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By William E. Dennis

"At the request of Philip S. Hoyt, the Eagle Mountain fluorspar deposits in section 35, block 68, Texas & Pacific Railroad Survey No. 9, were examined by an engineer3/ of the Bureau of Mines during March 1943: As a result of this examination; a preliminary program of trenching, sampling and core drilling by the Bureau of Mines was begun in. June 1943.4/ Eight deposits were explored.The Federal Geological Survey mapped the properties and furnished geologic information throughout the program, while the Texas .Bureau of Economic Geology6/ made preliminary maps that facilitated planning the exploratory work on section 35.The eight deposits explored by the Bureau of Mines are in block 68, Texas & Pacific Railroad Survey No. 9, Hudspeth County, Tex., approximately 7-1/2 miles south of. Hot Wells and 12-1/2 miles south of Allamoore, which is 10 miles west of Van Horn and 20 miles east of Gierra Blanco on U. S. Highway 80(fig. 1). A graded gravel road, built by the Public Roads Administration, leads from Hot Wells to the deposits in section 35, and there are unimproved roads, passable,by automobile, to all the other deposits.Carload shipments can be made to and from Hot Wells on the Southern Pacific Railroad, and complete railroad facilities are available at Allamoore on the Texas & Pacific Railroad.Van Horn and Sierra Blanca are the only local source of labor and. supplies. The prevailing wage for common labor is 60 cents an hour. Living accommodations are obtainable in Horn and Sierra Blanca.No electric power lines pass near the deposits. Diesel and gasoline motors are employed to furnish power for mining and milling operations.Two recently 'drilled water wells on section 35 furnish water for the mill. Other wells will be drilled if the present supply is inadequate for prolonged Peak operation.The properties lie in the foothills on the northeast side of the Eagle Mountains at 5,500 feet altitude. The Eagle Mountain Range rises sharply from a broad, flat valley known as Eagle Flat. Eagle Peak, the highest in the range, has an altitude of 7,510 feet. In the vicinity of the fluorspar deposits the topography is moderately reliefs ranging from 50 to 300 feet."

Jun 1, 1946

By P. J. Mackey

"The metal producing industries in Canada and in the U.S. in many ways owe their technical beginnings to Welsh technology. The first copper smelter in Canada started in 1849-50 with a Welsh reverberatory furnace and skilled furnacemen brought across from Swansea to operate the equipment1- The Rocky Mountain smelting industry began with the installation of furnaces resembling the Swansea smelters(2). It was only later that the North American industry developed newer and improved methods. This article briefly describes the metallurgical history of the Lower Swansea Valley, being that section of land stretching northward from Swansea along the banks of the River Tawe, and bordered by hillsides east and west of the river. The Valley has had a classic cycle of development and decline common to industrial areas of Britain which grew in the Industrial Revolution. Perhaps more than any other area today, the Swansea Valley represents a complete cycle of use from agricultural to the peak of industrial development through to decline and dereliction. The conditions which made the Tawe Valley attractive to certain industries were laid down nearly 300 million years ago with the Carboniferous swamps and forests which covered the area, later forming the south Wales coalfield. As a result of glacial action and the deposit of glacial outwater and alluvial gravels, the valley adopted the traditional configuration, with its rocky sides occasionally remaining exposed. Up to probably about the 16th century, the Lower Swansea Valley was similar in use and appearance to many other valleys in the region. From the 14th century, the natural oak woodland had provided a continuous source of timber to produce charcoal for iron making as well as for structural purposes. Agriculture was the main activity in the valley in this period, but it was likely that the farmers also mined small amounts of coal by way of ""bell pits"", and adits into the coal seams."

