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By J. M. Boxall

The electrolytic zinc plant at the Bartlesville division of Z.C.A. originally used lead anodes containing only 0.25% silver because only H.G. zinc and alloy was to be produced. In 1984, a small addition of strontium carbonate was made to control the surf ace morphology of the cathodes which reduced short circuits and increased current efficiency. Later a drastic reduction in colloid addition further reduced short circuits. Increased strontium additions allowed the production of some S.H.G. zinc. Increasing the silver content of the anodes reduced the lead in cathodes only slightly. More important, the anode scale that was not removed during anode cleaning, entered the electrolyte and contaminated the cathodes with lead. A piping system has been installed to purge the contaminated electrolyte directly back to the leaching section. This has allowed the plant to produce S.H.G. zinc consistently.

Jan 1, 1990

By Hongwei Wang, Hao Zhang, Panshi Xie, Xicai Gao, Youfu Zeng, Yongping Wu, Juan Dou

"The reserves and production of steep coal seams account for approximately 15–20% of the total coal reserves in China and 5–10% of the total annual coal production. More than 50% of the coal is high-quality coking coal and anthracite, based on reviews of steeply dipping seam mining at home and abroad. Fully mechanized mining is considered to be the only safe and efficient way to mine steeply dipping seams. To achieve fully mechanized longwall mining, four issues must be addressed: ground control, equipment stability control, development of complete sets of equipment, and safety. Engineers must focus on many factors of steeply dipping seams, such as the following: stope, roof asymmetrical structure, stability control, “roof-support-floor” system dynamics control, multi-region big-range ground control technology, adjustment false-inclined method, nonlinear layout method, control technology, safety, and development of complete sets of equipment. Engineering practices in China introduce these concepts systematically: fully mechanized mining of medium-thickness seams, fully mechanized high mining of thick seams, fully mechanized caving of extremely thick coal seams, fully mechanized mining of coal seams, roadway support, and so on. A steeply dipping seam is one with a dipping angle in the range of 35°–55°. This accounts for about 15%–20% of the total coal reserves in China. Because of the environment, more than 50% of the coal is high-quality coking coal and anthracite. Mining steeply dipping coal seams is known to be difficult in the mining industry. After the study conducted by Germany, Poland, and the former Soviet Union in the1970s–1980s, the fully mechanized mining basic theory, core technology, and key equipment research have not been a breakthrough. Being fully mechanized is a technological “restricted area” at home and abroad. A lot of steeply dipping seams are excavated by non-mechanized mining, leading to harsh job environments, labor-intensive work, low productivity, frequent accidents, and significant security hazards. The death rate per million tons of coal reaches above 4.0 and is far higher than slow inclined coal seam mining. Making steeply dipping seam mining safe and efficient is a global problem that urgently needs to be solved. Fully mechanized mining is the only way to overcome these problems (Wu, 2003; Wu, Yun, and Zhou, 2000). The realization of fully mechanized coal seam mining must solve four problems. (1) Examine the disaster mechanisms induced by overburden “key strata” mutation under steeply dipping seams as a basis for ground control methods and technical problems. (2) Work to solve the dynamic stability control technical problems of thousands of tons of complete sets of equipment under the steeply dipping seams. (3) Complete sets of equipment must be developed to adapt to the conditions of steeply dipping seams. (4) Overcome the steeply dipping face safety protection technical problems. In the last two decades, a number of universities, research institutes, and enterprises have realized the importance of safety and efficiency of steeply dipping seams in China. This is especially true of Professor Wu Yongping’s “hard coal layer safe and efficient mining” team at Xi’an University of Science and Technology. After several generations’ research and technical work, the team puts forward China’s longwall mechanized coal mining of steeply dipping seams strata control basic theory. This paper puts forth a series of key technologies of fully mechanized mining, develops and manufactures complete sets of equipment by independent innovation, and establishes a perfect face security system, forming a theory and technology system for fully mechanized longwall mining of steeply dipping seams."

Jan 1, 2017

By Swayn G. P, Floyd J. M

BHAS, together with the G.K. Williams Laboratory, and Ausmelt Pty. Ltd., has examined a number of possible alternatives to conventional silver cupellation in gas fired reverberatory furnaces. Of these alternatives, a Sirosmelt furnace was preferred for both technical and economic reasons. Small scale work on Sirosmelt processes was carried out at the G.K. Williams Laboratory at Melbourne University in 1985/86. Pilot scale work was carried out at the Ausmelt plant in Melbourne in July 1987. Board approval for construction of a full size plant was received in October, 1988, and construction of the plant is currently underway. The plant is expected to provide significant savings in operating costs and silver inventory. Hygiene around the furnace should also be greatly improved. An increase in cupellation rate of at least ten times is expected, compared with the conventional reverberatory furnace.

