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By A. Mular, T. Bond, G. D. Farnell, J. F. Kingston

"ABSTRACT:-A 60—cm semiautogenous grinding-(SAG) mill was designed -for continuous wet operation in closed circuit with a vibrating screen. The circuit is fully instrumented and automated. Part of the control system incorporates Foxboro SPEC 200 controllers around the pump box to maintain level. The rest of the control system involves an ACTION model BC2 basic direct digital controller that handles all other inputs/outputs and contact closure sensing:·The success of the control system is attributed to the uniqueness of specially designed mass flow metering systems. Various control strategies are possible. Employed to date has been control of a mill load set-point by cascading the deviation from load set-point as the set-point to the fresh feed rate ""controller"". Mill feed water is ratioed to total flow rate of solids fed to the mill. The performance of the circuit is being correlated with the performance of full scale SAG circuits in British Columbia to develop models, to assess the feasibility of scale up and to investigate the effects of design variables on responses such as circuit throughput. 1 INTRODUCTIONWhere ore types are appropriate, semiautogenous grinding and/or autogenous grinding (AG) has replaced conventional grinding. Secondary/Tertiary crushing plants are thereby eliminated so that capital expenditures are less and overall comminution costs can be lower (Barratt 1982). To establish the suitability of an ore for SAG circuits, pilot scale tests with 1.7 to 1.8 M diameter mills are often conducted. At least one test per ore type in a deposit has been necessary. Often, each test involves a large amount (up to 50 tonne) of sample, so that the testwork can be expensive and may discourage the use of SAG or AG circuits. However, MacPherson (1977) has employed a continuous 45 cm dry Aerofall mill circuit closed with screens and an air classifier to determine whether ores are suitable for SAG milling. The amount of sample per ore type is of the order of 60 kg or less. Thus an established SAG mill scale up procedure that employs small circuits is highly desirable.The purpose of this paper is to describe a small 60 cm wet continuous SAG/ AG mill pilot circuit assembled at the University of British Columbia to study the feasibility of scale up, to assess the influence of design variables, to develop mathematical models and to evaluate the effects of alternative control strategies on circuit performance"

Jan 1, 1988

By H. W. Mead, C. E. Birchenall

SELF-DIFFUSION of iron in austenite is a process which may play a significant role in some of the practically important reactions which occur in solid irons and steels. It also provides a system in which the interaction of vacancies on substitutional sites and atoms in interstitial positions may be investigated. For both purposes, accurate values of the diffusivities are required. Two sets of previous measurements exist, each containing features which require checking. In the course of an investigation to determine the rate of diffusion of iron in certain steels, Garwick and Rosenqvist' measured the self-diffusion coefficient of iron in a 0.10 pct* C steel and obtained a dicated that while at lower temperatures D increased with increasing carbon content, at the higher temperatures it first increased and then decreased. At the same time the activation energy decreased linearly from a value of 69,000 cal per g-atom for pure iron to 33,000 cal per g-atom at 1.06 pct C. This rapid decrease in activation energy implies that iron at higher temperatures within the austenite range would diffuse more slowly in high carbon steels than in steels of lower carbon content. Experimental Procedure The radioactive iron used in this work was obtained from Oak Ridge National Laboratory. The unit consisted of about 1 microcurie each of Fe7 and Fe It was allowed to decay for two years. At the end of this time most of the Fe with a half-life of 46.3 days, had decayed, while most of the original Fe(half-life 2.91 years) still remained. The iron as chloride was extracted into an ether layer, using the method of Ashley and M~rray. After evaporation of the ether the iron chloride was dissolved in 250 ml of aqueous solution containing 2.4 gpl Fe. The compositions of the steels used are listed in Table I. Disks of ¾ in. diam and about ¼ in. thick were machined from these steels for use in diffusion couples. The disks were surface ground on both sides to ensure parallel opposite faces. One face of each disk was polished through 000 emery paper. Half of the prepared disks were plated with radioactive iron. For this, 1 ml of the radioactive iron

