning plants there is a purification step which removes iron from the leachate before electrolysis. In these cases molybdenum would be coprecipitated with hydrous ferric oxide. In the case of the large leaching plant at Chuquicamata, Chile, where the leachate is reported to contain considerable molybdenum, it would be reduced to a lower oxidation state during removal of chlorine by means of cement copper. This may explain why the embrittling effect of molybdenum has not been reported previously. Moreover, in copper refining plants where soluble anodes have been prepared by a high temperature smelting process, it is most unlikely that any molybdenum would remain with the copper, whereas the calcines in this work were prepared in a fluidized bed reactor at a temperature not exceeding 720°C. Acknowledgment This work is part of an investigation on the roasting and extraction of copper from sulfide ores sponsored jointly by Mt. Morgan Ltd., Queensland, and the Mt. Lyell Mining and Railway Co. Ltd., Tasmania.

Jan 1, 1958

By E. S. Leaver

THE presence of soluble base metals in precious-metal ores usually precludes cyanidation as the best method of treatment. The laboratory experiments described in this paper show the possibility of cyanidation of gold-silver ores containing less than 0.5 per cent. of cyanide-soluble copper and the effect of zinc in the cyanide solution. PART 1.-COPPER IN CYANIDATION The proposed process of cyanidation of precious-metal ores containing copper is based on the regeneration of about 80 per cent. of the cyanide used in the dissolution of the copper. The method of regeneration involves a combination of the sulfide and the acid precipitation of the copper and silver. The resultant solution is made alkaline with lime, and the gold is precipitated with zinc dust.

Jan 1, 1929

El Indio, eighty percent -owned by Compania Minera an Jose Inc., a subsidiary of St. Joe Minerals, and twenty percent-owned by private Chilean investors, will open their new flotation and Carbon -In -Pulp plant near La Serena, 480 km (300 miles) north of Santiago, Chile, in early 1981. The underground mine, at an altitude of 4,270 m (14,000 ft), will produce about 1,270 mt per day (1,400 st per day) of ore containing about 2. 2570 Cu, and gold and silver to the extent of 16 and 254 g per mt (0.47 and 7.4 oz per st) respectively. While construction of the plant was underway, development ore, containing some 1,400 kg (45,000 oz) of gold, was shipped to the United States for recovery in a copper smelter. At El Indio the underground ore is rail-hauled to the mill a short distance from the portal where it is crushed, ground, and floated to produce a copper concentrate containing both gold and silver. The crushing plant also crushes waste for use as mine backfill. Approximately 92% of the copper and 60% of the gold and silver are recovered in the flotation concentrate. Flotation tailings, at a grade of 19 g Au and 116 g Ag per mt (3.4 oz Au and 0.55 oz Ag per st), are cyanided for 72 hours in six Pachucas with auxiliary air lifts for pulp transfer. Lime, cyanide, lead nitrate, and a phosphate are also fed to the cyanidation Pachucas. After cyanidation, the pulp is treated in four air lift Carbon-In-Pulp tanks equipped with external vibrating screens for carbon advance. Loaded carbon is stripped in two pressure vessels in parallel. Pressure stripping must be used because of the high altitude. Stripped carbon is dewatered in a spiral classifier before being reactivated in a rotary kiln and returned to the C. I. P. circuit. Pregnant solution is then sent to four banks of four Zadra cells for gold and silver recovery on steel wool. Zadra cell cathodes.are washed with the slimes being captured in a plate and frame filter. The filter cake and the washed cathodes are then dried on an infrared dryer and smelted in two stages to bullion bars. Published reports indicate that the total investment for this unique operation was in the neighborhood of $80 million.

Jan 1, 1981

By J. Wright, R. J. Tyler

Changes in the pore structure of hematite pellets during reduction to iron were investigated under a variety of experimental conditions. Pore structures were determined by a mercury penetration technique. The reduction temperature was found to be the major influence on structure, the pore system being bimodal below 7.50°C and unimodal above 900°C. Pore sizes ranging front 0.04 to 20 m were observed in complete metabolized pellets produced over the temperature range 650 - 1050°C. Gas composition and reduction rates also influenced the pore structure of the metabolized pellets, although to a lesser extent than the reduction temperature. An apparent activation energy of 24 kcal/mole was derived for the pore formation process.

