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By A. J. Lyon, J. B. Johnson, Samuel Daniels

The wrought alloys of aluminum with small amounts of copper and of magnesium have, with the development of the automotive and aircraft industries, sprung into prominence through the medium of duralumin; but the cast alloys remained in the background because their possibilities in the heat-treated condition were either not recognized or not fully investigated. In 1919, the Bureau of Standards published a preliminary survey1 of this field, which pointed out that, in general, the ultimate strength and the percentage of elongation of the alloys of aluminum with and without magnesium could be improved by so simple a treatment as heating at 932' F. (500" C.) for 2 hr., cooling in air, and aging at room temperature for several days. British investigators, too, working with chill-cast material, found that" considerable improvement in the mechanical properties of copper-aluminum alloys may be obtained from suitable heat treatment applied to castings."' Jeffries and Gibson8 made the first contribution to the literature dealing exclusively with the heat treatment of sand-cast aluminum-base alloys, describing the effect of small additions or of combinations of copper, magnesium, and iron on the response to a short heat treatment comprising heating at 932" F. in a nitrate bath for 1 hr., quenching in fish oil, and reheating at 300' F. for for 1 hr. Their results disclosed the facts that the ultimate strength and the percentage of elongation of sound alloys were usually, and quite uniformly, improved by heat treatment; that with the iron content ranging between 0.36 and 0.88 per cent. and the silicon content between 0.24 and 0.60 per cent. the effect of increasing the copper content (Tables 10

Jan 1, 1925

By W. L. Martin, J. N. Dew, H. B. Steves, M. L. Powers

This paper presents and discusses the results obtained during a pilot test in the Loco field in southern Okla homa. The test was conducted in a 2%-acre pattern that was part of a 20-acre conventional waterflood pilot area. The conventional flood was well past its economic limit when the hot waterflood was initiated to obtain technical information and operating experience. Temperature data from nine observation wells showed that about 75 percent of the pattern was affected by heat, and that heat losses were severe in the pilot pattern. About 60 percent of the injected heat was lost to overburden and underburden zones during hot-water injection. Wellbore heat losses were held at tolerable levels at the shallow depth of the test by providing a low-pressure air annulus between the injection tubing and well casing. Hot water provided water injectivity increases of 200 to 400 percent. The hot water channeled across the lower portion of the oil sand in three directions through zones of relatively high water saturation. There was no conclusive evidence that natural fractures or pressure partings aflected the flow of fluids in the pilot pattern; however response undoubtedly was aflected by localized pressure gradients and by injection-production rate ratios. The results showed that hot waterflooding increased oil recovery in a reservoir containing 600 cp oil. The total tertiary oil production from the pilot pattern area was 3,896 bbl or about 156 bbl/acre-ft swept by heat. The corre.sponding WOR was about 34:I. Description of the Hot Waterflood Process To hot waterflood an oil reservoir, water is injected that has been heated to a temperature substantially higher than the original reservoir temperature, but lower than the vaporization temperature of water at the prevailing pressures. In the reservoir the hot water flows continuously into cooler sand and rapidly loses heat to the sand until it has been cooled to the original reservoir temperature. Thus, a heated zone and a region or "bank" of cooled water begins to accumulate immediately after hot water injection is started. This bank of cooled water continues to grow ahead of the heated zone, which also grows, but at a slower rate. This occurs because heat transfer is almost instantaneous, and the ratio of heat capacities of water to rock is such that two or three unit PV of hot water must be injected to heat a given unit bulk volume of the reservoir. The primary displacement mechanism is the same for both hot and conventional cold waterfloods; i.e., "piston" displacement occurs at the original reservoir temperature. The incremental benefits of hot waterflooding usually occur long after the breakthrough of cold water at producing wells, and the increased oil recovery necessarily is accomplished with high WOR's (water-oil ratios). Heat decreases the viscosity and density of oil and water. These effects result in more rapid and increased recovery of secondary or tertiary oil. If the cost of the required heat is low enough, the ultimate oil recovery of a hot waterflood should be increased substantially over what would be expected at the economic limit of a conventional cold waterflood. The economic benefits of any hot-fluid injection project depend primarily upon the cost of the heat required to produce more oil at an increased rate. This cost depends in part upon the amount of heat lost to surrounding formations. Heat loss depends upon reservoir thickness, water injection rate and temperature, the depth of the formation, well spacing, and the characteristics of the reservoir rock and surrounding formations. In general, percentage heat losses decrease as injection rate and reservoir thickness increase. Although it is an old idea, hot-water injection has not received widespread field application in the oil industry as a drive process. Much of the original field work with hot water was done in Pennsylvania fields where water permeabilities and injection rates are low. In these cases, hot-water injection was used primarily as a means of increasing in-jectivities — not as a recovery process. Ramey' recently has published an excellent review of the development of hot fluid injection processes. Recent publications"' and our own experiences now indicate that hot water or steam injection also can effectively increase the oil recovery from reservoirs containing viscous crudes. However, the economics of these methods as displacement processes have not been established. Several theoretical predictive techniques have been published (for a review see Ramey'), but simplifying assumptions make

