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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 Błażej Doroszuk, Robert Król

Mill drums are essential equipment in various industries for the comminution of materials. To optimize their design and operation for efficient and cost-effective processing, it is necessary to understand the behavior of the mill drum under various operating conditions. This study investigates the behavior of a mill drum using a combination of experimental and numerical methods. The factors that can influence the behavior of the mill drum, such as the size and shape of the mill, grinding media, rotational speed, filling level, and material properties, are discussed. Experimental methods can be limited in obtaining measurements of internal processes, so numerical methods such as the Discrete Element Method (DEM) are also used to simulate the behavior of the mill drum. In this study, the experimental measurements of the motion of the grinding media are obtained using computer vision methods, and the numerical simulations are calibrated using the experimental results. A designed experiment approach is used to evaluate the performance of the mill drum under various operating conditions. The results of this study can be useful for the design and optimization of mill drums to achieve better performance and energy efficiency.

Jun 25, 2023

By Irving B. Cadoff, Alvin A. Machonis

The resistivity, Hall coefficient, Seebeck coefficient, and thermal conductivity were measured as a function of temperature for cation-rich alloy single crystals covering the composition range across the PbTe-SnTe system. Alloying of PbTe with up to 20 pct SnTe was found to have little effect on the energy gap. Above 20 pct SnTe the alloys were "p" type but below this range the sign could be varied by heat treatment. The lattice thermal resistivity of the compounds SnTe and PbTe is raised by alloying one with the other. Z values in the order the interesting values obtained. THE PbTe-SnTe system has several interesting features. For one, PbTe is a useful thermoelectric material and the possibility of improving its figure of merit by alloying with SnTe, an isomorphous compound, has been suggested since these pseudo-binary solid solutions generally have a more favorable ratio of electrical conductivity to thermal conductivity than either of the components.' Other interesting features relate to the conductivity mechanism, band structure, and stoichiometry of the compounds and their alloys. PbTe is a semiconductor with an energy gap of about 0.29 ev2 at room temperature whose conductivity sign and magnitude can be varied from "n" to "p" by controlling the proportion of lead and tellurium with respect to the stoichiometric ratio.3 Excess lead results in "n"-type conduction. SnTe is found to exist only as a "p"-type material of relatively high conductivity. This behavior is attributed to stoichiometric deviation by Brebrick4 but Sagar and Miller proposed that the behavior of SnTe must be due in part to the presence of an overlapped band. An investigation of alloys of this system, therefore, might give additional information which would permit one to evaluate which of the two proposals is the more appropriate one. Abrikosov et al.' studied the room-temperature electrical properties of these alloys and reported data for Seebeck coefficient and resistivity on poly-crystalline alloys. The present work is a more exhaustive survey of the PbTe-SnTe system. Re- sistivity, Hall coefficient, Seebeck coefficient, and thermal conductivity were measured over a wide temperature range for single crystals at 10-pct intervals of lead/tin ratio across the pseudobinary system. The relative concentration of tellurium was controlled so as to obtain metal-ion excesses in all cases. SAMPLE PREPARATION The crystals were prepared by melting elemental lead, tin, and tellurium in weighed proportions in evacuated Vycor capsules. The lead and tellurium were high-purity grades obtained from American Smelting & Refining Co. The tin was supplied by Comico. The proper calculated proportions of lead, tin, and tellurium were weighed and charged into prepared Vycor capsules prior to evacuation. The capsules were prepared from 15-mm Vycor tubing. A sharp point was worked on one end of the tube. A pyrolytic graphite coating was deposited on the Vycor walls by heating the tip to 800°C in an atmosphere of acetone-saturated argon. An additional coating of graphite was deposited on the pyrolytic coating from an Aquadag suspension. Above the coated tip the tube was reduced in diameter to form a constrictive neck. To avoid scratching the graphite coatings the charge was placed in the tube above the constriction. After a low-temperature bake, the evacuated capsule was sealed. On subsequent heating the charge melted down into the lower portion of the capsule. The crystals were grown by lowering the capsule through a Bridgman-Stockbarger furnace. The lowering rate was 1 in. per 8 hr. The upper portion of the furnace was set for 950°C and the lower portion for 800°C. In general the yield of single crystals was about 25 pct. The mixed compositions were, as expected, the most difficult to grow. The finished crystals were sectioned into 5/8-in. slices. The tip, end, and middle slices from each crystal were analyzed by X-ray fluorescence to determine the lead-to-tin ratio. The resulting values were used to plot a composition vs distance plot for each crystal. Slices were selected from each crystal, with the aid of the composition plots, to cover the complete range of compositions at 10-pct intervals. In general, the slices selected were taken from the seed end of the crystal where the longitudinal segregation (as determined from the X-ray fluorescence analysis) was a minimum. Laue single-crystal analysis and metallographic analysis was used to verify if a slice was single or polycrystal. Any grain boundaries were clearly visible in the as-cut and polished condition. In ad-

