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By J. M. Barker, V. T. McLemore

Most discussions of sustainable development (SD) in mining relate to metal mines in rural areas. Industrial minerals (IM) are closer to urban areas primarily owing to a high transport-cost component of delivered cost. This visibility to large populations overshadows IMs relatively small operation size and generally lower waste volume and toxicity. The IM industry needs to rethink its way of doing business to fit into the SD paradigm. Society and business in general will also undergo significant change to enable industrial SD along the lines of triple-bottom-line accounting with government pushing everyone. The benefit to IM mining from embracing SD is a renewal in their social license to operate and a rebirth of mining as a positively-viewed force in society. This positive view would lead to lower acquisition costs because society would not fight so hard to eliminate nearby mining. This would ease the current exploration woes of the IM industry by opening up new deposits for exploration and more readily allowing placement of mines nearer to markets. The earlier in the exploration process SD is initiated, the better.

Jan 1, 2004

By A Tawadrous, P D. Katsabanis, S Kim

A series of small scale tests, have been performed to establish the effect of powder factor and initiation timing on the grindability of rocks. The blasts were performed in high quality granite blocks, which were cut from stone prepared by dimensional stone quarry operations. Three different types of granite, Barre, Stanstead and Laurentian, were tested. The charge consisted of detonating cord placed in five holes with water as a coupling medium, or concentrated charges of equal mass of high exolosive, fully stemmed or air decked. Powder factor was varied from 0.39 to 1.17 kg/m3. Charges were detonated simultaneously or with a small delay and in some experiments, detonation waves collided in the middle of each borehole. All fragments were collected and screened, and drop weight impact tests were performed on selected size fractions, using different levels of impact energy, to calculate the resistance of the rock to impact breakage. The work index of some of the blasted samples was measured using standard Bond mill grindability tests. The experiments generally showed increased softening of the rock with powder factor, manifested by the increasing initial slope of the relationship between t10 and impact energy. The rate of change of the initial slope decreased with increasing powder factor. Changes in the work index were small and consistent in trend with the results of the single fragment impact characterization tests. Charge configuration, timing as well as in-hole collisions of detonation waves affected the initial fragmentation but had little effect on the impact breakage of the samples.

Jan 1, 2009

By W. M. Hart

The remote face loading and entry convergence monitoring systems, employed at some of the Cyprus Amax's underground coal mines, have been able to continuously record changes in face shield resistance across and entry convergence ahead of the longwall face at any moment during longwall mining. The instrumental data obtained from these systems plus in-situ observations during the mining operation can help us better understand the mechanism of overburden strata movement above a longwall face and verify the effectiveness of the current face and entry supports. This paper describes the functions of both the face and entry monitoring systems, and discusses our findings and the issues related to successful ground control for longwall mining.

Jan 1, 1996

By P Viljoen, A Holland, C Parken

In the raisebore industry, there are two known methods of removing large diameter reamer heads once the reaming of an underground shaft is complete: 1. Lower the reamer head to the base of the shaft and use a thermal lance to cut the drill string and drag the reamer away from the hole. This method exposes the person doing the job to falling rocks from the unsupported mineshaft, a hazard which is difficult to control and has caused fatalities. 2. Install breast plates in the backs above the reamer head at the top of the hole, and use air hoists to lift the reamer out of the hole. Steel beams are then placed across the hole which the reamer is lowered down onto, dragged forward away from the hole, disassembled, and removed. This method is time consuming, expensive, and complex in the controls required to ensure safe completion of the job At Carrapateena, the air hoist method had been successfully used, however, with the cost involved and the complexity of this process, development of an improved and fit for purpose reamer removal method was pursued. OZ Minerals collaborated with Pybar to develop an industry first application of Enerpac SL100 lift and shift technology; a track mounted system with hydraulic lifting units that can both lift the reamer out of the hole and tram it back along the tracks away from the hole. A key component of turning the idea into a reality was by establishing a truly collaborative relationship between client and contractor. When the SL100 mobile gantry arrived, it was set up in Adelaide and trialled with the whole raise bore crew attending the two-day workshop and risk assessment; a rare opportunity for drillers. There have now been five successful reamer removals with the SL100 mobile gantry at Carrapateena. It has proven to be an efficient process with the entire process from setup to demobilisation taking less than 48 hours. This new way has completely removed hot works and uncontrolled hazards from working at the bottom of a shaft. It has also saved the reamer stabiliser from being cut and thus wasted. Exposure to the open hole has been minimised significantly as assembly of the SL100 is completed well forward of the open hole. This new process is exponentially safer, quicker and cheaper.