Jan 1, 1969

By Ming-Kao Yen

A considerable amount of work has been reported in the literature in regard to the texture and earing behavior of copper strip. The rolling texture of copper has been confirmed as (110) [112] and (112) [111], which yields ears of a drawn cup at the position 45" from the rolling direction.1-3 The recrystallization texture has been established as the cubic or (100) [001] texture, where the earing positions are at 0" and 90" to the rolling direction.4-8 It has also been reported that in the development of cubically aligned grains of copper strips, the percentage of this cubic texture increased with an increase of final reduction and final annealing temperature.8,9 A comprehensive study on H.C. copper (British commercial copper of high-conductivity quality = Cu 99.95 pct, O2 0.03 pct, Ag 0.003 pct, Fe 0.005 pct and Pb < 0.001 pct) was made by Cook and Richards.6 They concluded that the recrystallization textures could be described as one or more of the following textures: (1) a single texture (100) 10011, (2) a twin texture (110) [112] and (3) a random orientation, depending upon the previous history of the specimen concerned. The effect of various alloying additions in copper was reported by Dahl and Pawlek.10 They found that certain alloying additions, such as 5 pct Zn, 1 pct Sn, 4 pct Al, 0.5 pct Be, 0.5 pct Cd, or 0.05 pct P suppressed the formation of cubic texture. Brick, Martin and Angierll reported that the cold rolled textures due to various additions fitted a rather simple pattern. However, the recrystallization textures were subject to very considerable variations. In the discussion of this paper, Baldwin stated that deoxidized copper containing 0.02 pct P gave a complicated recrystallization texture at lower temperature. When this copper was annealed at high temperature, a single texture appeared which was described as (110) [ill] but. according to a pri- vate communication from Baldwin, this orientation reported was in error and should have been reported as (110)[112]. He also reported that the earing positions of drawn cups were at 60" to the rolling direction.12 Recently, Howald, in his discussion on the paper by Hibbard and Yen,13 reported that the rolling texture of phosphorus deoxidized copper, containing from 0.006 to 0.020 pct phosphorus, was of the pure copper type. When these coppers were annealed at lower temperatures, they exhibited a random orientation, and when they were annealed at higher temperatures they had a mixed (111)[110] and (100)[001] texture, depending on the severity of the final reduction and annealing temperature. However, the specific influence of phosphorus and other impurities on the recrystallization textures and the deep drawing properties of copper strip has not been thoroughly reported. Therefore, an attempt has been made in the present work to determine the rolling and recrystallization textures and also the earing behavior of five types of commercial copper and thereby to evaluate the effect of phosphorus and some other significant impurities on the development of texture for cold reductions of about 87 to 89 pct. Materials Used The five types of copper employed in the present investigation were two phosphorus deoxidized coppers of different phosphorus content (0.007 and 0.013 pct P), an oxygen-free copper (OFHC), an electrolytic tough-pitch copper, and a fire-refined tough-pitch copper. These materials were subjected to a thorough spectroscopic and chemical analysis. The designations and the chemical compositions were as shown in Table 1. The coppers, FA1, FA2 and FA3. were hot-forged from 3-in. billets into a ½ X 6-in. plate and cold rolled to the ready-to-finish gauge indicated below. FA4 and FA5 were hot rolled and scalped to ready-to-finish gauge. The grain size of all the materials in the ready-to-finish condition was about 0.030 to 0.045 mm. Table 2 shows the last stage of the production schedule for each copper strip used. Experimental Procedure ANNEALING, GRAIN SIZE AND HARDNESS DETERMINATIONS Specimens of each type of copper were finally annealed in air for periods of one hour at temperatures ranging from 300 to 1600°F and were subsequently cooled in air. The average grain diameter of the annealed specimen was estimated by comparing with a standard grain size chart. Hardness was determined on the Rockwell 15 T scale. CUPPING TESTS Cups were made in a blanking and drawing set, in which blanks of 2-in. diam were drawn to a cup of 1.25-in. diam with an average depth of about 0.75 in. The clearance between the punch and die was about 0.032 in. The ears of the cup were measured with a special fixture which read the height of ears to one-thousandth of an inch on every ten-degree interval along the circumference of the cup. POLE FIGURES The usual transmission diffraction method with unfiltered copper radiation was employed to determine the pole-figures of the specimens cold-rolled or annealed at 900°F. All the pole-figures were derived from the positions of intensity maxima on 111 diffraction rings of the X ray photo-grams taken at 10 rotation of a

Jan 1, 1950

APPROXIMATELY 90 per cent of the steel that is melted and refined in the United States and poured into ingots is made in basic open-hearth furnaces, as shown in Table 1-1. The annual ingot capacity of the steel industry reported January 1, 1950, by the American Iron and Steel Institute3* was 99,040,160 tons, of which the basic open-hearth process accounted for 85,896,580 tons, 86.7 per cent of the .total. The capacity of the individual furnaces varies over a wide range, a large majority falling within limits of 50 to 200 tons (Fig 1-1). Most of the units constructed [ ] since 1942 are in the larger size groups, between 175 and 225 tons capacity.

