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By Rene Cane, V. I. Lakshmanan, John Starkey, Richard May

"The Agrium Kapuskasing Phosphate Operation mines and processes up to 6,000 t/d of phosphate ore. Run-of-mine crushing to minus 6 inches is done in a toothed roll-crusher, while grinding is done in a SAG/ball mill circuit with no coarse recycle classification of the SAG mill discharge. The primary grinding mill is an 18 ft. diameter x 9 ft. EGL, 1500 HP variable speed SAG mill and the secondary mill is a 13 ft. diameter x 21 ft. long, 2000 HP fixed speed, ball mill. When hard ore was treated, the SAG mill restricted the tonnage that could be ground. When soft ore was treated, the ball mill and spiral classifier overloaded and restricted throughput.Recommendations from a plant benchmark and an ore hardness testing program, on how to better utilize installed energy in the Agrium SAG mill when treating soft ore, and how to obtain best results when treating hard ore, were checked in a pilot plant investigation in the 3 ft. diameter Starkey-PRO Pilot SAG Mill. Correlation between the ore hardness testing and the results from the 3’ Pilot SAG mill was very good. In addition, the work indicated the results that could be expected for the different ore types when using either smaller SAG discharge grates and/or a vibrating screen with 10 mesh square opening cloth.INTRODUCTIONAgrium Kapuskasing Phosphate Operation consists of a 6800 t/d open pit mine and a concentrator, near the Town of Kapuskasing in Northern Ontario. The two major ore types are sandy ore and cemented ore. Sandy ore is fine, soft and easy to grind in the SAG mill while cemented ore is harder, more blocky and difficult to grind in the SAG mill. The ratio of sandy to cemented ore in the deposit is thought to be in the order of 80:20 but can range as high as 5050, so it was important for the grinding circuit to have the ability to treat sandy ore alone."

Jan 1, 2002

By Erick Lino

Como parte del proceso extractivo de la minería, se generan materiales de desmonte que deben ser almacenados en topografías bruptas y planas dependiendo de la capacidad de almacenamiento que requiera el proyecto. La geometría y los taludes de estas estructuras deben ser diseñados tomando en cuenta las propiedades de resistencia del desmonte de mina. Por ello, el presente texto describe la importancia de la caracterización geotécnica de este material con el objetivo de diseñar el depósito de una unidad minera en condiciones estáticas y dinámica.

Feb 1, 2017

By José R. Parga

In some parts of the Comarca Lagunera with population of about 2.5 million people, situated in the central part of northern México, chronic arsenic poisoning is endemic and severe adverse effects attributed to As exposure have been reported. Lifetime exposure to arsenic will develop classical symptoms of arsenic poisoning such as skin pigmentation changes, gastrointestinal disturbances and neurological changes. There are several methods available for removal of arsenic from well water in large conventional treatment plants, however a very promising electrochemical treatment technique that does not require the addition of chemicals or regeneration is Electrocoagulation (EC). The EC operating conditions are highly dependent on the chemistry of the aqueous medium, especially conductivity and pH. In this study, Powder X-ray Diffraction (XRD), Scanning Electron Microscope (SEM), and Transmission Mössbauer Spectroscopy (TMS) were used to characterize the solid products formed at iron electrodes during the EC process. This study will provide an introduction to the fundamental concepts of the EC method. Finally the results of this study suggest that the presence of maghemite and magnetite particles can be used to remove arsenic (III) and arsenic (V). In the field pilot plant scale study more than 99% removal of As species from groundwater is achieved.

