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[Adams, Douglas K General Electric Co., P.O. Box 2331; Denver 1, Colo. Allan, Andrew Thompson Creek Coal & Coke Corp., Glenwood Springs, Colo. Allen, Earl W Mine Safety-Appliances Co.; 257 Rio Grand St., Salt Lake City, Utah Alston, George L Mine Safety Appliances Co., 201 N. Braddock, Pittsburgh 8, Pa. Anderson, Nephi W Standard Colombine Coal Co., Scofield, Utah Anderson, W. J Texaco, Inc., 845 N. 5th E., Price, Utah Ankeny, Marling J U.S. Bureau of Mines, 1822 Metzerott Rd., Adelphi, Md. Archer, Bryon J., Jr Peabody Coal Co., 421 Glenwood St., Grand Junction, Colo. Arentzen, E. M Lee-Norse Co., P.O. Box 582, Charleroi, Pa. Arentzen, L. E Lee-Norse Co., 6030 W. 55th P1., Arvada, Colo. Arentzen, Rolf F Lee-Norse Co., Box 582, Charleroi, Pa. Artis, Cliff E Big Horn Coal Co., 247 Sherman Ave., Sheridan, Wyo. Atchison, Lee Atchison Construction Co., Drawer D, Aztec, N. Mex.]

Jan 1, 1965

By R. F. Sheldon

The discovery of the Carlin deposit was the result of discriminating geologic research and prospecting devoted to the objective of finding a gold deposit that could be mined by open pit methods. By the late 1950s, Newmont Mining Corp. was becoming increasingly concerned about the trend of rapidly rising underground mining costs unless bulk mining methods could be applied. The desirability of finding an open-pittable gold deposit was apparent. The attention of Newmont geologists was directed to Nevada by the publication of two papers: "Paleozoic rocks in North-central Nevada" (1958) and "Alinements of Mining Districts in North-central Nevada" (1960), both authored or co-authored by Ralph J. Roberts of the United States Geological Survey. John Livermore, geologist of Newmont Exploration Ltd., attended a talk given by Roberts in Ely, Nevada in the summer of 1961. These papers and the address by Roberts gave details concerning the Roberts Mountain overthrust, a 483 km (300-mile) long, shallowdipping fault which pushed clastic and volcanic rocks of early and middle Paleozoic age eastward over younger marine formations. Subsequent uplift and doming with consequent erosion had locally removed the upper plate strata exposing windows of lower plate carbonate rocks. Roberts noted that the principal mineral deposits of the region, including gold, were associated with these windows which exhibited a preferred alinement. In the summer and fall of 1961 Newmont's exploration geologists John Livermore and J. Alan Coope began a systematic examination of gold occurrences associated with these windows, aided by further discussion with geologists of the US Geological Survey. Attention was directed to the Lynn and Carlin windows, particularly to those formations lying immediately above and below the Roberts Mountain fault. It was appreciated at the time that many of the gold occurrences in the region were unusual in that no colours were obtained even when panning high grade samples. A large number of rock samples were collected and analysed using fire assaying procedures, and background values for gold were established. This systematic rock sampling resulted in the identification of a distinct area of anomalous gold values, and a block of claims was staked in late October, 1961. These claims, plus an adjoining optioned 32.4 ha (80 acres) of ground, cover the main area of the present Carlin gold mine. Prior to snowfall that winter, one of the bulldozed assessment pits required to establish a claim's discovery exposed 24 m (80 ft) of mineralization assaying 0.007 kg/t (0.22 oz per st) gold. In the spring of 1962 a program of trenching, sampling, and geological mapping followed by rotary drilling was underway. Other properties in the area were acquired. The generally undistinguished nature of the dolomitic siltstones and silty dolomitic limestones hosting the micron-sized gold particles, coupled with the lack of visually associated guide minerals, made identification of the gold bearing areas very difficult. The entire drill column had to be assayed to ensure that values were not overlooked, as very often sections that might be assumed to be waste turned out to be high grade. On September 10, 1962 a high grade intersection of 24 m (80 ft) assaying over 0.03 kg/t (1 oz per st) gold was encountered in the third hole drilled. An expanded program of both rotary and diamond drilling led to the further delineation of the ore body. By December 1963, the exploration program had established an initial reserve of 10 Mt (11 million st) grading 0.01 kg/t (0.32 oz per st) gold. A 1.8 kt/d (2,000 stpd) processing plant was constructed and the first gold bullion was poured in May, 1965, just two years and eight months after the discovery hole was drilled. REFERENCES Roberts, R.J., Holz, P.E., Gilluly, J., and Ferguson, H.G., 1958, "Paleozoic Rocks in North-central Nevada," Bulletin American Association of Petroleum Geologists, Vol. 42, No. 12. Roberts, R.J., 1960, "Alinements of Mining Districts in North-central Nevada," Professional Paper 400-B, US Geological Survey, Article 9.

