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By Paul M. Tyler

FULLY a decade ago, demand for ground mica began to exceed supplies of scrap mica from manufacturing operations and of waste block from feldspar and sheet mica mining in the United States, with the result that in recent years the market has absorbed larger and larger quantities of mica recovered as a by-product from clay washing,1 as a joint product with kyanite12 and by grinding schists. So far, only schists that contain a large percentage of mica have been utilized for making ground mica, but already there has been a good deal of wishful thinking with regard to recovering mica as an aid to the prospective working of low-grade gold ores and ores of certain base metals. With wet-ground mica selling at $60 to $100 a ton and even dry-ground mica worth over $20 a ton, many ores that could not be considered as otherwise workable under present conditions will show commercial possibilities if they can be made to yield even a relatively small percentage of salable mica. An interesting paper might be written on the coming role of industrial minerals as joint products of metal mines. It is largely a matter of markets. Many years ago I watched the mess left behind the gold dredges in the mountains of Colorado and thought of the Denver sand and gravel operators who prided themselves upon their canniness in providing devices for catching gold merely as an incident to their regular business of selling what we miners classed as "tailings." Nowadays almost every milling or mining operation near settled communities finds use for at least part of its waste for road material or concrete aggregate and several eastern iron mines have gone into the crushed-stone business. Promoters now mention fertilizer "filler" as potentially salable at all the way from $1 to $10 a ton, or whatever is needed to cover costs of mining and treating the "ore" and leave the metal values as "clear profit." In emulating the pork packers in their efforts to utilize "everything but the squeal,"

Jan 1, 1938

By R. J. Linney

Reserve Mining Co. produces 5 million tons of iron ore pellets per year. The finished product runs about 65.50 pct Fe, with 7.75 to 8.00 pct SiO2. Less than 12 pct arrives at the blast furnaces smaller than % in. This plant was originally designed to produce 3.75 tons annually—not 5 million. To handle the increased tonnage a fifth fine crushing line was added, and some very minor changes were made. When production can be stepped up 33 1/3 pct without extensive additions or changes—within the short period of a year—basic engineering has been sound. Operators of the plant consider it the best design job they have seen or heard of, and competent authorities have said that start-up was quicker and easier than for any previous large-scale minerals beneficiation plant. One big advantage is the elbow room allowed. Generally speaking, there is an eternal conflict between operator and designer over space. The designer has his costs per cubic foot, and he is keenly aware that if he permits the number of cubic feet to rise unnecessarily, total costs will be too high. But this was a new operation, and on a large scale. There were no proven standards to go by. Realizing this, Reserve's engineers provided ample and efficient access to machinery, allowing extra space for unanticipated situations. In three years of production this planning has paid off handsomely. Gravity Flow Saves Time and Money: This plant is on a rocky hillside, the car dumper about 270 ft above Lake Superior. In locating buildings and placing machinery the designers mad= good use of gravity flow, at considerable financial savings. In the concentrator building from rod mills to filters, gravity carries the material between all stages of the operation except one—the closed circuit pumping of concentrate between cone classifiers and ball mills. Since no machine is located directly above any other, an observer can stand on the truck ramp adjacent to the offices and laboratory and see practically every phase of activity in the huge building. The first mill section started in August 1955. After a day's run several modifications were made on auxiliary chutes and launders where spillage had occurred. This procedure was followed for the first

Jan 1, 1960

The research project on size reduction, sponsored by the mining industry through financial support from individual companies, is now embarking on a second three-year program. The first 3-year program had as its objectives assimilation of the literature on comminution and development of simulation programs encompassing the basic mechanisms of grinding. This involved investigating conditions under which a particle in a mill is captured, the stress to which it is subjected, its break probability, and the resulting particle size distributions. The results were used to obtain the most representative simulation models based on current knowledge of mechanisms. As part of this work a set of computer programs has been developed for the following cases: . Batch mill, by numerical integration for completely general selection and breakage functions. . Continuous open-circuit mill by same method as above. . Continuous open-circuit mill by general numerical integration method, including transient behavior