Jan 1, 1964

By AIME AIME

THE autumn is the time that nearly all the special groups within the broad field of the Institute&apos;s activitives chose for their own special meetings. The big annual meeting in New York in February has its own special advantages, but to realize them involves sacrificing others. A meeting of a smaller group with the subjects brought up for discussion much more closely limited to a single field of interest has some advantages that a big general meeting cannot possibly have. And so the custom has gradually developed of having these smaller meetings in the autumn, at con¬venient places, so the men will not have to travel too far from the job to be present. Sometimes the meetings have for their field of interest a single phase of mineral technology, and are called divisional, and at other times they cover a diversity of topics that are locally important in that geographical area, and are then called regional meetings. Of the five meetings to be held this fall, four are divisional and one regional. Two of the divisional meetings are to be held in the same place at the same time, but with programs so arranged that they do not conflict.

Jan 1, 1930

By C. W. Allen, L. C. Moore

THE Mather mine, of the Negaunee Mine Co., is within the limits of the City of Ishpeming, on the Marquette iron- range in Michigan&apos;s Upper Peninsula. It is named for William G. Mather, who has served as President, and Chairman of the Board, of The Cleveland-Cliffs Iron Co. for well over 50 years. The property comprises nearly the square mile of section 2, T.47 N. R.27 W., which, except for 20 acres in the northeast corner, was leased by The Cleveland-Cliffs Iron Co. to the Negaunee company. The latter is a partnership of Cleveland-Cliffs and the Bethlehem Steel Co., organized to equip the property, retaining Cliffs in charge as operator. Shipments of ore from this area are mainly through Marquette, which is 16 miles to the northeast on Lake Superior, or through Escanaba, a Lake Michigan port, 65 miles to the south. Cleveland-Cliffs, in 1943, operated 12 mines in the Lake Superior district, with shipments of 6,635,576 tons for the year. GEOLOGY The north footwall of the Marquette Range synclinorium outcrops near the north line of sec. 2, the iron formation dipping to the south at an angle of about 40º. The south side of the trough reappears at a distance of 4 ½ miles on sec. 26 and the surface geology between shows iron formation or intrusive diorites. The latter are harder than most phases of the iron formation and, as a result of glacial action, appear as buttes or low-lying hills. The general elevation of sec. 2 is about 1450 ft. above sea level, or 850ft. above Lake Superior, and the higher diorite outcroppings about 1600 ft. above sea level. The geological structure has been complicated by a series of major and minor faults, one of which may be illustrated by the diorite sheet, which tilts south with the iron formation in the north half of the section and then reappears near the center to repeat this sequence. Sectionally the geological series includes a nearly eroded capping of Goodrich quartzite followed by the considerable thickness of Negaunee iron formation, which is cut by intrusive dikes or diorite sheets, and underlain by the Siamo slates grading into quartzites. The ore deposits occur mainly in troughs formed by the intersection of impervious dikes and the footwall slate, but include locally enriched lenses or bands in the iron formation and also interbedded with the Siamo slate. The ore is a soft red hematite with an indicated dried analysis from the surface drilling to date of 61.50 per cent Fe, 0.134 per cent P, 5.40 per cent Si, 2.87 per cent Al and 0.015 per cent S. EXPLORATION A diamond-drilling campaign started in 1937 actually had its inception in 1917, when a number of shallow test holes were put down for the purpose of locating ore that might be quickly developed and mined