Jan 1, 1985

Mineral sands are different to almost any other commodity; however, it shares similarities with other commodity types, such as the importance of quality constraints of iron ore and coal or the importance of physical properties of diamonds. The mineral sands industry consists of two principal product streams; titanium dioxide minerals û in the form of rutile, ilmenite and leucoxene, and zircon. The principal valuable heavy minerals (VHM) include ilmenite, leucoxene, rutile and zircon. Variations of other titanium minerals occur between the end members of ilmenite and rutile, including pseudo rutile and anatase. Most mineral sands deposits are found in unconsolidated fossil shorelines several hundreds of metres to tens of kilometres and occasionally hundreds of kilometres inland from the present coastline. Mineral sands orebodies essentially fall into two categories based on the mode of deposition: alluvial or aeolian. Alluvial deposits are further split into marine beach placers (or strandlines) and lacustrine heavy mineral (HM) accumulations. Exploration for mineral sands involves the positive identification of key criteria leading to the focus of exploratory surface sampling, augering and drilling. Assaying is primarily focused around determining the percentage of HM contained within a given sample. Other results of interest include clay fines, sand and oversize. Metallurgical/mineralogical assessment is often undertaken via laboratory scale bench tests that replicate the wet concentration and dry mill processing routes. The most critical component in resource assessment for mineral sands is about quantifying HM grade, then mineralogical assemblage and then quality of those mineral species. This will determine whether a mineral sand final product is marketable or not. Mining of mineral sands is conducted either wet or dry. Wet methods are generally preferred for large tonnage, unconsolidated and low clay orebodies. Where ground conditions are hard and orebodies are small, high grade and discontinuous, dry mining techniques are generally employed. Concentration of mineral sands from the primary ore is carried out in two sections; wet, utilising sizing and gravity differentiation between HM, VHM, clay and quartz, and dry, exploiting the magnetic, electrostatic and to a lesser extent specific gravity (SG) properties of the minerals of interest. Mineral sands exploration, mining and processing faces the same operating challenges as the rest of the resource sector. Added to this are the issues that are unique to mineral sands and to which the industry devotes considerable resources to developing solutions.

Jan 1, 2009

By E. A. Loria

The Bureau of Mines analyzed the effects of oxygen, nitrogen, and carbon on the tensile properties of electrorefined vanadium in terms of the Petch equation. The effects of the different impurities and grain sizes on the yield and flow stress between 2730 and 770 K were studied. For specimens of like composition, the values of the Petch parameters 01 and ky obtained by two different methods, extrapolation and grain size, were nearly the same. For high-purity electrorefined vanadium, the 01 value at 2730 K was 10,300 psi (5,150 psi shear). The 01 values for this metal at 1950 and 2730 K were lower by factors of 3 to 4 than for three less pure samples of commercial vanadium metal of equivalent grain size. Nitrogen was more effective than oxygen in increasing 01' the increase being about twice as much for the same impurity level. Carbon produced a slight increase in 01 up to 0.024 percent and then a gradual decrease in 01 to 0.11 percent, the maximum carbon content studied. The corresponding ky values followed the same trend in regard to oxygen and nitrogen additions, whereas the addition of carbon produced no variations at either temperature. Dislocations in vanadium are anchored by carbon, oxygen, and nitrogen with increasing firmness in that order. The high purity of electrorefined vanadium was confirmed by low yield stress and high ductility throughout the low temperature range. The metal also exhibited profuse twinning with high ductility at 1230 and 770 K. The effective surface energy for fracture, y, was found to lie between 3.3 and 3.6 x 1()4 erg cm-2?