Jan 1, 1964

The Annual Meeting of The Rocky Mountain Coal Mining Institute was called to order by the President, John T. Atkins, at 2:00 p.m. Tuesday, June 26, 1984 in the Colorado Ballroom of the Marriott's Mark Resort in Vail, Colorado. There were 36 members in attendance. It was moved by Bob Diederich, seconded and carried to dispense with the reading of the 1983 Annual Meeting minutes as printed on pages 105 thru 109 of the 1983 Proceedings. In keeping with the printed agenda in the convention program/ticket booklet, the President called for reports of standing committees as follow: 1. Bylaws & Resolutions - Loren Linville, Chairman Linville noted that the committee had been working for two years on bylaws changes and concluded a corporate lawyer specializing in nonprofit organizations be engaged. He then introduced Mr. Peter B. Nagle of the Denver law firm of Sherman and Howard who outlined the major changes in converting the Institute's current status of 501(c)(6) to 501(c)(3). He pointed out that such a conversion would give the Institute a proper legal status, would give individual contributors to the scholarship or other programs a personal income tax deduction, would qualify the Institute for the postal nonprofit mailing rate and the Institute would achieve the advantage of limited liability inherent in a corportion and would insulate its members, officers and directors from liability for corporate obligations. Copies of the Articles of Incorporation as well as the Bylaws were available for all members. A discussion ensued and the following resolution was moved by Bob Diederich, seconded and carried: RESOLVED, that the unincorporated association known as the Rocky Mountain Coal Mining Institute be, and hereby is, dissolved and in its place, and to carry out its objectives, shall be formed a nonprofit corporation under the laws of the State of Colorado, which corporation shall also be known as The Rocky Mountain Coal Mining Institute, and to which shall be transferred all of the assets and liabilities of this association. FURTHER RESOLVED, that the proper officers of this association are authorized and directed to execute and file with the Colorado Secretary of State, Articles of Incorporation in a form substantially similar to those prepared for this association by its counsel, and the names of the individuals identified in such Articles of Incorporation as the initial Board of Directors of such corporation, which names were proposed and recommended by the Nominating Committee of this association, are hereby approved. FURTHER RESOLVED, that the proper officers of this association are authorized and directed to do all acts necessary or convenient to carry out the terms of the preceding resolutions. (Articles of Incorporation and Bylaws as adopted are printed following these minutes.) Additional resolutions were read by Linville as follows: Item 1: WE RESOLVE herein that in the year since the last convention of the RMCMI, the Institute and the coal industry have suffered the loss, through death, and the services of: Lloyd W. Taylor Charles T. Holland THEREFORE, BE IT RESOLVED, that we record in the minutes of the RMCMI, recognition of the passing of these members who have so greatly contributed to the advancement of the industry, both nationally and in the Rocky Mountain area, and that a copy of this resolution be sent to the next of kin as evidence of the sincere sympathy and deep regard of the members of the Institute. Item 2: RESOLVED, that in recognition of their long and faithful service to the coal industry that the status of Life Member be conferred upon: Bernard Bernstrom Roy Coulson Shirl C. McArthur Item 3: RESOLVED, that we commend and express our gratitude to our Executive Secretary, officers, chairmen and committees for an outstanding job in presenting. the 80th RMCMI meeting. Item 4: RESOLVED, that the RMCMI express its gratitude to K&P, INC. of Golden, Colorado for their generous donation of office facilities and furniture for the RMCMI office and to K&P SALES ENGINEERS, INC. also of Golden, Colorado for their assistance in telephone answering, other office procedures and the use of certain office equipment. Item 5: RESOLVED, that six scholarships be awarded annually to advance the education program of the Institute to worthy students pursuing career paths in mining or other related industry in the six states in which the Institute generally operates and has members and in accordance with the criteria set forth by the Scholarship Committee. It was moved by Jim Neal, seconded and tamed

Jan 1, 1984

By H. Tomlinson

The development of the Sevier Valley Coal Co.'s mine marks the opening of the Salina Canon coal district in Utah. The mine is favorably located with respect, to the existing markets of the West, and if the present estimates of the amount of workable coal in it are accurate it has many Important advantages in favor of its development. This mine is situated 18 miles east of the town of Salina, Sevier County, Utah, which at present is the nearest shipping point oh the Marysvale branch of the Denver & Rio Grande Western Railroad, and is about 152 miles south of Salt Lake City, which is the distributing center for much of the coal from the Carbon - Emery coal field and is the place from which nearly all Utah coal shipped to San Francisco and points on the northern Pacific coast has a common haul.

