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By D. L. Cramp

The Direct-Connected Electric Hoist In the latter part of 1931 a new electric hoist was installed at the Lake Shore mine. A number of interesting features were incorporated in this machine, which is the first direct-connected, direct-current-operated hoist to be placed in service in the Kirkland Lake district. Many factors requiring care and study enter into the selection of hoisting-engine drive. In this instance, conditions warranted the selection of the generator voltage control system (Ward-Leonard), utilizing a direct-current driving motor and a separate motor generator unit. It was further deter-mined that additional economies could be secured by the selection of a motor generator unit with a flywheel, and consequently the Ilgner-Ward Leonard flywheel set was adopted. Some of the advantages of this system are: 1) Greater accuracy of control, resulting in increased flexibility in hand-ling with increased precision and safety; especially desirable in large, important installations where the rope-speed is high or where frequent shifting is necessary. (2) Reduced energy consumption. (3) Feasibility of load equalization, without which, in some cases, the power system would be unable to carry the fluctuating load imposed on it by the operation of the hoist. ( 4) Elimination of reduction gearing in the construction of the hoist. The hoist is located on the surface and is used for skip operation at the main shaft, hoisting ore from a present maximum depth of 4,200 feet.

Jan 1, 1933

By P. B. Altringer

Searles Lake, CA, contains the largest known domestic tungsten resource. The brines in this near-dry lake bed, located in the Mojave Desert 130 mi northeast of Los Angeles, contain an estimated 135 MMlb of tungsten. However, the low concentration of tungsten (0.080 g/L W03) in the supersaturated brine makes selective tungsten extraction very difficult. Although many brine chemicals are extracted from the lake, tungsten recovery was an elusive goal prior to Bureau of Mines development of an ion-exchange technique. This publication describes the Bureau process. The following subjects are ad-dressed: ? The Searles Lake tungsten resource. ? General problems associated with tungsten recovery. ? Bureau research objectives. ? Synthesis of ion-exchange resins. ? Development of a two-stage ion-exchange tungsten extraction and concentration procedure. ? Product recovery and purification. ? Economic evaluation. Flowsheets for recovering three products (calcium tungstate, sodium tungstate, and tungstic oxide) are presented and evaluated, as well as a summary of the research effort and potential applications.

Jan 1, 1985

By E Wargem

The conventional method of underground mining involves a process known as drilling and blasting to excavate valuable ore bearing minerals through the load and haul process. Drilling and blasting processes are completed for both the development of the tunnels used to access the orebody and for mining out the orebody itself. The former being called development drilling and blasting, and the latter production drilling and blasting. In either development or production drilling and blasting, the effective recovery will be largely reliant on the extraction of the initial void created in each blast cycle. In development tunnelling, the initial void is commonly known as a Burn Cut. The initial void created in production drilling and blasting has many names such as a Slot, Winze, Rise or Raise. There are drill patterns and charge designs used to create these initial voids in the production cycle. This paper discusses common slot or winze designs for production drilling and blasting; and various critical design criteria for these. The document also aims to provide a process for completing desktop reviews of current patterns employed on-sites and provides an initial guide in a process for improving the recovery of a slot, winze, rise or raise in underground mines.

Mar 29, 2023

By A. Deneys

An experimental technique has been devised for studying the reaction kinetics of the reaction: ZnO + 2Fe0 = Zn(g) + Fe203, by which zinc evolution can occur from a molten slag bath in the absence of a carbonaceous reductant. When studying the reaction kinetics by heating from room temperature, solid state reactions and poorly defined melting points may obscure the starting point of the reaction. In order to eliminate this problem, an experimental technique was devised whereby two molten slags could be melted independently, and then mixed, to provide a well defined starting point from which to begin taking measurements. Two crucibles were mounted in a vertical reaction cylinder. The upper, mild steel, crucible had a bottom pouring plug which was sealed by a steel stopper rod. A fayalite mixture was contained in this crucible. The lower crucible contained an A12O3-CaO-SiO2-ZnO solution. Once the two slags had been melted, the steel stopper rod was removed from the upper crucible. The upper, reduced slag (fayalite in equilibrium with the upper steel crucible), drained into the lower crucible where the two slags were mixed by submerged argon stirring. Dip samples were taken to measure the zinc content of the final mixture as a function of time. The experimental procedure will be described, as well as results of experiments which were conducted on synthetic slags. The system investigated was similar to the lead blast furnace slag which is produced by the primary lead smelters of Missouri, USA.