Jan 1, 1989

By Sylvia S. Eskander, Sherif S. AbdelSalam, Mona B. Anwar

Expanded Polystyrene (EPS) geofoam has long been used in numerous geotechnical applications due to its lightweight and durability. Among these applications is the use of EPS as an alternative fill material to conventional roadway soil embankments. This study investigates the benefits of using EPS roadway embankments to reduce the induced stresses and strains on the underlying peat and soft clay. Firstly, mechanical and interface properties of the EPS were acquired from a companion study that also considered the cyclic loading and creep effects. In order to determine the actual stress-strain distribution within the EPS blocks, a physical prototype was developed in this study with overall dimensions of 1.2 x 1.0 x 0.5 m, with a density of 35 kg/m3. The prototype EPS blocks were subjected to cyclic loading with intensities ranging from 10 to 50 kN. Then, a numerical model was developed to simulate the behavior of EPS roadway embankments. The numerical model outcomes were verified against the prototype measurements. A parametric design study was also conducted to develop a correlation between EPS embankment height and the stress-strain that occur in the underlying strata. Results showed that EPS embankments could reduce stress and settlement in the underlying soil by about 40% and 54%, respectively. EPS APPLICATIONS AND BACKGROUND Expanded Polystyrene (EPS) geofoam blocks have been effectively used during the past several decades in numerous geotechnical applications. EPS static and dynamic properties as well as interface behavior were extensively characterized in various published studies such as that conducted by Horvath, 1994; Stark et al., 2004; Zarnani and Bathurst, 2008; Trandafir et al., 2010; Athanasopoulos et al., 2012; AbdelSalam and Azzam, 2016; and others. However, to be able to use locally produced EPS blocks, its physical and mechanical properties must be thoroughly determined. The lightweight and compressible material of EPS geofoam makes it suitable both as a backfill and for absorbing pressure in both the vertical and horizontal directions. Kim et al. (2010) adopted EPS as an alternative for soil backfill around buried pipes; their model showed a reduction in the vertical stress acting on the pipes by up to 36%. Oshati et al. (2012) presented a case study of a field-instrumented box culvert with EPS blocks on top, which reduced vertical stress by 58%. Meguid et al. (2017) used an EPS inclusion to surround culverts and achieved up to 72% stress reduction, depending on the EPS density and the culvert depth to width ratio. AbdelSalam et al. (2017) used a relatively thin EPS inclusion behind flexible cantilever retaining walls, verifying that the lateral pressure can be significantly reduced by 25%; they also developed a design chart for this application. Moreover, AbdelSalam et al. (2019) investigated the use of EPS as a lightweight alternative for soil backfill above shallow tunnels. It was found that EPS had created a differential movement between the wedge located above the tunnel and the neighboring soils, which accordingly reduced the vertical stress acting on the tunnel by 85%, due to the use of lightweight EPS material and positive soil arching.

Jan 1, 2019

By W. J. Slatosky

W. 0. Philbrook (Cairiegie Institute of Technologyogv—Mr. Slatosky has presented an interesting and constructive paper that represents another step along the way of converting steelmaking from an art to a science. I am confident that his computer will be practical and successful and that with a very few months of experience it will provide a significantly better degree of control than his record of 65 pct of heats within range obtained with the slide-rule calculator . A paper such as this, with a lot of symbols and condensed mathematics, is difficult to comprehend quickly. Since I have had an opportunity to study it carefully, perhaps my evaluation of its validity and accomplishments will save time for others. Mr. Slatosky has correctly used standard principles of stoichiometry and heat balances, which are available to anybody, but he has also brought to them two original contributions: 1) He has developed from operating data some empirical relations for predicting the final FeO content of the slag (at 0.5 pct C end-point) as a function of slag basicity, lance height, and scrap, ore, and scale in the charge. This improves the accuracy of prediction of temperature or scrap requirement compared with assuming an arbitrary, constant FeO content at the end of each heat. There is no assurance yet that exactly the same relations will hold for other furnaces or practices, but similar correlations can be expected. 2) He has combined calculations that are ordinarily carried out laboriously as a number of individual steps into a single, simple linear equation that can readily be fed into a machine. This involved a tremendous amount of painstaking detail work as well as the imagination to see the possibility and work out the steps. While his particular Eqs. [3] and [6] are valid only for the furnace design, charge weight, and blowing time used at Aliquippa Works, only a few numerical values have to be changed to adapt it for other conditions. In order to arrive at a useable solution, Mr. Slatosky had to make some basic assumptions about the process that are similar to those used by others. He neglected variation in some process variables and assumed an arbitrary average value for waste gas analysis and temperature for want of more exact information at the present time. All of these judgments are clearly stated. In addition, some thermody-namic data presently available are not adequate for the job, notably in relation to heats of formation and sensible heat in slag, and some expedient has to be adopted to get around the difficulty. Other people might prefer slightly different judgments about these details and hence obtain somewhat different numerical solutions. This is not of serious importance, however, because the errors accumulate in the "heat loss" term and are largely self-compensating for a constant heat time. Although the extended Eq. l(a) in Appendix I was set up as a rate equation originally, for convenience in using an analogue computer as stated in the paper, the time dependence was removed by later mathematical manipulations and assumptions about the process. The final result is an integration of element and energy balances from initial to final states; this procedure is as legitimate here as in any other form of heat-balance calculation. The formal handling of stoichiometry and thermochemistry appears to be correct, and it is assumed that any arithmetical errors would have come to light in applying the calculations to furnace practice. Mr. Slatosky's approach is not necessarily unique, in that other people might start with apparently different equations or prefer another form of final equation for another type of computer. However, he has presented an accomplished result that appears to be a theoretically sound and practically useful way of applying scientific principles and rapid computation for better control of steelmaking. His success will undoubtedly encourage himself and others to improve on the mathematical model and its use as better informatioq becomes available. John F. Elliott (Massachusetts Institute of Teck-t2ology)-The last comment by Mr. Richards that a calculator is quite unnecessary for an L-D operation ?-equi??es a rebuttal. The L-D furnace is a very high capacity process which places a premium on close control. When one is making steel at rates between 100 and 200 tons per hr, one cannot afford the luxury of an extra 5 or 10 min at the end of a heat correcting for an error that should never have been made in the first place. Mr. Slatosky's paper is a very sound application of the simple principles of stoichiometry and the energy balance. It is a satisfactory and valuable start, but only the start of the development of methods of control for this process. An analysis of the process shows that it should be very suitable to control by a computer. This is especially the case when various grades of steel are to be made. In fact, it would seem that the organizations who are planning new and bigger installations of L-D vessels should consider carefully the advantages that would stem from computer control of a vessel with the operator present to do little more