Jan 1, 1969

By Elmars Ence, Harold Margolin

The titanium-aluminum system has been investigated in the composition region 0 to 34 pct Al in the temperature range 800" to 1450°C. The phases encountered in this region were: a,ß, TiAl3, The reactions observed are as follows: 13 ; L - 6T. A1, 0 (15 pct Al) + 6 (18 pct Al) - yTi Al (16.5 pct Al) at -1170°C andP (7 pctpct Al) + y (9.2 pet Al) - ah(9 pct Al) at - 1060°C. A eutectoid reaction: y — a +d belm 700°C is postulated. THE Ti-A1 system has been the sub.ect of a number of investigations. Several of these1-3 have indicated extensive solubility of A1 in both a and ß titanium. Ence and Margolin, as a result of work on the Ti-A1-0 system:'5 have shown that the a solubility ofalumi-num is considerably restricted, and found that a compound, Ti2A1 existed in the region formerly designated as Q a This structure was identified as hexagonal with c/a = 0.803. Pietrokowsky confirmed the existence of this phase7 and pointed out that the structure was isomorphous with Ti3Sn.8 Although the existence of Ti,A1 has been confirmed by others,9-12 there has been considerable disagreement regarding the composition and mode of formation. A detailed reinvestigation of the Ti-A1 system has been published by Sagel et al.13 The tentative diagram indicated two phases, a, and e, occurring in the former all a region. The a, phase has the same structure as Ti2A1 and was indicated as having a broad region of solubility. Epsilon, a tetragonal phase, was shown to form within the a, solubility region. AS a result of unpublished work,14 Ence and Mugy3 olin disagreed with the interpretation of Sage1 et al., although some common features existed. The present study was conducted in an attempt to clarify prevailing uncertainties. EXPERIMENTAL PROCEDURE Alloy Preparation—Alloys were prepared from iodide titanium (99.9 pct Ti with less than 0.01 pct Zr) or Bureau of Mines electrolytic titanium (BHN-73) and high purity aluminum (99.99 pct). Alloys were arc-melted in argon or helium atmosphere as 15-g buttons in the composition range up to 34 pct Al. Depending on need, alloys were made in increments varying from 0.25 to 2 pct Al. Up to five 15-g buttons of some compositions were made. Weight losses during melting did not exceed about 1 pct of the charge and chemical analysis revealed that both titanium and aluminum were evaporated. In general, nominal and analyzed compositions agreed within 0.5 pct and therefore nominal compositions were used in plotting the data. Unavoidably, some overlap of composition occurred, particularly where small increments of A1 were used. The small increments were primarily used to obtain alloys which would fall into the narrow a -, r region and this purpose was achieved. In general, the data obtained from Bureau of Mines base titanium and iodide titanium alloys were in agreement. Forging—Both as-cast and forged samples were employed. Samples were heated for hand forging by an oxygen-hydrogen torch to the region 1200" to 1300°C. A titanium anvil and cover plate were used to prevent any iron pick-up and consequent formation of a low melting point compound. Frequent reheating kept the temperature in the desired range. Under these conditions, alloys containing up to 15.5 pct could be successfully forged. To check on the depth of contamination, forged alloys were heated slightlybelow the ß transformation temperature. Metal was removed to a depth below the contaminated surface so revealed, and the subsequent results obtained from these samples agreed closely with those secured from as-cast material. Heat Treatment—Prior to heat treatment, all alloys were vacuum annealed to remove hydrogen. Samples