Mar 15, 2021

By Robert Bakish, William D. Robertson

AFTER more than fifty years of investigation and continuing failure in service, a satisfactory mechanism to explain stress corrosion cracking in homogeneous alloys is still lacking. The relative chemical activity of grain boundaries and adjacent grains, inferred from the boundary structure, has been used to explain intergranular cracking in a brass.' This concept is probably correct as a general hypothesis, but it requires much more extensive development in order to explain satisfactorily the observed phenomena. For example, the specific function of the solute in an alloy must be considered, because intergranular failure does not occur in pure metals,' which are structurally similar to homogeneous alloys. Also, in this connection, an inter crystalline fracture path in binary copper alloys can be changed to a transcrystalline path by relatively small ternary additions.' The fracture path in ß brass is predominantly transcrystalline,' in contrast with an inter crystalline path in a brass. Whether the two different fracture paths are the result of a difference in composition, crystal structure, or the state of long range order is unknown and cannot be determined with ß brass, because only the ordered state can be obtained at room temperature. The study of chemical activity at structural sites in the Cu-Zn system is restricted by the formation of opaque films in an ammonia environment in which stress corrosion cracking is generally observed. Furthermore, both components of the alloy are chemically active and the interpretation of observations is therefore relatively complicated. The Cu-Au system is a better choice for the purpose of studying structural factors. In particular, CuAu is very susceptible to cracking in ferric chloride and opaque films are not formed in this solution; it may be obtained in different states of long and short range order; and, finally, only copper is oxidized in dilute ferric chloride, which considerably simplifies the chemical problem. To provide information necessary for a more general explanation of stress corrosion cracking in homogeneous solid solutions, the following structural factors have been investigated in a Cu-Au alloy having a composition corresponding to Cu,Au: 1) the effect of long and short range order; 2) the structural sites from which copper is removed from the annealed and cold worked polycrystalline alloy; and 3) the effect of annealing time and temperature on structural reactions in the cold worked alloy. Experimental Procedure A Cu-Au alloy of nominal Cu3Au composition was cast from cathode copper and high purity gold and fabricated into sheet and wire. Chemical analysis for copper and gold indicated 48.92 wt pct Cu and 51.01 pct Au. Quantitative spectrographic analysis indicated 0.01 pct Ag, 0.002 pct Pd, 0.001 pct Pb, and 0.002 pct Fe, with traces of magnesium, silicon, and zinc. Presumably most of the remaining 0.055 pct corresponds to the limits of analytical precision. Heat treatments of wire and sheet were conducted in argon or vacuum. Wire specimens, 0.064 in. diam, were heated for 1 hr at 850 °C to establish a definite

Jan 1, 1957

By George M. Fowler

DURING 1937 the subject of mining geology was probably given more attention and more mining geologists were usefully employed than at any previous time. Of the many contributing factors the most important were: (1) a higher degree of prosperity in many countries, (2) a. better understanding of geologic phenomena, and (3) the apparently diminishing reserves of several important minerals. Mining companies have come to look upon geology with a growing appreciation of its aid in the more effective exploration and exploitation of mineral deposits. Most of the larger and more active companies have one or more geologists permanently attached to their staffs and some have well organized and efficiently managed geological departments. Even the smaller companies, who feel they cannot afford to employ a geologist on a permanent basis, are seeking the advice of competent men in this field with increasing frequency. One of the most significant trends has been the growing tendency to base exploratory work on more thorough geologic studies instead of empirical "hunches,"on which so much off the work in the past has relied. Likewise, mining companies are apparently more willing to expend larger sums on the basis of geologic possibilities in a district.

Jan 1, 1938

By A. Burkert, A. E. Hajj

"An information system has been developed for the use of companies considering developing a new mine which may be liable to secondary sulphide ignitions or by existing operations. The system uses the Comdale X expert system to analyze information about a particular mine. The system will indicate whether the mine is likely to experience ignitions and what precautions can be taken to reduce or prevent their occurrence. The advantage of an information system using expert system and artificial intelligence principles is that it can recommend action when there is an incomplete knowledge of all the system parameters. By interrogation of the user it can determine the state of knowledge and react accordingly. It is considered that such systems could be of significant use in mine planning for new and existing properties, especially where operators have limited expertise regarding secondary sulphide ignitions. IntroductionSulphide ignitions occur in the underground mines of several countries, including Australia, Canada, Finland, South Africa and the United States of America (1·3) and there has been a significant increase in their number and severity over the last decade(2.4). The basic causes of such ignitions are not well understood and man incidents are not reported or documented.Sulphide ignitions can pose a serious threat to the health and safety of mine workers and to the cost and efficiency of the mining operation (1-5). Byberg(6) described the three major problems associated with sulphide ignitions as:1. the force of the initial explosion can endanger workers and equipment;2. the intense heat of the ignition's flame front can endanger workers and equipment and may also ignite combustible material resulting in mine fires; and3. the sulphur dioxide produced by the ignition may be of such quantity and concentration to endanger workers, as was tragically demonstrated at Geco. In this incident in October 1985, a worker was killed at a position 2400 metres from the ignition point of a sulphide ignition at Manitouwadge, Ontario."