Jan 1, 1999

By Ralph T. Buscarello

For practical purposes, the terms "natural frequency," "resonance" and "critical speed" are synonymous. A part passing though its "critical" or "critical speed" may also be said to be "resonant" or "resonating." The vibrations of a precisely balanced part can be magnified by the structure in which it is assembled. Anyone who has driven an automobile knows that it will vibrate more at a certain speed than at others. According to the formula for centrifugal force, vibration should increase steadily with a rise in rpm. What happens instead is that the vibration not only in- creases with r m. but suddenly rises at a much higher rate through the critical speed, and then apparently smooths out as the rpm passes beyond it. To visualize what happens, consider a simple flat spring with a weight mounted at one end. When the spring is deflected by pulling down on the weight and then letting it go, the spring will oscillate, and the spring-and-weight system will vibrate at its natural frequency. If the timing or direction bf the impulses imparted to the spring did not coincide with its natural frequency, the result would be an irregular vibration that would not build up. If, on the other hand, the timing and direction of the impulses did coincide with the spring's natural frequency, the result would be a larger and larger amplitude. If the impulses were continued indefinitely, the amplitude of the spring's vibrations would go on rising [while its frequency stayed the same] until the spring broke or the amplitude of the vibrations reached an equilibrium [because frictional or other forces would tend to damp it]. Resonance magnifies the amplitude of vibrations in a relatively undamped system between 100 to 200 times over that of nonresonant vibrations. Frictional or viscous damping often reduces the magnification [from 10 to 20 times], but even with this reduction the amplitude would still be large enough to cause excessive wear, and sometimes even fractures. For example, a blower might show only 0.001 in. of vibration on the manufacturer's test stand. Installed, the blower could pass into its critical speed [the speed at which the frequency of the vibrations caused by the rotor coincides with the supporting structure's natural frequency], and the 0.001 in. of vibration could be resonated to 0.010 in., or as much as 0.020 in. The cure for such a condition would not be to lower the vibration of the blower by further balancing or better alignment but to separate the two frequencies by at least 20 to 30%.

Jan 1, 1970

By C. D. Mann

INTRODUCTION The sublevel stoping mining method is usually applied to a relatively steeply dipping, competent ore body, surrounded by competent wall rock. Ore is produced by drilling and blasting longholes, which can range from 50 mm (2 in.) to 200 mm (7% in.) diam, with lengths up to 90 m (300 ft). Longholes can be inclined in any direction, but the ring or pattern usually forms a plane, and the holes are blasted as a unit. Recently developed mobile drilling and loading machinery, as well as new explosives products, blasting techniques, and cemented sand and rock fill have made sublevel stoping a highly efficient and versatile mining method. When designing a sublevel stoping production sys- tem, it should be kept in mind that production rates from conventional sublevel stopes vary widely through- out the life of the stope. Early production is at a low rate, coming only from the drawpoints near the slot, but increases as new drawpoints are reached by the stope face. As the stope nears completion, again, fewer drawpoints are productive. Enough drawpoints must be available at any time to provide required production. Drawpoint availability should be compared to equipment availability; plan for more drawpoints than are needed at any one time. Accurate, realistic scheduling is essential to smooth production rates. Also, initial recovery of ore in a stope/pillar block is normally from 35% to 50% in sublevel stoping. Planning of pillar recovery, representing the majority of ore tonnage in a production block, must be done during early mine planning. Since much of the development already done for primary stoping (access for drilling, drawpoints, and haulageways), can be used for pillar recovery, early production from pillars is highly desirable. The following description of components of the system is an attempt to highlight some of the most important features and requirements of mechanized sublevel stoping methods. Similar comments would apply to the use of older equipment (column-and-arm drill setups, slushers, etc.) in similar methods. As in any good mining system, maximum economic recovery of the resource in the ground is the primary consideration. STOPE DESIGN CHARACTERISTICS Length and Width The following are some of the factors which affect sublevel open stope length and width dimensions: ore body geometry, principal stress directions, competence of stope back, optimum drill pattern, and drilling drift layout. In new mines initial stope layout design may occur before the ore body is actually intersected by mine workings. Stope dimensioning is a critical decision, and assistance from as many knowledgeable people as possible at this stage is essential. Operators with past experience in similar ore bodies, rock mechanics experts, and others with mine design experience should participate at this stage of stope planning. Height The following are some of the factors which must be considered in determining stope height: competence of stope pillar and stope/fill walls; slenderness ratio of adjacent pillars; ore body dip; ore body thickness; hole depth capability of the drilling machine; fragmentation characteristics of the ore; and level intervals in existing mines. In competent ground, drill-hole length and accuracy are the most important determinants of stoping height. Frequently entire drilling sublevels can be eliminated because of the depth capability of sophisticated drilling equipment, resulting in significant development cost savings. Drawpoint Location and Design Some of the most important considerations of a good drawpoint system are optimum spacing of draw- points, within the constraints of stope dimensions, for uniform drawdown and maximum recovery; excavations designed for stability for the life of the ore block to be drawn-primary stope ore as well as subsequent pillar ore; floor or roadway design including type of surface, reinforcing, grade for water runoff; orientation with respect to the main haulageway, for optimum loader maneuverability and ground stability at the inter- section; and length, to allow articulated front-end loaders to work in a straight configuration. Careful drawpoint design and construction are keys to successful production. Extra care in development, such as smooth wall blasting, rockbolts or grouted rebar, wire mesh, and shotcrete usually will ensure long draw- point life. Human exposure during production loading is of longer duration than during development or production drilling, and consequently preparation of draw- points is easily justified, particularly when pillar ore can be drawn through the same drawpoints. Secondary blasting of boulders can weaken drawpoints, also justifying good ground control techniques. A smooth draw- point floor of poured, reinforced concrete, on a grade of +3% or +4% toward the ore pile facilitates water flow out of the drawpoint, and ease of loader bucket penetration into the muck pile. Slot Raising, Slotting A slot or other space for rock expansion is necessary in conventional sublevel stoping where vertical rings or rows of holes are blasted. The slot can be started at a slot raise driven by conventional raising methods, raise boring, drop raising (predrilling and blasting a raise from the top, using small diameter-less than 200-mm (7%-in.)-holes for relief), or crater blasting (similar to drop raising, but without relief holes). The slot usually extends from the extraction level to the back of the stope. It is normally expanded to full stope width by