Jan 1, 1962

Bill S.1431 for the creation of a Department of Natural Resources has received a propitious hearing before the Senate Committee on Government Operations. Interior Secretary Rogers C. B. Morton was one of the prominent witnesses who testified in favor. The department had been proposed, Morton said, with "the conviction that the development of all natural resources would be harmonized with the most practicable possible means to protect and preserve the environment, consistent with the social and economic needs of the Nation." Another opinion on the pro side came from Atomic Energy Commissioner Clarence E. Larson, who said that a "total picture" had become essential to the establishment of a policy for developing energy resources in the best public interest. Commissioner James T. Ramey, on the other hand, was one of the AEC dissenters. Ramey said he believed the reorganization of resource responsibilities would hinder efforts to deal with energy-environmental problems and prevent the use of AEC "expertise and facilities" to the greatest advantage. As Arnold Weber, associated director of the Office of Management and Budget, pointed out, however, retiring AEC Chairman Glenn T. Seaborg is in favor of the new department. Another supporter, and one who can cast a vote on the bill, is Sen. Frank E. Moss. Final Standards for emissions from new stationary sources will be published by the Environmental Protection Agency on October 27. The initial proposals, appearing in the Federal Register on August 17, called for limiting emissions per million Btu of heat input (maximum 2-hr average) to 0.2 lb of particulate matter, and, where solid fuel is burned, to 1.2 lb of sulfur dioxide and 0.7 lb of nitrogen dioxide. While the new rules are mainly for new plants, the EPA is also proposing them for "existing plants which are modified in such a way as to increase or alter. . . their emissions."

Jan 1, 1971

By A. K. Mukherjee, John E. Dorn

The effects of strain rate and temperature on the critical resolved shear stress for (321)[111] slip were determined for the silver-rich CsCl type of intermetallic compound AgMg. The flow stress increased only slightly as the test temperature was first decreased below room temperature; a rapid increase in the flow stress was obtained with yet further decrease in temperature from about 250' to 4OK. The effect of both the temperature and strain rate on the flow stress over the low-temperature range could be rationalized satisfactorily in terms of the Peiwls mechanism when the deformation was controlled by rate of nucleation of pairs of kinks. IN spite of the well-documented interest of metallurgists and engineers concerning the mechanical behavior of intermediate phases and intermetallic compounds, very little definitive information is available on this important subject. As recently summarized by westbrook,1 only limited and modest progress has been made thus far in elucidating the details of the mechanism of plastic deformation for the intermetallic compounds; the available information largely concerns phenom-enological descriptions of empirical observations and experimental facts, principally with reference to poly crystalline aggregates. The present report on the elucidation of the rate-controlling mechanism for slip in the bcc structure of AgMg is part of the comprehensive program of a systematic investigation on the mechanical behavior of intermetallic compounds. The intermetallic compound AgMg has a CsCl type of lattice structureZ and is completely ordered up to its melting point of 820c3 This material was selected for our present investigation because of its simple crystal structure, moderate and congruent melting point, ordered structure persisting up to the melting point, and some solubilities of the constituent elements, which could be expected to help in the growing of single-crystal specimens. Whereas previous investigations on the properties of AgMg include hardness,*"6 slip systems,7 tensile flow stress of polycrystalline specimens,' and electrical resistivity,4 the current investigation will be directed principally at elucidating the rate-controlling dislocation mechanism responsible for slip in single crystals at low temperatures. It will be shown that the strain rate is consistent with a model involving the rate of nucleation of pairs of kinks by the Peierls mechanism for plastic deformation. EXPERIMENTAL TECHNIQUES Single-crystal specimens of the AgMg intermetallic compound were prepared and tested as follows: 1) A master alloy ingot of AgMg was produced by melting and chill casting the high-purity silver (99.995 wt pct) and high-purity magnesium (99.997 wt pct) in an induction furnace under a helium atmosphere. 2) Sections of the above-mentioned ingot were placed in a graphite mold containing a spherical cavity in which a single-crystal sphere of 1 in. diam was grown under helium by the Bridgman technique. 3) The operative slip systems were investigated at room temperature on a singlecrystal of AgMg, from measurements of the angles made by the slip traces on the two surfaces of the specimen, which were at 90 deg to each other. The investigation confirmed that the AgMg compound undergoes slip primarily in the (321) plane and in the [Ill] direction, but a small amount of secondary slip was also noticed in the (112) (111) system. No slip was observed in the (110) planes. 4) The single-crystal sphere, mentioned earlier, was oriented in a graphite mold containing a cylindrical cavity of 3/8 in. diam above the spherical receptable to give the angle xo = 45 deg between the axis of the cylindrical bar and the normal to the slip plane (321), and the angle AO = 45 deg between the slip direction [Ill] and the axis of the bar. Oriented single-crystal bar seeds were produced by melting, under a helium atmosphere, a polycrystalline bar in the cylindrical cavity above the oriented spherical seed and by growing an oriented single-crystal bar from the seed. 5) Finally, oriented single-crystal specimens were grown from these cylindrical seeds. 6) The cylindrical specimens were machined in a High-Tension Spark Cutting unit to give a 2-in. gage length. The spark-machined gage section was elec-tropolished in a bath containing 200 ml of H3PO4,