Jan 1, 1979

By J. D. McBeth, G. M. Potter, Ramon Pizarro

Heap leaching is used by the Carlin Gold Mining Co., Carlin, Nev., to supplement conventional processing of gold ore by agitation-leaching and countercurrent decantation. Selected open-pit ore containing about 0.072 oz Au per ton is crushed to 3/4 inch, dumped by truck on an asphalt pad, and leached by sprinkling with dilute cyanide solution for 7 days. The heap is then drained and removed by front-end loader to tailings disposal. About 70 percent of the gold is dissolved. Leach solution from the heap joins the pregnant solution from countercurrent decantation for gold recovery by zinc precipitation. Four leach pads, each with a capacity of 1,400 tons of ore, are in use. About 10,000 tons of ore are heap leached monthly from about April through October. Heap leaching is recessed during freezing weather. Labor, equipment, and reagent requirements for heap leaching are presented.

Jan 1, 1974

By John Cogan

Field observations were made of rock phenomena in shale, dolomite, and fractured quartzite in the Burgin Mine, Eureka, Utah. Draft closure rates were measured with extensometers and convergence gages. In addition the effects of hydraulic props upon the retardation of closure was also monitored. Drift closure rates were found to be constant and to vary between 0.32 x 10-8 min-1 for dolomite and 23.8 x 10-8 min-1 for soft shale. The installation of props set at constant load served to retard the closure rates. General results show closure to occur at a constant rate for all cases, to be non-uniform around the drift openings, and to be increasingly retarded toward a zero rate by increases in prop load.

Jan 1, 1979

By R. B. Schluter, M. M. Fine

A process for winning copper from pelletized chalcocite concentrate without matte smelting and converting is under development at the Twin Cities Metallurgy Research Center of the U.S. Bureau of Mines. The new procedure includes a two-stage fluid-bed treatment in which the pelletized concentrate with 33% Cu and 7% S is first desulfurized to about 0.5% S in an oxidizing atmosphere, and then metallized by reduction with hydrogen. The reduced product is melted with flux to separate metallic copper from the gangue. The metal must be refined to remove residual sulfur and iron. The pelletizing of the fine concentrate into mini-pellets prior to fluid-bed treatment is an important feature of the process, because it permits shorter reaction time, decreases dusting, and facilitates subsequent processing.

Jan 1, 1972

By B. M. Blake, R. S. Reppert, A. F. Keiser, E. J. Trinkle

The southern West Virginia coalfield was formed in a rapidly subsiding depositional basin associated with deep-seated growth faults. Subsidence began to the southeast during deposition of the Pocahontas Formation, and then shifted to the northwest during deposition of the New River, Kanawha, and Allegheny Formations. Rocks of all the coal-bearing formations grade into massive sandstones to the northwest. For the No. 3 Pocahontas, Sewell, Campbells Creek, and No. 5 Block coals discussed in this study, the overall sulfur and ash values of each coal seam increase gradually toward the northwest. The thickness of each coal tends to increase toward the southeast. These trends are partly a reflection of the differential subsidence of the coal basin along with varying degrees of peat accumulation and degradation. The northwest side of the coal swamps underwent relatively less subsidence and, thus, more degradation due to increased aerial exposure, with a resulting increase in the sulfur and ash values of the coals, and a decrease in the thickness. The open sea existed to the west of the coal swamps, and this marine environment also contributed to the higher sulfur values to the northwest. Coals of the Pottsville Group in the southern West Virginia coalfield generally have sulfur values between 0.6-1.0% with some as high as 2-3% on the northwest edge of the coal basin.