Jan 1, 1982

By W. R. Reed, J. P. Rider

"The Pittsburgh Mining Research Division of the U.S. National Institute for Occupational Safety and Health (NIOSH) recently developed a series of models using computational fluid dynamics (CFD) to study airflows and respirable dust distribution associated with a mediumsized surface blasthole drill shroud with a dry dust collector system. Previously run experiments conducted in NIOSH’s full-scale drill shroud laboratory were used to validate the models. The setup values in the CFD models were calculated from experimental data obtained from the drill shroud laboratory and measurements of test material particle size. Subsequent simulation results were compared with the experimental data for several test scenarios, including 0.14 m3/s (300 cfm) and 0.24 m3/s (500 cfm) bailing airflow with 2:1, 3:1 and 4:1 dust collector-tobailing airflow ratios. For the 2:1 and 3:1 ratios, the calculated dust concentrations from the CFD models were within the 95 percent confidence intervals of the experimental data. This paper describes the methodology used to develop the CFD models, to calculate the model input and to validate the models based on the experimental data. Problem regions were identified and revealed by the study. The simulation results could be used for future development of dust control methods for a surface mine blasthole drill shroud. IntroductionIn surface mines, production drilling is an essential process in blasting to fracture the hard rock overburden for removal. During the process, considerable amounts of respirable dust are produced. Past sampling at the shroud area of blasthole drills had documented time-weighted-average respirable dust concentrations ranging from 8.68 to 95.15 mg/m3 (Organiscak and Page, 1995) and 1.04 to 52.30 mg/m3 (Listak and Reed, 2007). These high dust concentrations, which may contain silica, can lead to respirable dust overexposures for the miners working nearby. These overexposures can lead to silicosis, an occupational lung disease that has no cure and is ultimately fatal.Two basic methods are used to control drilling dust: a wet suppression system or a dry collection system (Cecala et al., 2012). Dry drilling techniques using a dust collection system are preferred in rural mining locations where a constant water supply is not available, and in the case of rotary drilling where water is associated with accelerated bit wear due to bearing degradation and hydrogen embrittlement. In addition, dry drilling eliminates freezing-related issues with water in colder climates. Dry dust collectors are highly effective at removing dust, especially the finer material, as long as they are properly operated and maintained (Listak and Reed, 2007; Organiscak and Page, 2005)."

