Search Documents

By Michael R. Cartwright

The Gold Acres outcrop was located by Lee Lakin in 1922, according to Mr. H. W. Treweek. There was no activity, however, until 1935, when prospect drifts and pits penetrated the deposit. It was then drilled by the Willow Creek Mining Company. Beginning in the late 1930s and continuing to the late 1940s more than 150 churn drill holes were drilled in the Gold Acres area. From this drilling came what is now known as the London Extension pit. Only oxidized ore amenable to cyanide vat leaching was mined during this time. Our records indicated that this period of mining removed some 1.4 Mt (1.5 million st) of ore at an average grade of 0.006 kg/ t (0.2 oz per st). Union Pacific and Newmont Mining Company worked in the area prior to acquisition by the Cortez Joint Venture. Union Pacific drilled a number of core holes in the area. The unoxidized nature of the ore, small tonnage, and low grade undoubtedly prevented mining at this time. From these indications of mineralization shown by previous companies, the Cortez Joint Venture developed and mined 2 Mt (2.4 million st) at 0.005 kg/t (0.15 oz per st) in the Gold Acres pit. This pit is located immediately southwest of the older London Extension pit. Cortez Gold Mines has continued to drill to the west and south of these pits along the downdip continuation of the former ore bodies. The downdip extension is unoxidized and more metallurgically difficult than the normal Carlin type ores, but, more than 2.4 Mt (2.7 million st) have been outlined. It depends now upon the gold price and milling costs to turn this mineralized area into an ore body. A possible lesson to be learned from this deposit is that continued mining success depends upon an ongoing adequately funded exploration-development-metallurgical research program. These three disciplines must always be present to locate the mineralization and to turn this mineralization into ore. [ ]

Jan 1, 1985

By E. H. Rose

The question is often pondered whether the flotation process is still an art or has become a science. The fact is that flotation is a science in so many variables that only art can blend them. It is demonstrable that no less than one hundred variables are involved in any flotation operation, most of them immeasurable on a routine basis. The behavior of every one of them is always governed by orderly and unchanging law, as no man can doubt. There is nothing truly indefinite about that behavior; the only indefiniteness is in our knowledge of it, either in general or from hour to hour in the circuit. In dealing with a simultaneous equation in, say, one hundred variables, all not only affecting the outcome to some degree but many influencing each other as well, it is small wonder that the mineral-dressing profession presents a case history of debate and controversial opinion. The statement of the subject of this symposium—a symposium on flotation machines—seems to delimit and define the matters to be considered and at least the general nature of the conclusions sought. Nevertheless, in such a meeting it is not possible to meditate upon all the variables that do in one way or another affect performance of flotation machines. To attempt it would be much like trying to operate a hundred-ring circus. Perverse variables being what they are, one could expect the elephant in ring 36 to squirt a trunkful of water at the tightrope performer in ring 37 to ruin that artist's act, just as our learning to depress pyrite in the early 1920's ruined the act of the formerly faithful old Callows. Honest Disagreements Selective flotation dates from 1922 or 1923, when several plants established their regular flotation operations on a selective basis. Since then, two decades of efforr have gone into the development of better and better flotation machines. Many have come and gone. The competition has simmered down to perhaps a dozen machines, each one of which is staunchly defended by some operators as being the best for their particular jobs, and as scornfully dismissed by at least a few others elsewhere as being positively the poorest for theirs. These disagreements are honest disagreements, honestly arrived at after plenty of sweat. In many a flotation plant, one machine has been tried out against another under conditions as fair as they could possibly be made. Sometimes the results are on a par but more often one machine shows a consistent and unmistakable margin of superiority over the other. In equally honest and enlightened tests, machine A may win over machine B in one place, and machine B over machine A by as great a margin in another. This paper purposes to identify some of the more prominent reasons why this is so,