Jan 1, 1957

By O. L. Kachor, A. V. Bogdanov

"In 2015 in the frames of federal special-purpose program «National system of chemical and biosafety of Russian Federation (2009-2014)» Federal State Budget Educational Institution of Higher Education ""Irkutsk National Research Technical University” (further–INRTU) performed works aimed at disposal of accumulated arsenical residues at industrial site of the former Angarsk arsenical plant (AAP) processing arsenious ores from Svirsk town in Irkutsk region. The plant in 1934-1949 produced arsenic trioxide for the defence industry. Prior to starting the works on disposal of industrial pollution focus the total amount of cinders stored on 13 ha of industrial site of AAP was 156 900 tons with 1% average arsenic content. Close to the cinders there were also remains of industrial buildings and process equipment in the amount of 6 600 m3 that contained 150 tons of arsenic. Additional source of pollution at the industrial site of AAP is contaminated soil amounted to 48 000 tons. Proven reserves of gold and silver in the cinders were 6.1 and 47 g/t respectively. In 2015 the source of arsenious pollution was detoxificated according to mineral matrix technology and stored in a special landfill with separate stocking of gold-containing cinder and other arsenious residues. This technology of waste disposal specifies extraction of valuable metals from cinder in future. The similar case is with Vershino-Darasunsky settlement in Trans-Baikal Territory where the abandoned industrial site of the former arsenic plant is located in the central part of the settlement. The concentration of arsenic and valuable components can be compared to the concentration at AAP. Since 2016 the demand for removal of such industrial pollution sources is legislated due to amendments in Federal Law No. 7-FZ on environmental protection. It results in a real possibility to remove such dangerous industrial residues deposits as the one located in Vershino-Darasunsky settlement. INTRODUCTION In the frames of federal special-purpose program «National system of chemical and biosafety of Russian Federation (2009-2014)» as well as the Irkutsk region program “Environment protection in Irkutsk region"" for 2006-2015, INRTU performed the works aiming at disposal of accumulated environmental deprivation–arsenical residues from the former Angarsk arsenic plant (further - AAP) (Figure 1) on processing of arsenopyrite ores of Svirsk town in Irkutsk regionThe plant produced arsenic trioxidein 1934-1949."

Jan 1, 2018

By Rod A. Levander, Kenneth B. King

INTRODUCTION The Panna Maria Uranium Project is located in South Texas, between San Antonio and Corpus Cristi, approximately 10 kilometers (6 miles) north of Karnes City, Texas. The tailings disposal facilities are located immediately south of the mill area and the central portion of the west pit. The mine and mill drainage pond is located north of the tailings pond. Location of the facilities is shown in Figure 1. The tailings pond is a disposal area for mill process wastes and has a capacity for storage of approximately 9 million metric tons (10 million tons) of tailings. This capacity is adequate for approximately 10 years of milling at a maximum rate of 2,700 metric tons (3,000 tons) per day. The tailings disposal facilities were designed to essentially eliminate discharge of contaminated waters to preclude degradation of surface streams or ground water. Water levels in the ponds will be maintained. through reuse in the mill and by evaporation. WASTE CHARACTERISTICS The sources of waste that is stored in the tailings pond include runoff from the mill area and tailings from the mill processing. The tailings consist of a sand slime slurry; the percentage passing the No. 200 sieve varies depending on the ore material, and it is generally in the range of 10 to 25 percent. The tailings have a density of approximately 1.2 metric tons per cu m (75 lb per cu ft) and an average specific gravity of 2.55. The liquid effluent from the mill consists of waste solutions from the grinding, leaching, extraction, and washing circuits of the mill. These solutions have an initial pH of 1.5 to 2.0 and contain the remnants of the leaching agents used in the mill process, sulfates, some silica as the primary dissolved solid, and trace quantities of soluble metals. This liquid is discharged with the solids into the tailings pond and is diluted by rainfall and runoff from the tailings dam. Runoff from the mill area, drainage from the mine area, and de-oiled water from truck washing facilities is discharged into the mine and mill drainage pond. Runoff from the mill area drains approximately 28 hectares (70 acres) from offices, shops, the mill site, the ore pad, haul roads, and adjacent areas. Mine drainage is collected in the mine area and pumped to the mine and mill drainage pond.