Jan 1, 1966

By Murray Malott

The Sheppard Subway extension is currently being constructed in Toronto, Canada for the Toronto Transit Commission (TTC). McNally International Inc., in a joint venture with PCL and the Foundation Company of Canada were awarded the Twin Tunnels contract of the Sheppard extension. The Twin Tunnels project is approximately 4.4 km in length and consists of two separate twin bore drives each 3.3 km and 650 m in length and approximately 400 m of cut and cover construction. This paper presents a case history of the construction of the soft ground tunnels through the complex Toronto glacial deposits using two 5.9 m outside diameter Lovat earth pressure balance (EPB) tunnel boring machines.

Jan 1, 1999

By Michael P. Davies

Mine waste facilities, and in particular tailings impoundments, are one of the highest risk aspects of any mining venture. There are far too many recent examples of how real and/or perceived physical or environmental impact from one of these facilities has had enormous economic impact on both the specific project and the overall parent mining company. In many cases, such dramatic impacts have followed a period where a due diligence of the mine had been carried out (e.g. in an acquisition, merger or similar) or an internal or otherwise audit had taken place. Consequently, it is suggested that there is a need for an audit/due diligence framework to assist those processes in terms of assessing mine waste facilities. This paper presents a recommended framework for completing either an audit or due diligence assessment of mine waste facilities under a number of typical scenarios. The framework will be of relevance to reviews carried out prior to mine construction, during operations and for mines that have been in a care and maintenance position and are to be re-opened. Concepts of risk management are introduced as an essential part of the recommended framework. The paper should be of value to current and prospective mine owners involved in stewarding an area of significant risk to viability of their mining venture. The paper should also interest prospective lenders to these ventures.

May 1, 2007

By E. T. Turkdogan, P. Grieveson

Experimental results are given for the rate 0.f solution of nitrogen in y iron in the temperature range 1000° to 1200°C. It is shown that, when purified reacting gas is used, the rate-controlling process is the diffusion of nitrogen into the iron. The difhsiuity of nitrogen in y iron is calculated from the results and the equation logD = —(8800/T) — 0.04 is recommended to represent the temperature dependence of the diffusivity. In some experiments using unpurified nitrogen, solution rates are found to be slower than those attributable to a solely solid-state diffusion process. Indications are that in the presence of impurities in the gas, e.g., H,O, chemisorption of nitrogen at the surface is hindered and becomes a rate-controlling process for the transfer of nitrogen across the gas -metal interface. ALTHOUGH the solubility of nitrogen in y iron at known nitrogen activities has been well-established,'-a the kinetics of the solution of nitrogen in iron are not known. The work of Darken, Smith, and Filer' indicates that at temperatures below about 1300°C the rate of solution of nitrogen in y iron is controlled by a chemical reaction at the metal surface, while at temperatures above about 1300°C the rate-controlling process is the diffusion of nitrogen in the metal. Fast and verrijps studied the denitro-genization of y iron using wet hydrogen at 950°C and their results clearly indicate that the rate of nitrogen removal is controlled by diffusion of nitrogen in the metal. In Part I of this series of papers, the kinetics of the solution of nitrogen in y iron are discussed in the light of the present experimental work. EXPERIMENTAL Apparatus. The reaction chamber consisted of a platinum wire-wound resistance furnace with a re-crystallized alumina reaction tube. The temoerature of the furnace, which was automatically controlled to approximately 1°C, was measured with a Pt/Pt-13 pct Rh thermocouple and a potentiometer. The temperature distribution in the working zone of the furnace was determined using a calibrated Pt/Pt-13 pct Rh thermocouple and the hot zone was found to be approximately 3 in. long with a temperature variation of * 3°C. A series of experiments was also carried out at 1000°C and pressures above atmospheric using a pressure vessel which was similar in construction to that described elsewhere.1° Materials. Ferrovac E grade iron used in all ex-