Jan 1, 1930

By John Birkinbine

The following data concerning the general dimensions and district-location of the blast-furnaces of the United States are intended to supplement a paper of similar title, which appears in volume xiv., pages 561 to 577, of the Transactions, and bring the record to the end of the calendar year 1886. The reasons for presenting the figures at so short an interval since the former paper are that the American Iron and Steel Association has lately issued a new directory of iron and steel works (edition of 1886) from which many of the plants which have been inactive are omitted and new industries added, to such an extent as to materially affect the figures presented in the first paper; and also that the improvement in business has encouraged a production of pig-iron in 1886 in excess of any previous outputs, and the results obtained in the industry at large as well as at individual plants form a better index of what the American furnaces can accomplish than those heretofore presented. The statement will not be extended beyond the production of pigiron, although it could well be made to include the manufactured iron and steel; and while the data given have been verified as far as possible up to the close of the year 1886, no attempt, except in special cases, has been made to note the changes which have taken place since that time—changes which within the next twelvemouth will affect some of the figures here presented. At the close of the year 1886 there were 579 blast-furnaces reported on the active list—an apparent decrease, as compared with my former paper, of 97 stacks; but this was not an actual decrease, since most of the furnaces dropped from the Directory were those which, owing to location, inefficient equipment, or fake of ore-supplies, etc., were considered likely to remain inactive, although a number of them will possibly be temporarily put into blast with iron advancing in price. On the other hand the new plants added

Jan 1, 1887

By B. H. Johnston, B. N. Doyle

Introduction In 1983-84 Kidd Creek Mines Ltd. in Timmins, Ontario, started up the second commercial zinc pressure leaching plant based on the Sherritt Gordon Mines Ltd. direct pressure leach process. The first plant using this process was that of Cominco Ltd. at Trail, B.C. (Parker, 1981; Martin and Jankola, 1985). The Kidd Creek plant was constructed to take advantage of the extra capacity that the existing purification, cell house, and casting sections had in comparison to the roasters and acid plant. In addition, pressure leaching offered flexibility in other areas such as treatment of low grade concentrates. The process design package and the detailed design of the autoclave system was carried out by Sherritt. The detailed design for the rest of the plant was done by Cambrian Engineering Ltd. under the direction of Kidd Creek. The construction and commissioning of the plant has been described earlier (Johnston, 1983). This paper describes, briefly, the application of the direct pressure leaching process for zinc concentrates and the integration of the pressure leaching plant into the main zinc plant at Timmins. It discusses some of the startup problems and changes made to the plant. The damage following a fire inside the autoclave is also described. The subsequent repair and recommissioning of the autoclave has enabled Kidd Creek to demonstrate the originally sought advantages of pressure leaching and to take advantage of their extra cell house and casting capacity to produce more zinc. Process principles The development of the direct pressure sulphuric acid leach of zinc sulphide concentrate at temperatures above the melting point of sulphur and at elevated oxygen pressure has been described in detail in a number of papers (Martin and Jankola, 1985; Johnston, 1983; Veltman and Bolton, 1980; Kawulka et al., 1975; Veltman et al., 1977). The most significant aspect of the process is that sulphidic sulphur is converted to elemental sulphur in reactions that can be simply described as follows. ZnS + H2SO4 + 1/2 O2 ZnSO4 + S° + H2O (1) FeS + H2SO4 + 1/2 O2 FeSO4 + S° + H2O (2) The rates of these reactions increase with temperature. However, at temperatures above the melting point of sulphur, 12°C, the presence of a surface active agent is necessary to prevent the occlusion of the sulphide particles by elemental sulphur and the consequent termination of the leach. The zinc that is dissolved according to reaction (1), after purification of the zinc sulphate solution, is recovered by electrowinning. The iron that is put into solution according to reaction (2) is further oxidized to the ferric state and subsequently, as the acidity decreases, precipitates as hematite or as complex basic sulphates. MINERALS AND METALLURGICAL PROCESSING Examples of this are shown below. 2FeSO4 + H2SO4 +1/2O2 - Fe2(SO4)3 + H2O (3) Fe2(SO4)3 + 6H2O - Fe2O3.3H2O + 3H2SO4 (4) 3Fe2(SO4)3 + 14H2O - 2(H3O)Fe3(SO4)2(OH)6 + 5H2SO4 (5) (hydronium jarosite) The conventional zinc process involves concentrate roasting to produce zinc oxide and sulphur dioxide. The zinc oxide is leached in spent electrolyte and the sulphur dioxide is converted to sulphuric acid. The pressure leaching process can, therefore, be used to replace conventional concentrate roasting, sulphuric acid manufacture, and atmospheric leaching with one unit operation. Integration of pressure leaching into the Kidd Creek flowsheet A block flowsheet of the Kidd Creek zinc plant, with the new pressure leach plant integrated, is shown in Fig. 1. The original plant, which was constructed in 1972, is based on standard electrolytic zinc plant technology. Concentrates are dead roasted to remove sulphur. Sulphur dioxide is treated to produce sulphuric acid. Calcine is leached in two stages using cell-house electrolyte. Iron is precipitated as jarosite and solutions are purified to remove copper, cobalt, nickel, and cadmium to provide feed for a partially mechanized cell house. Zinc is plated from solution, and the resulting cathode is melted, cast, and sold (Johnston, 1983). The integration of the pressure leaching plant into the main zinc plant is relatively simple. As shown in Fig. 1, a fraction of the zinc concentrate and spent electrolyte are fed to the pressure leach autoclave, while part of the calcine from the roaster is used as a neutralizing agent in the oxidation step along with. the jarosite leach solution. The main exit stream from the pressure leaching plant is the slurry discharged from the oxidation step, consisting of neutral zinc sulphate solution, unreacted calcine, and precipitated iron oxides. This slurry is transferred to the neutral leach step in the main zinc plant. The leach residue from the pressure leaching step, consisting of basic iron sulphate and elemental sulphur, is washed and impounded along with the jarosite residue produced in the main plant.