Jan 1, 1997

By C. Juers

Laboratory chromate coating was applied on AM50 and AZ91D alloys and compared to phosphate-permanganate one. Evaluation of the corrosion behaviour was performed using electrochemical DC and AC methods in ASTM water D1384-87. This electrolyte was chosen because it was moderately aggressive and found to give generally good correlations with long period exposure results in atmospheric conditions. Electrochemical impedance spectroscopy was used to determine more accurately the corrosion rates from the analyses of the high frequency capacitive loop. Anodic polarization was applied in order to evaluate the pitting resistance of the different coatings. These results were correlated with weight loss data evaluated from experiments performed by immersion tests in ASTM water D1384-87 and climatic cyclic chamber tests.

Jan 1, 2006

By Eric P. Bregman, Justin Zarrella, Ryan M. Coulson, Connor Johnson

Within a prominent Boston, MA suburb are a pair of historic stone and mortar structures built in the late 1800’s. Both the Main House and Carriage House have been maintained and occupied throughout their rich history, but renovations were needed to address structural concerns and building code compliance issues. The focus of this paper is to detail an innovative underpinning design that addressed ongoing settlement issues at the Carriage House. GZA GeoEnvironmental, Inc. (GZA) partnered with Phoenix Foundation Company, Inc. (Phoenix) to design foundation wall support systems for Consigli Construction Company (Consigli) that would effectively underpin the historic field stone foundations. Originally, conventional concrete pit underpinning and jet grouting were contemplated as methods for support. However, due to the age and condition of the existing foundations, worker safety concerns, and the risk of additional structural damage, GZA/Phoenix elected to pursue a less traditional approach. The alternate design consisted of micropiles used for combined temporary and permanent foundation support. The load transfer system included steel headers with needle beams cored through the field stone foundation walls and reinforced concrete grade beams/pile caps. To validate design performance and ensure worker safety, GZA installed a fully automated survey monitoring system to record lateral and vertical building deformations. This included real-time email alert notifications to inform the Project Team of any threshold exceedances. From the start of micropile drilling through full load transfer to the temporary support and permanent foundation systems, the underpinning performed as designed resulting in less than 3/16-inch of vertical and horizontal deformations observed. This magnitude of movement is less than half of what is generally accepted deformation tolerances for masonry structures and considered exceptional performance. The focus of this paper is to provide additional details on the unique design, sequencing, installation, and performance of the underpinning systems.

Sep 8, 2021

Falls of ground (FOG) still account for around 35 per cent of all fatalities in underground southern African mines. To help reduce FOG, appropriate support technology needs to be implemented, and a mindset and cultural change are required to make a significant change. Discipline plays a major role and needs attention, whilst appropriate first-world support technology is available to assist in the reduction of injuries and the improvement of productivity. Section 54 mine stoppages due to FOG are costly. Appropriate support should reduce these unnecessary stoppages. Coal mining has adapted leading support technologies over the years. Platinum and chrome mines have shown significant moves towards adopting leading and appropriate support technologies, seeing noteworthy changes in injury statistics for the better. In typical conventional underground mining operations, the use of temporary support components offers some protection to workers while the permanent and generally more elaborate support systems are installed or constructed. Nonetheless, FOG continue and still result in injuries to people. The author wishes to explore the options and practicality of making use of new-generation support products and technology to provide not only safer support operations, but to also enhance productivity at the working faces. This paper will examine the use of appropriate support technology as a means to making the working stope or development end a safer and more productive workplace.