Jan 1, 1947

By E. H. Rose

The question is often pondered whether the flotation process is still an art or has become a science. The fact is that flotation is a science in so many variables that only art can blend them. It is demonstrable that no less than one hundred variables are involved in any flotation operation, most of them immeasurable on a routine basis. The behavior of every one of them is always governed by orderly and unchanging law, as no man can doubt. There is nothing truly indefinite about that behavior; the only indefiniteness is in our knowledge of it, either in general or from hour to hour in the circuit. In dealing with a simultaneous equation in, say, one hundred variables, all not only affecting the outcome to some degree but many influencing each other as well, it is small wonder that the mineral-dressing profession presents a case history of debate and controversial opinion. The statement of the subject of this symposium—a symposium on flotation machines—seems to delimit and define the matters to be considered and at least the general nature of the conclusions sought. Nevertheless, in such a meeting it is not possible to meditate upon all the variables that do in one way or another affect performance of flotation machines. To attempt it would be much like trying to operate a hundred-ring circus. Perverse variables being what they are, one could expect the elephant in ring 36 to squirt a trunkful of water at the tightrope performer in ring 37 to ruin that artist's act, just as our learning to depress pyrite in the early 1920's ruined the act of the formerly faithful old Callows. Honest Disagreements Selective flotation dates from 1922 or 1923, when several plants established their regular flotation operations on a selective basis. Since then, two decades of efforr have gone into the development of better and better flotation machines. Many have come and gone. The competition has simmered down to perhaps a dozen machines, each one of which is staunchly defended by some operators as being the best for their particular jobs, and as scornfully dismissed by at least a few others elsewhere as being positively the poorest for theirs. These disagreements are honest disagreements, honestly arrived at after plenty of sweat. In many a flotation plant, one machine has been tried out against another under conditions as fair as they could possibly be made. Sometimes the results are on a par but more often one machine shows a consistent and unmistakable margin of superiority over the other. In equally honest and enlightened tests, machine A may win over machine B in one place, and machine B over machine A by as great a margin in another. This paper purposes to identify some of the more prominent reasons why this is so,

Jan 1, 1947

By James J. Scott

INTRODUCTION This paper presents the basic principles which must be followed to create a truly inherently safe mining system. The need to relate the support mechanism to be employed to the ground reaction curve and basic rock mechanics principles will be pointed out. The support system must function in concert with the deformations that are inherent in the selected mining system. Following the support selection, the mechanics for installing support must be developed. This is best done through a cooperative effort between mine engineers, mine operators, safety personnel, equipment sup- pliers, and government inspectors both state and federal. Accident statistics, near miss statistics, review of equipment used by others, geologic features, etc. will necessarily have to be studied. The goal of equipment development and selection should be to develop inherent safety by removing the miner from the hazard area or by providing absolute protection for the miner if he must work in close proximity to the potential fall of ground. The role of good productivity in reducing the numbers of men exposed to hazards is emphasized as a good safety measure ROCK MECHANICS PRINCIPLES A thorough understanding of the ground reaction curve surrounding openings to be created is essential in the

Jan 1, 1982

THE opening session was held on Tuesday evening, February 17th, in the house of the American Society of Civil Engineers. The President of the Institute, Mr. E. B. Coxe, after a few introductory remarks, congratulating the members on the general prosperity in all the branches of the professions and manufactures represented by the Institute, called on Mr. E. F. Loiseau, of Philadelphia, for the first paper, On the Successful Manufacture of Compressed Fuel at Port Richmond, Philadelphia. Professor Prime called attention to the fact that the work of the United States Testing Board had been left unfinished at a point when valuable results had just been reached, and offered the following resolutions: WHEREAS, The American Institute of Mining Engineers has seen with regret that the work of the United States Testing Board has been suspended through the lack of an appropriation by Congress, which appropriation dues not seem likely to be renewed, and WHEREAS, It seems most important to the vital interests of our iron and steel manufactures that the experiments commenced by the United States Testing Board should be carried to successful completion ; Resolved, That a committee of three be appointed, to act with a like, committee to he appointed by the American Society of Civil Engineers, to carry out to completion the work commenced by the United States Testing Board. Resolved, That the American Society of Civil Engineers be requested to cooperate with this society by the appointment of a like committee. Resolved, That the committee have authority to request, in the name of the American Institute of Mining Engineers, contributions to a fund to be devoted to the completion of the experiments aforesaid. Resolved, That the Council of this Institute be authorized, whenever requested by the committee so appointed, to petition Congress for the free use of the testing machine at the United States Arsenal at Watertown. Resolved, That a committee be appointed by the President, to select the afore-said committee of experts. Resolved, That the President of the Institute have authority to fill any vacancies which may occur in the committee. The resolutions were unanimously adopted, and President Coxe appointed the following committee in accordance with the resolutions : Messrs. A. L. Holley, James Park, Jr., and J. F. Holloway. The President also appointed Messrs. Heinrich, Mickley, and Spils-

Jan 1, 1880

By H. Eklund, R. T. Hukki

This article presents a simple arithmetic derivation for the relationship between sharpness of classification and circulating load, as well as world-wide industrial data which are compatible with the derived equation. It is shown that the sharpness of classification decreases with increasing circulating load. Based on the evidence available a comparison is presented between the performance of the present-day hydraulic classifiers. In a continuously operated closed circuit grinding system the tonnage and the fineness of the components may be indicated as follows: Eq. 8 expresses the relationship between sharpness of classification and circulating load in closed grinding circuits. If CL = 0, E = 100% according to the definition used. This, however, is a purely theoretical case because any apparatus to be called a classifier must produce the conventional two products. Furthermore, if CL = m, E =0%. In order to visualize the meaning of Eq. 8, the effect of factors d (fineness or percentage of under-size of classifier overflow at size x) and b (fineness or percentage of undersize of classifier sands at size x) should be studied. By the most rigid standard, factor d should be given the value of 100%. By so doing the size of classification is defined as that screen size x through which 100% of the classifier

