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By F. Hass, N. W. Lord

This paper gives the results of experiments conducted in the metallurgical laboratory of the Ohio State University with the objects of determining: First, the calorific powers of a number of coals in general use in Ohio; and second, how far ordinary analytical methods could be used as a basis for computing these values. Incidentally a test of the performance of the calorimeter was involved. The chemical work has been carried on in the department for some time, so that the whole investigation has extended over three or four years. The Mahler calorimeter was purchased from L. Golaz in Paris, and is practically identical with the one described and figured by Mahler in his monograph "Contribution a l'Etude des Combustibles," from the Bulletin de la Société d' Encouragement pour l'lndustrie Nationale. A figure and description is also given by William Kent in The Mineral Industry for 1892, page 97. There were furnished with the instrument three thermometers covering a range of from 7°.5 C. to 31° C. Thermometer "A" (17°.5 to 31°) was carefully calibrated by reading the ends of a thread of mercury in sixteen positions covering the whole scale, supplementing this by readings of longer threads and deducing the errors of scale by the method published by Prof. Holman in the Technology Quarterly of September, 1887. The maximum error of a single reading was found to be 0°.0042 C. Thermometer "B" (13° to 25°) was compared directly with " A " over the scale common to both, and a variation was found of 0°.0145 C. for 7°.5 (or 0°.002 for l°) of the scale. Thermometer " C" ('7°.5 to 19°.0) was omitted from the cornparison, as it was rarely used, being only required where measurements were made at low temperatures. From the

Jan 1, 1898

By J. W. Matousek

Cobalt behavior in pyrometallurgical processes is traditionally analyzed using slaglmatte distribution coefficients as functions of matte grades or matte iron contents. L(s/m) = (%Co) I [%Co] = f [%CuNi] or f [%Fe] The method is inconvenient to use in predictive mass balance work because of the interdependence of assays and mass quantities. An alternative procedure is proposed based on cobaltliron ratios for slags and mattes: Js = (%Co) / (%Fe), Jm = [%Co] / [%Fe], and J, = b + m [Jm]

Jan 1, 1993

By J. E. Lonergan

The in-pit placement of crushing and conveying equipment results in a flattening of the pit slope of a multiple-bench open pit mine. To maximize the economic benefit of lower haulage costs associated with in-pit conveying, one must minimize the cost of the flatter slope which is expressed in terms of additional stripping or tied up ore. A computer program that compares the tonnage aspects of various conveyor layouts will be discussed.

Jan 1, 1985

By D. Hnrrington

THE Dust Symposium was attended by approximately 100 persons, nearly all of whom remained from the starting hour (2 p. m.) until the use of the auditorium was demanded for another meeting at 5:30 p. in. In calling the meeting to order, D. Harrington made a few remarks as to the importance of the health dust problem in mining as well as in general industry and expressed gratification that the subject is now out in the open and that mining people are desirous of ascertaining the facts. B. F. Tillson acted as chairman of the meeting. The milling executive who desires to know what to do to prevent his workers from being unduly exposed to the dust hazard can get much out of the paper by D. E. Cummings, who has been doing much work in mines trying to keep -the dust in the air within safe limits. His recommendations indicate that the main protection of the worker in mines is by adequate ventilation, supplemented by the extensive use of water in drilling, as well as in water curtaining to condition air both in regard to dust content and to temperatures. Blasting was shown to be one of the chief causes for dust and it was recommended that whenever feasible all blasting should be delayed to the end of the working shift.

Jan 1, 1935

By A. Y. Senchenko, S. G. Rybkin, A. V. Aksenov

In gravity processing of gold-containing ore the target produced concentrate which is high gold grade primary concentrate contains a significant amount of cassiterite (SnO2). The plant attempts aiming at concentrate processing using a standard pyrometallurgical technology did not provide any positive results. The present paper is dedicated to development and testing of improved technology for pyrometallurgical processing of rich gold primary concentrate with high cassiterite content. According to material composition study primary concentrates produced at the plant contained 1.5–2.0% Au, 0.1–0.3% Ag, 55–65% SnO2, 20–25% Fe2O3, and 10–15% Fe3O4. The developed technology included alkaline nodulizing of primary concentrate mixed with sodium hydroxide and carbonate at 350–650°? with cake production. In blend nodulizing stannum and iron passed into sodium stannate (Na2SnO3) and sodium ferrate (NaFeO2) respectively. The blend of cake and fluxing agents was smelted at 1,150–1,200°? resulting in dore gold and slag production. Sodium stannate with 360°? melting point is well dissolved in slag system Na2OSiO2- B2O3-CaO. During plant tests 18.5 kg of high gold grade primary concentrate were processed, 1350 g of dore gold and 75 kg of slag were produced. Dore gold contained 16.5% Au, 0.95% Ag, 78.0% Sn, 3.1% Fe. Slag contained 12.4 g/t Au, 6.7 g/t Ag, 9.8% Sn, and 5.6% Fe. Gold recovery to target alloy was equal to 99.58 %, 87.8% of Sn passed into slag. This new technology was accepted for application at the plant.