Jan 1, 2012

By F. P. Haver

Because of increasing concern with air pollution, the Bureau of Mines has developed a hydrometallurgical procedure to obtain elemental sulfur from chalcopyrite (CuFeS2) by means of the following reaction: CuFeS2 + 3FeCl3 ? CuCl + 4FeCl2 + 2S. Under optimum conditions, 99.6 percent of the copper and 71.6 percent of the iron were extracted in 2 hours from a typical mill concentrate containing 75 percent CuFeS2 and 15 percent pyrite (FeS2). At the same time, 68.2 percent of the sulfur in the concentrate was oxidized to the elemental form. The leach was highly selective; gangue minerals, including pyrite, were not attacked, thus accounting for the relatively low recoveries of iron and sulfur. To obtain these results, the concentrate was ground to 98.9 percent minus 325 mesh, and 300 grams was treated with 1 liter of boiling ferric chloride (FeCl3) solution. The FeC13/CuFeS2 ratio was 2.7. Cement copper containing 99.9 percent copper was recovered from the leach solution by cementation on sponge iron. Excess iron, dissolved during leaching and cementation, was removed from the stripped solution by crystallization. Ferrous chloride (FeCl2?4H2O), obtained as a byproduct, was converted to high-purity ferric oxide (Fe203) and HCI by an oxidizing roast. The hydrochloric acid was then electrolyzed to obtain chlorine, which was used to regenerate the leach solution. The leach residue, which contained ~50 weight-percent elemental sulfur, was treated with an organic solvent to recover the sulfur and cyanided to reclaim gold and silver.

Jan 1, 1971

By Randall M. Wheeler

Pre-blast surveys are used to document the condition of a structure prior to the commencement of blasting activities. They provide the best means of determining the pre-existence of visible cracking and deterioration in a structure before blasting is initiated. A small investment in a good pre-blast survey, coupled with good public relations, could mean a large cost savings for a company in the future.

Jan 1, 1992

By Luis Hoyos-Reyes, Jesus Gonzalez-Trejo, Adriana Cervantes-Clemente, Marco Gutierrez-Villegas, Manuel Palomar-Pardave, Cesar Real-Ramirez

Recent improvements on Peirce-Smith Converters (PSC) have combined numerical and physical simulations. However, most of the physical simulations have been carried out in cold-water models with only one tuyere. Several authors have proved that 2D numerical simulations do not reproduce the hydrodynamic behavior observed in physical simulations. On the other hand, most of the 3D numerical models have used a PSC thin slice with only one tuyere, but symmetric boundary conditions on the virtual walls have been imposed in order to reproduce the behavior of the whole converter. Therefore, there is no coincidence with the rigid walls of the physical models. The aim of this work is to quantify the effect of the virtual walls boundary conditions of a PSC thin slice and to determine the minimum number of tuyeres to characterize the fluid flow behavior of the entire converter.

Jan 1, 2009

By J. E. Shearer

The state of Arizona produced 253,276.2 metric tons (279,000 st) of crude gypsum during 1986. This represents approximately 1.7% of the total U.S. production. Four mines provided Arizona's gypsum, three of these mines are in the San Pedro Valley. The uses of the gypsum produced in Arizona are for manufacture of cement, agricultural uses, and wallboard manufacture. The gypsum is used as a retarder in the manufacture of Portland cement. The use in agriculture is treatment of alkali soils to allow water penetration. Other deposits in the San Pedro Valley have production potential, providing there is sufficient market for additional wallboard. The author completed an evaluation of one property earlier this year.

Jan 1, 1988

By D. P. Goodall

THE development of the oil and gas resources of Western Canada is not, as many suppose, a new industry. For more than half a century an intensive search for these valuable products of nature extended from near the western edge of the Precambrian Shield to the Pacific coast, and from the International boundary north to the Arctic circle. The chief centre of interest to the oil prospector has, however, always lain in that region lying east of the first range of the Rocky mountains and confined largely within the boundaries of the Province of Alberta. Although there are many geological problems yet to be worked out in this Province, the oil-well driller has helped in no small way in determining sub-surface conditions and distribution of rock formations. The chief geological conditions as they affect the drilling of wells are generally re-cognized, and for this purpose the Province may be partitioned into three main divisions: first, the disturbed belt or foothills region extending along the eastern edge of the Rocky mountains; second, the Alberta geosyncline, a broad trough lying to the east of and parallel to the disturbed belt and from which the formations rise eastward to level off and form, third, the plains ? ? region, extending over the rest of the Province. A relatively small area in the northeastern part of the Province is underlain with Precambrian rocks and may be excluded from this classification, as it is of no interest to the oil prospector. The foothills, or disturbed belt, has presented the greatest difficulty to drilling, though it has produced the largest yield of oil within the Province to date. Within the foothills belt, the rock formations consist largely of hard shales interspersed with layers of abrasive sandstones, which wear out drilling bits at an alarming rate. The strata, in most flaces, dip steeply and are cut by numerous thrust faults. Although some o the faults have great displacements, their presence may easily be overlooked or the extent of the fault misinterpreted. Many a prospect well has been located on what appeared to be a suitable geological structure only to find, after drilling through a normal succession of rocks for 4,000 feet or more, that the drill had passed through a major fault plane to a younger formation than that in which drilling was started. Only a repetition of formations already drilled could be expected at greater depths.