Jan 1, 1987

By T. Maruyama

The Tamano Smelter of Hibi Kyodo Smelting CO. ,Ltd. went into operation in 1972. The Flash Smelting Furnace that was originally adapted was equipped with electrodes that were attached to the settler substitute of the slag cleaning furnace. Later the development of our own original coke combustion technology was promoted and a new Tamano Type Flash Smelting Furnace (T-FSF), without electrodes, was put into operation in 1988. The technology of coke combustion and combustion the burner have continued to improve. As a result the copper losses in the slag have been reduced as well as a stable operation has been achieved. Furthermore, this has been accomplished while increasing the concentration feed and the matte grade The current objective is to continue to strive for high matte grade operation and to raise production efficiency while paying attention to the important subject of copper losses in the slag. In this report, the data related to the copper loss in the slag reduction are presented along with the influence of the thickness, settling time and coke. The recent operation conditions of the T- FSF are also presented.

Jan 1, 2003

By Francisco Leite, Tomi Kouvonen, Bianca Saraiva, Vinicius Miranda, Auã Oliveira

One of the main challenges facing mining engineers is to draw up a plan that can be implemented in a timely manner, taking into account the operational variability of the mine’s production processes. These plans are often drawn up based on deterministic modelling (provided there are a set of known inputs that will result in a single set of outputs), when in truth we are talking about stochastic events (the cycles involved behave like random variables and their nature may be studied and reproduced). The result of using this kind of methodology can lead to non-optimized intuitive choices, which may harm the long-term targets defined by the planning team. Based on this principle, this document investigates the simulation of the rock fragmentation process in an open pit mine to assess the stochastic behavior of the loading and haulage operating cycles. The data assessed by the simulation process were the total costs of these operations in function of the powder factor. Based on a sample set data items, it was possible to check the probability distribution of the event as a Poisson distribution, applied to the loading and haulage cycles. The simulation consists of the repeated sequenced reproduction of these events (Monte Carlo simulation), in order to find the configuration that best fits the loading and haulage operations. By integrating these systems, it was possible to better understand the influence of the blasting in these operations, and as such to reduce the associated costs.

Feb 6, 2023

By David A. Keys

The object of this investigation was to determine with a vertical magnetic variometer the extent of the titaniferous ore deposit that occurs not far from Ivry in Terrebonne County, Quebec, about 70 miles north of Montreal. The Ivry mine, on lots 37 and 38 of range V, Beres-ford township, Quebec, is one of the largest and most accessible to transportation of the deposits of ilmenite in Quebec. The ore consists of an aggregate of ilmenite anhedra serving as host for numerous fine plaques of hematite. No magnetite is known from any occurrence near this lot, and in this way the deposit differs from many others that contain a large percentage of magnetite with the ilmenite. The deposit at Ivry thus has much lower magnetic susceptibility than those containing abundant magnetite, and a sensitive magnetometer was necessary for the work, although a body of magnetite with ilmenite about a mile from the Ivry mine was successfully prospected with a dip needle. The principal orebody is exposed in a quarry known as the Ivry mine, from which about 16,000 tons of ore have been shipped. Small exposures of ilmenite isolated by drift serve to show the presence at the surface of some of the bodies indicated by the magnetometer, but most of the hill is covered with glacial till and second growth hardwood. At the lower levels the till gives place to fluvioglacial sands and gravels from which the forest has been cleared. The increasing thickness of the sand and gravel as the hill descends makes the tracing outward of the deposits from the rocky knob of anorthosite a difficult matter. Within the orebodies too, some masses of anorthosite persist, which may account for some local irregularities. Fig. 1 is a photomicrograph, taken by reflected light, of a typical specimen of the ore. The dark patches are hematite and the light parts are ilmenite No trace of magnetite was found. Pig. 2 is a photomicrograph of a thin section of the country rock showing the texture of the anorthosite. The mean of four analyses quoted by Robinson1 is: titaninm, 19.26 per cent; iron 45.41 per cent. The ilmenite ore is not sufficiently magnetic to deflect an ordinary compass needle, but the effects of the induced magnetism in the ore by