Jan 1, 2005

By E. P. Wiechmann, J. A. Henríquez, P. E. Aqueveque, L. G. Muñoz, G. A. Vidal

"The original segmented intercell Optibar succeeded in reducing the magnitude and occurrence of short circuits. It also improved the specific energy consumption of the process. This paper expands the load current source concept introducing a triple segmented bar. This new bar adds immunity to dirt contacts and open circuits. Moreover, it further improves the standard deviation of current densities. As compared with industrial practices this development should accomplish an 8% specific energy reduction. Also, a plant implemented with BBS should be able to work with 12.5% higher process currents than Walker without compromising the operation.INTRODUCTIONIn 2011 the technology based on current mode was implemented at industrial scale. The success of this technology moved CIM Magazine (Lanktree, 2011) to write in the Jan 2012 issue an upfront story of this technological innovation. To cover the story, the technical reporter Graham Langtree was appointed by Peter Braul Editor of the upfront section. The article tells that “More than 40,000 tonnes of copper have been refined in a year-long test run of Optibar Inc.’s new technology at Barrick Gold’s Zaldívar copper plant in northern Chile. The new segmented intercell bars being used in the Compañía Minera Zaldívar refinery have slashed energy costs, reduced short circuits during refining and allowed the production of heftier, purer sheets of copper – gains that all promise to boost the plant’s margins. A copper refining tank house containing the cell circuits typically consumes 2,000 kWh of electricity to produce a single tonne of grade-A copper. Yet, monitoring the energy efficiency of the Optibar technology after increasing the weight of harvested copper sheets, the project’s engineers recorded a 92 kWh-per-tonne cut in energy consumption, a savings of 4.6 per cent. “Today, with the power prices we have, Optibar technology represents a potential savings for the whole plant of US$1.5 million per year,” says Robert Mayne- Nicholls, executive general manager of Barrick Gold’s Chilean operations.”"

Jan 1, 2012

By W. S. Cavender

The Round Mountain mining district, Nye County, Ne- vada, was discovered in 1906 on claims owned by Lewis D. Gordon. Initial mining operations uncovered gold veins of spectacular richness, and within a few days of discovery, Gordon sold his controlling interest for some $87,000. From this sale emerged the Round Mountain Mining Co., predecessor of Nevada Porphyry Gold Mines, Inc., the latter destined to become the major property owner in the area. Vein mining in the district continued sporadically into the early 1930s, yielding 9.3 Mg (330,000 oz) of gold plus substantial silver credits from approximately 626 kt (690,000 st) of ore. In addition to the lode deposits, the early miners recognized the placer potential in the alluvial fan material accumulated around the west and north sides of Round Mountain itself. Intermittent placer operations were carried out for a number of years, and in the 1940s and 1950s, Round Mountain Gold Dredging Co. worked the placers under a lease from Nevada Porphyry Gold Mines. The last placer operation terminated in 1959 when it, like some of its predecessors, proved uneconomic. Total placer production for the district is estimated at 3657 km (4 million yd) of gravel containing 59 Mg (210,000 oz) of gold and possibly 2.0 to 2.3 Mg (70,000 to 80,000 oz) of silver. Round Mountain is a small hill situated on the east flank of the Toquima Range in central Nevada. The hill is com- posed of relatively flat-lying Tertiary rhyolitic ash flow tuffs, which overlie Paleozoic metasediments and Cretaceous granites. Throughout the surrounding Round Mountain mining district, most of the known gold ores occur in the tuffs, although the metasediments and granites are also mineralized. Mineralization is structurally controlled, principally by a series of northwest-trending shears and shattered zones. Vein, stockwork, and disseminated ores occur, usually containing simple quartz-pyrite-gold mineral assemblages. The gold itself is electrum, having a silver content of 30% to 40%. In September, 1967, Elwood Dietrich, a prospector and mine promoter, obtained a purchase option on the 4452 ha (1 1,000 acres) of mineral rights held at Round Mountain by Nevada Porphyry Gold Mines. The original option had a buy-out price of $1 million and was established through Dietrich's friendship with officers of Nevada Porphyry. In April, 1968, Dietrich conveyed his option to Ordrich Gold Reserves Co., a partnership created by a group of west coast investors, mostly employees of the airline industry. There- after, Ordrich invested considerable funds in trying to test and develop the property, but soon recognized the need to seek financial and technical support from the mining industry. In December, 1968, Dietrich contacted Wayne Cavender, then Regional Geologist, Southwest, for Copper Range Exploration Company (CRX) in Tucson, Arizona, and made a data presentation. Shortly thereafter, Cavender was appointed Manager of Exploration and Chief Geologist for Copper Range Co. (parent company of CRX), New York City, and he asked C. Phillips Purdy, CRX Regional Geologist, Northwest, to make an initial property examination. Purdy's one-week field study took place in March, 1969, and resulted in a recommendation that CRX pursue its investigation of the property. The presence of low-grade gold mineralization in both the alluvial gravels and in the bedrock was verifiable, but the placer was deemed to have the greater immediate economic mining potential. At that time, gold was in the $1.41/g ($40 per oz) price range. Working from Purdy's information, Cavender decided to attempt acquisition of the property, and the first in a long series of negotiations was initiated with Ordrich. Basically, CRX felt that the placer had a promising potential for several reasons, including (I) past operators had recovered free gold but not the gold contained in the pebble fraction of the gravels; (2) past operations appeared to have been ineffectively designed or managed and not costefficient; and (3) the price of gold appeared to be poised for an upward move. Negotiations with Ordrich were prolonged and difficult, with CRX competing against several ma* mining companies, but finally an agreement was reached, effective June I, 1970. Gold was then back to $35. It is believed that, in