Jan 1, 1964

By L. E. Schultze

As part of the Bureau of Mines research program on recovering alumina from domestic resources, methods for extracting iron from aluminum sulfate solutions (obtained by leaching kaolinitic clay with dilute sulfuric ,acid) were investigated. In countercurrent extraction experiments at a l-gal/hr flow rate of aluminum sulfate leach liquor, the iron concentration 'was decreased from 1.4 g/l to 0.01 g/l after four stages of using a primary amine extractant. The iron-loaded amine solution was regenerated in five stages using 3[N] sulfuric acid or in four stages using 6[N] sulfuric acid. The aluminum sulfate solution obtained was suitable for the preparation of alumina meeting cell-grade standards of less than 0.01 pct iron.

Jan 1, 1979

By M. C. Fuerstenau

The zero-points-of-charge of apatite, calcite, and fluorite are pH 6.4, 10.8, and 10.0, respectively. Scheelite is negatively charged above at least pH 3. Flotation responses of these minerals in the presence of potassium oleate and sodium silicate are presented and compared with electrokinetic data. Colloidal silica appears to be the species principally responsible for calcite depression, while silicate anion is the species responsible for fluorite depression. Additions as high as 1 x 10-3 mole/liter silicate have no effect on the flotation responses of apatite and scheelite. Selective flotation of nonmetallic minerals is difficult to achieve with fatty acids or soaps by themselves. As a result, specific reagents are added to aid these separations, and one of the reagents commonly employed for this purpose is sodium silicate. Flotation separations of various calcium¬ bearing minerals such as fluorite from calcite (1,2,3) and scheelite from calcite(24) and apatite(2,5) almost always involve the use of sodium silicate, for example. The mechanisms by which sodium silicate functions as a depressant are still not understood; this fact is probably due to a number of reasons. That is, the dissolution process of sodium silicate is complex, giving rise to a number of ionic and colloidal species. (6) In addition, the type and concentration of these species depend on the ratio of Na20 to Si02, the concentration of sodium silicate, and the pH of the system. (6) At the present time, it is not known which species, colloidal silica or silicate anion, is responsible for depression. If colloidal silica is the species that is adsorbing, then adsorption must occur by electrostatic attraction between the colloid and the mineral surface. Silicate anion, on the other hand, may adsorb either physically or chemically. The objective of this paper is to determine first the active species of sodium silicate and then the conditions under which this species will adsorb and function as a depressant.

Jan 1, 1968

By F. Shigeyoshi

The development of pelletizing during the last ten years has continued, however, self-fluxing pellets have not been regularly and commercially produced for use in blast furnaces, except as Kobe Steel, Ltd., Japan. This paper tells about unique developments for self-fluxing pellets, and their effects when used in Kobe Steel's No. 3 blast furnace at the Nadahama Works (inner volume 1845 Nm3, hearth diameter 9.5 m)., Several phases of the pelletizing process are discussed individually, and also their relation to performance in blast furnace operation. Special emphasis is placed on proper preparation of the limestone added to the raw materials, and indurating at optimum conditions. Means of improving pellet quality by correct control of size distribution of the raw materials used in pelletizing, and indurating temperatures, are also discussed. As basicity (CaO/SiO2) of pellets increased 0. 1, coke consumption decreased 3. 8 to 4. 2 kg per ton of hot metal, when using 70% pellets in the iron burden.