Jan 1, 1983

By Alan Morris

In a previous paper1 the results of cutting tests on free-cutting brass rod were reported. Investigation was made of the effects of variation in lead content, microstructure and cold drawing. The author&apos;s reply to the discussion of that paper includes a study of the directional cutting properties of drawn and annealed rod. The present paper records the results of tests made on series of samples in which the tin, iron and copper contents have been varied. The apparatus used and method of testing are described in detail in the previous paper. The machine is of the pendulum-operated milling type used and described by Airey and Oxford.2 Angular readings of the height to which the pendulum rises after falling from a fixed position, both when a cut is made and when the pendulum swings freely, give the basis for calculation of the energy consumed in making a cut. Careful weighing of the sample before and after cutting gives a measure of the amount of metal removed. The results are expressed in footpoiinds per cubic inch of metal removed, and have been called "unit cutting energy." In the course of the discussion of the author&apos;s earlier paper, W. B. Price reported large-scale tests involving some 50,000 lb. of material. These tests indicated that the addition of as much as 0.6 per cent of iron and 0.25 per cent of tin had apparently no adverse effect on the highspeed cutting characteristics of the rod. The rod also contained a small amount of nickel. The test results reported here tend to confirm his observations. Effect of Variation in Tin Content The castings for this series were made 2 by 1 by 9 in. and were processed as follows: 1. Castings were annealed at 500" C. for 7 hr. 2. Cold-rolled to 0.700 in.

Jan 1, 1933

By N. M. Langdon

Spokane Meeting, September, 1909.) THE following corrections to the paper of Mr. Langdon (Bulletin No. 34, October, 1909, pp. 919 to 940), received after the publication of the paper in the Bulletin, will be incorporated in the paper as finally published in the forthcoming Vol. XL. of the Transactions, but are here presented for the information of members:&apos; Page. 921. Table I., furnace M2, item 11. For " 8801 " read "0.8801." 922. Table II., furnace C, item 2, under gas. For " 0.011 C " read " 0.011 0." Furnace G, item 13, under gas. For " 5.074 air " read " 3.074 air." Item 15, 114.269 " should be " 4.289." Furnace H, item 2, under iron. For " 0.04 " read ".0.01." 923. Table II., furnace J, item 1, under iron. For " 0.95 " read " 0.93." Furnace N2, item 6, under gas. For " 0.66 CO," read 0.066 C02." 924. Table III., furnace C, item 17. For " 379 " read " 397." 927. Table IV., furnace C, item 4. For " 0.0124 " read 0.0424." Furnace G, item 15.. For " 4,060 " read " 4,066." 928. Line 2. Items 19 and 20 of heat-supply, should be items 20 and 21. 930. Line 3. Expunge remainder of paragraph after the word " supplied," and substitute the following : "As compared with the fusion-zone, the loss by radiation n C was 25.5 per cent. of the total (2,278 units) supplied to that zone, while in D it was 22.5 per cent. of the total heat (1,824 units) supplied to that zone. Finally, as compared with the reduction-zone, C lost by radiation 36.8. per cent. of the 1,580 units supplied to that zone,

May 1, 1910

By R. C. Ruder, C. E. Birchenall

The method of decrease in surface activity was used to determine the rates of diffusion of Co" into cobalt and into nickel. Since the absorption curves for cobalt radiation were quite complex under the conditions of these experiments, the effects of geometry and surface preparation on the diffusion results were investigated. The cobalt self-diffusion coefficients obey the equation DCoco = e-61100/RT cm2 per sec. For dilute cobalt diffusing in nickel, DConi, == 1.46 e-/R&apos; cm2 er sec. AS part of the general program to measure the self-diffusion and dilute-solution diffusion rates of the iron-group metals, the diffusion coefficients of cobalt into cobalt and nickel have been measured using the radioactive isotope Co"&apos;. Self-diffusion studies in iron have been previously reported by Birchenall and Mehl.&apos; The technique of measurement used was the decrease in surface activity method described by Steigman, Shockley, and Nix.&apos; It has been determined that scattering of the measured beta radiation by the specimen matrix is significantly influenced by the condition of the surface and the counting geometry. Accurate diffusion coefficients can be calculated from the activity data provided that the specimens have plane diffusion surfaces and that the absorption properties of the emitted radiation are determined under geometrical conditions duplicating with external absorbers the absorption and scattering in a thin film at the specimen surface as nearly as possible. The cobalt used in this research was originally in the form of hydrogen reduced rondels 99.9+ pct pure. The nickel had been prepared by the carbonyl process and was found by spectrographic analysis to contain 0.03 pct Fe, 0.04 pct Co and smaller traces of other metallic elements for a balance purity of 99.9+ pet Ni. Both materials were vacuum melted and cast in rectangular ingots. Specimens 7/8x7/8x1/4 in. were machined from these ingots. The specimens were then ground to 4-0 emery paper and annealed in hydrogen for at least a day at 1200°C. Some of the specimens were used in preliminary experimentation after which their diffusion faces were machined to remove the radioactive material from the previous diffusion experiments. After grinding, the specimens were polished using an electrolytic lapping technique described by Mazia." This lapping was continued until at least 0.005 in. was removed from the diffusion surface. The final surfaces were metallographically smooth except for a few fine scratches caused by the glass cloth. A large percentage of the area was undisturbed as determined by microscopic examination. The grain diameters of the diffusion specimens varied from 0.1 to 5 mm. There were no effects due to grain boundary diffusion in the experiments