Jan 1, 2016

By A. B. Kinzel

ALTHOUGH vanadium is an important constituent of almost every brand of high-speed steel manufactured today, little is known as to its role in this series of alloys. The now standard 18 per cent tungsten and 4 per cent chromium constituents were the result of developments and improvements before 1903. In that year, vanadium was added to the steel by Dr. J. A. Mathews. This proved to be an outstanding con-tribution, so that the standard high-speed steel of today is generally referred to as the 18-4-1 type-18 per cent W, 4 per cent Cr and 1 per cent V. Various modifications have been attempted, since then, but with the exception of the introduction of cobalt, no striking improvements have occurred. Vanadium has been increased to 2 per cent in many of the common brands, and at least one brand contains 5 per cent V in addition to the usual tungsten and chromium. The carbon has been gradually increased so that it is not uncommon to find 0.80 per cent in the standard grade, although the carbon in the 5 per cent vanadium steel mentioned above is approximately 1 per cent, and even 1.2 per cent has been made experimentally. The role of vanadium has never been well understood, and although vanadium has been increased to 2.5 and even 5 per cent in certain high speed steels, it has not been generally established that the higher vanadium contents 'result in superior cutting properties. However, in all of the investigations, of which the writers are aware, the interrelation of the vanadium and carbon contents was not considered as a prime factor. The carbon content was held at 1.2 per cent or less-generally at the usual 0.65 to 0.75 per cent.

Jan 1, 1932

In west-central England, 183 m (600 ft) below the Cheshire Plain, lie the evaporate remains of an ancient sea. Stratified into nine beds of salt separated by marl bands, the deposit stretches roughly 21 x 12 km (13 x 7.5 miles) and contains reserves of rock salt estimated as high as 400 Gt (440 billion st). The Winsford Salt Works, one of England's largest underground mines, extracts the Cheshire salt and sells it for use as animal feed, for application in de-icing programs, and as a key ingredient in sodium fertilizer production and a variety of other industrial processes. Because road de-icing is an important market, demand for Winsford's salt varies with the severity of British winters. Design capacity is 2.25 Mt/a (2.5 million/year). However, production peaked at 2 Mt (2.2 million st) in 1987. Over the past five years, annual production has averaged 1.6 Mt (1.7 million st). Run-of-mine salt averages 92.5% NaCl, mixed with about 5% marl. The marl was formed by deposition of a reddish dust blown into the Cheshire basin from deserts in what is now central Europe. Association with the marl adds color to the salt, producing tea-colored crystals that resemble unrefined sugar. Adaptation of coal mining systems Rock salt mining in the Winsford district began at least as early as 1682. Mining at the current site dates from 1844 but was shut down from 1892 to 1928. In the early 1900s, a series of catastrophic collapses and floods plagued and ultimately shut down then-active Cheshire brine-salt producers. Subsequently, Winsford reopened and began adapting mechanized underground coal mining systems to rock salt extraction.

Jan 1, 1992

By Donald P. Tuchman, Robert P. Vinson, Steven J. Page

Researchers at the National Institute for Occupational Safety and Health (NIOSH) have long used stationary tapered element oscillating microbalances (TEOMs®) in laboratory settings. They have served to assess the mass concentration of laboratory-generated particulates in experimental , dust chambers and they provide a reference method for comparison with other particulate- measuring instruments. Current NIOSH research is focused on further adapting TEOM technology as a wearable personal dust monitor (PDM) for coal mining occupations. This investigation's goal is to help identify, quantify, and provide means for resolving certain TEOM- related error. The present research investigated bias caused by thermal effects on filter assemblies. New filters used in the PDM for 8 h tests show an average positive bias of 25.5 pg, while similar tests of equivalent filters used in two 1400A model TEOMs show an average positive bias of 34.3 µg. The derived bias values allow correction of previously collected biased data. Also, pre- heating the filters for 24 h at 46° C shows significant bias reduction, with PDM pre-heated filters subsequently averaging -3.3 µg and 1400A TEOM filters averaging 5.9 µg. On a single-point comparison to gravimetric sampling, a 25.5 µg bias is only significant at low mass loadings. At 2.5 mg, this bias represents a negligible 1 % of the mass measurement. If ordinary linear regression is used, the bias is still insignificant. However, if the more valid weighted linear regression is used, it gives more weight to the smaller dependent variable values, which are more impacted by the bias. Consequently, what is 1% bias on a single high-mass value can translate into a larger bias percentage at high-mass values when performing a weighted regression on data that include a large number of low-mass values.