Jan 1, 1969

By Robin J. Hickson

The Grasberg copper-gold deposit was discovered in the spring of 1988, one year before the planned closure of Freeport Indonesia's original Ertsberg openpit copper mine. In spite of Grasberg's remote location, 600 m (1968 ft) above the Ertsberg, atop a hitherto inaccessible mountain, Freeport-McMoRan brought its latest open-pit mine into production in the last week of 1989, less than two years after discovery. Freeport Indonesia's (FI) copper-gold mining operations are located in Irian Jaya, the easternmost province of Indonesia on the western side of the island of New Guinea. Fl's six ore deposits are all located within a 10- x 10-km' (3.9- x 3.9-sq mile) Contract of Work area. They are all between the elevations of 2900 and 4300 m (9515 and 14,100 ft) above sea level, some 96 km (60 miles) from the island's southern coast. A 118-km (74-mile) access route connects FT's port site of Amamapare on the Arafura Sea to the project area. The first section is an 18-km (11-mile) barge waterway. The final 100-km (63mile) is a two-lane gravel highway. The highway terminates 9 km (5.5 miles) beyond the FI townsite at the project's central mill, some 2850 m (9250 ft) above sea level. The mining areas are separated from the mill by a 780-m (2560-ft) sheer cliff face. The world's longest free span tramway system, with a total length of more than 1.6 km (1 mile), is the only means of access for men and materials between the mill and the surface mines. Weight restrictions for the tramway for equipment considered for the project are 15.4 t (17 st). The minimum clear width dimension of two tunnels along the portsite access road is 5.6 m (18.4 ft).

Jan 1, 1991

By E. H. Burdick

THE difficulties inherent in table concentration operations as applied to gold, silver, lead and zinc ores, are accentuated in the scheelite mill, which has a flowsheet that is similar in general principles. The narrow "spread" in the specific gravity between scheelite and the usual gangue in which it occurs often makes its clean separation extremely difficult and in some cases impracticable. The problem is further complicated by the extreme friability of the mineral and its inordinate tendency to "slime," particularly where fine grinding is necessary and wet grinding methods are employed. The physical aspect of the occurrence of scheelite in the ores, the kind and character of the rock in which the mineral is present, its hardness, friability and the specific gravity of its constituents, as well as like characteristics of the associated minerals-pyrite, pyrrhotite, sphalerite, and others, all have a bearing on the successful milling of crude ores and the production of a concentrate that is marketable with minimum price penalty or none at all. Most tungsten deposits are considered low grade when the limiting factor of low percentage recovery by gravity methods is borne in mind. The great majority of deposits developed or in process of development carry from 0.4 to 1.0 per cent tungstic oxide. Ore carrying I per cent is considered good grade; 0.7 to 0.8 per cent, fair, and below that figure, low grade. Steady mill feed averaging upward of 1.0 per cent W03 is exceptional. Inasmuch as buying specifications demand a 60 per cent W03 content, the mineral content of the ores as mined must be concentrated to at least that degree. A few companies have constructed chemical plants for the treatment of their complex concentrates and low-grade flotation product, but the smaller operator must rely on table concentration, borrowing the tools and to a large extent the methods of his predecessors in the gold, silver, lead and zinc mills. Whether these are the best tools and methods that can be devised is an open question. Except in rare instances, table concentration, however carefully conducted, does not yield a satisfactory final recovery of the mineral content in headings as indicated by chemical determinations of tungstic oxide content. A saving of 60 to 65 per cent is considered fair, and an average 80 per cent recovery, exceptional. Each scheelite ore body is a separate problem. Details of handling that give best results in one case may not be successfully operative at a property on the same contact zone and but a short distance removed. Supplementary treatment of tailings by flotation may be successful in producing a low-grade concentrate of around 15 per cent WO3, which is adapted to further treatment chemically at a given property and entirely unsuccessful at another. This problem constitutes another of the unanswered questions. The following paragraphs relate solely to the beneficiation of scheelite ores by table concentration and supplementary methods.