Jan 1, 1945

By Frank A. Wardlaw

MY subject covers the effect that recent metallurgy has had on operations at the Utah-Delaware mine. This mine is the old Highland Boy mine of Bingham Canyon, Utah, one that has now been in operation for about 30 years. It was first operated by Messrs. Weir and Newhouse as a gold mine, under direction of the Highland Boy Gold Mining Co. The ore was oxidized gold ore and it was treated on the property in a cyanidation plant. During the process of development, sulfide ore was found assaying about 5 per cent copper. It proved to be the discovery of what is yet the largest sulfide copper orebody in the State. The first shipment was made in December, 1896. A smelter was built in Salt Lake Valley to treat this ore and not long after the property was sold to the Utah Consolidated Mining Co. From 1900 to 1924, the mine was a large producer of direct-smelting ore which averaged for many years 4.5 per cent. copper, then later dropped to 2 per cent. copper. During the years 1912 to 1920, direct-smelting lead ore was also mined. In 1918&apos;and 1919 ore developments at depth showed a decided increase in silica content and a porphyry sill so closely associated with the ore that it was impossible to mine the ore clean and it was found necessary to build a copper flotation Eoncentrator at Tooele. The mill had a capacity of 1000 tons a day and made a recovery of 90 per cent., and would have been responsible for a continuance of profitable operations had it not been for a slump in the copper market and the necessary high cost of mining ore on the lower levels. RESULTS OF SELECTIVE FLOTATION In 1924 the mine was on the point of a shutdown which had all the earmarks of being permanent; it was being operated at a heavy loss and preparations were made to pull the pumps and salvage the machinery. The plant at Tooele was remodelled for selective flotation by the Inter- national Smelting Co. in order to treat custom ores, and incidentally to clean up a possible 10,000 tons of lead-zinc ore which it was thought possible to get in the Utah-Delaware before shutting the mine down. Results were so encouraging that an effort was made to develop more lead-

Jan 1, 1928

By L. S. Darken, H. A. Wriedt

THE constitution diagram of the iron-copper system derived by Daniloff&apos; exhibits, at the iron-rich end, phase fields similar to those of the iron-carbon diagram. At 1484° 1094°, and 850°C there are, respectively, a 6-y-L peritectic, a y-E-Lperitectic, and a a-y-E eutectoid. The phase designated E is copper-rich, containing 4 pct Fe or less according to the temperature. The composition of copper-saturated ferrite (a-solid solution) has been determined previously only twice as a function of temperature. By electrical conductivity measurements, Buchholtz and Köster2 found 3.4 pct Cu in ferrite at a eutectoid temperature of 810°C. They represented the composition of saturated ferrite by the equation, log (%Cu) = 4.32-4125 (1/T) Gregg and Daniloff3 reported that Köster later modified his opinion, raising the indicated eutectoid temperature to 833 °C and the copper content of eutectoid ferrite to 3.7 pct. The literature contains several indications4-7 that the work of Köster and his co-workers is not reliable. Norton,8 using an X-ray method, found that 1.4 pct Cu dissolved in ferrite at the approximate eutectoid temperature of 850°C and that the solubility decreased to about 0.3 pct at 700°C. Below 700°C his experimental data show almost no change in solubility. These findings, although inconsistent with the work of Ruer and Goerens4 and of Smith and Palmer,7 have been used generally since their publication. In investigations involving the observation of arrests in temperature during continuous heating or cooling the precise location of the eutectoid temperature has been obscured by a large hysteresis. The more reliable work 579 where the results are not effected by this hysteresis indicates a eutectoid temperature between 840° and 850°C. The foregoing discrepancies, along with a few observations in this Laboratory relating to the precipitation of copper from ferrite, prompted us to rein-vestigate the lower temperature portion of the Fe-Cu system. In this investigation the solubility of copper in ferrite was obtained by the chemical analysis of high-purity iron equilibrated with copper-silver alloys. This method eliminated the necessity for retaining by a quench the physical state of the specimen existing at the equilibration temperature. EXPERIMENTAL The reaction of the iron with the copper alloy was

Jan 1, 1961

Levi Holbrook was born in Westboro, Mass., March 7, 1836. He was a descendant of John Holbrook, who came from England in 1660 and settled in the Massachusetts Bay Colony. Mr. Holbrook was prepared for college at Williston Seminary, East-hampton, Mass., and after one year spent in the West, he entered Yale. He was a, member of Phi Beta Kappa and of Linonia. During the six months following his graduation he was in Boston, under medical treatment for his cyes; then he took a horseback trip across the Rocky Mountains to the Columbia River and back, and resided in Danville, Va., for a time. From 1860 to 1863, as his eyes permitted, he was a student of modern languages and literature in Cambridge, Mass., part of the time being a resident graduate at Harvard University. Finally, however, he was obliged to abandon the idea of a literary life and from 1864 to 1867 he was engaged in financial business in Boston. The following year, he spent in New York, and then traveled for a year in Europe, Egypt and Asia Minor. In 1871, he returned to New York City. Mr. Holbrook was a member of the New England Historic Gcncalogical Society, the Society of Colonial Wars, and the Sons of the Revolution, was registrar general of the Order of the Founders and Patriots of America from 1915 to 1920, and had been governor of the New York society of that order. He was also a member of the American Fine Arts Society, the Order of the Cincinnati, the American Geographical Society (of which he was a councillor and secretary). and the Geographical Society of America. He joined the American Institute of Mining Engineers in 1878, and served as manager from 1895 to 1897. Mr. Holbrook died on July 26, 1922, from organic heart trouble, at his residence in Center Harbor, N. H. He was married Dec. 27, 1871, in Newark, N. Y., to Viola, daughter of John A, and Elizabeth (Failing)Vowers, who died on Sept. 10, less than six weeks after her husband. His two children, Clark Holbrook of Red Bank, N. J., and Mrs. Julian P. Smith, survive him

