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By E. G. Joe

"This paper summarizes the main conclusions derived from twenty-two technical presentations made at the 1973 meeting of the Canadian Mineral Processors. Results of two surveys with response from over seventy mines were reported on Canadian milling practice for water re-use and on elimination of cyanide from mill effluents. Water and solution recycle in Canadian mills were detailed for copper, copper-zinc, lead-zinc-copper, nickel-copper, iron and gold mines. Cyanide elimination from mill effluents was reported under current practice and under requirements for further treatment by various other processes. Three recent process equipment entries into Canadian mills have been the on-stream particle analyzer, the new gravity concentration tables for recovery of fine tin and other heavy minerals, and solution-jet slurry equipment for reclaiming ground ore from storage. The principles of operation of this equipment were described along with the practical limitations that were found in their application to mineral processing. Capital costs and milling cost reductions were shown in some operations. Computer control of flotation mill processing was a major item on the program. Four major mining companies outlined their mill process control strategy, computer programs and on-stream analysis equipment. Some capital costs and personnel changes were presented, together with gains in concentrate grade and mill recovery. Recent development work has shown the effect of the type of grinding and of aeration on the recovery of sulphide minerals in flotation. One mining company reported a significant depressing effect on copper sulphide flotation while floating ore ground with mild-steel balls; also, pyrite appeared to be activated in the subsequent zinc flotation step. One company with extensive experience in the application of aeration to copper sulphide flotation showed significant improvement to copper recovery by pre-aeration. Both the effects of grinding and aeration were subject to the constraints of the chemical condition of the ore pulp. The metallurgy of silver recovery by combination grinding - gravity treatment and of gold recovery from a complex gold-sulphide ore were reviewed in two papers on precious metals. The latter paper further described the recovery of gold from the Cottrell collectors and the final disposal of arsenic from the dust collectors."

Jan 1, 1973

By H. S. Thomsen, W. C. Duesterhoeft, R. E. Hartline

The attenuation and phase shift which formations produce in the electromagnetic field of an induction-type electrical well-logging instrument are great enough to substantially affect the response of the tool to formation conductivity under normal logging conditions. The application of the general equations for the propagation of an electromagnetic field in a conductive medium to the transmitter-receiver coil pair of an induction-logging-tool coil system gives an expression for the response which properly takes the propagation effects into account. A comparison of a calculation of this type with the response computed from the commonly used geometric factor concept, which does not include the propagation effect, leads to a factor Go representing the ratio of the responses computed by the two methods. The value of Go decreases not only with increasing formation conductivity, but also with increasing transmitter-to-detector coil spacing. For a single coil pair with a 40-in. spacing, the value of Go is 0.972 in a 20-millimho/m formation. The value is reduced to 0.915 and 0.812 for conductivities of 200 and 1,000 millimhos/m. As a result, the basic signal generated by the induction-logging coil system is not linearly related to formation conductivity as expected from the geometric factor concept. A uniform conductivity scale can be obtained only by adding a suitable nonlinear element to the recording system. The addition of auxiliary coils, having spacings less than the main-coil span, to achieve focusing results in an even greater departure of the basic signal from linearity with conductivity. The solution of the field equations near the interface between two formations of different conductivity gives the curve shape on an induction logging tool in crossing the interface. The addition of auxiliary coils to achieve focusing can add anomalous character to the curve shape in crossing the interface, which might be mistaken for lithological detail. Preliminary calculations which include the propagation effects in the determination of the conductivity correction for thin beds lead to correction factors which are substantially smaller than those obtained from geometric factor considerations. It is apparent that thin-bed corrections derived from geometric factor calculations are of doubtful validity. INTRODUCTION Although the electromagnetic induction type of electrical well log has been used for a number of years and has gained general acceptance in quantitative well-log interpretation, no technically complete investigation of the mechanism and of its operation has been published. The studies which have been presented are based upon the premise that neither the shape nor the intensity of the induction field of the logging tool is in any way affected by the electrical characteristics of the formation. From the inception of induction logging it has been recognized that under some formation conditions the performance of the instrument could be expected to be substantially different from that predicted by the simplified analysis. This anomalous performance has been called "skin effect" since it is the result of the same phenomenon which causes high-frequency alternating currents to flow only near the surface of metallic conductors. In the range of conductivities encountered in the sediments, the formation actually does not exhibit the sharply defined conducting "skin" shown by metals. The gradual modification of the distribution of the currents in the formation, produced by the induction field of the logging tool, is the direct result of phenomena associated with the propagation of the electromagnetic field through the formation. Therefore, it can be much more appropriately described as the "propagation effect". One purpose of this paper is to review the results of a solution of the equations for the performance of a single transmitter-receiver coil pair induction-logging system which takes into account the propagation effect, showing the manner in which the tool performance established by this analysis differs from that derived from the less comprehensive analyses currently in use. PRINCIPLE OF OPERATION OF THE INDUCTION LOGGING INSTRUMENT The high-frequency alternating current which is maintained at a constant value in the transmitter coil