Jan 1, 1940

By Peter Stacey

This presentation provides comments on the increasing challenges associated with pit slope designs from the perspective of a design reviewer. Besides the technical issues related to the significant increase in current and proposed slope heights since the primary methodology in current use for pit slope designs was developed, other factors are coming into play. These include changes in mining equipment and associated operating practices, as well as a growing awareness on the part of mining executives and other stakeholders of the significance of stable, or at least well-managed, slopes. In the latter area, the associated clarification of responsibility has led to the increased use of either in-house or independent review consultants working on the behalf of management. From a technical perspective, since the mid-1970?s there have been significant improvements in the tools, particularly analytical methods, that are available to slope designers. There is also an increased understanding of the importance of a detailed geological model, with strong emphasis on alteration and/or structure to form the platform for the slope designs. However, the basic technology for determining the rock strength aspects has changed little, although an expanding body of experience is being accumulated in its use and there is a reasonable degree of comfort in its application for slope heights of up to at least 500 m. In addition, slope management is now viewed as a critical part of the implementation of slope designs and this has been supported by continuing improvements in slope monitoring systems. In recent years safety has become a primary concern, in part as a result of some major slope failures that have attracted the attention of regulators and the public, as well as mining executives, who are being held more responsible for unsafe conditions and associated events. There is therefore a growing demand for risk assessments to replace the deterministic design approaches upon which the current design technology is largely based. The focus of mine executives on safety has undoubtedly in part facilitated the recent long overdue revitalization of research into the design and stability of large pit slopes, the requirement for which has been recognized for many years by practitioners. This research, as well as the CSIRO Large Open Pit Study, will almost certainly provide advances beyond the empirical strength determination methods such as the Hoek-Brown failure criterion, which form the basis for current rockmass strength determination. At the same time, as mentioned above, there is a growing awareness of the requirement for a detailed structural model as a major component of every large slope design. In addition, other areas requiring further research include the impact of groundwater pressures on rock mass strength, particularly in rocks with low permeabilities, and the role of stress in high open pit slopes. From the perspective of design implementation, even with the recent significant advances in metal prices, mine operators remain under pressure to minimize mining costs. To address these constraints, mining equipment of ever increasing size is being introduced. There are, however, some disadvantages to this trend in the area of slope design. For example, the large electric shovels are not well designed for scaling bench faces, nor are they cost effective in this mode. As a result, specialized equipment may be needed in the mining cycle to perform the clean-up duties, which in turn increases operating costs. Further, where the large equipment is used in minimum width pushbacks to reduce the instantaneous stripping ratio, the advance rates can be high, but at the same time the associated changes in operating procedures are often not conducive to the concurrent use of such measures as controlled blasting, careful scaling and drain hole installation that generally improve stability. The resulting conflicts which may arise between the interests of production and those of slope stability are often exacerbated by the fact that these stabilization techniques actually increase the operating costs on which the operations manager is frequently judged, even though there is an overall increase in profit. Meeting the objective of developing slope designs which are practicable , i.e., achievable in terms of every aspect of the operating constraints in the specific pit, requires interaction and compromise between the geotechnical engineer, mine planners and operating staff during the formulation of the design criteria. In the current environment, it is often no longer sufficient to present slope designs in deterministic terms to a mine planner who accepts them almost without question. Increasingly, the requirement is that they be proposed within the framework of risk levels, related both to safety and to economic outcomes, to a decision maker who may not be a technical expert in the mining field. In this context, the mine executives must have sufficient information and understanding to be able to establish acceptable levels of risk for the company and other stakeholders; in this process the slope designers must play a major role. These changing requirements for presentation of slope designs necessitate clear communication of the basis for the design, which may be very complex, so that the implications of the designs may be thoroughly understood by all concerned. To ensure clarity at this stage, the slope designer must not only quantify the uncertainty in the input parameters, including the geological model, but must also be able to communicate and defend the resulting design recommendations. Where particular designs may have high associated risks, simply because of the degree of uncertainty in the data, this must also be recognized and proposed methods of reducing the uncertainty included in the presentation. In summary, from a reviewer?s perspective slope designs must not only be technically sound, but must also address the broader context of the mining operation as a whole, taking into account such factors as safety aspects, the available equipment to implement the designs, and the acceptable risk levels for the company. In addition, the designs must be presented in a way that will allow the mine executives, who are ultimately responsible, and the operators, who implement the designs, to fully understand the basis and short-comings of the designs and the risks associated with deviation from any constraints defined by the designer. It goes without saying that there must also be a well-defined monitoring system to confirm stability and detect and manage any variations in the design model or unexpected instability.