Jan 1, 1964

By H. J. Beattie, G. C. Gould

J. B. Seabrook' recently published properties of a low-alloy Ni-A1 age-hardening steel known commercially as "Nitralloy-N". He mentioned three possible mechanisms of age hardening, viz. order-disorder in the matrix, order -disorder in a precipitate, and a precipitate that goes in and out of solution reversibly without losing coherency. The following results indicate that the last - mentioned of these mechanisms is the cause of age hardening in Nitralloy-N. The reversibility of the age hardening with a sub-critical solution temperature was demonstrated strikingly by Seabrook, whose results are reproduced in Fig. 1. Aging for 22 hr at 975°F (525°C) can reharden the fully-tempered martensite to its original as-quenched hardness of R, 45. A Nitralloy steel was obtained from Seabrook' with the following analysis:

Jan 1, 1962

The Tuscarora mining district, which has been highly productive of gold and silver ores in past years, is situated about 50 miles north- west of Elko, Nevada. The town of Tuscarora lies at an altitude of about 6,000 ft on the west side of the Independence Valley. Precipitation, particularly the winter snowfall, is considerably greater than in many other parts of Nevada, and temperatures are lower. From Tuscarora can be seen-across Independence Valley in the mountains to the east, a distance of about 12 miles, the Freeport Minerals Jerritt Canyon mine, which seems to be-destined to become one of the great gold mines in the western hemisphere. Tuscarora, a gold and silver camp, was active from the early 1870s to 1903 and produced about $10 million in those 30 years. Ten major mines and several smaller ones were operated in an area of about 1.45 square kilometers (one-half squaremile). The deepest shafts were only 230 m (750 ft) but most of the ore came from about the 105 m (350 ft) levels. Mineralization above the water table, at about 30 m (100 ft), was mostly native silver and cerargerite with free gold. Below the water table the ore was rich in sulfides including argentite, stephanite, proustite, pyrargyrite, enargite, arsenopyrite, bornite, chalcopyrite, pyrite, sphalerite and galena. A number of the properties contain sizeable dumps of what was con- sidered waste rock when it was mined. These dumps aggregate over 408,000 mt (450,000 st) with one dump, the Navajo, containing over one-third of the total. A few grab samples taken in the summer of 1978 from some of the dumps indicated interesting silver values with minor gold content. AgAu Exploration Ltd., was organized to explore and develop these dumps into a heap leaching operation. In the fall of 1978 a campaign was started to lease, with front money payments on a royalty basis, three groups of claims on which were located a few hundred thousand tons of dumps. In September and October of 1978, a systematic program of sampling was started on the Navajo dump which was most favorably situated in that it was near to an area which would furnish good sites for heap leaching pads. At the time it was fully recognized that sampling of dumps was a most difficult problem. Because of the topography on which they were situated and the manner in which they were formed, each dump required its own special sampling technique.