Jan 1, 1985

By Shawna Munn, Carol Dounitch, Erika Acton

The EQ Bank Tower, a 24-storey office tower with 4 levels of underground parking, is currently under construction at 25 Ontario Street in the downtown core of Toronto, Canada. The tower geometry required an 18 m (60 ft) deep excavation inside an existing 3-storey 100-year-old façade, to be retained in place throughout construction. This presented a variety of design challenges that required several innovations and a customized monitoring program. In this paper, we will discuss two major challenges, and highlight the innovative and adaptable solutions carried out using performance-based design leveraging the observational method. The first was installation of a shoring system in a 500 mm (20 in) gap between the historical façade and a live gas main. The second involved excavating 10 m (33 ft) into expansive shale bedrock adjacent to a sensitive neighbouring building.

Oct 1, 2022

By Frank S. Shuri

INTRODUCTION The Calumet Pumping Station, located on the south side of the Greater Chicago area, is part of the Tunnel and Reservoir Plan to accomodate the city's wastewater. The Station consists of two parallel rock caverns, each 18 m high, 20 m wide, and 62 m long, located in Silurian dolomite approximately 91 m below the ground surface. The caverns are oriented N-S with a 24-m-wide pillar between them. Prior to construction, rock mechanics studies were conducted in an L-shaped test adit at the location of the west cavern. The purpo3e of the rock mechanic studies was to characterize the site, particularly the in situ stress field and deformation characteristics of the rock. This information, together with other physical properties of the rock and geologic data, was used to evaluate the stability of the preliminary cavern design, with particular emphasis on cavern orientation, rock reinforcement, and pillar stability. GEOLOGY OF THE SITE Stratigraphy The Chicago area is located on the Kankakee Arch, a broad, gently sloping arch of Paleozoic sediments. Rocks in the project area consist of Silurian dolomites unconformably overlain by Pleistocene glacial deposits. The formations of interest to this study are: - Racine Formation, inter-reef facies. This dolomite is gray- brown, silty, and compact. It contains occasional prominent bedding surfaces. The Racine Formation in the project area varies from 73 to 74 m in thickness.