Jan 1, 1952

The Ortiz Mountain property is located approximately 40 km (25 miles) southeast of Santa Fe and 13 km (8 miles) from the village of Cerrillos. The property embodies a Spanish mineral grant from which Gold Fields acquired about 2, 200 hectares (900 acres of surface and mineral rights. Following a period of exploration, metallurgical testing and development of a feasibility study in 1974, the decision was made, because of the higher gold price in the fall of 1978, to place the property into production. the gold is confined in a quartzite breccia; it is an unusual occurrence because the gold in this host rock is leachable at minus 19 mm (3/4 in. ). Total ore reserves are estimated at approximately 3.6 million mt (4 million st) averaging 1.89 g per mt (0.055 oz per st) based on a cut-off of 0.7 g (0.02 oz). The ore will be recovered by open pit mining at a stripping ratio of 1. 2:l. Production is scheduled for 13,600 mt (15,000 st) of ore per week or approximately 680,000 mt (750,000 st) per year. Primary crushing of pit-run material is by a 1,070 by 1,015 mm (42 in. by 40 in.) jaw crusher. Crushed ore is conveyed to stockpile and then to a 2.13 m (7 ft) standard crusher with the screened over- size crushed in a 1.5 m (5 ft) tertiary cone crusher. The screen undersize, the final product for leaching, is minus 19 mm (3/4 in. ). A 1, 360 kg (3,000 lb) sample is cut from the crushed ore stream, further crushed in a small jaw crusher and then sampled in two more stages by Vezin samplers to produce a final 13.6 kg (30 lb) sample for assaying. Pebble lime is then added to the ore on the belt as it is sent to the conveyor gantry spanning the leach pads. A tripper on this moving gantry stacks the leach ore without compaction. The entire crushing plant and conveying system is of open-air construction. Leach pads have a total length of 515 m by 60 m (1,700 ft by 200 ft) wide with usable dimensions of 490 m by 58 m (1,600 ft by 190 ft). The directional length slope is 3. 870 with 370 on the width. The pad is constructed of 175 mm (7 in. ) thick asphalt containing a rubberized asphalt membrane. It is divided by a 300 mm (1 ft) high curb into eight segments. Each segment is loaded with 13,600 mt (15,000 st) of ore, which is one week&apos;s production. Optimum working height of a pad is considered to be 3. 0 m (10 ft) with a maximum height of 4. 0 m (14 ft).