Jan 1, 1962

By J. H. Desautels

The modern coal mining industry is subject to a vast array of regulations affecting virtually all aspects of its operations. Probably the most significant over-sight role involves the Office of Surface Mining Reclamation and Enforcement (OSH) established by the Surface Mining Control and Reclamation Act, or SHCRA.1 Although complying and coping with SHCRA can be difficult for coal operators, this complex regulatory process has helped insulate the industry from ongoing efforts to extend regulation under the Resource Conservation and Recovery Act, or RCRA, 2 to aspects of the mining industry in general. The Environ-mental Protection Agency (EPA) has struggled with regulation of mining wastes under RCRA since the mid-1970's, but RCRA itself places most coal mining wastes firmly under the control of OSH and SHCRA. Despite the general exemption from RCRA regulation for waste materials that result from extraction and processing of coal itself, there may be circumstances under which coal mining operators could find themselves subject to regulation under RCRA or a number of other statutes administered by EPA. This paper will review briefly the history of mine waste regulation under RCRA, as well as touch on the policy decisions that led to the exemptions from RCRA regulation. The paper will summarize coal waste requirements under SHCRA, and also will review what operations at coal mines need to be controlled, in order to avoid potential liability under RCRA. In addition, it will mention requirements under other federal laws, including CERCLA, or the "Superfund" law,3 and describe how they may impose restrictions or liability on coal operations for waste or other materials commonly found at surface and underground coal mines.

Jan 1, 1993

By R. D. Cromarty, J. Du Toit

"Penetration of matte into tap-hole bricks causes detrimental refractory wear, which can lead to furnace breakouts. The ability of the tap-hole clay to form a protective layer on the brick, thereby limiting matte penetration was investigated by examining the interaction between platinum group metal (PGM) matte, tap-hole clay, and alumina-chrome refractory bricks on a laboratory scale.Samples containing clay and brick as well as samples containing clay, brick, and matte were heated to different temperatures to establish the clay-brick interaction and the extent of matte penetration. The greatest degree of physical contact between the brick and clay was achieved at curing temperatures of 600°C. Poor clay-brick contact was observed in the sample that was heated to 900°C.Matte displaced the clay in the clay-brick-matte sample that was heated to 1350°C, with significant matte penetration into the brick. Less matte penetration was observed when the clay-brick-matte sample was heated to 1500°C. Less matte penetration was also observed in the claybrick- matte sample in which the clay and brick were pre-baked at 800°C, and the sample then reacted with matte at 1350°C. IntroductionFloatation concentrate from ore from the eastern Bushveld Complex of South Africa, specifically the Platreef, Merensky, and UG2 reefs, forms the feed to the Polokwane platinum smelter (Hundermark, De Villiers, and Ndlovu, 2006). The ore from the UG2 Reef has a high chromite content, which requires high operating temperatures (1500–1700°C) to avoid chromium spinel build-up in the furnace hearth. These high operating temperatures lead to matte superheats in the vicinity of 500– 600°C, which put increased demands on the refractory materials (Nelson et al., 2005).Upon tapping, the matte moves through the tap-hole, which consists of an aluminachrome refractory brick-lined tapping channel (Hundermark, De Villiers, and Ndlovu, 2006). The flow of molten matte is stopped by the injection of tap-hole clay to form a plug in the tap-hole. The tap-hole clay hardens and strengthens during curing to obtain the required hardness and refractoriness against penetration, corrosion and erosion (Nelson and Hundermark, 2014)."