Jan 1, 1980

By V. L. Nkuna, S. R. Amos, T. Naidoo

"Ivanhoe Mines’ Kamoa copper project is a recently discovered high-grade copper sulphide deposit in the Katanga Province of the Democratic Republic of Congo. The mineralization identified to date within the resource is typical of sedimentary-hosted stratiform copper deposits and comprises three distinct units: supergene, mixed, and hypogene mineralization. Since 2010, various metallurgical test work campaigns have been conducted on drill core from different areas within the deposit in line with resource expansion. The most recent bench-scale flotation test work has shown positive repeatable results. The work was conducted on a composite sample of drill core from the southern part of the Kamoa resource. The sample is representative of the first four to five years of planned production, when Ivanhoe Mines intends to produce clean, highgrade copper flotation concentrate. A detailed test work roadmap developed from first principles for the campaign led to an optimum flow sheet for treating the Kamoa mineralization. Copper recoveries of 88.3%, at a concentrate grade of 39.0% copper, were achieved using a composite sample — an improvement on the previously published 85.9% life-of-mine average copper recovery indicated in the November 2013 Kamoa Preliminary Economic Assessment. IntroductionThe Kamoa project is a newly discovered, very large, stratiform copper deposit with adjacent prospective exploration areas within the Central African Copperbelt, approximately 25 km west of the town of Kolwezi in the Democratic Republic of the Congo (DRC) and about 270 kilometres west of the provincial capital of Lubumbashi. Ivanhoe Mines holds its 95% interest in the Kamoa copper project through a subsidiary company, Kamoa Copper SA Limited SPRL, and DRC government owns 5%. The project location is indicated in Figure 1.In January 2013, a new independent mineral resource estimate ranked Kamoa as Africa's largest high-grade copper discovery and the world's largest undeveloped highgrade copper discovery. At this time, Ivanhoe Mines had estimated discovered Indicated Mineral Resources of 739 Mt grading 2.67% copper, containing 43.5 billion pounds of copper, and Inferred Mineral Resources of 227 Mt grading 1.96% copper, containing 9.8 billion pounds of copper, as demonstrated in the Preliminary Economic Assessment (PEA) (AMC, 2013). A 1% copper cut-off grade and a minimum vertical mining thickness of 3 m were applied in each classification."

Jan 1, 2016

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

Jan 1, 1981

By Wallace Ralston

The East Texas district includes the northeast 38 counties of the State of Texas. It still remains one of the most important oil-and-gas-producing areas of the United States, since it includes within its boundaries the large East Texas field. Other fields of importance are: Van, Rodessa, Talco, Cayuga, Long Lake and Sulphur Bluff. Production is from the Nacatoch, Blossom, and Woodbine of the upper Cretaceous and the Paluxy and Glen Rose of the lower Cretaceous. Of these, the Woodbine is the most important, since it is the producing horizon iri the East Texas field. Fields East Texas Field.—There were 2471 oil wells, 1 gas well and 84 failures completed during the year of 1936 in the East Texas field; that is, 1565 fewer than in the year of 1935. The total number of producing wells is now 22,031, of which 50 are temporarily shut down; 18,435 wells flowing their production; 3150 pumping; 293 on gas-lift, and 82 operated by other methods. This shows an increase of 1642 wells on artificial lift over Jan. 1, 1935. The average daily salt-water production is estimated at 42,700 bbl. from 2327 wells—an increase of 1201 wells over the year of 1935. The field is now drilled to an average density of one well to 5.85 acres. Average bottom-hole pressure has dropped from 1193 Ib. to 1168 lb., approximately 1 lb. drop for each 6,600,000 bbl. of oil produced. During the month of December 1936, the estimated daily average production of illegal oil had dropped from 20,500 bbl. in December 1935 to 7500 bbl. The commercial gas production is of minor importance in the East Texas field, and figures for the amount of gas taken from the field are not available at the present time. Talco Field, Titus and Franklin Counties.—The discovery well was drilled by Housh, Thompson et al. on the C. M. Carr tract and completed in March 1936, as a pumping well making 552 bbl. per day of 24' gravity oil. This well was first completed at a total depth of 4208 ft., the top of the sand being 4190 ft. Later it was deepened to 4260 ft., which increased the production from 552 to 725 bbl. per day. This field is producing