Jan 1, 1965

By K. L. Komarek, P. F. J. Landler

Synthetic polycrystals of wustite were reduced ilz H2-H2O mixtures over limited roncentration ranges wiltun the wustite. field and the weight changes were continuously followed with a McBain quartz spiral balance. The reduction was. found to be controlled by a combined surface plus solid-state diffusion reaction. The solid-state diffusion had an activation energy of 38 kcal. Self--diflusioities of iron decreased with itzcreasing vacancy concentration. The activation energy of. the surface reaction runs different tor- different samples. The surface reaction-rate constan increased with increasirzg vacancy concentration. THE reduction of the oxides of iron has been studied for many years. Edstroem' has summarized the early work and has provided much information on the reduction of magnetite and hematite by hydrogen and carbon monoxide. More recent summaries are those by smeltzer2 and Themelis and Gauvin.3 Most investigators have reported that the iron metal which forms the outer layer during gaseous reduction is rather porous. According to Kohl and Engel14 the formation of such a porous iron layer depends on the properties of the underlying wiistite layer, on the reaction of the gas with the wustite surface, and on the transport of matter through the wustite layer. McKewan, Quets, Wadsworth, and Lewis, and Knacke 10 have found that, if a porous iron layer is formed, a surface reaction between gas and wustite at the bottom of the pores controls the rate of reaction. Seth and ROSS 11 and Warner 12 have pointed out that the influence of the gaseous pore diffusion through the porous iron layer cannot be neglected. It has further been shown that the wiistite layer formed during reduction is dense and that transport of matter across the layer occurs only by solid-state diffusion of iron ions.13,14 The solid-state diffusion rates observed by Edstroem and Bitsianis 14 and the surface reaction rates observed by Quets eta1.' show that an interaction of the two steps in controlling the reaction rate has to be considered. In this investigation the reduction of wustite was studied with respect to the interaction of the two steps, isolated from any other effects, by reducing wustite within the wustite field from higher to lower oxygen concentrations with H2-H2O gas mixtures. A similar study with CO-CO2 mixtures has been performed by Levin and wagner.15 Fused synthetic wustite polycrystals were reduced in steps of 0.4 wt pct from 24.4 to 23.2 wt pct 0 and over a temperature range from 700" to 1050°C. A few tests were also conducted with reduction steps from 24.4 to 23.6 wt pct 0 and 24.5 to 23.2 wtpct0. The Fe-O phase diagram is shown in Fig. 1 where the composition of the gas phase in equilibrium with the condensed phases is plotted as pct H2 vs the temperature. A careful study of the literature led to the conclusion that the fully drawn equicomposition curves (in wt pct O) in the wustite field used by Edstroem 1,16 and other authors needed some correction. Based on results by Darken and Gurry17 (shown as open circles) and on some of our own data, four corrected curves for 23.2, 23.6, 24.0, and 24.4 wt pct O were calculated (drawn as dashed lines in Fig. 1) which should replace the fully drawn equicomposition curves. MATERIALS AND APPARATUS Materials and Specimen Preparation. Wustite was prepared from iron sulfate (Fisher Scientific Co., New York) which had a purity of over 99.92 pct, with small amounts of manganese, copper, and zinc adding up to 0.03 pct. From the aqueous solution iron oxalate was precipitated and decomposed to magnetite according to Economos. 18 The magnetite powder was pressed into cylinders and sintered in vacuum at 900°C for several hours. The sintered compacts were placed in a platinum crucible of 2.5 cm diam which was fitted into a high-purity alumina crucible and then inserted into the center of the graphite heating element of an electric resistance furnace.lg The furnace was evacuated to below 1 p Hg and the magnetite sample was slowly heated. The evolution of gases during heating deter-

Jan 1, 1967

By Alexander Knoll, Dwight L. Baker

THE work herein reported shows that galena in certain organic solutions of fatty acids becomes coated with lead soaps, and that this coating is not only highly water-repellent but is also repellent to certain organic liquids, particularly those [ ] of high surface tension; i.e., nitrobenzene, diphenylmethane. WETTING PROPERTIES OF GALENA IN NONAQUEOUS FATTY ACID SOLUTIONS When a "captive bubble"' is withdrawn from a polished, plane galena surface under a nitrobenzene solution of stearic acid, the air bubble shows a decided tendency to "cling" to the mineral surface. The contact angle, therefore, was greater than zero, although it was not large enough to be measured quantitatively by the captive-bubble method.* [ ] When pure nitrobenzene was substituted for the nitrobenzene-stearic acid solution, no clinging of the bubble was observed; viz., contact angle = zero. Although the contact angle developed in air-galena-nitrobenzene-fatty acid solution was small, the organic solvent repellency of the galena surface was great enough to enable good recovery in a laboratory flotation cell.2 In Table I are given the results of flotation experiments using several different fatty acids dissolved in nitrobenzene.