Jan 1, 1981

By Renato G. Bautista, Robert A. Hard

A comparative study has been made of the ex-tractability of several of the transitim metals from thiocyanate sohtions using methyl isobutyl ketone as the organic solvent. Extractions were made of scandium (III), chromium (III), manganese (II), iron (III), and cobalt (II), of which only the scandium , iron, and cobalt were readily extracted. All extractions were made at pH 2 and at 25.5 °C. A series of extraction experiments was made on scandium, iron, and cobalt to determine the effect of thiocyanate cmcentration added as the ammonium and also as the calcium salt. Also determined was the effect of metal ion concentratim and ionic strength on the distribution ratio. Analyses of the organic extracts indicated the scandium and iron were extracted principally in the uncharged trithiocyanate form, while the extractable form of the cobalt contained the cobaltotetrathiocyanate anion. THE chemical separation of metal complexes in solution by solvent extraction has in recent years been developed for large scale separation as well as bench scale work in the laboratory. It has also found increased application in analytical chemistry. Its extension to extractive metallurgy as a unit operation is best illustrated in the recovery of uranium and vanadium from their leached liquor.5-8 The extractability of some thiocyanate systems have been extensively investigated. The rate earths have been separated from one another using butanol as a olvent. Scandium was effectively extracted in ethyl ether using ammonium thiocyanate as the com-plexing agent.'' Similarly, the separation of hafnium from zirconium has been reported using ether""' or hexone13 as the organic solvent. The extractability of various metal thiocyanates in ethyl ether is given by 0ck.l Sharp and Wilkinson15 prepared nickel-free cobalt slats by thiocyanate extraction. The mechanism of extraction of titanium (IV) thiocyanate with methyl isobutyl ketone was determined by elafosse.''" The characterization of cobalt (TI) thiocyanate complex using methyl isobutyl ketone as the solvent at near neutral aqueous solution and the spectrometric studies of the ex-tractable cobalt complex has been made by Hard." It is noteworthy to mention at this point that the fundamental characteristic of ketone (oxonium) solvent extraction is that the extractable species are uncharged ion pairs solvated by the ketone molecules, but probably still partially hydrated.

Jan 1, 1963

By E. F. Thomas

THE late George S. Rice was active in the inves--I- tigation of bumps, particularly in the last ten years of his career as chief mining engineer of the U. S. Bureau of Mines. Since most of his investigation was carried out in Great Britain, continental Europe, and—to a lesser extent—Canada, his thinking on prevention was influenced considerably by the experience of those countries. It is not surprising, therefore, that when he was called upon a few years before his retirement to investigate bumps in the U. S. and suggest ways to prevent them, he turned to longwall mining. A longwall method had been most successful in combating the bump hazard in mining coal under deep cover, especially in Great Britain, but the prevailing method there at the time was advancing longwall mining, which he knew was uneconomical under U. S. mining conditions. For this reason he proposed a modified retreating longwall system that he believed included the best features of the advancing method. As brought out by Rice,' if the cover is 2000 ft and 50 pct of the coal is extracted, the static load on the remaining pillars will be about 4000 psi, which exceeds the ultimate crushing strength in most instances. If the pillar coal is overloaded before a pillar line is established, then the abutment zone preceding a line of extraction is no place to split pillars or extract them by any method other than an open-end system. Rice therefore advocated open-end mining, preferably by longwall, but he was willing to compromise with long-face mining if the longwall method was not acceptable. Rice's system was put into operation in a mine in Harlan County, Kentucky,3 but subsequent experience has shown that it did not take into account two important factors—avoidance of pillar-line points and maintenance of adequate development in advance of the pillar-line abutment area. For ten years after Rice's retirement the USBM did little investigation and research on bumps, chiefly because so few were occurring that there was not much cause for alarm. But in 1951 there were three occurrences involving fatal injuries, and the Bureau began a statistical survey in that year. C. T. Holland, head of the department of mines at Virginia Polytechnic Institute, was retained as a consultant. The resulting study' of 117 case histories brought out these important conclusions: 1) Almost invariably the bump occurred in a locality affected by the abutment zones of one or more pillar lines. 2) In most cases the locality of the bump was influenced by the abutment zones of more than one pillar line. The term pillar-line point has been used for many years in the Appalachian region for such a situation. Point is used in the geographical rather than the mathematical sense. 3) In pillar-line extraction the following practices are safest in preventing bumps: a. The mine layout should provide for pillars of uniform size and shape along the extraction line. b. The mine layout should be planned so that no development need be done in the abutment zone of a pillar line. c. The layout should permit open-end extraction of pillar lines from the next goaf, so that it will not be necessary to resort to pocket mining, splitting pillars, or any practice that will involve driving in the direction of the goaf within the abutment zone. d. Pillars should be large enough to support area without undue roof and floor convergence before establishment of a pillar line. These are, of course, generalities, and while they are useful in laying out areas where bumps can be expected, they are of limited help in many mines that were committed to a system of mining before it was realized that they were subject to bumps. Under such conditions it becomes necessary to choose between the following alternatives: 1) Abandon the territory, except for pillars that offer no extraction problems. 2) Through experience select the pillars that are most heavily loaded, and, by augering, induce bumps from a safe vantage point so that impinged loads are relieved. This method was first developed at the Gary, W. Va. mines of U. S. Steel Corp. and later adapted to mining thick coal beds at Kaiser Steel's Sunnyside mine in Utah. No scientific method is available to determine where to drill within a loaded pillar. Although this method of unloading has worked very successfully at Gary—with one exception—