Jan 1, 2016

By D. H. Baldwin, R. E. Reed-Hill

DURING the course of experiments involving oxygen equilibrations with a high-purity Pd-5 at. pct Rh alloy, the appearance of a subscale was noted. Most of the heat treatments in a pure oxygen atmosphere were performed at 900°C. The presence of the subscale was noted when the oxygen partial pressure was in excess of approximately 95 mm, up to 1 atm, the highest pressure used. According to available thermodynamic information' on the free energy of formation of Rh2O3 (the only reliably reported solid oxide of rhodium), this oxide should not be stable under these conditions, nor, of course, any oxide of palladium. Extraction of the subscale from the matrix was affected by treatment with aqua regia, and the resulting X-ray diffraction powder pattern is shown in Table I. This pattern is not that of Rh2O3, nor that of any known rhodium or palladium oxide. It can be indexed on a hexagonal lattice with a. = 5.22Å and co = 6.0Å. In addition to this unknown material, there was a weak Rh2O3 pattern. An X-ray fluorescence analysis of the powder gave 3214 palladium counts per sec and 5381 rhodium counts per sec, thus indicating the likelihood of a double oxide of palladium and rhodium. In view of the fact that the above result was obtained on a single sample, it is perhaps best not to insist upon the above ratio of about 0.6, but possibly to tentatively assume 0.5 or 1/1. The small amount of Rh2O3 in the X-ray pattern might have occurred on cooling, and it seems possible that the double oxide, below some temperature level, is not as stable as Rh2O3. To examine this point, some of the extracted double oxide was heated on a sheet of pure palladium in air at 900°C, and allowed to furnace-cool over a period of several hours. The X-ray diffraction pattern of the powder showed approximately the same intensity of Rh2O3 and the "hexagonal" oxide as before, but, in addition, a rather faint pattern of another crystalline species made its appearance. This finding again suggests partial transformation of the major (double) oxide to other oxides on cooling from 900°C. Application of the phase rule shows that only one equilibrium oxide (subscale) could have been present at temperature. Vacuum-fusion, oxygen analysis of the extracted subscale showed 7.7 ± 0.7 pct O. Combining this information with that of the X-ray fluorescence analysis suggests the formula PdRhO. This compound should have an oxygen content of 7.07 pct, which is within the range of error of the oxygen analysis. However, until a more careful stoichiometric study is made, it is suggested that the formula be written

Jan 1, 1965

By Raj Rajamani

In general, many past works examine mill load, mill power, shell lifter design etc with a view to improving the mill capacity. On the contrary, only a handful of papers have addressed the transport of fine slurry through the pulp lifters. We present the flow dynamics of pulp lifters via laboratory scale experiments, high speed photography and modeling with computational fluid dynamics. Next, the flow behaviour in ArcelorMittal?s Mont-Wright autogenous mill discharge flow is examined. In an effort to improve mill throughput, a comprehensive simulation and plant scale study of the mill was taken up with a focus on the shell and pulp lifter design. The shell lifter was found to be at its optimum limit for breakage of ore particles given the constraints of wear life and occupied volume. Next, the pulp lifter study was taken up with computational fluid flow code and discrete element code. The simulations directly showed that there was considerable carry-over of slurry. This effect was traced to the area restriction in the pulp lifter channels. Further, crash-stop studies confirmed slurry pooling within the mill. Thus it was readily apparent that mill throughput can be increased if the pulp-lifter could be redesigned. Consequently, the discharger section of the pulp-lifter was redesigned. First, this paper outlines the modeling in laboratory scale studies and plant scale studies leading up to the discharger design. Next, the discharge flow in seven different published mill surveys is predicted with computational fluid dynamics (CFD) methodology shown here. Finally, Mont-Wright mill circuit performance before and after the discharger redesign is described. The usefulness of fluid flow simulations in the design of pulp-lifters is emphasized.