Jan 1, 1942

By K. S. McClure, S. B. Chitralekha, R. B. Gopaluni, Chmylek. T., S. L. Shah

This is a tutorial paper on industrially successful approaches for nonlinear process monitoring. Online process monitoring is essential to continuous operation of plants at high efficiency. The problem of process monitoring is one where an impending but undesirable process condition has to be identified and used to alert process engineers. This problem has traditionally been solved by practicing engineers using statistical techniques that detect a change in the underlying correlation structure of process variables. For linear correlation structures, algorithms such as principal components analysis, independent component analysis and partial least squares have been proposed. However, many industrial processes exhibit nonlinear correlation structure between the process variables. In this tutorial we describe various nonlinear monitoring techniques and elaborate on two recently developed methods, namely Support Vector Machine and Local Linear Embedding. The two methods are evaluated by application on two experimental case studies, a SAG mill and a polymer extruder.

Jan 1, 2014

By J Coombes

There is much debate on the Competent Persons qualifying criteria for providing classified resource estimates for public reporting. The CRIRSCO family of codes including the JORC Code, SAMREC, PERC and NI43–101, require Competent Persons to have at least five years ‘relevant’ experience and belong to professional bodies. These codes provide further, albeit vague, guidelines to assist individuals with their personal decision to take on the mantle of ‘Competent Person’. There is a risk, however, that these qualifying criteria can result in ill-prepared Competent Persons whose actions and recommendations would undermine their own reputation and that of the organisations that employ them; the integrity of the profession and the principles based reporting systems; as well as the stability of stock markets that rely on these estimates.Based on doctoral study on the development of professional competency within the mining industry, this paper presents an evaluation of the suitability of the current qualifying criteria as well as an analysis of alternative criteria. The research included a series of interviews with industry experts and an online survey of resource geologists. Semi-structured interviews provided a sense of experts’ expectations and interpretation of the current definition of ‘Competent Person’. The comprehensive online survey included a self-assessment, a series of situational questions designed to evaluate responses to typical resource geology scenarios and a set of open-ended questions relating to factors contributing to competency development.A summary of the data analysis, interpretation of the results and a set of recommendations for revised ‘Competent Person’ qualifying criteria are presented. This paper contributes to the conversation on what makes a person suitably qualified to report classified resource estimates and provides practical targets that can be used in the professional development of resource geologists.CITATION:Coombes, J, 2014. Redefining Competent Persons criteria for resource geologists, in Proceedings Ninth International Mining Geology Conference 2014 , pp 153–158 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

By A. C. Bigley

Before going into this subject, I would very much like to call to your attention some past history on patents granted by the United States Patent office that involved ammonium nitrate. On February 10, 1885, a patent was granted to Russell S. Penniman. "The object of this invention is to securely protect ammonium nitrate against deliquescent and to maintain it in substantially its normal disintegrated condition, so as to afford a new product suited for ready use in the manufacture of high explosives, and enable the immediate admixture thereof with any of the well known explosive substances or compounds heretofore employed. 'Whether the nitrate be thus granulated or otherwise reduced to a granular or divided condition, it is then thoroughly mixed with about twelve (12) percent of the fluid or semi-fluid carbonaceous protecting medium, which forms a readily applied jacket or coating for the granules of nitrate and also serves as a carbonaceous element in such explosive compounds as it may thereafter be employed in."