Jan 1, 2006

By Robert H. Merrill

In 1960, the Kennecott Copper Corp. and the U.S. Bureau of Mines entered into a joint research program to determine the changes in stress, strain, and displacement created by changes in slope angles. The field site was the Kimbley Pit, located near Ely, Nevada; the supporting laboratory investigations were made at the Bureau's Denver Mining Research Center. Following a study to determine the rock stresses in the mining district, the strike and dip of geologic defects (fractures) in the rock, and the laboratory-determined strength of the rock, a new slope angle and pit configuration was selected, and instruments were installed to measure changes in stress, strain, displacement, microseismic noise rate, and sonic velocities at several points in the west wall of the pit. Between February and November 1966, the west slope was changed from an average angle of 45° to a new slope angle that started at 45° on the south wall, gradually increased to 57 ½ at the center of the west wall, and to 61° on the north wall. The pit was deepened from about 500 feet to about 550 feet. About 2.55 million tons of waste and ore were removed by stripping and mining.

Jan 1, 1968

By David A. Keys

The object of this investigation was to determine with a vertical magnetic variometer the extent of the titaniferous ore deposit that occurs not far from Ivry in Terrebonne County, Quebec, about 70 miles north of Montreal. The Ivry mine, on lots 37 and 38 of range V, Beres-ford township, Quebec, is one of the largest and most accessible to transportation of the deposits of ilmenite in Quebec. The ore consists of an aggregate of ilmenite anhedra serving as host for numerous fine plaques of hematite. No magnetite is known from any occurrence near this lot, and in this way the deposit differs from many others that contain a large percentage of magnetite with the ilmenite. The deposit at Ivry thus has much lower magnetic susceptibility than those containing abundant magnetite, and a sensitive magnetometer was necessary for the work, although a body of magnetite with ilmenite about a mile from the Ivry mine was successfully prospected with a dip needle. The principal orebody is exposed in a quarry known as the Ivry mine, from which about 16,000 tons of ore have been shipped. Small exposures of ilmenite isolated by drift serve to show the presence at the surface of some of the bodies indicated by the magnetometer, but most of the hill is covered with glacial till and second growth hardwood. At the lower levels the till gives place to fluvioglacial sands and gravels from which the forest has been cleared. The increasing thickness of the sand and gravel as the hill descends makes the tracing outward of the deposits from the rocky knob of anorthosite a difficult matter. Within the orebodies too, some masses of anorthosite persist, which may account for some local irregularities. Fig. 1 is a photomicrograph, taken by reflected light, of a typical specimen of the ore. The dark patches are hematite and the light parts are ilmenite No trace of magnetite was found. Pig. 2 is a photomicrograph of a thin section of the country rock showing the texture of the anorthosite. The mean of four analyses quoted by Robinson1 is: titaninm, 19.26 per cent; iron 45.41 per cent. The ilmenite ore is not sufficiently magnetic to deflect an ordinary compass needle, but the effects of the induced magnetism in the ore by

Jan 1, 1940

By A T. Spathis, M L. Lawlor-O’Neill

A methodology for the inversion of a scaled charge weight superposition model to predict peak particle velocity (PPV) has been examined. The inversion involves two steps: first simulated annealing is used to get an initial estimate of the model parameters; and second, these parameters are used as input to a quasi-Newton non-linear optimisation method, used to refine the value of the parameters. The methodology recovers the three required parameters (two charge weight scaling parameters and a time window width) and does this directly from a single production blast for a given point of interest. It has been tested on a synthetic example and recovered the input model parameters successfully in the presence of introduced noise with a coefficient of variation ranging up to ten per cent. Another test used the vibration measurements of three tunnel blasts monitored during the construction of the Citybanan tunnel in Stockholm. In this case, the three model parameters are found from one blast and for a given point of interest, and the PPV at the same point of interest is predicted for the second and third blast fired in the same vicinity as the first blast. The three blasts used different firing sequences. The predicted PPV limits use a confidence interval determined from the scatter in the peak levels obtained from the single calibration blast. The methodology was able to predict an upper bound for the PPV in 21 of the 26 tunnel blast vibrations studied. These predictions improve on a previous study involving the same model but which did not use the inversion process above. Given this reasonable success, further testing of the method on other data sets and alternative forward prediction models is recommended.CITATION:Lawlor-O’Neill, M L and Spathis, A T, 2015. An inversion method for the prediction of peak particle velocity in blasting, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 97–108 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By D Reid