Jan 1, 1923

Levi Holbrook was born in Westboro, Mass., March 7, 1836. He was a descendant of John Holbrook, who came from England in 1660 and settled in the Massachusetts Bay Colony. Mr. Holbrook was prepared for college at Williston Seminary, East-hampton, Mass., and after one year spent in the West, he entered Yale. He was a, member of Phi Beta Kappa and of Linonia. During the six months following his graduation he was in Boston, under medical treatment for his cyes; then he took a horseback trip across the Rocky Mountains to the Columbia River and back, and resided in Danville, Va., for a time. From 1860 to 1863, as his eyes permitted, he was a student of modern languages and literature in Cambridge, Mass., part of the time being a resident graduate at Harvard University. Finally, however, he was obliged to abandon the idea of a literary life and from 1864 to 1867 he was engaged in financial business in Boston. The following year, he spent in New York, and then traveled for a year in Europe, Egypt and Asia Minor. In 1871, he returned to New York City. Mr. Holbrook was a member of the New England Historic Gcncalogical Society, the Society of Colonial Wars, and the Sons of the Revolution, was registrar general of the Order of the Founders and Patriots of America from 1915 to 1920, and had been governor of the New York society of that order. He was also a member of the American Fine Arts Society, the Order of the Cincinnati, the American Geographical Society (of which he was a councillor and secretary). and the Geographical Society of America. He joined the American Institute of Mining Engineers in 1878, and served as manager from 1895 to 1897. Mr. Holbrook died on July 26, 1922, from organic heart trouble, at his residence in Center Harbor, N. H. He was married Dec. 27, 1871, in Newark, N. Y., to Viola, daughter of John A, and Elizabeth (Failing)Vowers, who died on Sept. 10, less than six weeks after her husband. His two children, Clark Holbrook of Red Bank, N. J., and Mrs. Julian P. Smith, survive him

Jan 1, 1923

By C. K. Leith, W. J. Mead

In 1911, we published in the Transactions a brief account of the lateritic alterations of serpentine in eastern Cuba, producing the important iron-ore deposits of the Mayari and Moa districts.&apos; The special feature of that article was a quantitative treatment of the alterations, based on analyses, to show just what had happened in terms of chemical, mineralogical, and physical changes. Professor Kemp has published an interesting account of these ores, containing additional quantitative observations of his own, tending essentially to confirm the nature of the changes that had been worked out.2 His discussion suggests to us the desirability of publishing certain additional quantitative data which were not included in our previous article. This seems especially desirable because quantitative measurements of lateritic alterations are rare, and this particular case may be regarded as typical of a fairly wide range of alterations of this kind. At the time of our examination of Mayari and Moa deposits 29 complete analyses were supplied us by the chemists of the Spanish-American Iron, Co., representing a graded series from the unaltered serpentine rock below to the ore at the surface. These analyses are given in the accompanying table. The mineralogical composition of the ores and rocks is comparatively simple and can be easily calculated from the analyses. In Fig. 1 the mineralogical changes from the serpentine rock below to the lateritic iron ore above, based on the analyses, are expressed in terms of weight, assuming alumina to have remained constant during the alteration. It is not easy by mere inspection of this table of analyses to determine what the progressive chemical changes have been in development of the ore. In order to bring this out clearly, the analyses have been graphically represented in Fig. 2 in such a fashion as to show the progressive additions and losses of substances during alteration.

Jan 1, 1916

LIMA (PERU) Established May, 1952 (Official year begins in June) Luncheon meetings third Wednesday monthly Ernesto A Baertl, Chairman Arthur C Hall, Chairman-elect Raymond R Beard, 1st Vice-Chairman Edgardo Portaro, 2nd Vice-Chairman Walter E King, Secretary-Treasurer Av Santa Cruz 331 San Isidro, Lima, Peru Alberto E Brazzini, Asst Treasurer Daniel M Rodriguez, Asst Secretary Robert P Koenig, &apos;56, Section Delegate Arthur C Hall, Alternate Section Delegate

Jan 1, 1955

By L. A. Roe

The flotation of fine particles less than five microns in diameter and the flotation of liquids from liquids generally requires bubble generation methods different than methods used in dispersed air type machines used in mineral flotation. Conditions in the separatory zone in mineral cells are generally more dynamic as compared to the quiescent conditions often absolutely necessary for the flotation of fine particles and liquids. An example of quiescent flotation with very fine bubbles is the lithium flotation plant which operated at Trona, California for 34 years, until it ceased operations in 1978. A review of the process development and the commercial flotation operations at Trona is given in this paper. The clarification of oil field brines by flotation of the oil from the brine has been used for over twenty-five years. Recent installations related to the Prudhoe Bay oil field and the Alaska pipeline are described.