Jan 1, 1994

By Irving M. Symonds

Compania Minera de Penoles, S. A., at its Avalos unit operates a lead-zinc concentrator having a capacity of about 600 short tons per day. Lead and zinc concentrates are made by flotation methods. The concentrator is at Terminal de Provi-dencia, in the state of Zacatecas, some 70 miles south of Saltillo, Coahuila. The Penoles-Avalos railroad operating between the plant and Estacion Avalos, a distance of 6 miles, makes connections with the Coahuila and Zacatecas railroad, which in turn connects with the Nacionales de Mexico at Saltillo. The three railroads are each of different gauge, requiring the transfer of all supplies and concentrates at each junction. The mill operates on company ores, on ores from leased mines, and as a custom mill. The custom business is very small. The lead concentrates are shipped to company smelters at Monterrey, Nuevo Leon and Torreon, Coahuila. Since the European war, all zinc concentrates have been shipped to the United States. Ore Treated The major source of ore supply is a group of four mines owned or leased by the company, some three miles south of the mill in the mountains, from which the ore is delivered to the mill by a gravity aerial tramway. Custom ore comes in by rail. The ores occur in chimneys. The upper levels contained oxidized siliceous ores of lead, copper, silver and gold, with varying amounts of zinc. These ores were direct-smelted. In the first two years of operation a good deal of the tailing loss was caused by the oxidized content of the ores milled. Much of the ore in the transition zones between the oxides and sulphides was not mined. The ores now being milled are clean, hard sulphides in a limestone gangue, with only occasional contamination of oxides. The lead and zinc sulphides are of coarse structure and are freed even above flotation sizes. The galena is usually "steel" colored, easily and quickly floated. When there is any oxidation, it is usually on the coarsest crystals. The zinc minerals vary a good deal in color; a resin sphalerite predominates but there are varying shades of brown marmatites. On an average, the zinc sulphides contain 64 per cent zinc. The only common copper mineral is chalcopy-rite. Little is known of the occurrence of gold and silver. Waste minerals are mainly limestone and pyrite, both of which are barren. The ores vary from 20 to 35 per cent pyrite. Mill heads for the year 1940 and the first eight months of 1941 are given in Table I.

Jan 1, 1943

By Irving M. Symonds

Compania Minera de Penoles, S. A., at its Avalos unit operates a lead-zinc concentrator having a capacity of about 600 short tons per day. Lead and zinc concentrates are made by flotation methods. The concentrator is at Terminal de Provi-dencia, in the state of Zacatecas, some 70 miles south of Saltillo, Coahuila. The Penoles-Avalos railroad operating between the plant and Estacion Avalos, a distance of 6 miles, makes connections with the Coahuila and Zacatecas railroad, which in turn connects with the Nacionales de Mexico at Saltillo. The three railroads are each of different gauge, requiring the transfer of all supplies and concentrates at each junction. The mill operates on company ores, on ores from leased mines, and as a custom mill. The custom business is very small. The lead concentrates are shipped to company smelters at Monterrey, Nuevo Leon and Torreon, Coahuila. Since the European war, all zinc concentrates have been shipped to the United States. Ore Treated The major source of ore supply is a group of four mines owned or leased by the company, some three miles south of the mill in the mountains, from which the ore is delivered to the mill by a gravity aerial tramway. Custom ore comes in by rail. The ores occur in chimneys. The upper levels contained oxidized siliceous ores of lead, copper, silver and gold, with varying amounts of zinc. These ores were direct-smelted. In the first two years of operation a good deal of the tailing loss was caused by the oxidized content of the ores milled. Much of the ore in the transition zones between the oxides and sulphides was not mined. The ores now being milled are clean, hard sulphides in a limestone gangue, with only occasional contamination of oxides. The lead and zinc sulphides are of coarse structure and are freed even above flotation sizes. The galena is usually "steel" colored, easily and quickly floated. When there is any oxidation, it is usually on the coarsest crystals. The zinc minerals vary a good deal in color; a resin sphalerite predominates but there are varying shades of brown marmatites. On an average, the zinc sulphides contain 64 per cent zinc. The only common copper mineral is chalcopy-rite. Little is known of the occurrence of gold and silver. Waste minerals are mainly limestone and pyrite, both of which are barren. The ores vary from 20 to 35 per cent pyrite. Mill heads for the year 1940 and the first eight months of 1941 are given in Table I.