By J. W. Hawkins, K. B. C. Brady, M. W. Smith

The Pennsylvania Department of Environmental Protection (PaDEP) has been issuing surface mining permits since 1984 that authorize remining in areas that will affect preexisting pollutional discharges. These permits require the mine operator to employ best management practices to reduce pollution loading from abandoned mine discharges. A total of 110 completed mining operations with 233 discharges were evaluated to determine the impact to water quality. Overall, acid loading from these discharges was reduced by 7, 200 kg (15, 900 lbs) per day or 61%. Metal loadings were also reduced. For acidity, 49% of discharges improved or were eliminated, 52% showed no significant change and fewer than 1 % got worse. Loading reductions resulted from a combination of reduced concentrations and lower flows.

Jan 1, 2003

By Dermot Winters

The Surface Mining Control and Reclamation Act (SMCRA, or the Act) was enacted on August 3, 1977. The Act was written, in part, to put an end to uneven environmental standards and to establish minimum national standards. Certainly, one of the motivating factors for passage of the Act was the success of the Federal Coal Mining Health and Safety Act of 1969 in reducing deaths and injuries in the coal industry nationwide. Section 516 (b) (1) of the Act requires *that subsidence be prevented from causing material damage to the extent technologically and economically feasible except in those instances where the mining technology requires planned subsidence in a predictable and controlled manner. Section 516 (c) of the Act authorizes the regulatory authority to suspend underground mining if there is imminent danger to inhabitants of urbanized areas, cities, towns, and communities. On December 13, 1977, the interim program, which applies a portion of the performance standards of the Act, was promulgated. Then on September 18, 1978, the proposed permanent program was published for public comment. Finally, on March 13, 1979, the final permanent program was promulgated and the States were given a period of time to submit their programs for approval.

Jan 1, 1986

By Peter Willbrandt

The flash smelter at NA was commissioned in March 1972 with an initial throughput of 50 tons of dry copper concentrate per hour. As a result of various measures, for example, oxygen enriched process wind, extension of the radiation part of the waste heat boiler, etc., the throughput could be increased to the present 70 tons per hour of dry concentrate. NA's plant is located on the outskirts of the city of Hamburg. The flash smelter's location and the growing awareness of the German population as regards environmental issues necessitate again and again considerable capital expenditure being spent on environmental protection. Thanks to the numerous small and large measures taken to protect the environment the emissions of dust, heavy metals such as arsenic, copper, lead, etc., and as well SO, could be drastically reduced. This paper will be giving an overview on the present state of the flash smelting process and the impact of the increasing environmental awareness at NA.

Jan 1, 1993

By R. Gallinger

In 1997, a group of Canadian mining companies met to review the Mine Environment Neutral Drainage (MEND) program and to consider future activities. They concluded that (1) that Canadian industry and Natural Resources Canada (NRCan) should fund a small secretariat in Ottawa to support technology transfer in Canada and, (2) that it was important for industry to continue the momentum of MEND on an international scale. The need for the International Network for Acid Prevention (INAP) was strongly endorsed by consultants, company, university, and government experts from Australia, Canada, and the United States who convened in Vancouver in February 1998. The initial focus was the development of a solid knowledge base covering the collective experience of INAP?s members to promote the dissemination of successful practices, the identification of knowledge gaps, and the formulation of effective research proposals. The sharing of information on past successes and failures alone was viewed as sufficient justification for the formation of INAP. INAP was officially launched in October 1998, and is now an effective, industry-based initiative to globally coordinate research and development in the management of sulphide mine waste. Significant improvements in the management of sulphidic mine materials through knowledge sharing, research, and development of technology by INAP will reduce liability associated with acid drainage. This paper will explore the international efforts and activities of INAP