Jan 1, 1980

By Thomas E. Garnar

Zirconium and hafnium are curious elements because they are almost always found together in nature. Zirconium was discovered by Klaproth in 1789 and isolated 35 years later by Berzelius. Because hafnium&apos;s chemical properties are so close to those of zirconium, it was not discovered until 1923 when Coster and Von-Heresey detected it using x-ray spectrographic analysis. These elements occur most commonly in nature as the mineral zircon (ZrSiO4) and less commonly as the oxide baddeleyite (ZrO2). They also are found as a variety of other silicates. Zircon was identified as a component in alluvial and beach sands in 1895, but it was not produced in any quantity until 20 years later. Zircon is always a coproduct from TiO2 mining and processing. During World War I it was produced as a coproduct of beach sand mining for titanium minerals just south of Jacksonville Beach, FL, and was patented as a refractory. It was not until the 1930s when Zircon Rutile, Ltd. began mining at Byron Bay on the east coast of Australia that zircon was first used as a foundry sand. Later in the 1940s, NL Industries, Humphreys Gold Corp., and E. I. du Pont de Nemours & Co., Inc. began production of zircon sands from fossil beaches in northeast and northcentral Florida. Baddeleyite first became available as a commercial product in 1916, but never in the quantity of zircon. Zircon holds a unique position as an industrial mineral because it is used for both its physical and chemical properties as well as an ore of zirconium and hafnium metals. Australia is the major zircon producer followed by South Africa and the US. Zircon is also produced in India, Sierra Leone, Sri Lanka, Malaysia, China, Thailand, and Brazil. Undeveloped heavy mineral beach sand reserves containing zircon are known to occur in Egypt, Malawi, Senegal, and Tanzania. Locations are shown in Fig. 1. Mineralogy Zirconium and hafnium are always present together in naturally occurring compounds. Most commonly they form as the silicate sometimes containing iron, calcium, sodium, manganese, and other elements. Less commonly they are found as oxides in combination with titanium, thorium, calcium, and iron. Zircon is the most common form. It is extremely resistant to weathering, therefore, often it is present in ancient beach sands and placer deposits. Although resistant to alteration from external sources, it is vulnerable to internal alteration as the result of thorium and uranium substituting for zirconium either in the zircon lattice or in solid solution. Alteration to the metamict state takes place as radioactive emanations from these elements disorder the crystal lattice, accompanied by hydration, reduction in specific gravity, and changes in color. The mineralogy of zircon and baddeleyite and variation of physical and chemical properties are shown in Table 1. The effects of zircon&apos;s internal alteration and darkened color can be corrected by heating to 1000°C for 30 min or so. A certain time-temperature relationship makes it possible for decolorization and reordering of the crystal lattice to take place when heating at higher temperatures for shorter periods or at lower temperatures for longer periods. After heating, x-ray diffraction patterns are sharper, the grains may be harder, and except for particles of malacon, the grains will be white or colorless in the absence of grain surface coatings. The color change is only temporary and the particles slowly darken with time. Zircon fluoresces light yellow in shortwave ultraviolet light. If exposed to ultraviolet light for extended periods of time, white (calcined) zircon turns purple. With time the induced color begins to fade and eventually disappears (e.g., calcined Florida zircon exposed to ultraviolet light for 24 hr still retains faint, but discernible, purple color after six months).

Jan 1, 1983

By Horace V. Winchell

PERHAPS the matter of greatest interest to all mining men at the present time is the question of income and excess profits taxes on mines. Every producing mine in the United States is called upon to render to the government an account of its operations and to pay a graduated tax upon its earnings. It is manifestly a matter in which every mine owner, every stockholder in a mining corporation, every mining and metallurgical engineer, every person employed by a mining company and every one engaged in an industry which in any way depends upon the continued operation of our mines or who consumes their products is financially, to a greater or less extent, involved. In view of this fact it is surprising to see how little attention is paid to it and how few of our members have taken the pains to investigate and understand the provisions of recent enactments by Congress and the ways in which these enactments are being applied to mining industries.