Jan 1, 1942

By J. Norman Bulkley

As electrical power is used to a greater extent on the Rand than in any other mining center, it is thought that a short description of the methods used and results obtained may be of interest. In comparing Rand practice with that on other fields, several general factors should be borne in mind: First, the mines are all working on large tonnages of low-grade ore (average $6.50). The annual crushing rate is about 27,000,000 tons. Second, the mines are all grouped under the financial and engineering control of mining houses, so that the results of a group, rather than an individual mine, are open for study, and practice follows more uniform lines. Third, coal of fair quality (about 12,000 B.t.u.) can be obtained for $2 to $3 per ton. Good supplies of condensing water are, however, scarce, and have to be carefully stored. Power Supply Power is supplied by the Victoria Falls & Transvaal Power Co., Ltd., and its subsidiary, the Rand "Mincs Power Supply Co. Both companies, which are practically one from an engineering point of view, operate the following stations, all fecding into the same system. Capacity Installed Turbo Capacity Installed Steam-Tur- Station Alternators bine Dliven Centrifugal Compressors Brakpan............... Two 12,500-kw. sets Two 3,000-kw. sets Simmer Pan............... Six 3,000-kw. sets Rosherville................ Two 11,000-kw. sets Six 3,500-kw. machines Five 9,600-kw. sets Three 7,000-kw. machines Verckniging ............. Two 9,600-kw. sets Two 12,000-kw. sets 162,200-kw. sets Total steam-tkbine capacity installed, 204,200 kw. There are also at the Robinson Central air station six motor-driven centrifugal compressors, each of 3,500 kw., equal to 22,000 cu. ft. per minute at 120 Ib. capacity at 3,000 r.p.m. Power is supplied from these stations as 50-cycle, three-phase current generally at 40,000 volts,

Jan 1, 1916

By G. Wakefield, A. H. Daane, D. H. Dennison, F. H. Spedding

Preparation of pure scandium metal was accomplished by calcium reduction of the fluoride by two methods; a low-temperatzdre alloy process and direct reduction with subsequent distillation of the product. The following properties were determined: melting point: 1181OK; boiling point (calculated): 3000°K; lattice constants at 298°K (hexagonal lattice): a = 3.308 * 0.001 A, c = 5.267 * 0.003A; calculated density at 298°K, g per cm3: 2,990 + 0.007; electrical resistivity, ohm-cm: 299°K, 66.6 ± 0.2 x 10 -6; 373oK, 77.4 * 0.2 x 10 -6; thermal coefficient at 299°K, ohm-cm per deg: 5.4 X x; heat of sublimation at 298°K, kcal pel- mole: 80.79. The vapor pressure was determined as a function of temperature between 1505o and 1748°K, with the data fitted to a straight line shielding the equation: Log Pmm = -1.718 X 104/ToK + 8.298. SCANDIUM, element number 21, was first discovered by Nilson in 1879 and was recognized as the Ekaboron as predicted by Mendeleff. As it is in group III of the periodic table, the general properties are a little like aluminum and also resemble quite closely the properties of yttrium and the rare-earth metals, in both the metallic and ionic form. Although the earth's crust contains approximately 5 ppm of scandium (the element is as abundant as arsenic and twice as abundant as boron) it generally occurs so widely distributed that it has earned the reputation of being very rare. The one exception to this is the mineral thortveitite, which has been found in Madagascar (20 pct Sc2O3) and in Norway (35 pct Sc2O3). Scandium also occurs in small but distinct amounts in uranium and rare-earth ores; the recent larger scale processing of these materials has made some scandium available from these sources. As with other naturally-occurring monoisotopic elements (except Be), scandium contains an odd number of protons and an even number of neutrons. Scandium metal was first prepared by Fischer and coworkers1 in 1937 by electrolysis of scandium chloride in a molten eutectic mixture of lithium and potassium chlorides, using molten zinc as a cathode and collector of the scandium metal produced. The zinc was removed from the Zn-2 pct Sc alloy by vacuum distillation, leaving a product reported to be 94 to 98 pct Sc, with the main impurities being iron and silicon. They reported a melting point of 1400° C for this material. Scandium has also been prepared by the reduction of scandium chloride with potassium metal in a glass apparatus by Bommer and Hohmann in 1941,' resulting in a mixture of metal and potassium chloride; these workers did not isolate the metal proper, but the X-ray diffraction of the slag-metal mixture showed it to be hexagonal with a = 3.30A, c = 5.45A. petru3, 4 and coworkers have recently reported the preparation of the metal in a compact form by the reduction of either ScF3 or ScC13 with calcium metal and subsequent distillation of the product. This process probably yielded a metal of high purity, but they list no chemical analysis nor do they list any of the properties of their product. Previous related work in this Laboratory has been concerned with the production of yttrium and the rare-earth metals and the determination of their physical properties. Because of its similarity to these metals, scandium is being included in this study. PREPARATION OF SCANDIUM METAL The preparation of yttrium and the rare-earth metals may be accomplished by reduction of their fluorides with calcium metal in tantalum crucibles.5 This process leads to the introduction of tantalum (up to 0.5 pct) as an impurity in the higher melting rare earths, but since the tantalum occurs as dendrites, uncombined with the rare-earth metals, its presence is not objectionable in some cases. The preparation of scandium metal in this manner, however, was found to yield a product containing 2 to 5 pct Ta. To obtain a purer product, the following two methods were developed for the reduction of scandium fluoride with calcium metal: i) a low-temperature process utilizing zinc to form a low