May 1, 2005

By Michael Weaver, Dale L. Ramsey

The Pinkerton Tunnel open cut project is part of the National Gateway Clearance Initiative improvement project to achieve a minimum of 21 feet of vertical clearance along CSX Transportation rail corridor so that double-stacked intermodal railcars can be transported between mid-Atlantic states and their ports to midwest markets in the U.S. This project involved drilling and blasting of 1.2 million bank cubic yards of rock over a 14 month period while maintaining rail traffic of two passenger and 38 freight trains daily. See story and photos, page 6.

Jan 1, 2015

By D. Bernabe

The mine scheduling process defines a mining sequence across the life of the mine and stipulates the amount of ore, waste, and specific blocks to be mined each year so that the highest economic return can be obtained under a number of constraints. There are a number of mine scheduling software packages available. Each claims to be the best scheduler by maximizing Net Present Value (NPV) or by minimizing overall mining costs. These mine scheduling programs follow different procedures and scheduling philosophies based on either mathematical or heuristic principles. There was no study in which the mine production scheduling procedures applied by these packages were discussed and compared. This paper evaluates four open pit mine scheduling programs recently developed and compares them in order to determine if significant differences exist among them. The same input data, which consist of a 3-D block model, an economic set of parameters, and the initial surface topography for a operating copper mine in Southern Peru, was used to test each technique. The following programs were investigated with respect to their ability to optimize the NPV for the case study. 1) M821V1 Scheduler by Mintec; 2) NPV Scheduler by Earthworks; 3) Milawa Scheduler by Whittle Programming; and, 4) Fundamental Tree Scheduling Method (FTSM) developed at the Colorado School of Mines. The results from this case study demonstrate that the scheduling software packages, Mintec, Earthworks, and Whittle obtained similar NPV’s, while FTSM scheduler gave a six percent higher NPV than the other schedulers.

Jan 1, 2002

By R. E. Schilson, F. H. Poettmann

The direct determination of the stabilized performance behavior of low capacity, slowly stabilizing gas wells is extremely time-consuming and wasteful of gas. From both field experience and theoretical considerations, a test procedure has been evolved by which the stabilized hack pressure behavior of such gas wells can he predicted without having to revert to long time flow tests. The method consists of using the isochrona1 test procedure to establish the slope of the back pressure curve, "n", and the short time variation of the performance coeficient, "C", with time. From this short time transient flow data and theoretical considerations, the value of C at large times can he established. By assuming the radius of drainage of a well to be half the distance between wells, one can calculate the stabilization time for various well spacing patterns. Once the stabilization time for a given .spacing has been determined, the value of C can be calculated and the stabilized back-pressure curve can he. establbrhed. The calculated perfortnance coefficient as a function of tinze was cotnpared to the experintentally measured values for a number of gas wel1.e. The deviation of the calc~*lated from the cxperimental res1tlt.s vary depending on the set of short time experimental points used to evabrate the parcrttzeters of the equation. The longer the tirne far the flow test data user1 in the calculations, the better was the agreement with the experimental results. The time necessary to obtain this data from well tests varies consitlerably, depending on the physical natrcre oj the rt3scrvoir under consideratiot~. INTRODUCTION For many years, the U. S. Bureau of Mines Monograph 7' has served as a guide for testing and evaluating the performance of gas wells by means of the back-pressure method. The back-pressure performance of a gas well is expressed by the following equation: Q~ CW-W.........(I) where the characteristics of the back-pressure equation are determined by C, the performance coefficient, and n, the exponent which corresponds to the slope of the straight line when Q and {P* - PN2) are plotted on logarithmic paper. Q is gas flow rate at standard conditions, and P, and P, are equalized and flowing bottom-hole pressures, respectively. Prior to the development of the back-pressure tcst, the "open flow" capacity method of testing a well was common. By this method, the wells were flowed wide open to the air and the flow rate measured. Such procedure was wasteful of gas and did not provide information on the deliverability of the gas to the pipe line. Monograph 7 Procedure The back-pressure method of testing wells was developed to overcome these shortcomings. Although much has been learned regarding the laws of the flow of gas through porous formations, the original development of the back-pressure relationship was based entirely on empirical methods. The back-pressure behavior provides the engineer with information essential in predicting the future development of a field. It permits him to calculate the deliverability of gas into a pipe line at predetermined line pressures, to design and analyze gas gathering lines, to determine the spacing and number of wells to he drilled during the development of a field to meet gas purchasers' requirements, and to solve many other technical and economic problems. As described in Monograph 7, the flow-after-flow method of back-pressure testing, when applied to fast stabilizing and usually high capacity wells, correctly characterized the behavior of the well. However, as the value of the gas at the wellhead increased, small capacity gas wells having slow rates of stabilization became economically operable. The flow-after-flow method of testing could not be used to describe the behavior of these slowly stabilizing wells. The procedure of Rawlins and Shellhardtl for establishing the back-pressure behavior of a gas well was based on the rcquirement that the data be obtained under stabilized flow conditions; that is, that C is constant and does not vary with time. C depends on the physical properties of the reservoir, the location, extent and geometry of the drainage radius, and the properties of the flowing fluid. In a highly permeable formation, only a very short period of time is required for the well to reach a stabilized condition, and, consequently, the requirements for the test procedure described in Monograph 7 are met. For a given well, n is also constant