Jan 1, 2012

By James J. Dawd

The investigation to evaluate the reserves of coking coal is being made by the Bureau of Mines in three parts: (1) To estimate known measured and indicated recoverable reserves of all coking coal; (2) to upgrade marginal coals through effective preparation; and (3) to study carbonizing properties of coals and coal blends not now widely used for metallurgical coke making. This is the ninth of a series of reports giving results of studies by counties under part (1) of the investigation. (See Appendix.) This report covers Raleigh County, W. Va., which comprises parts of Bald Knob, Eccles, Beckley, Meadow Creek, Mullens, Flattop, and Big Bend quadrangles. (See fig. 1.) A base map for each bed in each quadrangle was made to the scale 1 inch equals 1,200 feet. Maps of mines, location of drill holes, bed and total coal thickness, and outcrop of the bed were plotted on the maps. With all available data plotted, isopach lines were drawn to limit areas of known unmined reserves in beds 0 to 14 inches thick, 14 to 28 inches thick, 28 to 42 inches thick, and over 42 inches thick. These areas of coal reserves also were divided into measured" and "indicated" categories. All areas in each thickness range and in each category, mined-out areas. areas excluded from the estimate but which may contain reserves based only on geologic inference, and areas outside of the outcrop were measured by planimeter on the base maps, Estimates of total reserves 14 inches and more thick and maps for individual beds were prepared from these data.

Jan 1, 1952

By N. H. Coates

The Bureau of Mines determined average film coefficients of heat transfer for helium flowing through smooth, water-cooled tubes at Reynolds numbers of 5,000 to 25,000, helium inlet temperatures of 1,600° to 2,600° F, average temperature difference across the helium film of 800° to 1,300° F, and heat flux to 215,000 Btu/(hr)(sq ft). Since helium was the only gas used, the ranges of values for physical properties of the fluid are limited--thermal conductivity from 0.15 to 0.20 Btu/(hr)(sq ft)(O F/ft), viscosity from 0.082 to 0.104 lb/(ft)(hr), and Prandtl number from 0.63 to 0.68. Four equations, three of them dimensionless, correlate these coefficients with other properties of the helium--bulk temperature, film temperature, dimensionless parameters based on temperature, and the ratio of wall temperature to bulk temperature.

Jan 1, 1966

By Y. Kusuda

Since 1981, the Annaka refinery has recovered zinc from zinc leach residues by the Waelz Process. The important considerations of this process are the prevention of the growth of coating accretions in the Waelz kiln and the improvement of the zinc volatilization rate. In this paper, simulation predictions and the results of the industrial operation regarding the prevention of coating accretions in the Waelz kiln and the improvement of the zinc volatilization rate at the Annaka refinery are reported and compared.

Jan 1, 2006

By John Campbell, Miller. W. E., G. W. Jones

"INTRODUCTION Because of its interest in accident prevention, the Federal Bureau of Mines has cooperated with city officials and utilities in investigating the causes of explosions in sewers, manholes, and other underground spaces and has aided in the development of means to prevent such explosions. For several years this Bureau has cooperated with the Boston Edison Co. and the Boston Consolidated Gas Co. of Boston, Mass.,for the purpose of detecting and eliminating combustible and physiologically harmful gases in manholes in Boston and vicinity.These investigations 5/ have shown that combustible gases from the usual sources — leakage from gas—distributing systems and underground tanks and gasoline wastes — can be located and the dangerous conditions remedied by a well—planned, adequately supervised, systematic, and continuous manhole survey. However, other types of gas contamination, termed ""soil"" gases, are present in manholes at times and are more difficult to eliminate because they have no fixed origin and because seasonal and barometric chances affect their production and passage into the manhole. Soil gases usually are abnormally high in carbon dioxide and low in oxygen and contain varying amounts of methane. They are especially dangerous to men entering manholes because they lack sufficient oxygen to support life."

Feb 1, 1940

By Timothy Batchler

"Pumpable roof supports are currently being used to provide a safe working environment for longwall mining. Because different pumpable supports are visually similar and installed fundamentally in the same manner as other supports, there is a tendency to believe they all perform the same way. However, there are several design parameters that can affect their performance, including the cementitious material properties and the bag construction practices that influence the degree of confinement provided. A full understanding of the impact of these design parameters is necessary to optimize the support application and to provide a foundation for making further improvements in the support performance.This paper evaluates the impact of various support design parameters by examining full-scale performance tests conducted using the National Institute for Occupational Safety and Health (NIOSH) Mine Roof Simulator (MRS) as part of manufactures' developmental and quality control testing. These tests were analyzed to identify correlations between the support design parameters and the resulting performance. Based on more than 160 tests over 7 years, quantifiable patterns were examined to assess the correlation between the support dimensions, cementitious material type, wire pitch, and single-wall vs dual-walled bag designs to the support capacity, stiffuess, load shedding events, and yield characteristics.INTRODUCTIONDeveloped in the 1990s, the first major use of pumpable supports systems in U.S. longwall operations was in the support of bleeder entries (Barczak, Tadolini, 2008). Since then, they have been utilized as yieldable concrete supports to provide a safe working environment for longwall mining gateroads, bleeders, and emergency escapeways while also maintaining adequate ventilation pathways. The basic structure of pumpable roof supports has remained unchanged over the years. Formed in place with a twopart fast-setting grout, the support material can be pumped into a containment bag from several thousand feet away often through surface boreholes. The containment bag then acts as a form to fill the support and provides confinement to the grout during loading and after failure. Pumpable roof supports provide full contact with the mine roof and floor, which eliminates the need for secondary material to establish proper roof contact (see Figure 1). They provide a high peak load capacity and a sustained, confinement controlled yield behavior while maintaining stable ground conditions, which is essential to underground mining operations."