Jan 1, 1943

By Irving M. Symonds

COMPAÑIA Minera de Peñoles, S. A., at its Avalos unit operates a lead-zinc concentrator having a capacity of about 600 short tons per day. Lead and zinc concentrates are made by flotation methods. The concentrator is at Terminal de Providencia, in the state of Zacatecas, some 70 miles south of Saltillo, Coahuila. The Peñoles-Avalos railroad operating between the plant and Estacion Avalos, a distance of 6 miles, makes connections with the Coahuila and Zacatecas railroad, which in turn connects with the Nacionales de Mexico at Saltillo. The three railroads are each of different gauge, requiring the transfer of all supplies and concentrates at each junction. The mill operates on company ores, on ores from leased mines, and as a custom mill. The custom business is very small. The lead concentrates are shipped to company smelters at Monterrey, Nuevo Leon and Torreón, Coahuila. Since the European war, all zinc concentrates have been shipped to the United States. ORE TREATED The major source of ore supply is a group of four mines owned or leased by the company, some three miles south of the mill in the mountains, from which the ore is delivered to the mill by a gravity aerial tramway. Custom ore comes in by rail. The ores occur in chimneys. The upper levels contained oxidized siliceous ores of lead, copper, silver and gold, with varying amounts of zinc. These ores were direct-smelted. In the first two years of operation a good deal of the tailing loss was caused by the oxidized content of the ores milled. Much of the ore in the transition zones between the oxides and sulphides was not mined. The ores now being milled are clean, hard sulphides in a limestone gangue, with only occasional contamination of oxides. The lead and zinc sulphides are of coarse structure and are freed even above flotation sizes. The galena is usually "steel" colored, easily and quickly floated. When there is any oxidation, it is usually on the coarsest crystals. The zinc minerals vary a good deal in color; a resin sphalerite predominates but there are varying shades of brown marmatites. On an average, the zinc sulphides contain 64 per cent zinc. The only common copper mineral is chalcopyrite. Little is known of the occurrence of gold and silver. Waste minerals are mainly limestone and pyrite, both of which are barren. The ores vary from 20 to 35 per cent pyrite. Mill heads for the year 1940 and the first eight months of 1941 are given in Table I. [ ]

Jan 1, 1942

By Irving M. Symonds

COMPA[N]IA Minera de Penoles, S. A., at its Avalos unit operates a lead-zinc concentrator having a capacity of about 600 short tons per day. Lead and zinc concentrates are made by flotation methods. The concentrator is at Terminal de Providencia, in the state of Zacatecas, some 70 miles south of Saltillo, Coahuila. The Pe[n]oles-Avalos railroad operating between the plant and Estacion Avalos, a distance of 6 miles, makes connections with the Coahuila and Zacatecas railroad, which in turn connects with the Nacionales de Mexico at Saltillo. The three railroads are each of different gauge, requiring the transfer of all supplies and concentrates at each junction. The mill operates on company ores, on ores from leased mines, and as a custom mill. The custom business is very small. The lead concentrates are shipped to company smelters at Monterrey, Nuevo Leon and Torréon, Coahuila. Since the European war, all zinc concentrates have been shipped to the United States. ORE TREATED The major source of ore supply is a group of four mines owned or leased by the company, some three miles south of the mill in the mountains, from which the ore is delivered to the mill by a gravity aerial tramway. Custom ore comes in by rail. The ores occur in chimneys. The upper levels contained oxidized siliceous ores of lead, copper, silver and gold, with varying amounts of zinc. These ores were direct-smelted. In the first two years of operation a good deal of the tailing loss was caused by the oxidized content of the ores milled. Much of the ore in the transition zones between the oxides and sulphides was not mined. The ores now being milled are clean, hard sulphides in a limestone gangue, with only occasional contamination of oxides. The lead and zinc sulphides are of coarse structure and are freed even above flotation sizes. The galena is usually "steel" colored, easily and quickly floated. When there is any oxidation, it is usually on the coarsest crystals. The zinc minerals vary a good deal in color; a resin sphalerite predominates but there are varying shades of brown marmatites. On an average, the zinc sulphides contain 64 per cent zinc. The only common copper mineral is chalcopyrite. Little is known of the occurrence of gold and silver. Waste minerals are mainly limestone and pyrite, both of which are barren. The ores vary from 20 to 35 per cent pyrite. Mill heads for the year 1940 and the first eight months of 1941 are given in Table I. [ ]