By Kia Riihioja, Mike Higgins

A major difference between electronic and pyrotechnic detonators is that, in almost all but the simplest of blast layouts, computer software is required to plan a blast using electronic detonators. While it may be possible to program delay times without software, using only the manufacturer’s control hardware, the full potential of electronic detonators can only be realized through software. Until recently, the only software available has been provided by the suppliers of the electronic detonators. These usually have just enough features to create a hole layout and to assign times to the detonators in the holes. The software is also able to communicate only with the manufacturer’s detonators. This means that the client may be compelled to change several parts of the blasting process when a different delay system is adopted, from design through to implementation and reporting, even though these may be sub-optimal at the client’s mine. While many authors have demonstrated improvements in blasting outcomes using electronic detonators, very few have discussed how the delay times were calculated. Where the calculations are discussed, it is also obvious that the delay times must be monitored and re-evaluated as conditions in the mine change. For example, the principles of superposition and scaled distance, burden relief and energy cooperation between holes should be checked constantly for different locations within the mine, changes in rock properties, or even for differences between design and actual hole positions. The most common methods in blast design software to calculate delay times are inter-hole interval, relief contour and direct editing. Although they are simple in concept, they can result in fundamental errors if they are not applied correctly. The software for electronic blast designs should be capable of not just a timing calculation, but also of showing, to some degree, the impact and interaction of the calculated times.

Jan 1, 2004

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By H. J. Levinstein, H. J. Guggenheim, C. D. Capio