Jan 1, 1981

By Robert Franke

(Butte Meeting, August, 1915.) IN view of the prominence which the conical mill has attained in the fine-crushing field within the few years since its introduction, the following comparison with its more mature forerunner, the Chilean mill, based on extensive tests, is submitted in the interest of the milling profession. Soon after designing the concentrating plant of the Miami Copper Co. in 1909, the Hardinge conical mill made its appearance in the milling forum. Its possibility as a suitable crushing device for the plant was well recognized, but in view of the lack of commercial demonstration, at the time, as to capacity, efficiency for desired product, and the still more uncertain factors of cost of maintenance and power consumption, it was deemed that the immediate adoption of this machine throughout the plant would be a hazardous undertaking. For these reasons it was decided to equip the majority of the imme diately required units of the plant with Chilean mills, the fine-crushing proficiencies of which were better known, and one section with Hardinge mills, to serve as a test unit for the guidance of future installations and replacements. Thereby, after 1.5 years&apos; operation with both types of mills, a thorough test as to metallurgical efficiency and cost economy has been obtained. The conical mill used in these tests is the 8-ft. Hardinge pebble mill, having a cylinder 22 in. in length. The cylindrical portion of this mill is lined with cast-iron liner plates, and the conical extensions with silex bricks bound together by cement. Each liner plate carries a projecting lifter, the function of which is to increase the height of drop of the lifted material. Danish No. 5 pebbles, obtained from the coast deposits of Denmark, are used for the grinding charge. The Chilean mill used is a fast-running, 3-roller, 6-ft. Saturn mill, with screens of 0.037-in. opening. The feed to these mills is the oversize of Callow screens having 0.029-in. openings, which follow rolls crushing to 0.5 in.

Jan 7, 1913

The following is an incomplete list of members and guests who called at Institute headquarters during. the period Mar. 10, 1919,. to Apr. 10, 1919. F. T. Agthe, Hannibal, Mo. Jay Lonergan,. Denver, Colo. W. F. Almy, Canada. R. R. London, Cebu, P. I. P. G. Bandy, Mexico. W. D. Mainwaring, Philadelphia, Pa. H. R. Bank, Humboldt, Ariz. _ W. G. Mitchell, Montreal, Can. Alan M. Bateman, New Haven, Conn. Francis Nicholson, Joplin, Mo. C. A. Brooks, Lead, S. Dak: C. H. Palmer, Los Angeles, Calif. Louis J. Brunel, Oakland, Calif. D. A. Powell, Columbus, Ohio. W. Burns, Sante Fe, N. M. W. H. Reber, Dallas, Tex. J. B. Carlock, Major, 1st Gas Regt. Capt. J. B. Richards, Spokane, Wash. Louis S. Cates, Hayden, Ariz. T. C. Roberts, E. Orange, N. J. W. M. Corse, Mansfield, Ohio Max Roesler, Washington, D. C. E. 0. Daue, Philadelphia, Pa. C. 0. Sanford, Los-Angeles, Calif. John Davenport, W. Norfolk, Va. Gustave Sessinghaus, Denver, Colo. James T. Dixon, London, England, . E. H. Shriver, Milford Center, Ohio. S. K. Dahl, Messina, Transvaal, S. Af. H. M. Spoor. W. G. Eberhardt, Tucson, Ariz. W. H. Staver, New York. H. S. Emlaw, Grant Haven, Mich. D. B. Sterrett, Washington, D. C. I. W. Farrand, Camp Merritt. W. G. Swart, Duluth, Minn. Oscar Gotsch, Jr., St. Louis, Mo. A. W. Stickney, London, England. James 0. Greenan, Oakland, Calif. E. C. Templeton, Downey, Calif. Gilbert Hart, Philadelphia, Pa. . . R. T. Walker, Pioche, Nev. K. L. Hsueh, China. A. T. Ward, Bellefonte, Pa. Zay Jeffries, Cleveland, Ohio. Joel H. Watkins, Washington, D. C. R. W. Kern, Verde, Ariz. J. Wertheimer, Cleveland, Ohio. H. D..Kinney, Easton, Pa. C. R. Wilfley, Ouray, Colo. E. L. Lasier, Washington, D. C. J. P. Wood, Engrs., U. S. A. Carl Zapffe, Brainerd, Minn. Leland D. Adams has left the employ of the Weedon Mining Co., Ltd., at Montreal, Quebec, and is going to Oakland, Calif., where his address will be 1180 Mandana Boulevard.