Characterisation of non-plastic segregating tailings presents a number of challenges: the material cannot typically be sampled in an undisturbed state, remoulded samples exhibit different behaviour to undisturbed samples, penetration test interpretation methods suffer from uncertainty in their application to many tailings and the tailings gradation changes considerably over relatively small lateral distances and depth. In addition, the process liquor of many Western Australian tailings contains large quantities of salt and other dissolved solids, which can further complicate laboratory test programs. To provide an insight into the variability at a typical gold tailings facility, four gradations of the material were tested to obtain critical state parameters. Historic surficial tube samples were collated and organised by gradation to establish the range of densities typical after settling and air drying. These were then compared to laboratory-accessible densities. Finally, tube samples were tested in the laboratory to observe if the large-strain material behaviour was consistent with the inferred critical state parameters for similar gradations. Site pore fluid was used in the testing of the tube samples in an attempt to assess if the pore fluid modified the critical state conditions of the material.The tube samples tested all exhibited strongly dilative behaviour in undrained shear. Of the four samples tested, only one exhibited behaviour consistent with the inferred location of the critical state line relevant to the material’s gradation. The reasons for the unexpected behaviour of the other three samples include layering within the samples, the potential influence of large quantities of salt in the pore fluid and likely differences in the gradation of the fines components of samples taken from mid-beach and the decant area.CITATION:Reid, D, 2015. Observations on the behaviour of a gold tailings with hypersaline pore fluid, in Proceedings Tailings and Mine Waste Management for the 21st Century , pp 191–200 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 27, 2015

By D. Harrington

During the 5-year period 1906-10 inclusive, 84 major disasters (a major disaster being one in which 5 or more lives are lost) occurred in the coal mines of the United States, or an average of 17 per year, and the total number of fatalities from coal-mine gas and dust explosions (major plus minor) for the 5 years aggregated 2,338 or 477.6 per year. The monotonous continuance of coal-mine disasters month by month aroused the people of the United States to such an extent that it was felt necessary for the Federal Government to try to do something looking to the avoidance of the admittedly excessive loss of life in our coal mines, and this was one of the deciding factors in the establishment of the United States Bureau of Mires in 1910. That results have been obtained in this particular cause of coal-mine accidents may be seen from the fact that for the year 1933 but one major explosion disaster occurred in the coal mines of then united States and the total number killed in that disaster was 7. Unquestionably the coal-mining industry through use of oaf or mining practices and aided by the various kinds of work of the Bureau of Mines looking to the prevention of mine fires and explosions is in large part responsible for this very creditable change. If the gas and dust explosion fatality rate of 1.059 per million tons of coal produced during the period 1906-10 inclusive had continued to the first of January 1934, instead of the 6,373 deaths which actually occurred due to gas and dust explosions, there would have been 13,480. Hence, during the 23 years existence of the Bureau of Mines there has apparently been a saving of about 7,102 probable deaths, or an average of more than 300 per year, from gas and dust explosions as against the number which would have occurred had the 1.059 gas and dust explosion fatality rate for the period 1906-10 continued. Thirty-nine persons were killed in all kinds of explosions (both major and minor) in the coal mines of the United States in 1933; if the 1.059 gas and dust fatality rate of the 1900-10 period had been in effect in 1933, about 400 persons world have been killed. Unquestionably, the various teachings of the U.S. Bureau of Mines looking to the avoidance of explosions and fires in mines have had a vital influence in reducing the number of explosions and fires and also in decreasing their destruction in case they occur.

Jan 1, 1934

By Magnus Evertsson, Johannes Quist

HPGRs have since introduced in the 1980s (Kellerwessel, 1993; Schönert, 1988), now become integral comminution units at several minerals processing plants around the world. Even though a lot of the initial issues concerning performance, availability and wear have been addressed and solved there are still phenomena and aspects which can be improved and understood further. Schönert and Sander (2002) presented experimental results and a mechanistic model including the internal force dynamics between the rollers and the compacted bed. They derived expressions for the pressure distribution involving most of the influencing terms and effects. However, in order to solve the differential equations several of the terms vanish by assumptions. The Schönert and Sander model reflects the experimental data in many aspects but the predicted shear force distinctly deviates from the experimental results at the angular position of the peak pressure level. This finding suggests that the particle to particle interaction and breakage process in the different zones seen in Figure 1 are complex and difficult to model using analytical models. A numerical modelling approach using DEM does not require the same amount of assumptions and simplifications of boundary conditions. Hence such modelling may provide further knowledge and vital understanding. In this paper the breakage process is simulated in order to investigate the shear slip behaviour and how it relates to different operating variables. A laboratory scale HPGR corresponding to the experimental setup used by Schönert (2002) have been modelled in EDEM (DEM-Solutions). The bonded particle model (BPM) is used for modelling quartz particles (Potyondy, 2004). The breakage model strength parameters are calibrated against interparticle breakage experiments in an Instron 400RD compression device. The contact model parameters are calibrated using calibration methods developed by Quist (2015). The floating roller dynamics are modelled using a spring-damper system. Results present insights regarding the interparticle slip behaviour as well as the frictional tangential force distribution on the rollers. The simulated roller pressure distribution shows good correspondence with previously reported findings in the literature.