Jan 1, 1941

By H. P. Rassbach, E. R. Saunders

MUCH progress has been made in recent years in the theory and practice of making stainless steel. By effective utilization of oxygen for decar-burization and more suitable alloying agents, it has been possible to attain consistent production of very low-carbon stainless steel. In order to facilitate economical production of stainless steels, the Electro Metallurgical Co. has carried out an extended experimental program that has clarified some of the complex interrelations of temperature and composition under decarburizing and reducing conditions. These results'-:' have been founded in large part on small experimental heats, and in order to confirm their validity and significance under commercial conditions, a survey of stainless steel melting practice has been made with the cooperation of stainless steel producers. Conditions required for decarburization and associated oxidation of chromium, manganese, and iron have been fairly well established in relation to the beneficial effect of the highest practicable temperature. The recovery of chromium and manganese from highly oxidized slags by reduction with silicon has been indicated with somewhat less precision and this study has shown significant deviation in large commercial heats from the small experimental heats. In spite of an unsatisfactory degree of accuracy in estimating the conditions affecting reduction of metallic values, it has been possible to calculate a slag weight in reasonably good agreement with the results observed in large heats. While there still is much to be learned about both the qualitative and quantitative aspects of stainless steel melting, this survey has indicated the manner by which the efficiency of the production process may be measured. Oxidation Period In order to establish practices for the recovery of chromium and manganese from the slag the amounts of these metals oxidized should be known. Moreover, since reduction of oxidized chromium and manganese must necessarily be accompanied by similar reduction of iron, knowledge of the total quantity of metallics oxidized is essential. The relations between carbon, chromium, and temperature under oxidizing conditions were developed by Hilty in 1948.' While it had been realized previously that retention of chromium in the metal during carbon oxidation is favored by high temperatures, the Hilty relation provided quantitative information useful for evaluating melting procedures. For example, it was shown that decarburization to moderate carbon levels can be achieved while retaining substantial amounts of chromium. However, in decarburizing to the low-carbon level necessary for producing 0.03 pct maximum carbon stainless steel, very little chromium remains in the bath at the temperatures generally employed. For this reason, Hilty, Healy, and Crafts' later projected an extension of the chromium-carbon relation to the low-chromium region. Although this chromium-carbon-temperature relation defined the composition of a chromium steel bath at the end of the oxidizing period, it did not provide a direct means for estimating the total amount of metallic oxidation occurring during decarburization of a stainless steel heat. This subject was investigated by Crafts and Rassbach³ and, later, by Hilty, Healy, and Crafts" who demonstrated that metallic oxidation is a function of the chromium content of the initial furnace charge, the minimum carbon content attained, and the temperature. It was further demonstrated from data on 1-ton heats that an empirical relationship exists between the ratio of chromium plus manganese to iron in the slag (S) to the corresponding ratio of these components present in the metal bath. The following expression was derived to permit the calculation of the amount of metallics oxidized during decarburization of a given charge: W-2000 (Cr1 + Mn1)-(Cr2 + Mn2) 1 100s/s+1-(Cr3 + Mn2) where W is the pounds of chromium, manganese, and iron oxidized per ton of charge; Cr1, the percentage of chromium in the charge; Mn1, the percentage of manganese in the charge; Cr2, the percentage of chromium in the bath after oxidation; Mn2, the percentage of manganese in the bath after % Cr + % Mn oxidation; and S =%cr+ % Mn/%Fe in the slag after %Fe oxidation. In order to establish whether the slag ratios of commercial heats are consistent with those found in the experimental heats, the slag and metal analyses shown in Table I were evaluated. These data represent samples taken after the oxidation of two 1-ton and seven 15 to 25-ton heats of types 303, 304, and 304L stainless steel made in chromite, acid, and basic hearths. Fig. 1 illustrates that the results for the commercial heats correlate reasonably well with the relationship originally established for experimental 1-ton heats. It will be noted that in the range of metal composition of these heats, the slag values generally lie above the line. As was suggested by