Jan 1, 1942

By Rick Piovesan, Pramod Rao, Nihal Bohra, Mark Wahler

"This paper discusses the approach followed for the geotechnical design of 355 mm and 457 mm diameter driven open-ended steel pipe piles used for a gas plant at a green field site in Pilbara region in Australia. The principal bearing layer at the site comprises of calcareous Red Bed soils and weak rock, which is the typical geology along a significant length of the northern Western Australian coastline with carbonate content up to 70%. The uncertainty associated with the mechanical behavior of the calcareous soils and cemented soils (weak rocks) presented design challenges in the selection of appropriate pile design method and design parameters for pile capacity estimation during Front End Engineering Design (FEED). The American Petroleum Institute (API)-based design methodology was adopted. The representative design parameters for the Red Bed weak rock were established by treating the weak rock as clay with undrained shear strength as one-half of the characteristic unconfined compressive strength (UCS) value. Both statically load tested and CAPWAP (Case Pile Analysis Wave Program) estimated ultimate axial pile capacities on 25 early pre-production indicator piles demonstrated that the API design methodology for siliceous soils provides reasonable prediction of axial pile capacity in the red bed soils/weak rock and no reduction to API computed axial capacity due to carbonate content is warranted. The results also demonstrated that penalizing axial pile capacity solely based on carbonate content is not appropriate.IntroductionA gas processing facility was proposed to commercialize the gas reserves located 130 km offshore from the north-west coast of Western Australia (WA). The project site is a green field site located in the Pilbara region along the northwest coast of WA. Specific information regarding the site has been withheld due to confidentiality reasons. The site layout is shown in Figure 1.The site comprises low-lying tidal flats, clay pans and sand dunes covered with low scrub. The onshore site levels vary from elevation 1 m to 10 m AHD (Australian Height Datum), with the largest proportion of the area (excluding sand dunes) between elevations of 1 m and 3 m.The onshore facilities are to comprise a liquefied natural gas (LNG) plant site, a construction camp, the land-based portion of the materials offloading facility (MOF), roadways, and borrow pit areas.The existing ground levels within the process train area, where pile load testing was performed, vary from over 7 m AHD where the summits and peaks of the central dune exist to typically 1 m to 2 m AHD towards the east and west edges of the dune. The proposed finished grade for the process train is set at 7 m AHD."

Jan 1, 2015

By D. W. Smith, J. J. Koskinen, T. A. Enwald, M. E. Ball

"A full-scale demonstration unit of Tampella's LIFAC desulphurization process was tested at SaskPower's Poplar River Power Station Unit No. I. The test program included a series of parametric tests, followed by a 1500 hour operational evaluation. This paper summarizes the results of the test program, and SaskPower's economic evaluation which led to selection of the LlFAC process for SaskPower's 300 MW Shand power station. IntroductionOne of the major influences on the coal mining industry today is public and legislative response to acid rain. When sulphur-bearing fuels such as coal are burned, the sulphur oxidizes to sulphur dioxid e (S00 which, if uncontrolled, enters the atmosphere as one of the by-products of combustion. S02 then combines with water in the air to form an acid and acid rain results. Combustion also produces nitrogen oxides (N0J which can combine with water in the air forming nitrogen based acid , also contributing to acid rain.Acid rain has become a significant international concern. Newspapers have reported on dying forests, dead lakes and crumbling statues. Papers also include articles professing that the physical results of acid rain are vastly overstated. In the United States, a IO-year, Y2-billion dollar study called the National Acid Precipitation Assessment Program (NAPAP) was commissioned to evaluate the effects of acid rain. Even this extensive scientific study did not resolve the controversy. Despite disagreement over the quantitative effects of acid rain, there is no doubt that public concern over the issue is widespread and deeply felt.In response to this concern, legislators throughout the world have reacted with programs that restrict the emission s o f SO, and NO,. To industries involved in the supply and use of coal, the impact of these legislative programs has been extensive. United States power plants will be spending 10 billion dollars in the next decade for flue gas clean-up equipment. Even all this new equipment will not stem the tide of change to low sulphur coals and to “clean “natural gas. The following summarizes legislative actions taken for Canada and the United States."