An optical incestigation of the phase transformations in RbFeF, has been conducted. Details of the ferromagwetic phase transition and the metamagnetic state are disczrssed. The three-dimensional magnetic domain structure existing in bulk crystals of RbFeF3 is described, as well as the effect of the magnitude and direction of the applied magnetic field on the domain struclure. The strong magnetoelastic interaction in RbFeF, is demonstrated. The results of the direct observation of the role of dislocation subboundaries and slip bands as nucleation sites and as pinning sites for magnetic domain walls are reported. THE perovskite compound RbFeF, has recently been shown to exhibit three magnetic states.''' Above 102°K it is paramagnetic, between 102" and 87°K it is antiferromagnetic. and below 87°K it is ferromagnetic (i.e., exhibiting remanence) with a modification of the magnetization occurring at 40°K. The changes in magnetic structure at 102o, 87°. and 40°K are accompanied by shear transformations to successively lower crystal symmetry classes.3 At 102°K a second-order phase transformation to a tetragonal crystal structure occurs. The c/o ratio increases with decreasing temperature to a value of 1.0034 at 87°K where the crystal undergoes another shear transformation to an ortho-rhombic crystal structure. The magnetic modification at 40 K is also accompanied by a shear transformation to a lower crystal symmetry. RbFeF3 is unique in that in the ferromagnetic state it is transparent in the bulk to visible light, has a low saturation magnetization, a large magnetic rotation. and has good optical properties.' All these features make it an ideal material for the investigation of magnetic domain structures in bulk crystals. In addition a dislocation etch has been developed which reveals the point of emergence of dislocations with the (100) surfaces of RbFeF3,5 making it possible to determine the dislocation arrays in the material. As a result domain wall dislocation interactions can be observed in the bulk crystal. In this paper we report on 1) the crystallography of the phase transformations in RbFeF3. 2) the domain configurations as a function of magnetic field and crystal orientation. 3) the interaction of dislocations and magnetic domain walls in RbFeF,. EXPERIMENTAL PROCEDURE Since the temperature range of interest is below room temperature the dewar shown in Fig. 1 was employed. It was designed to fit on the stage of a Leitz panphot metallograph, and permitted examination of the sample at a magnification up to 150 times. The dewar consists of a double-walled glass cylinder bent into an L shape. The space between the walls is evacuated. The viewing wirldows were made of four optically flat quartz discs aligned parallel to each other and sealed to the sides of the concentric cylinders. The specimens were mounted in a holder attached to a flexible plastic shaft. Various holder designs were employed depending upon the type of observations to be made. For studies of the phase transformation between the antiferromagnetic and ferromagnetic state the holder shown in Fig. 2(d) was employed. This holder permitted accurate temperature control at 82" to 88°K by balancing the heat input from the carbon resistance heater against the heat loss to a heat sink immersed in liquid N2. Magnetic field studies were conducted by employing the holders shown in Figs. 2(a) and (b). The holder shown in Fig. 2(n) has a solenoid imbedded in it. such that the magnetic field direction is in the same direction as the incident light. The magnetic field provided by the holder in Fig. 2(b) is perpendicular to the incident light direction. The holder in Fig. 2(r) was employed when observations of the specimen were desired while an elastic bending stress was applied to the sample. The stress was applied to the sample by pulling a wire from outside of the dewar. The wire was attached to a lever pivoted on the holder. which caused the knife edge of the lever to push against the sample. The crystal growth and the sample preparation were described previously.' PHASE TRANSFORMATIONS The first phase transformation in this system is from the cubic perovskite structure to n tetragonal

Jan 1, 1970

By John W. Cole

The Butte district has been the principal producer of metallurgical-grade manganese in the United States since World War I. Manganese dioxide nodules are produced from rhodochrosite ore by concentrating the rhodochrosite and roasting the concentrate in rotary "cement-type" kilns. The Anaconda Copper Mining Co. is the only producer at present. Under the stimulus of a purchasing program administered by Metals Reserve Company during World War II, however, several independently owned or operated mines (including the Norwich) produced manganese carbonate ore. These mines were examined in Juno 1949 by John W. Taber, mining engineer, Bureau of Mines. A core-drilling program to develop ore in the Norwich mine was recommended and approved. drilling was completed early in 1950. This report gives factual data obtained.