Nov 1, 2016

By V Glenn, P McKenna, H Vickers, P Audet, M J. Gillespie, A Gravina, D R. Mulligan

A requisite corollary of mineral sand mining on North Stradbroake Island (Queensland (Qld), Australia) has been the progressive rehabilitation of post-mined lands for the revegetation of native subtropical Eucalypt communities. Since 2005, rigorous environmental monitoring has been undertaken to verify the annual development of post-rehabilitated sites. From these long-term analyses, a chronosequence of land revegetation (ranging from three to 26 years postrehabilitation) was developed to examine the rehabilitation trajectories of various post-mined sites in relation to proximal undisturbed locations. Here, we identified that some sites may be deviating from the intended natural analogues as evidenced by the opportunistic colonisationand otherwise æinvasiveÆ behaviour of a regionally native species (Black Sheoak - Allocasuarina littoralis), which appear to be displacing the intended Eucalypt varieties. Although it is recognised that the broader data trends correspond to distinct periods of adaptive modifications to the implementation of rehabilitation (ie pre- and post-1995), this analysis focuses on further identifying potential above- and belowground feedback mechanisms underlying these divergent ecological outcomes.CITATION:Audet, P, Gravina, A, Glenn, V, McKenna, P, Vickers, H, Gillespie, M J and Mulligan, D R, 2012. Above/belowground feedback may facilitate and sustain mono-dominance on rehabilitated North Stradbroke Island, in Proceedings Life-of-Mine 2012 , pp 15-24 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 10, 2012

By H. J. Rose

Sixty-five million net tons of coal were carbonized in the by-product and beehive coke ovens1 of the United States during 1924. This tonnage represented 13.4 per cent. of the bituminous coal which was mined in that year. Table 1 shows production and disposition.2 Table 1.—Production and Disposition of Ocen Coke in 1924 By-product Beehive Total Coal charged into ovens, tons................ 49,061,339 15,914,310 64,975,649 Coke produced (exclusive of screenings and breeze), tons............................. 33,983,568 10,286,037 44,269,605 Average yield of coke, per cent ............. 69.3 6.46 68.9 Disposition of coke, per cent.: Blast furnace.......................... . . 82.6 Foundry............................... . 13.3 6.6 Domestic.............................. . 1.4 6.7 Water gas manufacture and all other uses.. 4.6 . 4.1 Cokes having somewhat different combinations of chemical and physical properties are desired for the various uses listed above, and these properties depend primarily upon the types of coal used. American coals have a wide range of purity and coking quality, and the chief subject of this paper will be the selection of coals for the manufacture of blast furnace, foundry, domestic, and water-gas coke, in by-product coke ovens. It will obviously be necessary to include a discussion of those physical and chemical characteristics that are generally considered essential or desirable for each type of coke. Suitable Coals The examination and development of new sources of coking coal are important because of the irregular and limited distribution of high-grade

Jan 1, 1927

By Kousik Sinha

The advantage of using a clean coal as a feedstock in a direct liquefaction plant is improved performance of the plant by several factors e.g., increased throughput of coal; reduced wear and tear of valves, pipes, and pumps; increased liquid yields, etc. This paper examines the technical feasibility for cleaning Illinois #6 coal as a feedstock to a direct coal liquefaction plant via selective agglomeration using a coal-derived heavy distillate, Wilsonville V-1074 as agglomerant. Effects of addition of Kerosene or coal-derived Naphtha as conditioner, coal particle size, slurry pH, oil to coal ratio, and operating temperature on ash reduction and clean coal weight recovery have been studied. Results show that about 70% of the ash can be removed from this coal and the clean coal weight recovery over 85% can be achieved by using this technique with suitable selection of coal particle size, slurry pH, mixing intensity and oil to coal ratio.