Jan 1, 1961

By H. E. Sadler, Prof. Silliman

The research here recorded was undertaken early in the present year, and has been prosecuted steadily for about eight months. While the work has been under my constant supervision and advice, the labor has all been performed by Mr. Sadler, to whom is chiefly due the development of the successive steps which have led to the final result. The process is better than I dared to hope, and we are able to place at the command of chemists and engineers interested in the matter an apparatus of continuous and uninterrupted action, and methods of analysis by which the daily averages of the two variable and inconstant factors of illuminating gas, ammonia and sulphur, may be determined with all needful accuracy, even by those little skilled in chemical manipulation, and with very little loss of time. This research has been carried out in the laboratory of the New Haven Gaslight Company.

By O. W. Shimer, D. C. Helms, C. E. Brown

The early history and progress of anthracite stripping, from the first known operation at Summit Hill in 1821 through 1917, was covered in 1917 in a paper by J. B. Warriner,1 then chief engineer, now president of the Lehigh Navigation Coal Co., and stripping methods between 1917 and 1931 were the subject of a paper in 1931 by H. H. Otto,2 of the Hudson Coal Company. The present paper is confined to methods used today on pitching beds in the western middle and southern anthracite fields. The anthracite industry, not unlike other large industries, had been affected for more than a decade by the adverse economic conditions of the country. High development and operating costs and the competition of other fuels caused some collieries to be closed temporarily, while others were permanently abandoned. In order to obtain a larger tonnage at a lower development cost, a larger stripping program was put into effect by many companies. In 1932, stripping operations accounted for 3½ million tons, or 7 per cent of the total anthracite production of the year, and in the next 10 years this increased until more than double that amount, or over 14 per cent of the total production, was from stripping operations. These figures represent the total production of the entire anthracite region; unfortunately, the figures have not been separated from the different fields. There are some operations in the southern and middle western fields where the tonnage obtained from stripping amounts to more than 30 per cent of the total colliery production. Stripping operations on pitching beds present difficult problems. Owing to the various geological formations, a different situation is encountered in nearly every stripping pit; and because of the proximity of an overlying bed, it is very often necessary to load the entire overburden into trucks and haul it out of the pit to the dump area, although back filling and casting is done whenever possible. Beds in the areas described in this paper pitch from 20" to 90° and some are on the invert in many sharp folds and faults. Planning The more successful strippings have had adequate engineering analysis. Some work is being done without much preliminary planning, but many strippings would have failed completely if close supervision by the engineering department had not been maintained. Locations of stripping areas are determined as follows: I. By constructing cross sections over the areas and drilling sufficient diamond-drill or well drill holes to study the strata and to determine the location, thickness and pitch of the beds.

Jan 1, 1946

By R. S. Dean

[During t!~c past f isc.~l y~:-.r the Xetzllvz-gicd 3ivi oion has made im- portant contribution-, don,- tlie lines rlictnteil by the Bureau policy of con- servation of domestic rnker~l rcsoul-ces throui;h use. During t!~c past f isc.~l y~:-.r the Xetzllvz-gicd 3ivi oion has made im- portant contribution-, don,- tlie lines rlictnteil by the Bureau policy of con- servation of domestic rnker~l rcsoul-ces throui;h use.]