Jan 1, 1937

By K. Huttu, A. van Wageningen

"The Kittilä underground mine in northern Finland is extracting one of the largest known gold deposits in Europe. Kittilä achieved commercial production on May 1, 2009, becoming Agnico Eagle’s first mine to open outside of Canada. The Kittilä orebodies were initially mined from two openpits – Suuri and Roura – and underground operations were added in October 2010. The pits were mined out in November 2012, so mining is now entirely underground in the Suuri and Roura deposits. Additional, small openpits will mine any remaining near-surface reserves in the future.The Kittilä Mine is located in the Lapland region of northern Finland, approximately 900 km (560 miles) north of Helsinki and 150 km (93 miles) north of the Arctic Circle. With a mine life estimated through 2034, Kittilä is Agnico’s longest-life mine.The mine’s proven and probable reserves contain 32 Mt (4.7 million oz) at 4.6 g/t (0.13 oz/ st) gold. Ore has been mined from underground since 2010. The original 3 kt/d (3,300 stpd) operation underwent an expansion project to increase the throughput capacity of the mill by 1 kt/d (1,100 stpd), starting by the end of 2014.The region around the Kittilä Mine is underlain by mafic volcanic and sedimentary rocks of the Kittilä Greenstone Belt oriented nearly vertical. This greenstone belt is similar to those hosting Agnico Eagle’s Canadian deposits in Quebec’s Abitibi region and Nunavut. At the contact point between ironrich and magnesium-rich rocks lies a 100- to 200- m (330- to 660-ft) thick structural zone known as the Suurikuusikko Trend. This trend hosts the Kittilä deposit, where multiple mineralized zones have been traced over a strike length of more than 25 km (15 miles).Most of Agnico Eagle’s work has focused on a 4.5-km (2-mile) segment of the Suurikuusikko Trend that hosts the five main zones of the known gold reserves and resources. From south to north they are: Ketola, Etela, Suuri, Roura and Rimpi.The ore at the Kittilä Mine is refractory, making gold extraction relatively difficult because the gold is generally locked inside the two main sulfide minerals – arsenopyrite and arsenic-rich pyrite. Only about 4 percent is very fine-grained “free gold” (not locked in sulfide grains) and easier to extract. Most of this free gold is found in the outer, oxidized or eroded sections of the ore."

Jan 1, 2015

By Paul R. La Pointe

GEOTECHNICAL OBJECTIVES OF UCG PROGRAM Underground coal gasification (UCG) is one of several in situ extraction technologies for utilizing deep, otherwise unmineable coal. This technology involves the injection of air or oxygen into the target coal seam, ignition of the coal and recovery of product gases. Geotechnical characteristics of the coal and enclosing rock strata strongly influence the success of this Process. The flame front follows the coal cleats (Hill, 1979), and cleating and bedding control the manner in which coal collapses into the cavity, thereby affecting recovery efficiency (Ramirez and others, 1982). Moreover, coal seams typically behave as fracture dominated aquifers in which cleat orientations govern subsequent contaminant dispersal. For these reasons, the Anaconda Minerals Co. undertook a detailed study of fracturing within the Wyodak-Anderson Coal in the eastern Powder River Basin (PRB) near its Rocky Hill UCG test site (Fig. 1). STRESS HISTORY AND FRACTURING IN THE POWDER RIVER BASIN Vertical jointing, or cleating, in coals may arise from shrinkage shortly after deposition under low confining pressures, or later in time, soon after burial but after the coalification process is well advanced (Ting, 1977). There is much evidence to support the formation of cleating after coalification (e.g. Hawkes and Mellor, 1970; McCulloch and others, 1974), which suggests that the stress state existing at that time strongly influences cleat formation. McCulloch and others (1974) argue that the typically better developed face cleat forms parallel to the plane containing [ol] and [o2]. Where the relation between cleats and stess has been studied (McCulloch and others, 1974; Stone and Snoeberger, 1977; Ward, 1977; Kulander and Dean, 1978), the highly systematic face cleat strongly parallels the [o1] direction. Stone and Snoeberger (1977) report face and butt cleat measurements of N70E and N29W for the Felix Coals 220 m stratigraphically above the Wyodak at the Hoe Creek UCG site (Fig. 1). Structural trends also indicate the past state of stress in the PRB. Glaze and Keller (1963) suggest that the NE-SW regional compression during the late Permian formed major N-S fold lines and basement magnetic anomalies. Mapel (1959) describes folds having N60W and N30W rends in the northern PRB, as well as a large normal fault trending northeast. Glass (1976) also describes normal faults trending northeast. Similarly, Olive (1957) mapped normal faults