Jan 1, 1984

By X. S. Yuan

With a demand for increasing energy efficiency and a need for improving bitumen recovery, a highly integrated solution for sustainable lease development of Aurora North of Syncrude has been proposed and implementation plans are underway. Flocculation and thickening technology as an important component of the integrated solution provides Syncrude an opportunity of rapid heat recovery from tailings and more flexible tailings closure and reclamation options. Flocculation of fine-particle suspensions by polymers is a non-equilibrium process whose performance is strongly influenced by physical and chemical variables. While polymeric flocculants are frequently used to establish the chemical conditions favorable for flocculation, the actual performance of a flocculation process generally depends more on the physical variables, such as the hydrodynamic conditions in mixing the flocculants and slurry. In order to study the effect of hydrodynamic conditions on flocculation of Syncrude fine tailings, a standardized laboratory test procedure called power flocculation has been developed in comparison with the conventional manual mixing methods. It is shown that the agitation speed and time in conditioning the flocculant solution and the slurry have a profound effect on the flocculation performances measured by initial settling rates, supernatant clarity and sediment density. It was found that the optimum bench test conditions for power flocculation of Aurora fine tailings with a sand-to-fine ratio (SFR) of about 1:1 are conditioning of the flocculant solution and the tailings slurry at 250 rev/min to 300 rev/min in 20 s to 40 s. The test results are inter-preted with a flocculant adsorption/redistribution kinetic model that helps understanding of the effect of hydrodynamic conditions on flocculant adsorption, floc formation and floc growth.

Jan 1, 2014

By C. W. Chen

MARTENSITE transformations in ß Cu-Al alloys have been studied by Greninger1 and other investigators. According to Greninger, the parent phase ß1 an ordered body-centered-cubic structure obtained from ß phase by suppressing the eutectoid decomposition, transforms into an ordered hexagonal-close-packed phase in composition containing 12.9 to 14.7 pct Al. The M, temperature decreases with increasing aluminum content; for the alloy containing 14.5 pct Al, for example, the ß1??' transformation occurs below room temperature. More recently, Kurjumow2 studied the transformation in ß Cu-Al alloys with the addition of nickel. His report stimulated new interest in the subject due to the observation of completely reversible transformation without hysteresis in the transformation temperature ranging from 10° to —10°C. In the present paper some characteristics are described of the transformation of Cu-Al-Ni alloys that were partly studied by Kurjumow. Experimental Procedure High purity copper (99.999 pct) and aluminum (99.99 pct) and electrolytic nickel were used in the preparation, by the Bridgman technique, of single crystal specimens which contained aluminum and nickel of 14.5 and 0.5 to 3.0 pct, respectively. Polished surfaces were prepared mechanically. Specimens were then chemically etched to remove distorted material, homogenized at 1000°C for several hours, and quenched drastically to room temperature in a 10 pct NaOH bath to produce the parent phase ß1. The transformation was studied under a microscope and, in some cases, recorded by means of motion pictures. A device similar to that designed by Greninger and Mooradian3 was used to cool and reheat the specimens. Results and Discussion When the specimens were cooled below room temperature, the ß1 to ?' transformation began at 10°C with the appearance of ?' crystals In relief, Fig. la. As the specimen temperature dropped further, the transformation continued, either by the growth of the ?' crystals, with the ß1 — ?' interface moving into the ß1 phase, Fig. 1c and 1d, or by the formation of new ?' crystals, Fig. 1b. As a consequence of the former process, banded structure is observed as a common feature of the low temperature phase. According to the theory of the formation of martensite by Wechsler, Lieberman, and Read,' the bands of ?' phase are probably twin-related, as is the case in the diffusionless phase change of In-T1 alloys,5 but this was not revealed by X-ray tech- niques. New ?' crystals, in needle form, often emerged suddenly across the ?' bands during the transformation. These acicular crystals then grew, both in length and in width, see Fig. 2a through 2d. The transformation on cooling is completed at about -35°C. Upon heating, the reverse transformation started at —10° C, in a manner nearly opposite to the transformation on cooling, and completed at 35°C. There was no noticeable change in the transformation temperature when the nickel content was varied within the limits previously mentioned. Through control of the specimen temperature, the transformation can be started, stopped, or reversed at will. This phenomenon has frequently been observed in the martensite transformation of many nonferrous alloy systems. Other systems are Au-Cd6 and In-Tl.5 ow-- ever, in the latter systems, the transformation is accomplished by single interface motion if the specimen composition is homogeneous and the temperature gradient in the specimen is uniform and sharp, whereas in the Cu-Al-Ni specimens, only multiple interface transformation is observed. The speed of the interface motion appears to be a functionof the rate of temperature change and the temperature gradient across the specimen length. In one case, in which the temperature increased at the rate of 10°C per min and there was no temperature gradient along the specimen axis, the speed of the disappearance of a ?' plate was determined, by the study of the motion pictures made, to be 26 µ per sec. Quench markings were observed on the polished surfaces of specimens. The markings were grouped into one or more sets of different orientations, and were parallel in each set. The ?' plates formed in subsequent transformation were parallel to the markings, indicating that the ?' plates and the quench markings had the same geometric relation-ship to the ß1 matrix. The quench markings on two intersecting surfaces of a specimen were therefore used in the determination of the habit plane of transformation, by the trace method suggested by Barrett.' Results obtained from five sets of markings in three specimens indicate that the habit plane is an irrational plane about 2" from one of the {221} planes. This is very close to the habit plane (3" from 221 planes) of ß Cu-Al alloys containing more than 13.0 pct Al.1 The martensite transformation of Cu-Al-Ni alloys is reproducible. No sluggishness was found between consecutive transformation cycles, although a slight difference in the distribution pattern of the ?' plates was observed, compare Figs. Id and 2d. The transformation can be strain-induced. This characteristic has been tested by a simple method. When a specimen was elastically strained slowly in a vise, ?' plates were gradually produced in the same fashion as during transformation on cooling, This test was done at room temperature, and thus above the M,