Jan 1, 1954

By Euzebio De Oliveira

ONE of a recent batch of samples from the Serra da Baliza, in the state of Parana, Brazil, contained asphalt and a dark heavy oil; and workmen on the railway from Porto União to Uruguay discovered asphalt, coming from eruptives that outcrop along the Rio de Peixe. The occurrence f asphalt in the triassic eruptives of southern Brazil, however, has been known a long time, according to Dr. Gonzaga de Campos. It is generally believed that the Botacatú sandstone is always a hard vitrified rock, from the metamorphic action f. the overflowing eruptive contacts. In this region, however, the contact metamorphism is almost nil; the sandstone is slightly hardened in a narrow zone about 20 to 30 cm. wide. In many places, the sandstone is so friable as to be easily reduced to sand, which is used-.in, mortar for building in Guara-puava and Palmas. South and west from Porto União,. this bench of sandstone is about. 50 m. thick, and is capped by a heavy bed. f basic eruptives, many of which are amygdaloids. Dr. Geo: P. Merrill, after studying the triassic eruptives collected by the Coal Commission, reached the conclusion that. "All these rocks are of typical basalt-diabases, not in any essential different other than in structure. An interesting mineralogical phase is its paucity in olivine, which in many cases is completely lacking. Professor Hussak, who carefully studied these rocks and their accessory minerals, decided that in the dikes they are granular (diabase) and, that in the lava sheets, porphyritic (augite-porphyrite or melaphyre), and that the latter pass evidently to normal diabases and are always typical of effusive rocks. The examination f many slides from dikes and sheets leads us to adopt the opinion of Professor Hussak; the rocks of the dikes are of ophitic structure, while that of the sheets show a great variety f structure and may vary from almost granular to basaltic. The great paucity in olivine had been noted by Hussak, who classes as melaphyres, the porphyritic triassic rocks in Brazil which contain olivine.

Jan 4, 1921

By J. T. Shapland

When externally mounted sliding-gate valves are used to control the flow of metal from ladles, none of the valve components are immersed in steel. In general, the refractories used to date in sliding-gate valves have been made of a high-alumina composition. Because these refractories replace the fireclay sleeves and clay-graphite heads used to construct conventional stopper rods, they represent a major change in the use of pouring-pit refractories, Changes in the well and lining of the ladle as a result of conversion to external flow control probably will not be as severe; however, the sliding-gate valve makes it possible to preheat to a higher temperature and to decrease the turnaround time, and materials more refractory than those currently employed, such as basic brick, might be used in these areas. Another area of pouring-pit refractories that has became quite important is the tundish used with continuous casters. The use of external flow-control devices on tundishes makes it possible to use the tundish for prolonged periods of time and makes the use of better refractories necessary. In this application, high-alumina and zirconia are finding acceptance.

Jan 1, 1972

By N. J. Themelis

What is the "best" copper smelting technology? When a future Agricola examines the development of copper smelting in the 20th century, he will be amazed at how little took place in the first half of the century and how much in the second. In a period when the technology of transportation, communications and chemical processing advanced many fold, copper smelting remained practically static. The result of tightening energy supplies and stringent environmental controls in recent years has been a spate of development activity that not only is challenging the supremacy of the reverberatory furnace, but threatens to wipe out pyrometallurgical techniques entirely.

Jan 1, 1976

By B. Egeberg, H. B. Smith

Britannia metal is a white alloy consisting primarily of tin and antimony, the tin greatly predominating. The alloy usually contains a small amount of copper and occasionally very small amounts of one or two other metals, such as bismuth, zinc, and lead. It is capable of being cold-rolled, it takes a high polish, and it is fairly noncorrosive; properties which make it a popular metal for many household and traveling utensils. Britannia metal is similar in many respects to pewter and has largely replaced pewter in the arts, but it should not be confused with that alloy, because pewter is essentially a tin-lead alloy, while Britannia is a tin-antimony alloy. In view of the fact that so little has been published concerning Britannia metal, the authors feel justified in presenting the results of their investigation, particularly since the behavior of this alloy under cold rolling and heat treatment differs so greatly from that of other metals and alloys. The old craftsman, engaged in the making of ornaments or tableware, heated the cold-rolled sheet metal or his finished product to a straw color corresponding to about 375" I?., claiming to produce thereby a harder and more rigid article. The fact that possibly led him to this heating was that a reheated article or sheet made from Britannia metal has a distinct metallic ring when struck a sharp blow, whereas an unheated article or sheet has a leady, dead note. No doubt he has associated the metallic ring with a greater rigidity and strength, and justly so, as will be seen later. The approximate analysis of the metal we investigated was 91 Sn, 7 Sb, 2 Cu, and the molten metal was bottom-poured into an iron mold yielding an ingot 7 3/4 in. wide by 11 in. long and % in. thick, the casting temperature being 800" F. The ingot was cold-rolled to a sheet 0.032 in. thick, sufficiently large samples for experimental purposes being cut from the sheet at intermediate stages. The samples were all tested for hardness with scleroscope and Brine11 machine, then heated to 212" F. for 30 min., and again tested after cooling to room temperature. The same samples were heated successively at 300°, 400°, and 440" F., followed