Jan 1, 1993

By B. A. Kutlin, S. A. Ropeyko, N. A. Yurlova

"The aim of NPO RIVS research and practice activities is to minimize impact of the mineral processing plants (MPP) on environment by using new concentration methods, high-technology equipment, implementation of engineering solutions directed to reduce emission of polluting substances (PS) into atmosphere, application of low-waste technologies, decrease of water withdrawal from a surface water or groundwater source for process needs. In our studies environmental safety of mineral flotation is provided through technologies using highly selective chemicals, that leads not only to reducing reagent consumption, and production of saleable products of high content of extracted minerals, but also to reducing pollutant emissions, decreasing supply of surfactants (SAA) in natural geosystems, primarily, in ground and surface waters. Decrease of water withdrawal from a surface water or groundwater sources for process needs can be achieved by using closed circuits and waste water treatment for further use in process operations, recycled water conditioning to obtain optimum quality parameters for process operations of concentration. Concentration in water (wet processing), tailings dewatering and semidry and dry (filtration cake) stockpiling in tailing ponds, efficient dust-collecting systems contribute to pollutant emission reduction. Use of mineral wastes – concentration tailings as a source of valued components, facilitates efficient use of natural resources and reduces wastes and its impact on natural ecosystems due to smaller areas of damaged lands during waste products disposal. Process and engineering solutions of NPO RIVS that allow diminishing impact of MPP on natural environment (NE) have been implemented in mining and processing complexes of Russia, Kazakhstan, Ukraine, Kirghizia and Mongolia. Compliance with environmental safety requirements contributes to enhancement of mineral processing and sustainable mineral development.INTRODUCTION Mining and processing complexes have a significant impact on environment as a result of extraction and recovery of mineral resources, landscape changes, waste disposal and, as a consequence, pollution of soils, surface and underground water sources. Therefore, the aim of the operation of the mineral processing enterprises in its research and practical activities is to reduce impact on environment due to new beneficiation technologies, high-technology equipment, implementation of engineering solutions directed to reduce emissions of polluting substances (PS) into atmosphere, use of low-waste technologies, decrease of water withdrawal from a surface water or groundwater source for process needs. Compliance with the requirements of environmental safety encourages advancement of mineral processing technologies and sustainable mineral development. Concentration in water, tailings dewatering and semidry and dry (filtration cake) stockpiling in tailing ponds, efficient dust-collecting systems contribute to pollutant emission reduction. Use of mineral wastes – concentration tailings as a source of valued components, facilitates efficient use of natural resources and minimizes waste formation and its impact on natural ecosystems due to reduction of damaged land areas during waste disposal."

Jan 1, 2018

By S. K. Kawatra

"A typical steelmaking facility produced 50,000 tons per year of slag in their hot-metal desulfurization process. This slag had a sufficiently high iron content that it could be beneficiated and recycled in a blast furnace, provided that the slag could be agglomerated into pellets suitable for use as blast furnace feed. In this work, beneficiated slag was agglomerated using both a conventional bentonite binder, and a fly-ash-based binder (FBB) that was formulated by the investigators. The FBB consisted of a combination of highcarbon fly ash with calcium hydroxide activator, allowing the FBB to exhibit binding properties when moistened. It was determined that beneficiated slag agglomerated with FBB had superior compressive strengths compared to slag agglomerated with conventional bentonite binder. The FBB is therefore a satisfactory method of binding the agglomerated slag. Since the fly-ash component of the FBB is commonly discarded as waste, this approach also provides a means for utilizing high-carbon fly-ashes that currently have very little commercial value.IntroductionIn 2006 approximately 72.4 million short tons of fly-ash were produce in the United States [1]. The majority of fly-ash which was used that year was consumed concrete, structural fills, and cement. It is also used in road base, aggregate, and other applications. However, 55.2% of the fly-ash produced in 2006 was not utilized and ended up being sent to a landfill. Essentially all of the land filled fly-ash could be a suitable binder for pelletizing desulfurization slag and other steel plant waste oxide materials. Fly-ash based binders (FBB) have been shown in laboratory studies to be suitable for producing iron ore concentrate pellets of equivalent strengths as those bonded with bentonite [2,3,4]. The utilization of highcarbon fly-ash within the market is important for reducing the volume of currently unmarketable and land filled fly-ashes. FBB would be a much cheaper alternative to the current bentonite binder or organic binders used for iron ore concentrate and would help to reduce pellet costs. Currently, transportation more than doubles the final cost of bentonite since it is produced hundreds of miles from the pelletizing plants, while fly-ash is produced from power plants located nearby pellet plants and steel mills. FBB’s also show promise for use in other high volume applications such as foundry sand mold binders."