Jan 1, 1951

By R L. Chaney, M Mahoney

Mine and ore beneficiation wastes and smelter contaminated soils often cause phytotoxicity and threaten adverse environmental effects if not remediated. Science has clarified both the risks from soil metals and methods to alleviate those risks that can be applied at low cost. Phytoremediation is a family of plant and agronomy based technologies to deal with environmental aspects of sites requiring remediation. If valid risk assessment indicates that metals comprise risk, phytostabilisation may be used to alleviate that risk for most contaminated sites. In particular, Zn-Pb, Cu and Ni mine and smelter sites are readily remediated by making the soil calcareous and fertile, with added organic and metal sorbent amendments that include soil microbes to ‘revitalise’ the soil. In each case, the soil fertility and causes of phytotoxicity should be clarified by agricultural soil extractions, and the combination of amendments required to improve fertility and reverse phytotoxicity identified and found locally. In contrast with phytostabilisation, phytoextraction/phytomining uses rare metal hyperaccumulator plants and agronomic management to maximise annual phytoextraction of soil elements into plant shoots for harvest, ashing and use as a metal ore. Plants that accumulate more than two per cent Ni in shoots, and yield 10–20 t ha-1 dry biomass yield 200–400 kg Ni ha-1, which has value greater than all common agricultural crops. Considering the value of elements, existence of hyperaccumulator plants and demonstrated agronomic management for Ni phytomining, it is clear that mine sites and Ni mineralised subore-grade serpentine soils can be phytomined economically as an agricultural mining technology. A team developed agronomic methods and improved cultivars of Alyssum, which can be profitably grown on temperate zone Ni mineralised or contaminated soils. Plant species for tropical ultramafic soil materials have been identified. Phytoextraction is also strongly needed to remove Cd from rice soils to prevent human disease from soil Cd.CITATION:Chaney, R?L?and?Mahoney, M, 2014. Phytostabilisation and phytomining – principles and successes, in Proceedings Life-of-Mine 2014 , pp 3–16 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By P. Jurovcik, D. Yantek, D. Ingram

The National Institute for Occupational Safety and Health (NIOSH) identified hearing loss as the most common job-related disease in the United States in its 1996 National Occupational Research Agenda. Previous studies by NIOSH have shown operators of surface drill rigs without cabs are commonly exposed to A-weighted sound levels exceeding 90 dB. During hammer-drilling, A-weighted sound levels exceeding 100 dB were recorded at the operator station. NIOSH collaborated with a local drilling company to design and install a partial cab between the operator and the drill steel in an effort to reduce the sound level at the operator station. Sound level measurements in the field show the partial cab reduces the A-weighted sound level by 5-9 dB at the operator’s left ear and 2-5 dB at the operator’s right ear while hammer-drilling. This paper will briefly discuss the preliminary design of a cab and the testing of materials for construction of the prototype. Explanation of the ideas behind the prototype design and fabrication will be covered. Finally, the field test method and results will be discussed in detail.

Jan 1, 2007

By A. E. Bellis

ROBERT J. ANDERSON, Cleveland, Ohio (communication to the Secretary *).-The paper by Messrs. Bellis and Hardy was interesting to me and has led me to make a few remarks concerning some of the points brought out. The question of employing correct hardening temperatures is one that is sadly neglected, particularly in many large plants that forge and treat their machine tools for private use. As is most aptly remarked in this paper, "The custom of using one, high-speed temperature for all tools is very poor practice;" still, such practice is more common than otherwise. An instance came to my attention some time ago, in connection with some work with the standardization of high-speed steels for machine tools, where there were no less than six different brands of alloy steels being used for machine tools such as lathe, shaper, planer, boring mill and cutters of like nature. All of these different steels received identical heat and forging treatment-or at least that was the aim of the tool hardener. That the results were highly unsatisfactory goes without saying. That the hardening department of many plants is conducted on a rule-of-thumb fashion is the reason for a great deal of the trouble and disappointment which seems to be the lot of many tool-steel users. Exceedingly crude execution is the rule rather than the exception, and it is no small wonder that the results are as good as they sometimes happen to be. It would seem that at the present price of alloy steels the subject should receive a closer scrutiny by purchasers than it does. One not familiar with the workings of souse plants that disdain to lend an ear to scientific heat treatment would scarcely give credence to some of the methods employed, not only in the treatment of the steel but also in its selection and purchase. Heating for hardening should be conducted under close pyrometric control only after the correct hardening temperature has been determined.