Jan 1, 1940

By A. Uribe-Salas, J. A. Finch, B. J. Huls

Gas holdup profiles in the collection zones of a pilot flotation column (0.91 m x 13.5 m) and a laboratory flotation column (0.057 m x 9.5 m) operated with air and water only were determined from pressure measurements. The profiles were determined discretely and continuously. For the pilot column, a minimum distance between pressure measurements was necessary to give reliable results. For the laboratory column, a third-degree polynomial fit to the pressure vs. height data was required. Water manometers proved to be less sensitive to sensor and process noise than pressure transducers. The profiles showed that gas holdup approximately doubled over a height of 8 to 10 m. The profiles were not linear, and the gradient increased towards the top of the column, especially for gas rates greater than 1 cm/s. The profiles cannot be predicted by a simple bubble expansion model. The implications of this new finding of a significant gas holdup increase with height in the collection zone on various aspects of flotation column design and operation are discussed.

Jan 1, 1996

By Raghavan Srinivasan, Y. V. R. K. Prasad, William G. Frazier, Steve C. Medeiros

"Aerospace gas turbine disks operate in an environment of relatively high stresses caused by centrifugal forces and elevated temperatures. These severe conditions necessitate the need for materials with good high temperature strength and low cycle fatigue resistance. One class of alloys used for this task is the nickel base superalloys, out of which, lN 718 is the most widely used in the aerospace industry [I]. The properties of IN 718 are attributed to the combined effects of the chemistry, heat treatment, and microstructure. The chemistry is tailored not only for solid solution strengthening but also for precipitation hardening developed during heat treatment, which combined with a fine grained microstructure lead to excellent mechanical properties such as low cycle fatigue resistance and elevated temperature strength. The properties of a gas turbine disk are sensitive to the microstructure, in particular the grain size, which is dependent on the processing history. The ability to precisely control the microstructural development during forging is dependent on controlling the process so that the workpiece is deformed within a ""safe"" region where no microstructural damage or flow instabilities occur. The microstructural mechanisms during deformation may themselves vary within the ""safe"" region and it is desirable to determine them within the range of parameters that are commonly used in industrial processing. The objective of this work is to establish a relationship between the grain size and the process control parameters i.e., temperature and strain rate, in the hot working of IN 178.BackgroundIN 718 belongs to a class of nickel-iron base superalloys with a face-centered cubic (FCC) austenitic matrix strengthened by precipitation of ordered intermetallic or carbide precipitates and is a nickel-rich alloy strengthened by ordered body centered tetragonal (BCT) precipitates, Y"". The alloying of nickel and iron results in the formation of the austenitic matrix, y, while the addition of chromium and molybdenum cause solid solution strengthening. Also, alloying with niobium causes precipitation of the metastable hardening constituent Y"" (Ni3(Nb,Ti)). Extended aging treatments in the temperature range 650 °C - 980 °C allow the more stable orthorhombic 8 phase (Ni3Nb) to form. Trace amounts of boron are also added to serve as grain boundary strengtheners for high temperature creep resistance. Hot deformation of IN 718 is performed between the Y"" and 8 solvus temperatures to facilitate recrystallization and to control grain growth. While deformation below the Y"" solvus (""""962°C) produces no change in grain size, deformation above the Y"" solvus but below the 8 solvus will lead to a dynamically recrystallized microstructure. The presence of the 8 precipitates will pin the grain boundaries after recrystallization and restrict grain growth so that a fine grain microstructure is obtained. Deformation above the 8 solvus ("""" 1038°C) will lead to a microstructure which has undergone grain growth during deformation."

Jan 1, 2000

By E. H. Rose

This paper, describing the concentrator practice of the Moctezuma Copper Co., is one of a series being prepared by the Bureau of Mines. The Nacozari concentrator of the Moctezuma Copper Co. (the Mexican subsidiary of the Phelps Dodge Corporation) is located about 75 miles south of Douglas, Ariz., in the Moctezuma District of the State of Sonora. Nacozari is the southern terminus of the Nacozari Railroad, ever which all concentrate is transported to Douglas for smelting.

Jan 1, 1930