Jan 1, 1991

By Paul H. Johnson

The CH Leaching process is a new, more sophisticated, form of gold, silver, uranium and sulfide copper heap leaching wherein each layer of a multi-layered heap is divided into a number of independent cells, each of which are successively formed and leached in the top layer of the heap to their maximum economic recoveries, and then rinsed and sealed from further leaching. Triangular earth berms form the outside walls of the heap and a gridwork of rock-filled berms on top of the bottom pad form independent under- heap solution storage zones. The high degree of solution recirculation and staging that occurs in this process produces a relatively small volume of very rich pregnant solution. The cellular heap design and its mode of operation result in a uniform, intensive, long-term leach for all of the ore particles in the heap. Besides improved metal recoveries, the process appears to have several cost and operating advantages. In the CR2MR process, sulfuric acid and a soluble sulfide are added to a gold and/or silver pregnant leach liquor to precipitate most of the contained silver, mercury, copper, etc. as sulfides at a pH below 2.5. The other product of the reaction is solubilized HCN. Following a liquid-solid separation step, cyanide is regenerated in the supernatant solution by raising the pH to a level above 9.0. This process regenerates a portion of the cyanide used in the leaching step, removes contaminant ion species (silver, copper, mercury, etc.) from the leach solution and, in a subsequent step, recovers the gold in an enriched form. It is believed that substantial cost savings may be experienced by the use of this process in some gold and/or silver leach systems.

Jan 1, 1984

By Irshad A. Rana

In a modern copper smelter, the capital cost of a sulfuric acid plant constitutes the largest single cost item of the total investment. It is thus essential to properly size the acid plant in order to reduce its capital cost and enhance the economic viability of a new smelter project. Whereas the off-gas from a smelting furnace is continuous and generally uniform, the gas flow from the Peirce-Smith converters is intermittent with significant fluctuations. The converters off-gas volume is dependent upon a number of interrelating factors. The most significant factors are the matte feed grade, the system air infiltration, the converter's size and blowing rate, the maximum number of converters simultaneously blowing and the gas cooling system employed. The acid plant size, therefore, is heavily influenced by the converter off-gases. Several of these factors are discussed, and their impact in determining the acid plant size is explored.

Jan 1, 1993

By Orica Peru, Pedro Lozada

Maintaining a continuous mining process during thunderstorms is one of the primary challenges in Peru's open-pit mines. Safety protocols necessitate the establishment of exclusion zones to minimize the risks of unplanned detonations. During such conditions, loading and hauling equipment must be relocated or, in the worst case, halted, directly affecting productivity (t/h). Implementing a wireless initiation system mitigates the risk of unplanned detonations during electrical storms by eliminating the use of electrical cables that are susceptible to electrical discharge. This study documents the implementation of this system in a copper mine located over 3,500 meters above sea level in Peru, where electrical storms between December and March necessitate modifications to blasting schedules and limit the number of drills fired per event. Through a test using wireless detonators, a key performance indicator (KPI) was established to ensure safe and continuous operation during thunderstorms. The results were significant: 20 blasts were successfully executed, firing approximately 5,000 holes and moving around 10 million tons of material, all without safety incidents or critical operational deviations, achieving zero misfires and no scheduled blast suspensions. Shovels SH001 (ore) and SH002 (waste) continued to operate together for an additional 100 hours continuously during thunderstorms, generating a potential saving of USD 3.3 million by continuously sending ore to the plant during the testing period under stormy conditions. Additionally, the blasting initiation time was reduced by 25%, eliminating two activities that involved the use of surface connection cables. This reduction led to over 150 hours of decreased personnel exposure to interactions with vehicles and environmental hazards such as dust and noise. Furthermore, the number of holes fired in a single event increased by 50%, enhancing the blast tonnage capacity and reducing equipment stops due to evacuation and reentry from two to one.