Jan 1, 1929

By H. B. Smith, B. Egeberg

Britannia metal is a white alloy consisting primarily of tin and antimony, the tin greatly predominating. The alloy usually contains a small amount of copper and occasionally very small amounts of one or two other metals, such as bismuth, zinc, and lead. It is capable of being cold-rolled, it takes a high polish, and it is fairly noncorrosive; properties which make it a popular metal for many household and traveling utensils. Britannia metal is similar in many respects to pewter and has largely replaced pewter in the arts, but it should not be confused with that alloy, because pewter is essentially a tin-lead alloy, while Britannia is a tin-antimony alloy. In view of the fact that so little has been published concerning Britannia metal, the authors feel justified in presenting the results of their investigation, particularly since the behavior of this alloy under cold rolling and heat treatment differs so greatly from that of other metals and alloys. The old craftsman, engaged in the making of ornaments or tableware, heated the cold-rolled sheet metal or his finished product to a straw color corresponding to about 375" I?., claiming to produce thereby a harder and more rigid article. The fact that possibly led him to this heating was that a reheated article or sheet made from Britannia metal has a distinct metallic ring when struck a sharp blow, whereas an unheated article or sheet has a leady, dead note. No doubt he has associated the metallic ring with a greater rigidity and strength, and justly so, as will be seen later. The approximate analysis of the metal we investigated was 91 Sn, 7 Sb, 2 Cu, and the molten metal was bottom-poured into an iron mold yielding an ingot 7 3/4 in. wide by 11 in. long and % in. thick, the casting temperature being 800" F. The ingot was cold-rolled to a sheet 0.032 in. thick, sufficiently large samples for experimental purposes being cut from the sheet at intermediate stages. The samples were all tested for hardness with scleroscope and Brine11 machine, then heated to 212" F. for 30 min., and again tested after cooling to room temperature. The same samples were heated successively at 300°, 400°, and 440" F., followed

Jan 1, 1929

By S. R. Dunbas, D. C. Jillson, E. A. Anderson

THE present work was undertaken to furnish information, lacking in the literature, on the deformation mechanisms active in pure titanium at room temperature. Since it was started, Rosi, Dube, and Alexander' have indicated that slip can occur in coarse-grained (2 to 6 mm diameter) titanium on the {1010} and {1011} planes and twinning can take place on the {1072}, (1122) and (1151) planes. Liu and Steinberg,' working with titanium flakes produced by fused salt electrolysis, found evidence of twinning on the {1032$, {1121), {1122}, (1123}, and (1154) planes. The crystals used in the present work were substantially longer (up to nearly 2 in. in length) and possibly purer. Production of Large Crystals It was the purpose of this part of the work to provide single crystal or large grain specimens for use in the deformation studies rather than to study the mechanism of crystal growth. Accordingly, all further exploratory studies on methods were abandoned when it was found that cold-rolled bars sealed in vacuo in individual Vycor tubes and given three or more cycles of 4 hr at 1200°C and three to five days at 850°C, gave large grains. Eighteen specimens about 0.2 in. square x 2 in. long yielded 33 useful grains, ranging from Va in. to nearly 2 in. in length, with an agreeably wide range of orientations, Fig. l. The 0.25 in. thick strip from which the starting specimens were cut had been reduced 67 pct on 18 in. diameter rolls from a scalped, hot-forged (959" to 1000°C), arc-melted, iodide titanium ingot. Thc bar was at room temperature, but the rolls were at 200°C. Reductions of about 0.03 in. per pass were used. The bar was cut into four segments, each of which was rolled in a somewhat different manner. No effect on the growth of large grains was noted. The composition of the titanium, determined on turnings from the ingot scalping operations, is given in Table I. The Rockwell A hardness of the ingot was 26. The standard starting specimens, about 1/4x1/4x2 in., were milled on all four sides, hand polished on 280 and 400 mesh emery papers, etched for 3 min in a 50 HF-50 glycerine solution, rinsed, and dried in alcohol and ether just before sealing them in Vycor tubes. The average specimen weight was 7.5 g, and the tube volume was 6.5 cc. A vacuum of less than 1 micron pressure was applied. The specimen tubes were placed in an evacuated Inconel muffle to minimize the danger of collapse of the glass tubes at temperature. In practice the specimens were first heated to 1200 °C for 4 hr and then transferred fairly quickly to an 850°C furnace where they were held for three to five days. This cycle was repeated at least three times before the specimens were cooled, removed from the tubes, lightly polished and etched on one surface for examinstion. Those in which large grains had not developed were resealed in Vycor and returned to the cycling. Satisfactory specimens (about one out of every four cycled) were polished with emery paper on all four side;, then metallographically polished using ctchinz between wheels to remove distorted metal and, finally, given a very light etch to reveal the grain boundaries. This was necessary because the surface was roughened by the sublimation of titanium under high temperature vacuum conditions. Specimens were identified by TiXl numbers using supplementary letters to designate specific grains.