Jan 1, 2008

By Joseph R. Stephens

Deformation substructures of' polycrystalline tungsten, W-2, 9, and 24 pct Re, and W-3 pct Ta were studied by tra?zsrnission electron microscopy. The stress-strain curve for unalloyed tungsten showed gradual yielding followed by work-hardening. Electron nzicrographs indicated a gradual increase in dislocation density with increase in strain up to 5.0 pct. Dislocations, although frequently jogged, were straight over moderate distances and were in a randorn array. Stress-strain curves for alloy specimens of W-2 and 9 pct Re and W-3 pct Ta exhibited a drop in stress at yielding followed by only slight work-hardening. Electron micrographs of these specimens after strains of 0.05, 0.1, and 0.5 pct revealed no change in dislocation substructure from the unstrained specimens. After 2.0 pct strain, the three alloys exhibited dense networks. W-3 pct Ta was characterized by straight, frequently jogged dislocations comparable with the dislocation structure in unalloyed tungsten after a similar amount of strain. In contrast, W-2 pct Re exhibited dislocations that contained widely spaced jogs, while W-9 pct Re had developed a cell structure after the relatively srnall strain of 2.0 pct. The W-24 pct Re alloy contained a few dislocations after 0.1 pct strain, while after 0.5 pct strain twins were evident. Dislocation slip bands apparently preceded the twins. The stress-strain curve for the alloy indicated that twinning commenced after approximately 0.25 pet strain. These results indicate that the primary effect of low rhenium concentrations (2 and 9 pct) in tungsten is to increase dislocation multiplication after macroyielding by reducing the Peierls-Nabarro force (lattice resistance to dislocation motion). The dislocation bands that precede twins in W-24 pct Re may be caused by localized internal stresses resulting fro a metastable structure, for example, clustering of rhenium atoms. The effect of high rhenium additions (22 a 65 pet* Rproperties of tungsten. Klopp, Witzke, and Raffo5 reported bend transition temperatures as low as -100°F (200°K) for dilute electron-beam-melted W-Re alloys tested in the worked condition. Recrystalliza-tion increased the bend transition temperature, but alloys with 2 to 4 pct Re were still markedly superior to unalloyed tungsten. Fractographic examinations of tungsten and W- 3 pct Re and W- 5 pct Re alloys by Gilbert 6 revealed that these low rhenium alloys showed a greater tendency toward cleavage failure than did tungsten. Garfinkle 7 showed that rhenium additions, up to 9 pct, to (100) oriented tungsten single crystals increased the proportional limit stress and decreased the flow stress and the rate of work-hardening. In addition, while deformation in unalloyed (100) oriented crystals apparently involved both (110)100) and (112) slip, crystals with rhenium contents of 5 andpct or more deformed primarily by (112) slip. The mechanism by which high and low rhenium additions affect the mechanical properties of tungsten is still not well-established. The present investigation was undertaken to determine by transmission electron microscopy the effects of low rhenium additions, 2 and 9 pct, and a high rhenium addition, 24 pct, on dislocation substructure in the early stages of deformation of polycrystalline electron-beam-melted tungsten. Unalloyed tungsten and a W- 3 pct Ta alloy were included for comparison. EXPERIMENTAL PROCEDURES Materials. Triple electron-beam-melted tungsten, W-2, 9, and 24 pct Re, and W- pet Ta were used for this investigation. Chemical analyses of the cast ingots are given in Table I. A description of the starting metal powders and melting and fabrication pro-cedures for unalloyed tungsten and the W-Re alloys is reported.5 The W-3 pet Ta alloy was processed in a similar manner. Compression specimens measuring 0.300 in. (7.6 mm) in length by 0.130 in. (3.3 mm) in diam were machined from swaged rods. All alloy specimens were annealed in a vacuum of 8 x 10- 6 Torr (10'2iVper sq m) for 1 hr at 3600°F (2255°K). The recrystallized grain size ranged from 0.06 to 0.08 mm diam for the alloy specimens. Unalloyed tungsten was annealed at 2400° F (1589°K) for 1 hr to produce a recrystallized grain diameter of approximately 0.12 mm. Specimens were electropolished in a 2 pet NaOH solution to a diameter of 0.125 in. (3.18 mm) to remove surface notches resulting from grinding and to improve reproducibility of t data. The ends o the compression specimens were ground flat, parallel to each other, and perpendicular to the longitudinal axes with 4/0 emery paper. Compression Tests. The compressive stress-strain apparatus used for compression tests is described in detail by Stearns and Gotsky.9 Room-temperature compression tests were conducted at a crosshead speed of 0.01 in. per min (0.25 mm per min).

Jan 1, 1969

By Benoit Clermont, Jan Lux

"Magotteaux and Anglo Platinum started a co-operation program with the main aim of establishing the optimum milling configuration measured in terms of throughput, fineness and energy consumption.For this purpose, a fully instrumented pilot milling plant, supplied by Magotteaux (Pty) Ltd, and equipped with a 0.9 x 1.0 m ball mill, was erected at Frank concentrator (Merensky Pt ore). The primary ball mills at Frank concentrator were surveyed in order to set the conditions for the base case. This data was then replicated in the pilot mill under the same configuration. Thereafter, a series of different parameters were varied in order to attempt to optimize the milling system.Some of the grinding parameters that have been studied are:•,Gate versus overflow discharge•,Open versus closed circuit•,Different filling degrees•,Different shell liner profilesOnly grate versus overflow discharge and open versus closed circuit will be discussed in this paper.It should be noted that this paper does not attempt to give a universal solution as to what milling configuration is optimum in all cases. The conclusions apply for this particular case and probably most others, but it would be advisable to consider each case on its own."