Jan 1, 1958

By A. C. Burr

This progress report summarizes operating results on the gasification of natural lignite from five runs of the commercial-scale pilot plant at Grand Forks, N. Dak., during December 1946 and calendar years 1947 and 1948, when the plant was operated for 1,996 hours. A primary specific objective was to determine the useful life of the outer metal retort tube as it is influenced by corrosion and other factors under a wide range of operating conditions. A secondary objective was to develop the technology of production when generating gas with a high ratio of hydrogen to carbon monoxide in both the divided and continuous annuli. Results are summarized in graphs and tables that show hydrogen-carbon monoxide ratios from 2.3 to 9 at gas-production rates from 6 to 12 M cu. St. per hr. (thousand cubic feet per hour), or approximately 30 to 60 cu, ft. per square foot of heated retort-wall surface per hour. Gas yields ranged from 36 to 54 M cu, ft. per ton of natural lignite or 63 to 91 M cu. ft. per ton of moisture- and ash-free lignite, Production rates were higher in the divided annulus, but these represent the maximum capacity of the plant only at high hydrogen-carbon monoxide ratios when the high concentration of excess steam and low furnace temperature limit gas production. A ratio of 6 was the highest produced in the continuous annulus because mechanical Limitations of the auxiliary equipment prevented attainment of higher ratios. The lignite feed rate appears to influence the production rate more than any other individual operating variable. Heat balances for four periods when gas with ratios from 2.3 to 7.2 was produced with a lignite feed rate of about 325 pounds per hour suggest that the potential heat in the gas is lower at higher ratios, owing to increased heat losses in the excess steam. Heat-transfer rates through the retort tube in these four tests were between 1,500 and 3,000 B.t.u. per square foot per hour. These rates are much lower than those reached in many earlier tests, but no attempt was made to achieve maximum rates of heat transfer.

Jan 1, 1953

By Nikhil Dhawan, Sanjay Pindar

The limited lifespan, ever-growing demand, and significant lithium and cobalt content are responsible factors for the immediate recycling of discarded lithium-ion batteries. Discharged batteries were dismantled, segregated, and size reduced to recover cathode and anode material. The active cathode material comprises LiCoO2 (65.8%) and LiMn2O4 (34.2%) and was carbothermal reduced with recovered graphite in a microwave furnace followed by water leaching and magnetic separation. A Box Behnken statistical experimental design is used for process optimization. Both cobalt and manganese report to magnetic fraction, whereas graphite and lithium carbonate report to non-magnetic fraction and solution crystal respectively. The final product comprises of cobalt 73% and manganese of 14% with process yield of 36% and saturation magnetization of 76 emu/g.

May 16, 2020

By William Daly

THE application of artificial refrigeration, or air-conditioning, to the ventilation of deep, hot mines has long been a subject of interest to the operators of such properties. Artificial cooling of the air in theaters, railroad coaches, and various industrial plants has been so successful within recent years that it has been natural to assume that the adoption of similar methods would afford a ready solution to the difficult problem of mine ventilation. It should however, be obvious, that although there may be nothing technically impossible in the application of this process to mine ventilation there is a great economic distinction between the maintenance of a high standard of comfort for a large number of people in a relatively small space such as a theater and the cooling of air to improve working conditions for miners scattered throughout the extensive workings of a large mine. Many methods of air-conditioning have been considered by the management of the Anaconda Copper Mining Co. at various times, but until the development of the process to be described in this paper none has seemed economically practicable.

Jan 1, 1934

By James J. Dowd

The investigation to evaluate the reserves of coking coal is being made by the Bureau of Mines in three parts: (1) To estimate known measured and indicated recoverable reserves of all coking coal; (2) to upgrade marginal coals through effective preparation; and (3) to study the carbonizing properties of coal and coal blends not now widely used for metallurgical coke making. This is the tenth of a series of reports giving the results of studies by counties under part (1) of the investigation. (See appendix.) This report covers Knott County, Ky., which comprises parts of the Salyersville, Prestonsburg, Troublesome, Hindman, Pikeville, Cornettsville, and Whitesburg quadrangles. (See fig. 1.) A base map for each bed in each quadrangle was made to the scale 1 inch equals 1,200 feet. Maps of mines, locations of drill holes, bed and total coal thicknesses, and the outcrop of the bed were plotted on the maps. With all available data plotted, isopach lines were drawn to limit areas of known unmined reserves in beds up to 14 inches thick, 14 to 28 inches thick, 28 to 42 inches thick, and over 42 inches thick. These areas of coal reserves also were divided into "measured" and "indicated" categories. All areas in each thickness range and category, mined-out areas, areas excluded from the estimate but which may contain reserves based only on geologic inference, and areas outside of the outcrop were measured by planimeter on the base maps. Estimates of total reserves 14 inches and more thick and maps for individual beds were prepared from these data.