Jan 1, 1920

"Acta Metallurgica" William C Ellis, Dec &apos;57 Alfred Noble Joint Prize Committee Same as Rossiter W Raymond Award Committee AIME American Association for the Advancement of Science 0 B J Fraser, Dec &apos;57, J T MacKenzie, Dec &apos;58 American Geological Institute H A Meyerhoff, Oct &apos;57, Charles F Park, Oct &apos;58 Committee on Government Relations J L Gillson, &apos;56 American Power Conference M Tenenbaum, &apos;57 American Standards Association Mining Standards Board E R Borcherdt, Jan &apos;57, Otto Herres, &apos;57, J K Richardson, Dec &apos;56, Alternate E W Bullard

Jan 1, 1957

By P. G. Winchell, J. K. Jackson

The strtcture of body-centered Fe-Ni-C alloys (0 to 16.5 wt pct Ni) containing less than 0.015 wt pct C was investigated by measuring the carbon-diffusion peak at low frequencies with a torsion pendulurrr. By determining the Snoek damping in ferrite and in quenched alloys of identical composition, the effects of nickel content and heat treatment were separated, and the existence of a structure in which Snoek damping is greatly reduced was revealed in the quenched ternary alloys. The relative amount of this structure increases with increased nickel content and cooling rate. It is proposed that this structure may be highly imperfect or tetragonal. In Fe-Ni-C alloys containing only ferrite the addition of nickel (0 to 5.1 wt pet) decreases tke area of the carbon-diffusion peak and shifts its position to lower temperatures but broadens the peak only slightljl. The decrease in peak area with increasing nickel implies that short-range interactions with the nickel atoms prevent some of the carbon atoms from participating in the relaxation process. The shift in the peak position indicates a long-range effect of nickel on the movement of carbon atoms so that at 300°K the diffisivity of carbon in -ferrite is increased by about 35 pct by the addition of 5.1 tut pct Ni. The results are presented as distributions of relaxation strength 072 the activation energy under the assumption that the frequency factor is invariant. BODY-CENTERED iron-base solid solutions dissolve carbon interstitially in the octahedral sites. The occupation of an octahedral site produces a tetragonal distortion of the surrounding lattice. These sites fall into three sets according to the orientation of the tegragonal distortion associated with each and may be identified by their lattice coordinates as: type 1/2 0 0, type 0 1/2 0, type 0 0 1/2. At low carbon levels in cubic ferrite the carbon atoms are distributed at random1 among these sites. In high-carbon body-centered iron-base solid solutions, which are unstable with respect to several two-phase mixtures, a tetragonal structure occurs as the product of the diffusionless transformation of austenite. This transformation is commonly accompanied by appreciable microstrain and its product is called martensite. The tetragonality of martensite, a supersaturated solution of carbon in a iron, is maintained prior to decomposition even though local diffusion of carbon occurs as evidenced by carbide precipitation. The tetragonality is taken as evidence that one set of octahedral sites is preferentially occupied by carbon. (This preferential occupation is called ordering below.) Support for this view is presented by the absence of strain-induced carbon-atom diffusion in high-carbon solid solutions (martensite). Such diffusion occurs in ferrite and produces the 40°C internal-friction peak observed at 1 cps frequencies.&apos; Thus, there appears to be qualitative evidence favoring the view that at sufficiently high carbon levels tetragonality arises spontaneously as a result of the ordering of the carbon atoms; however, no experimental evidence exists concerning an equilibrium between cubic and tetragonal body-centered solid solutions. Moreover, attempts to estimate the energy of ordering7-&apos; predict a value too low to maintain the tetragonality observed at medium carbon levels, 0.3 to 0.6 wt pct C. In the present work an attempt using internal-friction measurements is made to establish the presence or absence of tetragonality in low-carbon body-centered solid solutions formed by quenching austenite. Fe-Ni alloys containing about 0.01 wt pct and 0 to 15 wt pct Ni were selected for the study. While the work was in progress Gilbert and Owen10 reported that the transformation which occurs in these alloys during quenching at rates up to 5000°C per sec is not martensitic but is more properly categorized as massive. Thus, the solid solution

Jan 1, 1964

By Michael J. Callow

THE PROPERTY of Mount Isa Mines, Ltd., is located at Mount Isa, Northern Queensland, Australia, at roughly 20" south latitude. It is 600 miles from Townsville on the east coast, the port for Mount Isa, 1200 miles from Brisbane and approximately 1600 miles from Sydney in New South Wales. Living conditions in the camp, though wild and crude at the commencement of building, are now excellent and little has been left undone to provide comfort and attractiveness. The climate being warm and dry, many of the buildings are built on posts, 8 to 10 ft. off the ground, with spacious screened-in verandas. Club houses, stores, school, hospital and ice plant have all been provided by the company as well as modern conveniences, such as running water, shower baths, electric lights, septic tank arrangements, tennis courts and amusement grounds.