Jan 1, 1984

By Laxman M. Amaratunga, Glen G. Hein

"Most mining operations utilize the tailings coarse fraction for backfilling while the fines have been traditionally disposed on surface in tailings ponds. Recently, paste fill has been developed which uses a fraction of the fine tailings (10-30% w/w of <45 µm) in the fill (Brummer and Moss, 1991). This development allows a mining company to use a portion of the fines for backfill, but still requires surface impoundment of the unused tailings. Reactive tailings have the added problem of generating acid after being oxidized in the presence of air and water. Treating acid mine drainage (AMO) from these ponds has proven to be a costly environmental problem and a long term liability.One practical method to reduce the environmental impacts associated with the tailings pond is to modify the physical and chemical prope1ties of the fine tailings so that they can be utilized in an environmentally safe manner. This was accomplished by agglomerating the tailings into pellets and using theses pellets as a coarse aggregate in a total tailings paste fill. This technology does not require any supplemental sand or gravel to be added to the fill.This paper presents the ongoing research into utilizing these tailings agglomerates as a coarse aggregate fraction, in order to create a total tailings paste fill which has supe1ior strength characteristics.IntroductionThe Canadian mining industry produces more than 500 million tonnes of tailings and waste rock annually. The waste rock is usually returned underground for use as rock fill in underground support with the tailings remaining to be disposed of in tailings ponds. Tailings consist of a coarse and fine fraction, and can be either reactive (acid generating) or non-reactive. In an effort to minimize tailings disposal, the coarse fraction has been used underground in backfill along with alluvial sands. This diversion of the coarse fraction, however still requires that the fine fraction be disposed of in tailings ponds. This disposal creates the potential for acid mine drainage (AMO), since in the presence of air and water, reactive sulphide tailings can generate acid. It is therefore an important goal in the mining industry to minimize the disposal of fine tailings into tailings ponds since the treatment of AMO is a costly and long-te1m environmental liability."

Jan 1, 1997

[Abstmct of remarks by R. II. Sweetser arid suhscqueut discussion at the New York Meeting, February, 1926. The Transactions mill ultimately contain a report of work resulting from these discussions. Need for a standard of ualuration.—There should be an accepted method for evaluating the different kinds of bituminous coals based on their carbon just as that for iron ores is based on iron. Variation in the classes of coals and their desirability fol the different industrial uses is analagous to the case of iron ores. The iron people have developed a standard. In the past coals have not been intelligently distributed because the public has not been informed as to its uses. Many operators have known nothing about the final disposal of the coal shipped from their mines. Sales agents and retailers have had no means of evaluating coals except to make the price as high as the market will stand. Producers and distributors need to know how much more valuable the coals would be if prepared so as to contain less objectionable material. Mr. Swectser cnlletl on the Goal and Coke Committee of the Institnte to organize a study through producers and users to develop a method scientifically and comnereially sound E. A holbrook said that the information exists but the public is unwilling to investigate. He praised Ashley&apos;s1 nomenclature of the different kind of coal which has been accepted by the Coal Mining Institute of America and recommended to the American Enginecring Standards Commitee. G. S. Rice pointed out the wider variety and distribution of consumption of coal compared to Iron ore. C&apos;. M. Young approved Mr. Sweetser&apos;s remarks and hoped lor a research institution, probably endowed, to study utilization of coal after it comes from the ground. Mr. Sweetser cited incidents to prove his former statements. He said that a reduction of 1 per cent. ash in the coal would cut down the cost of pig iron at a blast furnace 30 c. a ton, that the coal men should tell what the iron men would have to pay to get better coal. H. J. Prryn endorsed the need of blast-furnace operators to know the kind of coal from which their- coke is made and the latler&apos;s ash content, as this affects coke consumption. A. C. Fieldncr explained that the difficulty of evaluating coal on the basis of chermical analyses for calorimetric tests is: (1) representativr