Jan 3, 1917

By M. R. Geer, H. F. Yancey

THE performance of coal-washing equipment, that is, the effectiveness with which coal and impurity are separated, is a subject of undisputed interest to all concerned with the preparation of coal. Through the years much has been written in both American and European literature on the performance of coal-washing equipment and its interpretation, but until recently no effort has been made to consolidate this wealth of information into a single, comprehensive treatment of the subject. A splendid effort in this direction was made at the First International Conference on Coal Preparation held in Paris in June 1950 in which representatives from virtually every coal-producing country participated. Although the scope of the conference covered the entire field of coal preparation, many of the sixty-odd papers and a great deal of the discussion dealt with the subject of evaluating washery performance. Prominence was given this subject by the Centre d'Etudes et Re-cherches, des Charbonnages de France, which sponsored the conference. This research organization, generally referred to as Cerchar, was formed in 1948 following the nationalization of the French coal industry. Since that date it has conducted an intensive study of coal preparation. Perhaps because of a fresh viewpoint, Cerchar realized immediately that complete information on the performance of French washeries and a thorough understanding of the mechanism of gravity separation were the fundamental prerequisites in a program to improve preparation practice. For this reason Cerchar concentrated on obtaining performance data and on a comprehensive study of methods of evaluating such information. Some idea of the magnitude of this work will be appreciated from the fact that 175 performance tests were made in 44 plants. The work performed by Cerchar emphasized a condition that is recognized by many of those concerned with coal preparation, namely that the lack of uniformity in presenting data on washery performance, the diversity of methods used in evaluating such information, and the frequent confusion of the terms related to performance have been prime factors militating against the progress from art to science. Thus, Cerchar feels that standardization of the methods employed for evaluating washery performance, not only in France but in the entire world, is essential to progress, and the prime objective of the Paris conference was to seek international standardization of the evaluation methods advocated by Cerchar. At the conclusion of the conference the question of standardization was referred to the International Standards Organization. A careful study of the dozen papers on the subject of evaluating washery performance presented at the conference reveals that none of the methods is new; most of them date back over a decade. Nevertheless, the intensive scrutiny of these methods brought about by the conference unquestionably has resulted in a valuable exchange of ideas and a clarification of thought that could scarcely be achieved in any other way. The object of the present report is to make available to the American coal industry the ideas developed at the conference which seem most important from an American viewpoint and to provide a critical analysis of some of the evaluation methods advocated by Cerchar. The various methods of evaluating washery performance are illustrated partly with data that have been published in American literature, and some new data have been obtained as the result of experimental work undertaken since the conference to provide information on certain controversial points. An English translation of the proceedings of the conference, including the discussion as well as the papers, will be available,' and previous to the conference Coal Age2 published a very brief description of some of the principal evaluation methods used by Cerchar. Americans will appreciate as well as do the French the desirability of standardizing both the termi-

Jan 1, 1952

By John Tyssowski

In 1925, Harry Howard Stout, then metallurgist for Phelps Dodge Corporation, while investigating the cleaning of cathode copper by various gases at elevated temperatures below the melting point of the metal, observed in the laboratory that particles of pure copper heated in a closed tube filled with hydrogen seemingly "stuck together." By further investigation he found that there was actual crystal growth across the boundaries of the particles of pure metal; in other words, there was a complete new crystal realignment and growth throughout the newly formed copper mass. It was coalescence, not adhesion; the several particles had become one unit. From this discovery has been developed the coalescence process, whereby copper of cathode purity is squirted out in semifabricated form from an extrusion press, by the application of about 3000 tons pressure on a 10-in. piston. Stout decided that cathode copper produced by electrolytic deposition was the logical source of the pure copper required for the coalescence process, but as there was no way of reducing the ordinary tough commercial cathodes to the small particles required he was confronted by a serious problem at the outset. Under his direction, this problem was solved by William H. Osborn and Harry Howard Stout, Jr. They developed the process and technique for producing cathode copper in brittle, friable form, and to them must be given the greater part of the credit for this phase of the development program as well as for that relating to the gas cleaning of the cathode particles. Concurrently, the three metallurgists worked on the problem of effecting a complete union of the cathode particles. Adequate washing was all that was needed to free the particles from electrolyte solution and slimes. A press was designed to produce briquettes of the proper porosity and density, and a method was devised of removing surface oxidation from the particles of cathode, by treatment in an atmosphere of deoxidizing gas at an elevated temperature. The temperature was determined at which the surface cleansing and preheating could be done in one operation, and for this a charging and preheating furnace in which there is maintained a positive pressure of deoxidizing gas atmosphere was designed. And finally, a method of feeding the heated briquettes of deoxidized cathode copper into an extrusion press without contamination from the air was evolved. For the final stage, a press was required because not only heat but pressure was required to obtain complete growth of the new crystals between the cathode particles as well as the elimination of all voids or non-coalesced areas. All essential stages of the process are protected by patents or patent applications in the United States and principal foreign countries, and the copper is produced on a commercial scale.