Jan 21, 2025

By C. Y. Sun

China is a country with abundant resources of tungsten, taking one of the world?s top-positions regarding both the total reserve and the amount of industrial production. China?s tungsten resources occur mostly in multi-metallic deposits rather than in single-metallic deposits. China?s multi-metallic tungsten ores have the characteristics of coexistence with other valuable components such as molybdenum, bismuth, copper, sulfur and fluorite. These ores are usually low in grade and difficult to process. The technological challenge in mineral processing lies in separation of multi-metallic components in these complex ores and a comprehensive utilization of natural resource. Several cases of industrial practice of processing China?s typical multi-metallic tungsten ores are presented.

Jan 1, 2014

By Les A. Beckett

The Bureau of Mines has developed MULSIM/BM, an improved computer program using the boundary-element approach for geomechanical analysis of mine plans in single or multiple coal seams. This Information Circular presents the improved features of the MULSIN/BM program. The four major improvements are the following: (1) Ability to specify up to 26 material property sets. (2) A new material model capability for gob and inserted materials such as packwalls or cribs. (3) Ability to model a large area with a grid of undivided blocks and to subdivide only a specified fine-mesh portion of these blocks into elements. (4) Specification of extraction ratios for the undivided blocks. In addition, an interactive graphics mesh generator program was written to accompany MULSIM/BM. A complete user's guide and program listings are included within the appendixes. The MULSIM/BM and mesh generator programs are available on tape from the National Technical Information Service.3

Jan 1, 1988

By Peter Zhang

Room and pillar mining under shallow depth is usually conducted with a high extraction ratio, which leaves small pillars with a low safety factor in some old workings in abandoned mines. Pillar sloughage, squeezing or even collapse may have occurred in those old workings with small pillars, either at the time of mining or later. The conditions of the pillars in abandoned mines can affect mining above them. Subsidence fractures and abutment pressure are indications of multi-seam interaction caused by squeezing or collapse of small pillars. To mine over areas with potentially unstable small pillars, the panel layout in the upper seam should be such that multi-seam interaction can be minimized and proper roof and rib support should be installed in the influence zones. This case study presents an example of multi-seam mining above small pillars under a shallow depth of cover. Though the interburden is relatively thick and competent, interaction is still present over the influence zones. Pillar stability and multi-seam interaction analysis, numerical modeling and in-situ monitoring are presented to demonstrate how the upper seam can be mined safely over the small pillars.