Jan 1, 1954

By Joseph F. Shirley

A froth flotation plant to recover byproduct molybdenite must be designed specifically for the selected separation process and the type of ore being treated. However, some general rules concerning byproduct molybdenite plant design can be suggested that would be valid for most processes. The term byproduct molybdenite in this paper is limited to the recovery of molybdenite from copper porphyry ores, which essentially amounts to selectively separating molybdenite from copper and iron sulfide concentrates. In addition, most of the discussions would apply only to the almost universally used approach of floating the molybdenite and depressing the copper and iron sulfide minerals. Thickener It is considered essential that the feed to a byproduct molybdenite plant be thickened prior to being treated for molybdenite recovery for the following reasons: 1) Thickening allows the rougher and any preconditioning stages to be operated consistently at the optimum percent solids. 2) Generally, the size of equipment used for pretreatment and the rougher flotation is considerably smaller and less expensive than if the copper flotation concentrates were treated directly. 3) The removal of as much water as practical by thickening removes flotation reagents added in the copper flotation circuit, especially frother which can be quite troublesome in a byproduct molybdenite plant. Although thickening of the molybdenite plant feed has long been recognized as necessary to optimum metallurigcal results, very often this area has been designed poorly. The thickener should be sized so that it can produce a consistent thickened product containing 60 to 65% solids without the use of flocculants. Some plants are only able to thicken to 45 to 50% solids consistently while other plants let the percent solids vary from 45 to 60% over a 24-hr period. While the former is more desirable than the latter, in many cases better metallurgical results could be obtained if the percent solids could be held consistently close to 60%. Probably the simplest and most satisfactory system of handling the thickener underflow is to spigot it into a sump, using a sphincter type valve for the orifice, controlled by a density meter. Another satisfactory approach is to use an ODS type pump to meter the underflow. There are other methods of handling thickener underflow which work satisfactorily as long as the amount of copper concentrates entering the thickener is relatively uniform. However, they often fail to operate satisfactorily when there are large fluctuations in the thickener feed rate due to fluctuations in ore tonnage and/or copper grade. If the percent solids in the plant feed could be held close to 65%. There is the possibility the molybdenite plant tailing would not have to be thickened prior to the copper filter plant. This means that the molybdenite rougher float would have to be conducted at 45 to 50% solids with all the cleaner tailings returned to the rougher for dilution. Admittedly, in some situations this approach would be impossible, but if capital or plant area is limited, it should be considered.

Jan 1, 1981

By E. P. Barrett, T. L. Joseph

The large iron-ore producers on the Mesabi Range are able to maintain the silica in their shipping products at from 8 to 10 per cent. by mixing ores of various grades, some assaying 4 per cent, silica and some as high as 18 per cent. silica. For each ton of 4 per cent. silica ore mined, a ton of 14 per cent, silica ore can be mined and mixed with it and the average grade maintained at 9 per cent. This mixing operation has made possible the utilization of large quantities of ore too lean to be used alone, and herein lies the advantage of the producers who operate a number of mines at the same time. As the grade of the ore from one of the properties falls off, a little high-grade ore from another property can be mixed with it and the average maintained, while if one property alone were being operated all of the ore below grade would be left behind. This mixing operation, however, can continue only as long as there is high-grade ore available for mixing with the high-silica ores and, while the tonnage of iron ore in the Mesabi Range is still enormous, the tonnage of high-grade ore suitable for mixing with any quantity of low-grade ore is being rapidly depleted. Obviously, if an 18 per cent. silica ore can be reduced by concentration methods to 12 per cent. silica, considerably more of it can be utilized with a given amount of high-grade ore. The operators on the Mesabi Range recognize this situation and are energeticdly endeavoring to reduce the silica in their concentration plants because every unit of silica that can be removed by concentration is a conservation of their high-grade ores. More Elaborate Processes Will Be Needed At the present time there are thirty-seven concentrating plants on the Mesabi Range, and about one-third of the total ore shipped is beneficiated in these plants. In about one-half of these plants no reduction is made in the silica but the moisture or structure of the ore is improved. The washing process is by far the most important method for the removal of silica, and it is interesting to observe that the tonnage of "wash" ore during the past five years has shown practically no increase although the total tonnage of ore shipped has increased. This is due to the fact that the available tonnage of wash ore is being depleted and the amount