Jan 1, 1932

By J. C. Allen, E. V. Herring

CHILE EXPLORATION CO&apos;s. oxide plant at Chuquicamata, the largest of its kind, started in 1915 with initial operations at the rate of 10,000 tons of ore per day. Anaconda Copper Mining Co. acquired the controlling interest in 1923 and in the next few years modernized the plant and mine, extending operations to a nominal capacity of 490 million lb of electrolytic copper yearly by 1942. Plant production averaging 495.7 million lb of copper per year was maintained for the eight-year period extending from 1941 to 1948 inclusive. During the World War II production reached its highest level when in 1944 the output was 531.8 million lb. The maximum production for any one month was 49.6 million lb, achieved in May 1947. E. A. Cappelen Smith, consulting metallurgist for M. Guggenheim&apos;s Sons, worked out the first process for the treatment of Chuquicamata copper oxide ore about 1913, and directed a staff of 20 young engineers operating a pilot plant at Perth Amboy, N. J., on three shifts for an entire year. The copper is extracted from the oxide ore by a hydro-metallurgical treatment described in detail by T. A. Campbell in AIME Transactions, Vol. 106, Copper Metallurgy, 1933. The Oxide Plant is divided into five main operations as follows: (1) Crushing, (2) Leaching, including dechloridizing and sulphur dioxide treatment, (3) Tailing disposal, (4) Electrolytic precipitation, (5) Smelting and melting. Crushing The present Primary Crushing Plant has been in operation 25 years, during which time 4,342,112 cars of 70 tons net capacity have passed through the dumpers. The ore as received from the mine is frequently blocky and up to 5 feet in section. It is reduced to 9-in. size by two 60-in. gyratory crushers, and transported to the east and west ore bins by 60-in. conveyors. In the secondary crushing operation the ore is first reduced to 2 to 3-in. size by five standard 7-ft cone crushers and two No. 10 gyratories. This product is then reduced to 3/8-in. size by fifty-two 48-in. vertical Symons disc crushers and is transported by belt conveyors to the loading bridges from which it is charged° into the leaching vats. Conveyor belt No. 14 which serves the eight vats in the east section of the leaching plant has transported over 125 million tons of ore, which is a record for any type of conveyor. This 60-in. belt travels at a speed of 600 fpm and carries loads up to 5500 tons per hr. Leaching The ore is leached in 14 vats, averaging 13,500 short tons net capacity each. These vats are lined with a natural mastic which adequately resists the sulphuric acid leaching solutions. The leaching process as presently conducted comprises soaking the ore with successive volumes of leaching solution and spent electrolyte and withdrawing successive volumes of enriched solution from the vat to a total of approximately 3.2 million gal. The first 1.6 million gal of this volume withdrawn is called strong solution and contains about 35 parts of copper per 1000. The balance of 1.6 million gal subsequently withdrawn is stored in a

Jan 1, 1952

By B. Russell, J. R. Tuffley

The stability regions of the normal sulfate and the various basic sulfates of lead in 02-SO2 and PhS-SO2 gas atmospheres were calculated from available thermodynamic data over the temperature range 600o to 1200oC. Roasting experiments were carried out on samples of synthetic lead sulfide at various temperatures and for various roasting times. Compounds in the oxidized surface layers were identified by X-my techniques. The order in which the various reactions occurred was determined by considering the experimental results in conjunction with the thermodynamic calculations. The results of these investigations were applied to the sintering of lead sulfide concentrates, and the extent to which each of the various reactions occurred in a sinter bed was determined. PRODUCTS which may be formed during the roasting of PbS are PbS04, PbSO4. PbO, PbSO4 . 2PbO, PbSO4 . 4PbO, PbO, and Pb.1, 2 The formation of some sulfates has been studied by a number of investigators3-&apos; whose experiments have mostly been confined to the temperature range between 500" and 850°C. Culver, Gray, and spooner3 were able to isolate the reactions which formed PbSO4 and PbSO4. PbO by carefully controlling the compositions of the roasting gases. The kinetics of these reactions were studied within the temperature range 673" to 807°C. Gardner4 and Merrick5 also studied the kinetics of oxidation of PbS. Gardner disregarded the existence of basic sulfates and, as a result, misinterpreted some of his experimental data. Merrick recognized that most of the products of oxidation were mixtures of basic sulfates, but was unable to identify the individual components. Hoschek6 oxidized PbS in air at temperatures up to 800°C. The products were analyzed after roasting had proceeded for some minutes, but no attempt was made to follow the reactions in the initial stages. Consequently, no roasting mechanism was derived. He also showed that PbSO4 and PbSO4 - 4PbO decomposed slowly in air at temperatures in the region of 800°C. The stability regions of sulfates in the temperature range 500" to 850°C have been only partially determined, while practically no information is available for higher temperatures. In the lead sintering process, PbS concentrates are roasted at temperatures within the range 1000o to 1200°C. In order to investigate sulfate formation, it is necessary to determine the effects of temperature and gas composition upon the various oxidation reactions and upon the regions of stability of the various lead sulfates. Kellogg and Basu2 introduced an equilibrium diagram similar to those in Figs. 1 to 3 which provides a convenient medium for represent-

Jan 1, 1964