Jan 1, 2016

By Stephen C. Tadolini, Steve Bessinger, James Pile, Christopher Mark

The San Juan Coal Mine, located near Farmington, New Mexico, supplies the San Juan Generating Station with more than 6 million tons of coal annually. To replace dwindling surface mine production, San Juan installed a longwall in 2002. The roof rock in much of the underground reserve consists of interbedded coals, carbonaceous shales, and mudstones. Roof support consists of fully-grouted roof bolts, both tensioned and non-tensioned, supplemented by modified cable trusses. Short-encapsulation pull tests conducted by San Juan’s geotechnical engineering department showed that some of the immediate roof layers provided variable bolt anchorage. To ensure that the roof bolts were obtaining the necessary anchorage, additional tests were conducted in a variety of locations. The results have been used to determine the optimum support design for different areas of the mine. Another series of tests evaluated the performance of several bolt and hole diameter combinations. The experience at San Juan Mine shows that short-encapsulation pull tests can provide the information necessary to improve ground control in variable roof conditions. It also illustrates how a pro-active geotechnical engineering program can be an essential component of a state of the art underground coal mine.

By Hansheinz Portisch, Harold Margolin

A surface expansion was found to occur sometime after etching in Ti-A1 alloys containing 9.5 to 12.5 wt pct Al. The structure formed, grew, and disappeared with tzrrze. The surface expansion was followed by microscope observations and interferometric and lattice parameter measurements. Activation energy measurements for the growth of the "expansion structure" and chemical analysis indicated that the phenomenon occurred as a result of hydrogen pzckup during etching. It is proposed that hydrogen initially enters octahedral sites of Ti3Al coherent with a Ti and later shifts to the tetrahedral sites. It is postulated that expansion occurs when hydrogen enters the tetrahedral sites. The expansion structure disappeared, it is proposed, because of diffusion of hydrogen from the surface into the body of the alloy and because of loss of coherency of Ti3Al. In examining Ti-A1 alloys, Ence and arolinl observed markings on the surface of specimens. These markings did not appear after electropolishing only, but rather appeared only after etching. The markings appeared to grow as a function of time after etching and later seemed to disappear. Although the markings had some similarity to precipitates, showing, frequently, a Widmanstatten type of arrangement, the observation that other microstructural markings continue to be seen within the new structure suggested that it was actually not a precipitate. The source of this structure was unknown and an attempt was made in the present investigation to develop some understanding of its nature. It has been labeled expansion structure. 1) EXPERIMENTAL PROCEDURE A) Alloys. Alloys in the range 6.5 to 14.5 wt. pct A1 were studied. Bars, 4 in. sq, forged from Bureau of Mines (73 Bhn) titanium consumably arc-melted 4-lb ingots,' as well as 50-g arc-melted buttons of desired aluminum contents were used. The 50-g buttons also used Bureau of Mines titanium (73 Bhn) and aluminum of 99.99 pct purity. Buttons containing up to 8.5 wt pct A1 were hot-rolled from a furnace at 900' . Those with higher aluminum contents were hand-forged on a titanium anvil, and heated with an oxygen-hydrogen torch in the region of 1200" to 1300°c. Frequent reheating kept the samples at the desired temperature range. After a reduction of about 30 pct, the samples were water-quenched. To eliminate any contamination picked up either during hot rolling or forging, at least 1 mil of the surface of the sample was taken off. B) Heat Treatment. Prior to heat treatment all alloys were vacuum-annealed to remove hydrogen. Samples were annealed at 900' until a vacuum of 10'5 mm Hg was established. After this treatment, the samples were wrapped in molybdenum sheet and heat-treated in argon-filled quartz capsules, which were broken under water or iced brine at the conclusion of the heat treatment. All heat treatments under dynamic vacuum were performed in a rapid-quench furnace. This consisted of a molybdenum-lined quartz tube attached to a vacuum system and through a stopcock to a beaker of water. At the completion of the heat treatment the vacuum stopcock was closed, the furnace shut off, and dropped below the quartz tube. Then immediately the inlet stopcock was opened and water admitted until the tube was filled. The steam formed was allowed to escape through the inlet stopcock. This method was used in heat treatments up to llOO°C. C) Metallography. Specimens for metallographic examination were ground, then electropolished using a Disa Electropol machine with a perchloric acid electr01te. Specimens were etched with R-etch.3 A standard etching time of 3 min was used, with the specimen being agitated during immersion. D) Sample Preparation for X-ray Analysis. Samples 0.2 to 0.5 mm in diam were produced from heat-treated rods which were turned to 2 mm diam on a lathe and then rotation-etched. An etchant consisting of 1 pt HF, 7 pt HN03, and 12 pt HzO was satisfactory. With the rod rotating in a vertical position at 50 to 100 rpm, a needle with uniform dimensions could be obtained. Care had to be taken to insure that the rod was at the center of rotation, otherwise cavitation developed. If a small unevenness developed, it was possible to grind it off with a fine emery paper. E) X-ray Diffraction. All X-ray work was done on a North American Phillips X-ray diffraction apparatus and a Jarrell Ash microfocus X-ray unit. The Phillips unit was used with a copper target and nickel filter. The Jarrell Ash unit was fitted with a cobalt target and iron filter. For the Phillips unit 35 kva and 20 ma were used, whereas for the microfocus unit with the 100-p fixed-focus gun 40 kv and 1.5 ma were used. It was found that alignment of cameras on the Jarre11 Ash unit was very critical. The X-ray beam contains an intense area which is not the beam center. The cameras were aligned with the intense region by monitoring the beam coming out of the camera with a Geiger counter. Adjustments were made until a maximum intensity was obtained. A Phillips diffractometer with a Brown chart recorder furnished some of the lattice parameter data. The divergence slit up to 80 deg 28 was I deg; above 80 deg, a 4-deg opening was used, while the scatter slit was1 deg and the receiving slit had a 0.003-in. opening. The general scanning rate was 1 deg per min, while peaks of special interest were rescanned at -j deg per min. For elevated-temperature X-ray diffraction a Uni-cam High Temperature Camera with a film diameter