Jan 1, 1927

By G. A. Croes, N. Schwarz

This paper reports the results of a series of model displacement experiments carried out for measuring the efficiency of the water-drive process. This series forms a continuation and extension of that described by Engel-herts and Klinkenberg. Detailed information has been obtained on the influence of the oil/water viscosity ratio. The results are represented in the form of a diagram from which both the oil recovery and the water/oil ratio can be easily read off as a function of the total production (oil plus water) for all values of the viscosity ratio M between I and 500. These resu1ts are viewed in the light of the two existing theories on water flooding, viz. that of Buckley and Leverett and that of Dietz. It is shown that if the Buckley-Leverett theory applies, the relation between the relative permeability ratio and saturation, which plays a major role in this theory, (.an be calculated with a high degree of accuracy from the results of the model experiments. Dietz&apos;s theory is found to be in agreement with the experimental results in a certain range of circumstances. A discussion of the physical background of the Buckley- Leverett and Dietz theories makes it clear why the latter fails to agree with some of the experimental results. INTRODUCTION For an economical exploitation of oil fields both a rational control of natural (primary) recovery processes and a proper application of artificial (secondary) recovery methods are essential. These require detailed information upon the dependence of the relevant flow processes on the variables by which they are governed. Three methods are available for obtaining this information: (1) (statistical) analysis of actual field data; (2) appropriately scaled model experiments; and (3) mathematical analysis of the physical problem. If sufficient data were available, the first method would, in principle, be the most reliable one, although it has the drawback that the basic relationships between the variables involved are to a large extent obscured by the natural inhomogeneities of most reservoirs. The most profitable procedure seems to be: (a) determining the fundamental relations for idealized reservoirs by methods (2) and (3); (b) checking the consistency of the results obtained by these two methods; and (c) establishing the range of applicability of these methods to actual reservoirs by comparison with the data accumulated under (1). In this paper items (a) and (b) are dealt with for the case of homogeneous loose sands initially saturated

Jan 1, 1956

By Charles Wurms, Robert Steinitz

DURING an investigation to determine the compatibility of several metals with tungsten, the degree of reaction between tungsten and lead was determined. A literature survey showed considerable disagreement between the results of various earlier investigators. Information presented by M. Hansen1 states that, according to S. Inouye,2 tungsten has very poor resistance against attack by molten lead, even at the low temperature near the melting point of lead. A phase diagram in the same reference indicates a solubility of 30 wt pct W in molten lead at 1300°C. Contrary evidence was shown by Parkman and shepard3 in their investigation of materials for use in heat transfer systems containing liquid lead alloys; they performed experiments at 980°C for times up to 50 hr without noticing any appreciable attack on the tungsten. Tungsten sheet was immersed in molten lead contained in Vycor tubes evacuated to at least 10-4 mm Hg. R. Bernard4 and R. I. Jaffee5 also indicate that no reaction occurs between molten lead and tungsten. However, no experimental data were included in these reports. In order to settle the question of the solubility of tungsten in molten lead, tungsten powder with an average Fisher subsieve particle size of 6µ was pressed in a double-acting die at 6 tsi to small compacts. They were then sintered in an atmosphere of hydrogen for 2 hr at 1700°C, resulting in a sintered density of 64 pct of theoretical with final dimensions of 0.260 in. in height and 0.245 in. in diameter. For all tests, the tungsten pellets were immersed in molten lead. Table I is a summary of the experimental procedures showing atmospheres, temperatures, and times. The first experiments were performed at 650°C; at this temperature, the liquid lead should wet the porous tungsten better than at its melting point, and penetrate into the pores. The lead was heated in a graphite crucible with a graphite cover. No protective atmosphere was used, as it was expected that the formation of CO would be sufficient to prevent oxidation. This, however, proved to be wrong, and all tungsten samples showed a slight oxide film after treatment. The tungsten samples were immersed in the molten lead for times of 5 min, 10 min, 1/2 hr, 1 hr, and 2 hr. After this time, the melt was poured into a form and the