Jan 1, 1941

By John Tyssowski

In 1925, Harry Howard Stout, then metallurgist for Phelps Dodge Corporation, while investigating the cleaning of cathode copper by various gases at elevated temperatures below the melting point of the metal, observed in the laboratory that particles of pure copper heated in a closed tube filled with hydrogen seemingly "stuck together." By further investigation he found that there was actual crystal growth across the boundaries of the particles of pure metal; in other words, there was a complete new crystal realignment and growth throughout the newly formed copper mass. It was coalescence, not adhesion; the several particles had become one unit. From this discovery has been developed the coalescence process, whereby copper of cathode purity is squirted out in semifabricated form from an extrusion press, by the application of about 3000 tons pressure on a 10-in. piston. Stout decided that cathode copper produced by electrolytic deposition was the logical source of the pure copper required for the coalescence process, but as there was no way of reducing the ordinary tough commercial cathodes to the small particles required he was confronted by a serious problem at the outset. Under his direction, this problem was solved by William H. Osborn and Harry Howard Stout, Jr. They developed the process and technique for producing cathode copper in brittle, friable form, and to them must be given the greater part of the credit for this phase of the development program as well as for that relating to the gas cleaning of the cathode particles. Concurrently, the three metallurgists worked on the problem of effecting a complete union of the cathode particles. Adequate washing was all that was needed to free the particles from electrolyte solution and slimes. A press was designed to produce briquettes of the proper porosity and density, and a method was devised of removing surface oxidation from the particles of cathode, by treatment in an atmosphere of deoxidizing gas at an elevated temperature. The temperature was determined at which the surface cleansing and preheating could be done in one operation, and for this a charging and preheating furnace in which there is maintained a positive pressure of deoxidizing gas atmosphere was designed. And finally, a method of feeding the heated briquettes of deoxidized cathode copper into an extrusion press without contamination from the air was evolved. For the final stage, a press was required because not only heat but pressure was required to obtain complete growth of the new crystals between the cathode particles as well as the elimination of all voids or non-coalesced areas. All essential stages of the process are protected by patents or patent applications in the United States and principal foreign countries, and the copper is produced on a commercial scale.

Jan 1, 1941

By Ian W. Wark, Keith L. Sutherland

IT has been shown' that in the absence of heavy metal salts, the nature of the alkali used to promote differential flotation-whether caustic soda, lime or sodium carbonate-is unimportant. The hydroxyl ion is the active depressant upon which differentiation depends, and the pH value determines whether a film of collector forms on any given mineral surface. Even when sodium cyanide is used in conjunction with alkalies, it is the pH value and not the nature of the alkali or its concentration that is important, for upon the pH value alone depends the fraction of the added sodium cyanide that is present as cyanide ion, which is the true depressant. When heavy metal salts become dissolved from partly oxidized ores, or are added intentionally as activators, conditions are different, and metallurgists have found that lime and sodium carbonate do not give identical results, even when added in amounts sufficient to produce the same pH value. In the few instances in which we had attempted to confirm this view, we had failed to do so, for, with nothing to guide us, we had inadvertently chosen solutions that gave contact curves of a "standard type," which correspond to conditions of ready floatability. Now, however, we find that, in the presence of copper salts, and within a restricted range of reagent concentrations, the nature of the anions present is important, and that sodium carbonate does influence the results, apart from the usual effect due to its influence on pH value. With the recognition of this anion effect, and with the control of temperature already introduced,2 it appears that the factors of major importance in connection with this set of reagents have been taken into consideration, and that the results obtained and the conclusions drawn from contact tests will be of more immediate value to the operator. It is not difficult to suggest an explanation for the influence of the carbonate ion in the presence of copper sulphate. The works of Picker-

Jan 1, 1939