Jan 5, 1919

By H. J. French

WITHIN the past few years, particularly because of the rapid growth of the airplane industry during the war, considerable attention has been paid to the manufacture of light-wall cold-drawn seamless and welded steel tubing for structural purposes. While the manufacture of such tubing is not, in itself, a difficult matter, some difficulties have. been encountered in producing the same under specifications requiring definite physical properties. It is the purpose of this paper to describe the methods of manufacture and the properties that may be obtained by such methods for tubing coming within about ½ to 2 in. (12.7 to 50.8 mm.) outside diameter and 10 to 22 Birmingham wire gage (0.134 to 0.028 in. wall thickness). Such tubing is largely used as structural material in the aircraft industry, and it is of the utmost importance that the physical and chemical properties should be known, and also that the various factors entering into the manufacture should be under such control as to allow these properties to be varied and held within definite limits. In general, tubing may be divided into four classes: namely, specialty tubing, such as riveted-sheet or plate, lock jointed, etc.; brazed tubing; welded tubing; and seamless tubing. Welded tubing may be divided into hot product, such as lap welded and butt, welded, and cold product, such as acetylene and electric welded. Seamless tubing may be hot rolled or cold drawn.

Jan 9, 1919

By W. A. Foster

NINETY years after the Riddle nickel deposit N was discovered in Oregon in 1864, Hanna Coal & Ore Corp. began mining operations. Until 1954 much prospecting and preliminary development work had been done on the property, but no ore had been shipped other than for metallurgical testing. The deposit is being explored by churn drilling and trenching, and a few test shafts have been put down for the purpose of studying the ore in place, determining volume and moisture factors, and gathering bulk samples for pilot plant and smelter tests. In churn drilling, 6-in. casing is driven down 2 or 3 ft behind the drill bit and bottomed at each 5-ft interval. All the sludge for each cased 5-ft interval is dumped through a splitter and one eighth is saved and analyzed. Drillholes are stopped after going through the ore zone and penetrating fresh peridotite in which the nickel content is less than 0.5 pct.

Jan 8, 1957

By G. L. Oldright, F. W. Schroeder

Very great changes were made in the dimensions of the smelting hearths of the furnaces in the period from about 1800 to 1906, the length increasing from about 11 to 116 ft., and the width from 8 to 19 ft. During this time the tonnage smelted increased from about 9 to 250 tons per 24 hr., and the tons of charge smelted per ton of coal increased from 0.77 to 4.3. Dr. Mathewson3 noted no great difference in the "fuel ratio" (i. e., the amount of charge smelted per ton of coal) as the furnace at Anaconda was increased in length by stages from 85 to 116 ft. (The percentage of copper in the slag, however, was lowered from 0.42 to 0.36 per cent. The other components of the slag were not given.) Lump coal fired on grates was the fuel used during this period. The practice in operating the furnaces over such a length of time naturally varied considerably, and there were differences in composition and nature of the charge, yet it may he considered that the improvements made were due largely to increases in the size of the furnace used. Doubtless some credit should be given to improvements in operating technique, like changes in the methods of firing the lump coal. However, the greater part of the advance was probably due to factors influencing the

Jan 1, 1928

By H. E. Wilborg

This chapter deals with those sands employed by the foundries for the manufacture of cores and molds used in the casting of such common metals as steel, gray iron, ductile iron, aluminum-based alloys, and copper-based alloys. Most metal is cast in "green sand" molds in which the sand mixture, consisting primarily of silica, clay, and water, forms a plastic material. This plastic material is rammed around a pattern in a split box called a flask, then the box is separated, the pattern is withdrawn, and the metal poured. These green sand molds are broken apart when the metal has solidified. The sand is then remixed and a new mold made for the next casting. To a lesser extent silica sand is mixed with resins or oils and, after forming, the mixture is hardened either by heat or chemical reaction. In any case, the foundry mold or core is used just once before being broken up and remixed for manufacturing another core or mold. By far the greatest consumption in the foundry is of those types of sand high in silica, low in clay (under 0.5%) and occurring in numerous locations thoughout the United States.