Jan 1, 1962

By R. S. French, W. R. Hibbard

M. Balicki—As one who some years ago spent much time searching for an alloy with high thermoelectric power that would be suitable for heat energy-electric energy converter based on the principle of a thermoelectric pile, I would like to make a few comments touching upon that aspect of this excellent paper. The specifications which one of the authors has formulated so aptly do not appear to me as complete, and I would like to add to them the following two demands: (1) Melting point not less than 500°C and (2) elongation on 2 in. at least 5 pct. Condition 1 would insure good thermodynamic efficiency of the pile by permitting a reasonably high temperature of the hot junction and at the same time making it suitable for such application as salvage of vast amounts of energy which is lost in the form of sensible heat of waste gas leaving industrial and other furnaces. Condition 2 is also a very important one; particularly, because of the fact that all elements and alloys which exhibit attractive thermoelectric properties seem to be as a rule very brittle. The design, erection, and maintenance of a pile incorporating a brittle material in one of the most crucial parts of the converter would be a very difficult and costly affair indeed. To illustrate this point I may mention great difficulties I encountered when trying to make thermocouples of silicon. Casting of 1/4-in. rods in iron or graphite mould yielded only broken pieces which had to be welded into a bar. Such a bar had to be handled very carefully during the determinations. Raising the molten silicon by vacuum into a silica tube and solidifying it there solved the problem of stability as well as that of electrical insulation of a bar. On tests this bar (98.5 pct Si) showed dE/dT = — 275.5 microvolts per degree C vs. copper. Thermocouples made of this grade of silicon and pure nickel developed 0.39 v when temperature of the hot junction was approximately 1000°C. The possibility Dr. Gonser has mentioned of trying to impart to metals good thermoelectric properties by alloying them with silicon, has been considered. The preliminary survey which I have made was, however, disappointing. Perhaps, it will be advisable to record the results obtained (table VI), as this might save repetition of search in fields already covered.

Jan 1, 1951

By J. R. Martson

The arrangement of the larger furnace at the United Verde smelter, together with the boiler layout and connecting flues, is shown in Fig. 1. The important dimensions are given in the figure, and cross-sectional areas of combustion space, flues, etc., are recorded. The location and extent of the baffling in the boilers also are shown because of their importance in the interpretation of the data. Two boilers serve each reverberatory furnace. They are of the class M-26 Stirling type (rated at 713 boiler hp. each). Steam is generated at 178-lb. gage pressure at a total temperature of 528" F., from feed water at a temperature of 113" F. The feed water is condensate from turbogenerators plus make-up softened by the lime-soda process. The make-up is rather large in amount as it must match the amount of steam consumed in reciprocating blowing engines and the like, which are not served by surface condensers. As a result the boiler feed water has a hardness of 1.5 to 3.0 grains per gallon, but this results in the formation, during six months to one year of boiler operation, of scale only of "eggshell " thickness or less. The boilers are set at the same elevation as the tapping floor of the furnace, which results in the introduction of the waste-heat gases at the top of the front bank of tubes and necessitates the inversion of the usual baffling arrangements. The boilers are equipped with valve-in-the-head soot blowers, which keep most of the dust blown off the surface of the tubes, but it is necessary to supplement the soot blowers by high-pressure air lances, inserted by an attendant through side doors in the boiler setting. Deposits of coal ash and of dust from the furnace charge accumulate in the flues between the reverberatory furnaces and the boilers and this necessitates cleaning of these connecting flues by hand during the greater part of one of the three 8-hr. shifts during each 24 hr. This cleaning operation requires the opening of clean-out doors, which, of course, has some effect on the yield of steam. Revebberatory Furnace Practice The United Verde furnaces are fired with pulverized coal of sub-bituminous type of which the screen and chemical analyses are given in Table 1.