Jan 1, 1958

By D. Harrington

This circular presents part I of a review of the literature on effects of breathing dusts, with special reference to silicosis, largely in the minim; and allied industries. The literature on this subject is so extensive that no attempt has been made to include all published material but only to assemble some of the more outstanding data on several aspects of the subject. Parts II and III will deal with prevention and treatment of dust diseases and some of the economic and legal aspects of dust diseases in industry, with special reference to mining. Dust diseases arc much before the public at present, and there is good reason to believe that the information on the subject available to the layman is extremely meager; even those rated as experts usually have a more or less theoretical conception of the diseases, their causation, remedies, and prevention, rather than actual experience with them. Until recently little attention has been given to prevention, and the usual attitude of the engineer, upon whom the main reliance for prevention ultimately must devolve, has been to let "sleeping dogs lie" or, more definitely, to "let well enough alone." However, the "laissez faire" attitude no longer is tenable, and in the past few years numerous conditions have caused those engaged in industry, including mining, to bring the subject before the public with a view to ascertaining the facts upon which to formulate remedial measures, primarily preventive.

Jan 1, 1935

By John E. O’Rourke, Kevin M. O’Connor, Pamela H. Rey

INTRODUCTION The resurgence of coal mining activity in the United States, brought on by the spiraling costs of fossil fw1 energy in the Seventies, has come at a time of intense public concern for the quality of the environment. Notwithstanding pressure on our economy to develop alternate sources of fue1 energy to the import of oil, the legislatures of several states have reacted to public concern over the environment by passing strict regulations aimed at con- trolling the subsidence effect s of underground mining. Agencies of the federal government charged with assistance to the mining industry, including the Department of Energy and the Bureau of Mines, have sponsored a number of instrumentation and field measurement projects aimed at the development of subsidence prediction models that can aid the mine operator's task of subsidence control. There are good empirical models developed in Europe for subsidence prediction, but they were made possible by a large body of mining-induced subsidence data collected there over a long period of time. No com- parable subsidence data base exists in the United States, and consequently empirical modeling of subsidence is not a realistic approach for our near term needs. Moreover, the geologic and topographic diversity of the several coal regions in the United States is expected to necessitate the development of several empirical models, each one expected to be relevant to its own region. Because of the time and costs that are likely to be involved in an empirical modeling approach, it is considered more expedient and cost effective to develop a general, mechanistic model for subsidence prediction purposes. In order to develop such a model, it is necessary to investigate and quantify the mechanics of the subsidence process from the mine level up to the ground surface. The series of projects discussed in this paper are designed to achieve this objective and include the following work: (1) the identification of geotechnical instrumentation that will pro- vide mine level overburden and surface subsidence data. (2) a field demonstration of selected instruments, and (3) documentation of case histories for complete subsidence mechanics, using the demonstrated and preferred instruments. An identification of feasible instrumentation and monitoring techniques was completed by Woodward-Clyde Consultants (WCC) in 1977 (O'Rourke and others). This paper discusses a demonstration of those instruments at a mine in Utah, and at two subsequent projects, currently underway at longwall mines in Colorado and West Virginia. 