Jan 1, 1959

By M. S. Burton, J. E. Ritter

The sessile-drop technique was employed to study the effect of atmosphere and alloy additions of titanium and chromium on the surface tension and contact angle of nickel on sapphire substrates at 1500oC. Shear tests were conducted on the solidified sessile drops. The liquid-gas surface tension of pure nickel on sapphire was found to be 1747 ergs per sq cm, with a contact angle of 109.6 deg. The surface tension and contact angle were not affected significantly by the four atmospheres, vacuum, impure argon, pure argon, and pure hydrogen. Small additions of titanium and chromium decreased the contact angle and for the same alloy content, contact angles were lower in pure argon than in pure hydrogen. On the basis of shear-test results , the most probable cause for adhesion in the Ni-Al2O system is the formation of NiO:Al2O3 spinel at the interface; in the titanium alloy-Al2O3 system adhesion is thought to be due to a mechanical interlocking effect; in the chromium alloy-Al2O3 system adhesion is believed to be the result of solid-solution formation involving Cr2O3 and Al2O3 at the interface. In the last few years there has been considerable interest in reinforcing metals with ceramic whiskers. It has been postulated that, by reinforcing nickel with sapphire whiskers, high-temperature strengths su- perior to those of the best current nickel alloys can be achieved.' To fabricate a whisker-metal composite having optimum density and strength, understanding of wettability and adherence in the system is of prime importance. Although there has been some previous research in this a a need for further studies has been cited4-8 to understand better the effects of impurities, alloys additions, and atmosphere on the wetting and bonding of metals to metal oxides. The purpose of this research was to investigate the factors involved in the wetting and adherence of nickel/nickel alloys to sapphire (single-crystal a Al2O3). Sessile-drop experiments were conducted in vacuum, impure argon, pure argon, and pure hydrogen, to determine the effect of atmosphere on the contact angle of the metals on sapphire, and the liquid-gas surface tension of pure nickel. The solidified sessile drops were sheared from the sapphire substrate to measure the bond strength between the metal and ceramic. The alloy additions chosen were titanium and chromium since both enhance the wetting of nickel on sapphire.2, 3c Also, titanium is commonly used when brazing metal-to-ceramic seals to improve adherence. Since chromium forms an oxide that is completely soluble in sapphire, it was thought that the oxide could serve as a transition phase to promote bonding between the Ni-Cr alloy and sapphire. I) EXPERIMENTAL APPARATUS AND PROCEDURE The sessile-drop apparatus was a horizontal furnace, with an induction-heated, cylindrical molybdenum susceptor in a water-cooled Pyrex chamber. This apparatus was patterned after that of Williams and Nielsen.7 A flat Pyrex window at each end of the furnace allowed temperature measurements to be made at one