Jan 1, 1934

By R. A. Bergier

The Transactions of the Institute contain nothing, as yet, on the subject of pocket-compasses; and in the belief that American miners, explorers, geologists, and engineers will gladly welcome any infor mation about new forms of apparatus, simple in construction and adapted to rapid use, I offer a short description of an improved French combination pocket-compass, which, if once introduced into use, would seem likely to supersede similar instrumens of older pattern. The instrument in my possession I owe to the kindness of M. G. de.La Bouglise, of Paris. The frame of the compass is a double-bottomed box, 2 1/2 inches square and 9/16 of an inch high, made of white metal, and with its sur faces very carefully planed. The sides of the box make true right angles with each other. Two sides are parallel with, and two are normal to, the north and south line of the principal graduation. The bottom is accurately planed, so that the instrument will lie true on a plane surface. A glass, 2 3/8 inches in diameter, is set in the top plate. The graduation for the needle is engraved on the upper face of a ring, connected with the flat disk, in which the center-pin is in serted. Between this disk and the bottom of the box there is a hollow space a quarter of an inch deep. The needle is 1 7/8 inches in length, and can be balanced by means of a small sliding piece of metal. The graduation extends from 0° to 360°, reading from right to left. The stopping of the needle is effected by a screw, A, on the side of the box near the handle, H (Fig. 1). There is also a clinometer attachment, the graduation for which is engraved on the upper face of the silver-plated disk. The compass has two sights. The north sight, B, is a simple pin, a little more than an inch high, which, when not in use, is kept folded down, and is partially protected by the exterior sliding-plate, C. When raised for use, the sight is stopped in a vertical position by a projecting pin attached to the side of the box.. The south sight, D, is connected with the handle, H, and the sight and handle

Jan 1, 1890

By B. D. Cullity, J. T. Norton, M. Telkes

M. Balicki—As one who some years ago spent much time searching for an alloy with high thermoelectric power that would be suitable for heat energy-electric energy converter based on the principle of a thermoelectric pile, I would like to make a few comments touching upon that aspect of this excellent paper. The specifications which one of the authors has formulated so aptly do not appear to me as complete, and I would like to add to them the following two demands: (1) Melting point not less than 500°C and (2) elongation on 2 in. at least 5 pct. Condition 1 would insure good thermodynamic efficiency of the pile by permitting a reasonably high temperature of the hot junction and at the same time making it suitable for such application as salvage of vast amounts of energy which is lost in the form of sensible heat of waste gas leaving industrial and other furnaces. Condition 2 is also a very important one; particularly, because of the fact that all elements and alloys which exhibit attractive thermoelectric properties seem to be as a rule very brittle. The design, erection, and maintenance of a pile incorporating a brittle material in one of the most crucial parts of the converter would be a very difficult and costly affair indeed. To illustrate this point I may mention great difficulties I encountered when trying to make thermocouples of silicon. Casting of 1/4-in. rods in iron or graphite mould yielded only broken pieces which had to be welded into a bar. Such a bar had to be handled very carefully during the determinations. Raising the molten silicon by vacuum into a silica tube and solidifying it there solved the problem of stability as well as that of electrical insulation of a bar. On tests this bar (98.5 pct Si) showed dE/dT = — 275.5 microvolts per degree C vs. copper. Thermocouples made of this grade of silicon and pure nickel developed 0.39 v when temperature of the hot junction was approximately 1000°C. The possibility Dr. Gonser has mentioned of trying to impart to metals good thermoelectric properties by alloying them with silicon, has been considered. The preliminary survey which I have made was, however, disappointing. Perhaps, it will be advisable to record the results obtained (table VI), as this might save repetition of search in fields already covered.

Jan 1, 1951

By Thierry Bernard

Today, experimental measurement of the shock energy/gas energy pair, or pool test, has now become virtually standard. But pool" energy per m(exp 3) of bedrock has no significance for explosives used in any medium other than water.

Jan 1, 1994

Year of Election 1944 Walter Hull Aldridge, New York, N.Y 1955 Louis S Cates, New York, N.Y 1946 William Fraser (Lord Strathalmond), London, England 1917 Herbert Hoover, New York, N Y 1955 George M Humphrey, Cleveland, 0

Jan 1, 1959

By Henry E. Colton

Very little information has been published concerning the Tennessee coal-field. The State never appropriated over $600 per annum for a geological survey, and that was discontinued about 1870. Yet even with that small sum Dr. J. M. Safford made a

Jan 1, 1886