'he latter two projects when complete will provide documented case histories of subsidence mechanics. The process of optimizing subsidence instrumentation and monitoring techniques to the conditions encountered during installation and monitoring for these underground mines is shown to be an evolving one, and one which has had some notable successes to date. INITIAL DEMONSTRATION The initial design and demonstration of selected monitoring systems was carried out at the SUFCO No. 1 mine, near Salina, Utah. The instrumented panel was approximately 152 m wide. 640 m long. and was 290 m to 320 m deep. The mined height of coal. seam averaged 2.4 m. The mining method was room and pillar using continuous mining machines. This method allowed some monitoring of the supported condition during development, and eventually allowed monitoring of a caved system when both chain pillars and room pillars were extracted on retreat. The two instrument systems shown on Table 1 were selected from the earlier feasibility report for demonstration at SUFCO No. 1. Collectively, the two systems, one for a fully-supported mining method and the other for a fully-caved method, incorporate most of the instrumentation to be found within all five systems listed in the earlier feasibility report. The instrumentation includes surface, subsurface and mine monitoring installations. All of the SUFCO instruments selected to meet the specifications of the general instrument types listed on Table 1 were manually operated. That is, data from the installed system could only be obtained while a person was there to physically observe or operate the system readout. Automatic data recording equipment was available for some installations, but the objectives were to keep the systems as simple as possible for the demonstration project. A complete description of the surface, sub- surface and mine level instruments, and the demonstration project results are given in WCC (1982), and selected features are discussed in this paper.

Jan 1, 1982

By C. D. Hall

This paper may be divided into two main parts: (1) analysis and data to show that Eq. 6 adequately correlates the authors' experimental data, and (2) the assertion that Eq. 20 "should permit scale-up with data taken from one pipe diameter for Newtonian, inelastic non-Newtonian, and viscoelastic fluids, such as dilute polymer fracturing fluids, which produce considerable pressure reductions." The authors might well recommend that this broad grouping of fluid types could be correlated by using Eq. 6 (requiring data from at least two pipe diameters). On the other hand, Eq. 20 is only shown to be adequate for conditions within the range of the present data set. Specifically, predictions by Eq. 20 for diameter scale-up to casing sizes would give larger deviations than implied in Table 2, and scale-up for some other fluids that give large drag reduction could be grossly in error. The description of a class of fluids "which produce a diameter family of straight parallel lines on a log (d?p/4L) vs log (8 v/d) plot. . ." defines a necessary condition for using Eq. 6; i.e., that s shall be constant for each fluid. The proposal that Eq. 20 is applicable to turbulent flow of such fluids is equivalent to saying that the spacing of these parallel lines, with A being constant, will be found to correspond to m = -1.20 for all such fluids. Table 1 shows the values of m derived from data on 16 non-Newtonian fluids. The key statement of this paper is that the diameter exponents "do not vary significantly from . . . -1.20." In statistical terms, this says that the m value for each fluid, which was obtained from all the measurements made on that fluid. has not been shown to be different from the arithmetic mean of the m values for the 16 fluids tested. This implies that the group mean, -1.20, is a more reliable value to use for an individual fluid than is the m value derived from the data on that fluid alone. No statistical test is described to support the key statement that the diameter exponents do not vary significantly from -1.20. Such a test would have to take into account that each individual m value is shown to correlate the pressure gradients for its fluid, with only about 3 percent standard deviation over the diameter range covered. That is, the m value is not the result of a single independent measurement. Thus, it would not be sufficient, or valid, to compare the deviation of one of the m values with the variance of the 16 values of m about their mean. An alternative interpretation of the key statement is that the diameter exponents do not vary appreciably from -1.20; i.e., that using -1.20 does not cause serious inaccuracy. Table 2 seems to be offered in support of such an assertion. However, such a statistical treatment (1) is unnecessary and (2) cannot lead to any valid scale-up generalization. The amount of error introduced by using a 1"wrong" value of diameter exponent — i.e., -1.20 instead of m — is determined by Eq. 6. The ratio of the pressures predicted by Eqs. 6 and 20 for scale-up from ds in small pipe to d, in casing is The pressure for 30-percent calcium carbonate slurry (m = -1.112) in 10-in. ID pipe would be predicted 30 percent higher by Eq. 6 than by Eq. 20, if based on tests in 0.5-in. ID pipe. The data set tested in Table 2 represents only a limited sample of pipe diameters (ratio of about 4:l) out of the population of diameters that would