Jan 1, 2016

By Mario Giavani

"IntroductionBy the 1980s, the continuous casting process had almost replaced the ingot casting of steel in Japan, Europe and North America. While economic benefits were probably initially the prime motivator, the transition rapidly proved that continuous casting in conjunction with ladle metallurgy offered a much greater degree of process control and made possible routine steel specifications not previously available.This new technology demanded improved procedures for the quantitative addition of the required metallurgical reagents, especially as regards the more volatile or toxic elements such as calcium , boron, carbon, sulphur, lead and tellurium. The injection of a thin steel tube containing the additive into the metal bath of a ladle refining station offers the best procedure for this. In the industry this has been referred to as ""Cored Wire "".Operators using this technology find that they are faced with balancing the injection speed against the small cross section of the tube to feed the required amount of reagent within the short time span available to the operation. Large-ladle shops requiring major quantities usually resort to multiple machines and high speeds. Many machines approach a maximum feed rate of 400 m/min. Higher speeds exceed the melting rate of the system, and the ferro static buoyancy of the tube will cause it to coil up the ladle wall and possibly resurface into the slag layer, and even eject itself above the surface. Under these conditions, exact control of element recovery becomes difficult, especially with the commonest addition, calcium silicon alloy.Many experts are persuaded that the final melting position of the carrier tube has major importance in establishing consistent element yield to the bath.For instance, the vapour pressure of calcium at l500 °C is close to I bar. Reaching a melting depth sufficient to accommodate this value is considered to be critical to achieving an acceptable and reproducible yield.Many studies have elaborated semi-empirical formulae to establish the melt zone depth, connecting this value essentially to the diameter of the tube , the injection rate and the superheat of the steel. The author's experience in steelmaking could not confirm such formulae in many cases, and he set about to develop a mathematical model which would be useful in addressing this problem, taking into account the several parameters which will regulate the behaviour of a filled tube immersed in molten steel."

Jan 1, 1991

By P. Middendorp, Gerald E. H. Verbeek

"INTRODUCTIONStatic Load Testing (SLT) is generally considered the pile testing method that provides the de-facto answers regarding pile capacity and load-deflection behavior. However, as this testing method is time consuming and costly, alternative pile testing methods are applied, such as High-Strain Dynamic Testing (HSDT), commonly referred to as Dynamic Load Testing (DLT), and Rapid Load Testing (RLT), which includes Statnamic Testing (STN) and the Fundex Pile Load Tester (PLT). Extensive descriptions of these pile testing methods as well as comparisons of the results from these two methods have been widely published, but little is written on the choice between HSDT and RLT. This article will address this issue for concrete piles by discussing five different aspects that should be considered when making a choice between HSDT and RLT for either cast-in-situ piles or precast driven piles. TIMINGThe timing of the pile load test is independent of the testing method. For cast-in-situ piles load testing is performed some time after the pile has been installed (to allow the pile to obtain the compressive strength required to withstand the test load), while for precast driven piles load testing is performed after a so-called set-up period following pile installation (to allow the soil to recover from driving induced disturbances and, at least in most cases,regain strength).In other words, the timing of the test does not affect the choice of the testing method.ACCURACY OF THE APPLIED LOADWhen a pile is tested by RLT the applied load is measured by a calibrated load cell placed on the pile head, just as it would be during SLT. This means that the error in the applied load value is less than 0.1% of the maximum range of the load cell.However, when a pile is tested by HSDT the load is calculated from the measured strains, which means that the outcome is dependent on the pile material properties. In this case the applied load (F) on the pile head is calculated by multiplying the measured strain (?) with either the calculated or estimated elastic modulus of concrete (E) and the pile cross section (A).The elastic modulus is generally back calculated from the (average) stress wave velocity (c), which in turn is derived from the reflection time (T) it takes for a stress wave to travel over the pile length (L) from the pile head to the pile toe and back to the pile head."

Jan 1, 2006

By I. N. Zavarine

THE present paper is a continuation of the work on the relationship between the magnetic and the phase transformations in carbon steels during quenching. An account was given by the author in a previous paper1 of an investigation of this relationship during a continuous quench. The experimental difficulties then encountered prevented a conclusive proof of the existence or absence of such a relationship. The conclusions of the previous paper were derived indirectly. The present paper deals with the same problem, but the method of attack is modified. The technique developed by Davenport and Bain2 in their well-known study of austenite transformation at subcritical temperature was adopted in the present investigation.. They have shown that the austenite to martensite transformation can be caused to proceed at a relatively slow rate by quenching steels into a hot metal bath. Slowing down of the rate of transformation, together with an observation by Lewis3 that a eutectoid steel can be preserved in the nonmagnetic state by such quenching, pointed out a possibility of a study of both transformations .under more favorable conditions than previously described. The preliminary steps were connected with the technique of the hot bath quench. To study this phase of the present work a modification of the apparatus previously described1 was used. The modification consisted in the use of a hot quenching bath instead of brine, and in fixing the secondary and the primary coils of the apparatus in a stationary position in and around the bath. Tubular specimens, as in the previous investigation, were heated in a suitable furnace to a temperature in the austenitic

Jan 1, 1935