Jan 1, 2003

By G. Song

This paper reports the anodizing voltage, gas evolution and sparking behaviors of an anodized coating formed on AZ91D in a silicate containing solution. According to their responses to an anodizing current, a four-stage anodizing mechanism is proposed. At the same time, the formed coating is characterized in terms of its microstructure, composition, thickness and corrosion performance. It was found that deposition of MgSiO3, Mg(OH)2 and MgO as well as dehydration of Mg(OH)2 are involved in the anodising process; the vigorous gas evolution accompanied by intensive sparking and the corresponding negative anodising efficiency in the later stage of the anodising process are closely associated with thermal decomposition of water due to the high temperature resulting from the intense sparking on the surface of the anodized alloy; the thickness, porosity, and composition of the anodised coating on a magnesium alloy depend on the final anodising cell voltage; and the corrosion resistance of a thick but porous coating containing more silicates appears to be more corrosion resistant than a thin but compact coating containing less silicates.

Jan 1, 2006

By W. A. Crosby

The ability of open pit mines to steepen overall final wall angles allows significant cost reductions as a result of reduced waste movement requirements. Final wall angles are principally controlled by three factors: geological conditions; hydrological conditions; and wall control blasting practices. The first of these factors is not controllable by the mine operator, the second can be influenced to some degree, but the last one certainly can be controlled, hence the current interest in final wall blasting practice at many mining companies. This paper discusses common blasting techniques employed to steepen final pit walls and outlines the attractiveness of moving from small hole high cost systems to the use of full size production drills as a means of reducing wall control blast costs. Attempts to make this change have been limited at many mines by the physical size of the drills and their inability to reverse drill. In the past, pres pi it, pre-shear or trim holes have had to be drilled vertically, at considerable distance from the upper bench toe, with the result that inter-berm face stability is reduced and final wall angles have to be made somewhat more shallow than desirable. The paper concludes with a description of how these problems can be overcome by utilizing the new Bucyrus 39R which has the capability of drilling at a ?15 degree angle underneath the machine. It is concluded that the resulting wall control blast designs will offer large mines improved final wall angles with resulting significant overall cost reductions.

Jan 1, 1998

By L. Larson

SMELTERMEN in the copper industry know that punching the tuyeres of a copper converter is a difficult, disagreeable, and at times a hazardous job. Knowing this, many men in the industry have given serious consideration to punching by mechanical means. As evidence of such consideration, a great many patents have been issued covering various mechanical devices or machines for doing this task. In 1942, as war-created manpower shortages became more acute, it became increasingly difficult to get men to punch tuyeres. Faced with this situation at the McGill smelter, it was decided that an all-out effort should be made to develop some sort of a mechanical punching device. The various patents were studied over thoroughly but none of the devices seemed suitable, all were discarded and it was decided to develop one, using the ideas of the personnel at McGill. Initial test work indicated that a separate punching device should be attached to each tuyere. This meant that an adapter would have to be developed which would accommodate such a device; the adapter to be so constructed that it could be substituted for the regular Dyblie valve head, and be punched either by mechanical means or by hand. Also a method for quickly attaching or detaching the punching unit would have to be devised. If the punching unit was attached to the adapter, then the punch rod and tip would have to remain inside of and reciprocate within the tuyere pipe. This last requirement meant that the punching device, the adapter, the piston rod, and punch rod must be in perfect alignment. The whole arrangement thus would become an integral part of the converter and would rotate along with the tuyere line to all the positions required for converter operation. If the puncher was to rotate with the converter, there was also a problem of limited clearance. The construction of a device to meet all of these conditions imposed many problems. In the latter part of 1942, two pieces of equipment were constructed and tried out under test-stand conditions. Both of these devices were very crude, but test-stand operation indicated that progress was being made. Neither of the units was considered good enough to be attached to and tried out on a tuyere of an operating converter. In May 1943, work was started on another device and in August of that year it was attached to and actually punched a single tuyere of a converter on an experimental basis (fig. 1). This unit was essentially a cylinder 3 in. in diam and about 15 in. long inside. Valve ports were so located as to provide a 1-in. cushion on each end of the stroke. The cylinder ports were piped to a 3/4-in. four-way disk valve. High pressure air (90 psi) was delivered to the valve by means of a long hose, thus making it possible to supply air to the device at all operating positions of the converter. The forward and return punching strokes were accomplished by shifting the valve handle as required. With the 3/4-in. valve, stroke velocities from 15 to 18 fps were achieved. Punching with this experimental arrangement was carried on for some time, but many difficulties were encountered. The punch rod tips often stuck in the tuyere accretion, indicating that more energy of some sort was necessary to avoid sticking the punch rod. To overcome sticking, experiments continued with a speeded up unit and although progress was made, troubles of various kinds continued to appear. On this unit, piston cup seals passed the valve ports which proved to be of very poor construction. Also the piston, driven forward at approximately 28 fps, would often bottom metal to metal in the forward end of the cylinder causing both mechanical and operational trouble. Sometimes the punch rod and at other times the piston rod would break loose and be shot through the tuyere pipe into the converter. Bottoming and poor timing of valve reversal caused a bounce and a hesitation at the end of the forward stroke, with the result that the punch rod

Jan 1, 1951