Jan 1, 1970

By Mark Jolly, Xiaojun Dai

"Foundry engineers in the traditional foundry usually regard the quality of casting component as the most important issue and leave the energy saving or energy efficiency as the subsidiary one. This frequently causes disproportionate energy consumption as a result of the inefficient casting processes used and increases the production costs. This paper presents the novel CRIMSON aluminium casting process and compares its facility and melting process with traditional melt furnaces and aluminium alloy melting process. A real example is investigated to demonstrate quantitatively how the traditional foundry wastes energy and what the improvement of energy efficiency can be achieved using the new CRIMSON method. The results of this investigation will help the foundry engineer recognize the importance of energy saving and demonstrate how to use this new technology to reduce production costs and carbon footprint without decreasing the quality of the cast component.IntroductionAluminium melting in metal casting industry is an energy intensive process where it is estimated that the energy consumption is in the order of 1,700 ~ 4,700 kWh tonne-1 in using crucible and natural gas [1]. Different countries may use different energy resources for aluminium melting: in US the main energy resources are natural gas, electricity and coke [2] and in UK most of the foundry use electricity, gas and oil as energy resources [3]. Due to the pressure of rising energy price and the strict environment law, it is anticipated that under the right socioeconomic conditions, the efficiency optimisation of industrial process can be an important step toward increased industrial sustainability [4]. In metal casting industry the energy efficiency of a casting facility depends largely on the efficiency of its melting and heat treating performance. In association with the two performances, over 60 % of the total process energy costs are represented in a typical casting facility [5] where there are huge opportunities for metal casting industry to adopt the best energy practices for potential energy saving. How to ameliorate the current processes for increasing energy efficiency will have a vital effect on reducing the production costs and increasing the competitiveness. For instance, by implementing some state of the art technologies such as the CRIMSON method in aluminium alloy casting will make use of such opportunities."

Jan 1, 2011

By B. W. Gandrud

[The Bureau of Mines is conducting an extensive investigation of coal preparation in the Alabama coal fields in cooperation with the University of Alabama and coal-mine operators. In this investigation the drainage characteristics of typical coals from the Warrior and Cahaba n fields were studied. Tire main purposes of this study has been to evaluate no nearly as possible relative influences of certain factors on rate of drainage and to determine whether the drainage characteristics of coals from different beds or from different mines on the soma bed differ materially. A further objective was the development of a laboratory drainage test that would predict approximately the rate at which any coal will drain after having been loaded into a railroad car in a saturated condition. The study has been confined to samples of slack coals - that is, coals from which none of the fine sizes have been removed. The samples ranged in upper size limits from 7/16 inch to 3 inches. It did not seem necessary to study nut and egg sizes, that is, sizes from which the fine coal has been removed by screening, as these lose water so rapidly by natural drainage that they present no serious problem in dewatering after treatment in a washer. The significance of drainage characteristics of Alabama coals is apparent in view of the fact that not less than 70 percent of Alabama?s total coal product is subjected to wet washing before being materials.]