Jan 1, 2014

By J. D. Embury, R. M. Fisher, A. S. Keh

An investigation of the defect structure and properties following large strain deformation has been carried out using transmission electron microscopy and mechanical testing for a range of ferrous materials and for copper. It is shown that, for a variety of ferrous materials, a cellular substructure is developed during the initial stages of working and on further deformation the dimensions of this substructure are reduced. Quantitative measurements of the flow stress and the scale of the substructure indicate that the strength of heavily worked materials is largely determined by the spacing of cell walls. These cell walls act as dislocation barriers in a manner analogous to grain boundaries. FCC materials do not harden as extensively as bcc after cold working and the present observations on copper indicate that the dimensions of the substructure in fcc materials are not markedly reduced on deformation. The differences between the fcc and bcc structures produced by large plastic strains are ascribed to differences in the extent of dynamic recovery. It is tentatively proposed that the greater stability of substructural barriers in bcc structures results from strong interstitial Pinning effects. It is proposed that for large plastic strains the work-hardening process may be considered in terms of the reduction of substructural-barrier spacing during the working process. This pvovides a simplified hut useful analysis of the strain hardening occurring during mechanical processes. THE mechanisms of strain-hardening processes have occupied the attention of numerous investigators over the past three decades and a multitude of theories have emerged in this field. The majority of previous work has been concerned with the development of a description of the plastic behavior of single crystals in terms of dislocation theory.'" In contrast, scant attention has been given to the strengthening mechanisms operative during mechanical working processes such as rolling and drawing. The large plastic strains involved in these processes make it difficult to correlate the observations with any elementary dislocation theory. Further, as Bullen and coworkers3 have opined, for large plastic strains both hardening and recovery occur simultaneously and thus it is extremely difficult to estimate the effect of plastic strain on the ambient internal stress field. It is probable that for large plas- tic strains any theoretical estimate of the internal stress field must be made in terms of complex dislocation groups. Although some valuable theoretical work has been done toward calculating the elastic properties of dislocation groups4 any detailed self-consistent model for large plastic strain is as yet impossible. With these limitations in view it is still of value to examine the mechanical properties of highly strained materials as a function of their substructure. This relationship is explored in the present communication for a variety of materials subjected to rolling, drawing, or swaging. The primary object of this investigation has been to establish the validity and the limitations of the substructural strengthening mechanism proposed by Embury and eisher' for drawn pearlitic wire and by Meieran and Thomas for drawn tungsten wire. No attempt has been made to delineate the basic interaction mechanism which occurs between a slip dislocation and the substructure and the authors believe that a great deal more experimental evidence must be compiled before these fundamental aspects can be explored. Also, no specific attention has been given to the detailed formation of substructure during working. Experimental Details. A variety of materials were used in this investigation, the compositions and heat treatment of which are tabulated in Table I. The commercially pure iron is referred to throughout this report as Ferrovac E. The mechanical working processes were all performed at room temperature. The wire drawing was performed on a manual drawing bench. Reductions of between 10 and 20 pct in area per pass were used in both swaging and drawing. The experimental parameter used to evaluate the mechanical properties of the materials was the 0.2 pct proof stress. This was determined for wire samples using an Instron tensile machine at a crosshead speed of 0.02 in. per min. The techniques for preparation of samples for transmission electron microscopy have been described previo~sl~.~ Specimens from wire samples were prepared from longitudinal sections and in some cases from transverse sections. To evaluate the substructural characteristics of rolled materials, thin foil samples were prepared of both the face and edge sections. Results. One of the more familiar ways of expressing work hardening in mechanical working is by the flow stress as a function of reduction in cross-sectional area. Fig. 1 shows such a plot for Ferrovac E eutectoid carbon steel and for copper. The curve for Ferrovac E is characterized by an initial stage of rapid work hardening and a final stage of rapid work hardening separated by an intermediate stage with a low rate of work hardening. This behavior is typical of many ferrous materials and the drawn pearlite wire previously investigated by Embury and Fisher is a more pronounced example. Copper shows an initial

Jan 1, 1967

By E. F. Ferrell

Gold was recovered from scrap electronic 60-40 (tin-lead) solders by drossing with aluminum or zinc at elevated temperatures. The gold, together with other metallic impurities, collected in the dross phase which could then be separated from the molten solder by skimming or filtering. Drossing upgraded the gold concentration by a factor of 10. Further upgrading was achieved by a combination of steps including an oxidation roast and either a combined cyanidation-amalgamation treatment or an aqua regia leach. The gold was then recovered in highly concentrated form from the aqua regia or cyanide solutions by cementation and from the amalgam by volatilization or dissolution of the mercury. The solder was purified by reaction with NaOH to remove dissolved aluminum or with SnCl2 to remove dissolved zinc.

Jan 1, 1976

By Lj. D. Andric

By the micronization-milling of the material, the size and the structure of its particles are changed and it comes to the increasing of the potential energy of the system as well as to the chem-ical activity of the material what is the result of an increased free surface and of the concentration of the grate defects, that is of the internal energy of the system. All these changes that are reached, show that new characteristics of treated materials are got by micronization process. Controlling the process of the micronization and by purpose directing of the new characteristics of the activated materials, we can have the positive effects and use them in more technologies. Talc is important nonmetallic mineral raw material that is widely used as either the basic raw material, or as filler. This has stimulated its technological investigation, as well as the investigation of its physical-chemical characteristics. In this paper are shown experimentally reached results of the physical-chemical characteristics of the talc during the process of the micronization in high energetic mill. In this paper, the optimum conditions of miconization of talc in high-energy-speed rotary mill and its influence micronization on a degree of Fe2O3recovery by hydrometallurgical methodes, are presented. The X-ray analysis of the obtained products showed that high quality talc with low content of can be obtained by these physical-chemical processes In this paper, the principles and application or fine and ultra-fine grinding high-energy attriton mill with small balls, enabling fast, efficient grinding, low energy consumption, short grinding time and easy handling will be described.