Jan 1, 1930

By A. Alonso, J. Daubenspeck

In the nickel extraction process at Nicaro, Cuba, dried and ground nickel bearing laterites are reduced at 1300-1400F in multiple-hearth furnaces. Nickel is leached from the reduced ore by aeration in an aqueous solution of ammonium carbonate. Nickel carbonate is precipitated from the liquor and calcined to yield nickel oxide. The following are the more important modifications of the Nicaro process since plant rehabilitation and expansion. Ten percent increase in tonnage was achieved without decreasing the fineness of the ore by minor mechanical modifications of the grinding plant. Another five percent increase in tonnage was attained by intentionally overloading the ore prepration and leaching circuits in an effort to control cobalt extraction. Modifications at the sintering plant allowed it to handle the entire output of the expanded plant. A fifty percent increase in leach liquor cooling facilities alone provided an increase in nickel production of five percent. Stronger fresh leach liquor for increased nickel washing efficiency was obtained through improved distribution of water to the ammonia absorption towers. Advantage was taken of the reduced calcine's capability to precipitate cobalt, and modifications in the leach liquor circulation provided controlled cobalt solubility with minimum harm to nickel extraction. Several process changes at the drying plant brought about a ten percent reduction in the fuel oil consumption at this pint. Controlled air addition to the reduction furnaces' upper hearths burned formerly wasted combustibles thereby reducing oil consumption by twenty percent. The nickel extraction process successfully emplayed on the extensive but lowgrade lateritic ores of Eastern Cuba at the U. S. Government owned Nicaro plant, has been previously described in various publications. The large mine and plant operated by the Nickel Processing Corporation, located on the northeastern shore of Oriente Province, have been recently expanded to a yearly capacity of about 50 million pounds of nickel. Essentially, the operation involves open pit mining of a limonite-serpentine ore mixture containing 1.3 to 1.4 percent nickel and 34 to 38 percent iron. The comparatively lumpy, wet (25 to 30 percent moisture) ore is first dried and ground prior to chemical treatment, to which the entire 7,000 tons per day feed must be subjected. In the initial chemical action the dried and ground ore is reduced with carbon manoxide and hydrogen at 1300-1400° in multiple-hearth furnaces. The resulting metallic nickel is selectively leached from the reduced calcine by aeration in an

Jan 1, 1961

The Howe Memorial Lecture, in memory of Henry Marion Howe, Past President of the Institute, was authorized in April, 1923, as an annual address to be delivered by invitation under the auspices of the Institute by an individual of recognized and outstanding attainment in the science and practice of iron and steel metallurgy or metallography, chosen by the Board of Directors upon recommendation of the Iron and Steel Division. The following were the Howe lectures and lecturers for the years indicated

Jan 1, 1929

By C. W. MacGregor

In the testing of materials there exist various methods of recording graphically the behavior of a material subjected to tensile stress. Probably the most common method is to plot the tensile stress So, obtained by dividing the load P by the original area of cross section A0 at the start of the test, as a function of the strain e0, which is obtained by dividing the ?Lo or the change in the gauge length by the original gauge length Lo. A list of the symbols used here is given in Table 1. This method, while it has proved useful in the routine testing and comparison of various materials, does not have a very sound physical basis, as was first pointed out by P. Ludwikl, and consequently does not afford a very deep insight into the true physical behavior of the material under stress. Among the objections to this method may be mentioned: (1) that by dividing the load P at any stage of the test by the original area A. a fictitious stress is obtained, which does not actually exist; (2) that the strain e0 = ?L0 is not an exact expression for the true strain for large values of ?LO in per cent of Lo; (3) that when necking starts the strain varies considerably along the gauge length and the above expression no longer gives the true strain existing at any point on the outside of the bar; and (4) that the axial strain varies over the cross section of the necked portion and e0 does not represent an average of these strains over the cross section. Stresses taken from such a diagram would be only approximate and, if used to compare results of a test on a given material under simple tension with results obtained under combined stresses, certain errors would be thus introduced. In order to overcome some of these objections, various methods have been suggested, both as to the quantities to be measured and as to their graphical representation. In some cases the so-called "true stress," obtained by dividing the load P by the actual area of cross section A existing when the load P was applied, is plotted as a function of the strain

Jan 1, 1937