Jan 1, 1968

(This list. includes only those who have entered military service within the past month, or whose entry has only recently become known to us; it also includes a few names of those whose titles or assignments have been changed within the month. A complete list will he found in the Year Book mailed with this Bulletin. Another Honor Roll is maintained in the reception room of the Institute.) ALDEN, HERBERT W., Lieutenant-Colonel, Ordnance Dept. ALDER, ALFRED; Engineers' Reserve Corps. ALFORD, NEWELL G., Sergeant, Bureau of War Risk Insurance, Headquarters 84th Division. AMBLER, HARRY A., Lieutenant, 30th Infantry. ANDERSON, GEORGE K., JR., Sec. C, 1st Co.; 3d Officers' Training Camp. AYER, -FRANK A., 1st Lieutenant, 43d Squadron, 3d Prov. Regiment, Aviation Camp.

Jan 3, 1918

By C. S. Smith, E. W. Palmer

IN 1934, when Gregg and Daniloffl wrote their excellent monograph on the alloys of iron and copper, the most recent literature on the constitution of the alloys indicated a narrow single-liquid area for 20" above the liquidus with a closed liquid mis-cibility gap above this. Since that time, however, a number of investigators have confirmed the much earlier studies of Mushet2 and Stead3 who, among others in the nineteenth century, had described the true state of affairs. Liquid copper and iron are completely miscible in the absence of carbon, but small amounts of carbon cause liquid segregation. In 1934, one of the writers' mentioned evidence for the absence of liquid separation in carbon-free alloys at temperatures up to 1700°C. Two years later, Mad-docks and Claussen5 unequivocally established the absence of liquid segregation and published a revised diagram, reproduced in fig. 1. They also studied the limits of the two-liquid area in ternary alloys of iron, copper, and carbon as a function of carbon and iron content. In the same year Simpson and Bannister" and Schumacher and Souden7 described in some detail the properties of alloys containing 50 to 75 pct Cu. More recently Hodges et al.16 have developed the copper-rich alloys for use as high-strength conductors. In 1937, Iwase, Okamoto, and Ameniya8 published data on the miscibility gap in ternary alloys of iron, copper, and carbon that supported the conclusions of Maddocks and Claussen, and definitely located the two-liquid area at temperatures of 1450" and 1540°C. They also reported that additions of 1 pct Al, Ni, Pb, Sn, or Zn caused no segregation in 50 pct Cu alloys at 1540°C. The use of copper in amounts approximating 1 pct has found industrial employment to an increasing degree in both wrought and cast steels, for such steels are capable of being precipitation-hardened following normalizing, and have high yield strengths. This is summarized by Gregg and Daniloff,1 Sallitt,9 and Lorig and Adams,10 while Alexander" gives additional information on the heat treatment of copper-steel castings. Lippert12 has described Digby's two-phase steels containing copper and chromium. The work described in the present paper was done in the years 1932 to 1934, at which time much of the information was new. The excellent work subsequently done in other laboratories but more promptly published makes the present paper in part only a confirmation of what is now well known. There has not, however, been published any unified study of

Jan 1, 1951

ON FEBRUARY 13, 1953 the Miami Copper Company signed a contract with the United States Government, through the Defense Materials Procurement Agency, whereby it under- took, with a base price of 27.359 per lb (subject to escalation based on wage scales and wholesale-price indices), to exploit a new ore body in its mine in Gila County, Arizona. Recoverable ore was estimated at 23,000,000 tons, averaging about 0.677 per cent copper. Moreover, the proportion of oxidized copper, difficult to recover by flotation concentrating, was so high that an average net recovery of only ten pounds per ton (0.5 per cent) was expected. The company proposed to mine the ore by the underground method known as undercut block-caving. Unfortunately the ore body, as has been true of all Miami's ore bodies, lay so deeply buried that no possibility existed of adopting opencut mining as has been done in recent years at Miami's neighbors, Ray and Inspiration. Probably the most distinctive characteristic of Miami's operations has been the extent to which it has developed and improved methods and equipment for underground mining. True, necessity was a potent stimulant, because its available ores since 1926 have been exceptionally low-grade; so, if the enterprise were to survive, it was necessary to keep costs at the minimum. This fact does not detract, however, in any way from the remarkable achievements of Miami engineers in conducting efficient and low-cost underground mining.