Jan 1, 1952

By I. C. Edwards, W. J. Dengler, D. F. Lillie

By 1959 the milling plant of Bicr oft Uranium Mines Ltd. had been in operation for two years. During this time many changes, both physical and chemical, had been made in an effort to improve plant efficiency. Investigational work had been facilitated greatly by the fact that the Bicroft mill consists to a large extent of two parallel circuits; however, the number of interrelated variables is great, and it is difficult to measure, let alone control, all of them simultaneously. The Bicroft mill staff feels that much progress has been made, but there is much yet to be done and many questions still to be answered. The Addendum outlines noteworthy developments to the fall of 1960. The property of Bicroft Uranium Mines is located in Cardiff Township in the Highlands of Haliburton County, southeastern Ontario, some 140 miles northeast of the city of Toronto and 70 miles due north of Lake Ontario. The ore-bearing dikes are classed as mafic. nonsegregated, syenite and granite pegmatite. Microline, microline perthite, quartz, sodic plagioclase, and pyroxene or biotite are the main constituents with uraninite, uranothorite, fluorite, titanite, cryolite, calcite, pyrite, pyrrhotite, and molybdenite as important accessories. The mill was designed for a nominal capacity of 1000 tpd with all sections except grinding, leaching, and clarification capable of 1500 tons. Feed was entered to one mill unit on Oct. 29, 1956, and to the second parallel line on November 13. The total feed rate was at 1000 tpd on November 15. The first precipitate was produced on November 12. By August 1958 the average entry had been increased to 1410 tpd without the installation of any additional major equipment. The overall recovery has been greater than 93.0 pct on several months, but a recovery of 92.5 pct appears to be most economical. Operating time for eight months in 1958 has been 97.1 pct although all grinding mills were relined during this period. A generalized flowsheet is shown in Fig. 1 and the plant layout in Fig. 2. The ore is crushed in three stages to -3/8-in. and then is wet ground in two stages to approximately 50 pct -200 mesh. The pulp is thickened to 68 pct solids and agitated by air for 33 hr at a pH of 1.9 as maintained by the addition of 93 pct sulfuric acid. Separation of the uranium-bearing pregnant liquor from the residue is accomplished by five-stage countercurrent decantation (CCD), following which the liquor is clarified for removal of fine slimes. The clarified liquor is treated in the ion exchange equipment in which both the upgrading of the solution and the removal of some impurities is accomplished. Addition of magnesia to the high-grade eluate from ion exchange precipitates the uranium as sodium-magnesium diuranate. After being filtered and dried this precipitate is packaged in drums for shipment. The residue from CCD and the barren solution from ion exchange are neutralized by the addition of lime and pumped to the tailings disposal area. CRUSHING All crushing is done on surface in a conventional three-stage unit consisting of a 36x48-in. Birdsboro Buchanan jaw crusher, a 4 1/4-ft standard Symons cone crusher, and a 4-ft Symons shorthead. Both the two primary and the two secondary screens are 4X 8-ft Symons rod decks. Moisture content of the feed ranges from 3 to 4 pct. In the initial installation the primary screens were equipped with 5/16-in. rods at 1 1/4-in. spacing and the secondaries with 5/16-in. rods at 1/2-in. spacing. With this arrangement the average crushing rate was better than 180 tph. To reduce the work required of the grinding circuit as tonnage was increased, several changes were made in the crushing circuit. By September 1957, both secondary screens had been reduced to 3/8-in. spacing, and rod size had been reduced to 3/16-in. to provide increased open area. Average throughput was reduced to 160 tph. Later, 50 pct of one primary screen was decreased to 1/2-in. opening to better balance the load between the two cone crushers. In May 1958,

Jan 1, 1961