Jan 1, 1969

By E. H. Mohr

AT the McGill concentrator of the Nevada Consolidated Copper Co., all flotation operations have been carried out in Forrester cells since November, 1926. In respect to cost of operation, the new cell has proved to be superior to the Callow, which it supplanted, and its incorporation in the flow sheet has resulted in a marked improvement in the metallurgy. Before citing the advantages of the new machine over the old, it should be explained that the Forrester has not only replaced the Callow, but it has made possible the elimination of a number of tables as well. When Callow cells were in use, table concentration preceded flotation. The ball mill pulp was elevated, by means of air lifts, to roughing Wilfley tables situated two floors directly above the flotation floor. On an intermediate floor were additional tables which cleaned the concentrate from those above. The finished table concentrate constituted approximately 60 per cent. of the total mill production, and contained about 40 per cent. of the recovered copper. The table tailings were returned to the classifiers, the overflow from which went to the flotation cells and the oversize to the ball mills. As soon as the Forrester cells were installed, this tabling operation was discontinued. The ball mill pulp now discharges into the classifiers direct, without the intermediate tabling. It was impracticable to employ the present flow sheet when Callow cells were in use, for the reason that unless the greater part of the heavy mineral was first removed, it accumulated on the mats and adversely affected the flotation operation. By tabling ahead of flotation, it was practicable to maintain the concentrator at 10 to 20 per cent. greater capacity (depending up on the pyrite content of the ore) than was possible by flotation alone. This was due to the necessity for finer grinding in the latter case. Even though the table concentrates contained only 6 to 8 per cent. copper con- tent, the low smelting costs made it advantageous to operate at high mill capacity, at the cost of low grade of product. STANDARD SECTION EQUIPMENT With the old-flow sheet, a standard section consisted of four ball mills and classifiers, one air lift, 20 roughing tables, 10 cleaner tables, 12 primary