Jan 1, 1957

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915; when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps: Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible: Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn-drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blasthole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated' by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Such shots were fired from either end .by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1952

By Gregory L. Gullord

In September, 1975, production was begun at the new milling facility of ASARCO Incorporated, the Young Mill. The mill is located adjacent to the Young Mine 35 km (22 miles) east of Knoxville, Tennessee. The Young Mill replaced the obsolete Mascot Mill which had been operating since 1913. The basic flow sheet of Mascot was retained for the construction of the Young Mill, with a 20% increase in capacity included in the design. The Young Mill is capable of processing 7710 - t/d (8500 stpd) of crude ore from three underground mines (Young, Immel and Coy) yielding 300 - t/d (330 stpd) of 63% zinc concentrate. The Young Mill is unique in that 92.5% of the ore milled is recovered as saleable products, including zinc concentrate, construction aggregate, sand and agricultural limestone. Processing at the Young Mill is divided into five basic stages: two-stage crushing, heavy media separation, grinding-flotation, concentrate filtering and drying, and limestone by-products recovery. The ore bodies are located in the Knox formation, which consists of primary dolomite, secondary dolomite and limestone with silicas in the form of chert and jasperoid. The zinc mineralization occurs as an exceptionally pure sphalerite. The ore contains an occasional abundance of pyrite and traces of fluorite and barite. TWO STAGE CRUSHING Ore enters the crushing section from two points. The Young Mine ore is hoisted into a 816 ton (900 st) bin which feeds a 1.52m (60 in) conveyor via a 1.52 m (60 in) vibrating feeder. The Immel ore is trucked 10 km (6 miles) by an independent contractor and weighed on automatic recording truck scales at the entrance to the property.

Jan 1, 1977

By Albert Sauveur

Floris Osmond, Honorary Member of the American Institute of Mining Engineers, born in Paris, March 10, 1849, died at Saint-Leu near that city, June 18, 1912. Taken suddenly ill with congestion of the brain while on his way from Paris to his home, he reached it in time to write on the typewriter his last wishes, one of which was that no biographical notice of his life should be published. Holding Osmond's memory sacred, as those who have known him will, is it our duty to conform with his desire ? Shall no mention be made in our Transactions of the beautiful life and great achievements of one of our most honored members? Is it not, on the contrary, incumbent upon us to disregard Osmond's instruction in this matter and to place on record our admiration for the nobility of his life, for his talents and for the importance of his work, that younger men may be inspired by it with a desire to be guided by worthy motives in the path of unselfish pursuits ? I cannot feel that in following this course I am disloyal to Osmond's memory and I am strengthened in this belief by the opinion of men of high honor whom I have consulted before undertaking the writing of this obituary notice. One of them, Professor Henry M. Howe, justifies my decision better than I could do it myself when he writes : " Let us look frankly at the matter. Here we have a most extraordinary life, a life of rare beauty and power of inspiration to the rest of mankind. In my opinion we should not deprive mankind of the stimulating influence of the example of M. Osmond simply because he shrank from publicity. Our young men see every day how greed and vulgar push succeed, and with what material rewards they are crowned. They see too rarely the more valuable crown which comes to self devotion. A permanent and public recognition of the beautiful qualities of Osmond's life cannot fail to appeal to the imagination of young men and to help to deflect them toward leading a noble rather than a sordid life. Therefore I feel that the Revue de Métallurgie and other bodies owe it to the public to give prominence to his career by means of biographical notices."

Jan 1, 1914

By J. S. Blakemore, E. O. Hall

SERRATED yielding, commonly called the Portevin-Le Chatelier effect, is known to occur in a wide range of alloys where the moving dislocations are considered to be locked by the diffusion of solute atoms to them. From the work of Cottrell,1 an approximate relationship between strain rate, E, dislocation density, p , and solute diffusion coefficient, D, giving rise to serrated yielding is: e =Dp

Jan 1, 1969