Jan 1, 2014

By Manuel Gomez

Published experimental coke data from a pilot plant slot-type oven were examined by statistical regression analysis. Quantitative relationships were found to exist between the coking rate, bulk density, and fine grinding of the coal, sole-oven expansion, volatile matter, moisture, ash, and carbon con-tent of the coal and coke parameters indicative of size and strength. Regression equations developed from coke data were shown to yield computed values that were in good agreement with experimental results for 11 coke properties. Prediction equations developed were evaluated with an independent set of data. Present results indicate that the variables cited undergo much interaction and the sum of these interactions determines coke size and strength.

Jan 1, 1967

By Shoji Taniguchi, Lifeng Zhang, Kaike Cai

"In the present paper, based on the two-phase (Eulerian-Eulerian) model, the 3D fluid flows in gas-stirred systems, i.e., air-stirred water vessel and argon-stirred liquid steel ladle, are simulated. In the Eulerian-Eulerian two-phase model, gas and liquid are considered to be two different continuous fields. The phases are assumed to share space in proportion to their volume fractions. The exchange between the phases is represented by source terms in conversation equations. Turbulence simulated by the k- e model is assumed to be the property of the liquid phase. The following effects of mathematical treatments and operational factors on the fluid flow are discussed, such as interphase drag force, turbulent model, the size of bubble, and gas injection mode.Some interesting results are derived. Except the interphase drag force, the interphase lift force should be taken into accounted in order to exactly simulate the fluid flow in gas-stirred systems. Suitable turbulent model and drag coefficient have effect on the mathematical simulation. Injecting small bubbles can realize a well mixing flow condition. The distance between the two gas-injection nozzles has effect on the fluid flow."

Jan 1, 2000

By J. P. Capp

The Federal Bureau of Mines and the Alabama Power Company conducted an experiment in underground gasification of coal at Gorgas, Ala. involving the use of hydraulic fracturing to increase the air acceptance of a coalbed for the introduction of combustion air and removal of product gases. Air acceptance is defined as the flow rate of air which can be injected into the coalbed at 65 p.s.i.g. This was the second time the hydraulic-fracturing technique was used at Gorgas. Sand suspended in a heavy oil served as the fracturing fluid in the first experiment; in the second the sand was two to six times larger and was suspended in a water-starch gel. In addition to providing further tests of the underground-gasification process, the purpose of the second experiment was to determine the effects of the larger sand and greater volumes of a water-base fluid on (1) the extent of the fractured area, (2) the air acceptance of the coal, and (3) the rate at which a gasification path could be carbonized (linked) through the coal, and to compare these results with those of the first experiment. The coal was fractured by injecting 11,000 lb. of sand and 33,580 gal. of water and water-gel at 200 to 600 g.p.m. and 900 to 1,000 p.s.i. The fluid fractured the coal and the sand served as a propping agent to hold the cracks and crevices open. Pressures during the fracturing operation at test boreholes indicated that the fracturing fluid extended only about 150 ft. from borehole H, the point of application of the fracture. Injection of air revealed that the air acceptance of the coalbed within this e t was materially increased. On the other hand, increased air acceptance also was noted at boreholes up to 280 ft. from borehole H and at the electrolinking area 300 to 400 ft. away, indicating that the fracture effects extended considerably farther than was evident from observation. Similar effects were noted at approximately the same distances after the first hydraulic fracture.

Jan 1, 1961