Jan 1, 1975

By E. H. Rennhack

Phase equilibria of the zinc-rich portion of the Zn-Fe-A1 system containing up to 20.0 wt pct Fe and Al have been investigated at 600°C (1112°F), 450°C (842°F), and room temperature by metallographic and X-ray diffraction techniques. The phase relationships were found to involve only those phases present in the three binary systems. THE presence of small quantities of aluminum (approx. 0.2 wt pct*) in zinc galvanizing baths is known to suppress the formation of Zn-Fe alloy layers within the coating.&apos; As part of an extended investigation of this suppression effect and the alloy reactions involved, the zinc-rich corner of the Zn-Fe-A1 equilibrium system containing up to 20.0 pct Fe and Al was determined at 600°C (1112°F), 450° F (842° F), and room temperature. The present paper reports the findings of this study. The constitution of the Zn-Fe system has been reported by Raynor2 based mainly on the work of Schramm3-3 and Truesdale et a1.6 The solid solubility of iron in zinc is extremely small and lies between 0.0009 and 0.0028 pct in the range 150° to 400°C (302°to752°F). Within the temperature and composition range of interest, a eutectic reaction involving nearly pure zinc, liquid = Zn+FeZn13, is observed at about 419°C (786° F) and 0.018 pct Fe along with a peritectic reaction, FeZn7+ liquid e FeZn13, at 530°C (986° F) and 6.2 pct Fe. The latter reaction is sluggish due to the large proportipn of solid phase involved in the transformation. Halla et al.,7 identified the compound FeZn13 as monoclinic with a = 13.68, b = 7.63, c = 5.07, and B = 128°44&apos;. The structure of FeZn7 was reported by Bablik, et al.,8 to be of hexagonal symmetry, belonging to either the C6v = C6/m mc or D46h = C6/mc space group. Divensions of the unit cell are a = 12.81A and c = 57.6a. The Zn-A1 diagram as composed by Raynor9 and later Gebhardt10 is characterized by a eutectic decomposition, liquid 382°C (720°F) and 95.0 pct Zn and a eutectoid decomposition, at 275° (527°F) and 78.0 pct Zn. The eutectoid reaction cannot be suppressed permanently by quenching. Above 275° (527° F), a and a&apos; co-exist in a closed miscibility gap between 31.6 and 78.0 pct Zn. Both phases are fcc and have nearly the same lattice parameter. The maximum solubility of aluminum in zinc (B) is approximately 1.0 pct at the eutectic temperature. PROCEDURE Materials. The experimental alloys were prepared from both massive and powdered material. Cast alloys were composed from three zinc-base master alloys containing 8.0 pct Fe, 16.0 pct Fe, and 50.0 pct Al which were made from C.P. zinc, Armco iron, and

Jan 1, 1962

By Harold Roast

THE object of this investigation was to compare the arsenical antimony-lead alloy with some of the regular bearing-metal alloys. With this end in view, the following tests were made: 1. Chemical analyses to show the composition. 2. Thermal analyses. 3. Macroscopic examination of fractures. 4. Microscopic examination. 5. Tensile tests. 6. Crushing tests. 7. Scleroscope tests. 8. Brinell tests, at temperatures from 80° to 400° F. Tests 3 to 8 were conducted on both sand-cast and chill-cast specimens. For the purposes of this paper the alloys have been numbered from 1 to 6; their composition is as follows: LEAD, ANTIMONY, TIN, COPPER, ARSENIC, NUMBER PER CENT. PER CENT. PER CENT. PER CENT. PER CENT. 1 82.64 12.95 4.40 0.01 none 2 27.54 9.94 59.87 2.65 none 3 77.65 20.75 none 0.13 1.47 4 82.95 16.85 none 0.20 none 5 83.78 15.35 none 0.05 0.82 6 0.15 7.87 84.13 7.85 none The analysis, in the case of the antimony, was made by solution in sulfuric acid and titrated with permanganate. For tin, the ferric-chloride method was used; for copper, the alkaline separations, and the thiocyanate methods; for arsenic, volatilization of arsenic and titration with iodine. The lead, with the exception of alloy 6, was taken by difference.

Jan 2, 1922