Jan 1, 1928

By Howard Martens, Pol Duwez

XPERIMENTS were performed on a Cr-Fc alloy Econtaining 44.7 pet Cr in order to determine the heat evolved during the transformation of the a solid solution into the s phase, and the change in volume accompanying the transformation. Since previous measurements on the rate of formation of a from s indicated that the reaction was quite fast as soon as the equilibrium temperature of 820°C was exceeded by some 50°, standard thermal analysis experiments were performed on heating specimens of s at the rate of approximately 40°C per sec. A definite thermal arrest was observed at 870"C, Fig. 1. This result indicates that the reverse seaction of into a is exothermic. From the heating curve of Fig. 1, taking 15°C per sec as the heating rate in the neighborhood of the transformation, 2.4 sec as the time for transformation, and assuming that the specific heat of the alloy at 850°C is an average between iron and chromium, namely 9.5 cal per mol per "C, the latent heat of transformation is approximately 342 cal per mol. The order of magnitude of this latent heat was also obtained by comparing the heating curve of Fig. 1 with a similar curve obtained with pure iron under identical experimental conditions. It was found that the length of the thermal arrest for s was about 1.8 times as long as that for iron. Since the latent heat of the a to 7 iron transformation is 203 cal per mol1, 2 the latent heat of it to a would be 365 cal per mol, which is in relatively good agreement with the value determined from the heating curve. For the determination of volume change, thermal expansion curves were recorded using a specimen containing 44.7 pet Cr and completely transformed to cr by previous treatment. The curves were recorded at a rate of heating of about 4°C per min and at approximately 840°C a sharr, increase in length of about 0.08 pet took place. Fig. 2. It is therefore established that the specific volume of a near the transformation temperature is about 0.24 pet smaller than that of a, and consequently, when the s phase forms in an a matrix, it would tend to shrink away from the matrix. Since the s phase is known to be extremely brittle, these results offer an explanation for the microscopic cracks often observed in the a phase under the microscope. The results of this study are in qualitative agreement with previously published data on stainless steel 19-9DLa and on a Cr-Fe alloy containing 45 pet Cr.' The increase in density due to the transformation from n to s given in ref. 4 is 0.90 pet. These measurements, however, were made at room temperature. whereas the presently reported 0.24 pet is the change at the temperature at which the transformation takes place during heating. In the graph of Fig. 2, the percentage change in length that would be observed at room temperature is given by the length X, which is about 0.27 pet. This would correspond to 0.81 pet in volume change. Acknowledgment This work was done at the Jet Propulsion Laboratory, California Institute of Technology, under Contract No. DA-04-495-ORD-18 with the Army Ordnance Corps, Washington, D. C. The authors wish to thank this agency for permission to publish the results of this investigation. References I H. Klinkhardt: Measurements of True Specific Heats at High Temperatures. Annalen der Physik (1927) 84, pp. 167-200. - H. Esser and W. Runghardt: Influence of the Method, of Dctcrmination on the Measured Heat Content of Iron. Archtvo ffur Dct Eisenhiittenwesen 11934) 8, pp. 37-38. J. J. Gilman, P. K. Koh, and 0. Zmeskal. Delta Ferrite Formation and its influence on the Formation of Sigma in a Wrought Heat Resisting Steel. Trans. ASM 1 1949) 11, pp. 1371-1388. 1 J. M. Thomas: Discussion of ref. 3. Trans. ASM (1949) 41, pp. 3 390-1392.

Jan 1, 1957

By Charles A. Beasley

Increased technical and legislative analysis by all parties of interest could result in more efficientand better directed regulatory legislation in the mining industry. Efforts to forestall or frustrate legislation through massive lobbying and public relations efforts are seldom effective over time and often lead to more stringent and far reaching legislation than necessary. The Federal Coal Mine Health and Safety Act of 1969 is perhaps an outstanding example of such insufficiences; a future example may well be regulatory surface mine legislation. Although a lack of extensive hard data seldom enables rapid or sophisticated analysis early in the legislative process, even preliminary analysis may be sufficient. Simple statistical analyses of available data, assessment of the legislative rationale, and subjective appraisal of other non-quantifiable constraints to improved mine safety. An average of 115 underground fatalities at an exposure frequency rate of 0.67 can be reasonably expected in the immediate future, under existing levels of technology, degree of regulation, and economic conditions. If gross unrefined estimates of 6 preparation plant, 5 surface, and 30 strip mining fatalities are assumed, the total fatality industry level could be expected to approximate 156 Serious professional, economic, and technical constraints within the the industry probably preclude any major decrease of these accident levels within the near future

Jan 1, 1972

By E. Stansfield

Coal samples are sometimes taken from a wet spot in a mine, or from an open car or pile shortly after rain, and submitted for analysis in a dripping wet condition. Other samples are taken from a dry spot in the mine, or from an open car or pile after a period of dry weather. Comparison of the analyses obtained in the two cases would be absurd, although the samples might be of essentially the same coat apart from the surface moisture on the one. The analyst therefore attempts to reduce all samples to an equal basis by the operation known as air-drying. Air-drying is practised by the chemist for three reasons: (1) To make it possible to grind the coal to powder. (2) To bring the coal sample to a condition in which it will not change in weight rapidly whilst portions are being weighed out for analysis -in other words, to bring it into equilibrium with the air in the laboratory. (3) To determine the water-retaining power of the coal. Any kind of drying is satisfactory for (1); the coal should be dried more or less according to the humidity at the time in the laboratory for (2); whilst it should be always dried to the same extent for (3). In other words, no method can be satisfactory for both (2) and (3). In this paper the only question considered is (3) -that is, air-drying the coal under standard conditions so that strictly comparable results can be obtained for the water content of coals after air-drying. After a sample of wet coal is exposed, it loses moisture until it is in equilibrium with the air at the time: thus the coal will loose distinctly more water in air of 20 per cent relative humidity, for example, than it would in air of 80 per cent.

Jan 1, 1923