Jan 1, 1941

By Mitsutaka Hino

The contents of gangue minerals, such as A1203, P205, crystalline water and so forth in the iron ore imported to Japan tend to increase year by year, Though the increments of such gangue contents are approximately 0.1-0.2 mass%, some problems for the blast furnace operation in ironmaking process have already been reported, In order to clarify the mechanism of harmful effect caused by the small increment of gangue minerals, mainly A1203, on the blast furnace operation, the dripping behavior of FetO-CaO-SiO2-AI2O3-MgO slag from the iron funnel, which simulated the micro-porosity in sinter ore, has been observed with the wide range of basicity, Al2O3, FetO and Mgo contents. The effect of MgO on the dripping behavior of slag has been summarized in this paper, It was found that the addition of 2mass% MgO resulted in large increase of hold-up on the most of the conditions. When FetO content was less than 20mass% and the basicity was lower than 0.8, MgO lowered the hold-up even if the slag contained 10mass% Al2O3. Every slag hold-up showed the minimum at the basicity of less than unity on the conditions of constant FetO and A1203 contents. This trend became remarkable when the slag contained 2 mass% MgO.

Jan 1, 2003

By Craig M. Jensen

A laboratory assessment of prompt and delayed neutron activation analysis with 252Cf and 14MeV neutron sources, for use as an in-situ method of determining low-level copper concentration has been performed. A simulated borehole using porphyry ore was constructed and varying copper concentrations were simulated using copper rods. Prompt neutron activation analysis experiments were consistant with earlier work and do not appear to be feasible fur copper concentration of 1.5% or less. Delayed neutron activation analysis experiments indicate that copper ore concentrations of less than 0.1% are detectable using either the 511-KeV gamma-ray of copper-64 or the 1039-KeV gamma-ray of copper-66, with high resolution detectors such as Ge (Li) or intrinsic germanium.

Jan 1, 1977

By G. V. Henderson

Large reserves of high calcium limestone have been outlined in the Redwall formation at Nelson, Arizona. Geologic mapping, core drilling and shallow seismic refraction studies have outlined more than 100 million tons of high quality, chemical and metallurgical grade limestone in the Mooney Falls member of the Mississippian Redwall Limestone. Well developed Karst zones in the Redwall were delineated by field mapping, seismic profiling and core drilling. Development of large reserves of high quality, thick beds of limestone near the present mine and rail facilities was the basis for the current mine and plant expansion. When completed in 1978 the planned 1800 Tpd production will represent the largest lime operation in the Western United States.

Jan 1, 1978

By Xinlei Jia, Wenming Guo, Yanling Zhang

A large amount of stainless steel dust rich in Fe, Cr, and Ni has been being generated in China. Smelting treatment is accepted as a practical method for the recovery/reutilization of these second resources, during which Fe, Cr, and Ni is reduced into one metal phase and then separated from the residue slag. Based on the typical compositions of stainless steel dust, the residual slag can be termed as CaO-SiO2-MgO-Al2O3 system. In this paper, the activity coefficients of Cr in Fe-Cr-Ni-C (-Si/Al) melt and the properties of CaO-SiO2-MgO-Al2O3- 1%CrO system thermodynamically approached. Then, in laboratory experiments, the recovery ratios of Cr and its distribution behaviors between Fe-Cr-Ni-C melt and CaO-SiO2-MgO-Al2O3 slag system were examined.

Jan 1, 2015

By B. Ladanyi

An adequate knowledge of the overall behavior of a rock mass at a given site is fundamental to the satisfactory design of a rock slope. The be- havior is controlled by a variety of factors, in- cluding rock type, geological structure, stress and ground water conditions. There are essentially three lines of approach that may be used to obtain information on the in-situ rock mass behavior for the purpose of the design of a rock slope: (1) Direct measurements of rock mass properties by means ol laboratory and field tests. (2) Evaluation of rock mass properties from a study of slope failure case histories and sta- tistical information on stable and unstable natural rock slopes. (3) Evaluation of in-situ behavior of rock mass by using a physical or mathematical model technique.

Jan 1, 1980

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada's Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989