By M. L. Mellish

In 1989, for the third consecutive year, both the production and consumption of coal in the United States reached record levels. Domestic coal consumption totaled 806.5 Mt (889 million st) in 1989, a 5.4 Mt (6 million st) increase from the year before (Table 1). US coal exporters also had a banner year. For the first time since 1982, exports ex¬ceeded 91 Mt (100 million st). US coal production in 1989 rose by a surprising 27.3 Mt (30 million st), reaching 889 Mt (980 million st) for the year. To a certain extent, the sizable difference between production and demand growth was due to smaller consumer stock withdrawals. Coal consumers withdrew only 12.7 Mt (14 million st) of coal from their stocks in 1989, com¬pared to 22.7 Mt (25 million st) in 1988. Despite strengthening coal consumption, exports, and production, coal prices to all domestic sectors declined slightly during the year. That essentially continued the trend of the past several years. On the labor front, the United Mine Workers of America struck the Pittston Coal Group in 1989. It was a bitter work stoppage that was not settled until February of 1990. This was the first major coal strike since 1985. The strike also precipitated a series of wildcat outages by UMWA members in June and July. At their peak, more than 40,000 UMWA members participated in the wildcat strikes. Except at Pittston, however, most miners were back to work by August. It is estimated that perhaps 4.5 to 9.1 Mt (5 to 10 million st) of coal production was lost due to the Pittston and associated wildcat strikes. However, some of this production may have been recouped by other coal companies not affected by the wildcat strikes, and by the affected companies after the strikes ended.

Jan 1, 1990

By Joseph Mcavoy

El presente artículo señala a la excavadora terrestre más grande del mundo, como el tributo tecnológico de maquinaria moderna, el cual fue. construida por la empresa Bucyrus-Erie para la compañía Ohio Power, que forma parte del American Electric Power System, y costó 24 millones de dólares. La 4250-W, apodada "~ran Muskie", por la mina Muskingum en que trabaja, entró en servicio el año 1969, en las aderas de Cumberland, cerca de Zanesville, Ohio.

Jan 1, 1971

By C. H. Elder, Fred N. Kissell, S. D. Maksimovic

The Bureau of Mines evaluated the hydraulic stimulation of a gob-degasification borehole to determine if this procedure would aid in degasifying a longwall panel. The stimulated borehole did not degasify the longwall panel as expected either before or after mining of the panel started. It reduced the underground methane emission only 11 pct, considerably less than the reductions effected by two unstimulated boreholes in the same panel. Some probable reasons for the reduced methane flow from the hydraulically stimulated borehole are very low reservoir gas pressure, location of hydraulically stimulated zones within the grouted casing, a casing break caused by lateral shifting, possible vertical movement of rock strata during caving, grouting of a 4-inch pipe blocking three stimulated zones, plasticity of shales, and possible hole damage by swelling of water-sensitive minerals. The borehole was also located between two second-mined areas, which may have allowed partial degasification of the overburden and the coalbed in advance of mining. Finally, a period of 3 months is probably not long enough for the borehole to degasify the overburden before mining.

Jan 1, 1977

A simulation model of the load and haul process where the loader cycle has been subdivided into its individual components is used to consider the benefits derived from improved drill and blast methods. Field data from a Hunter Valley mine is u

Jan 1, 1998

By Carl F. Dietz, Dyke V. Keedy

The problem of drying ores is one that most mill-engvineers are sooner or later called upon to meet, and it may be timely to point out some of the difficulties resulting from such operations from purely a metallurgical and economic standpoint. We mean, in the following pages, to discuss primarily the effect of such operations on the ore. itself, without particular reference to the mechanical or cost features of the subject. That an appreciable tonnage of the ore fed to the usual form of drier, either rotary or stack, is lost, is fairly well appreciated, but we believe that only a few mill-managers know just what their drier-losses amouilt to, and we have found that in most cases they console themselves with the fact that the drier-smoke is not very much discolored, and it is generally believed that the draft is not sufficient to carry away more than a negligible percentage of the values.

Jan 1, 1913

By James R. Yancik

Economical and environmental pressures necessitates the recovery of the minus 0.6 mm x 0 coal generated by todays mining machines. This size fraction, dependent on the type of production unit, maintenance of the production unit, and the Hardgrove Grindability Index, can be a substantial portion of the Preparation Plant feed. Historically, dewatering of the collected 0.6 mm x 0 product has often been rather expensive and marginal with respect to surface moisture reduction. The Freeman United Crown II Preparation Plant more than adequately demonstrates that the continuous scroll type centrifuge can dewater 0.6 mm x 0 coal to acceptable surface moisture levels.

Jan 1, 1981

By R. Bruce Graham

"For the purpose of this paper, the Chibougamau area is taken as lying between lats. 49°20’ and 50°10’ N;-longs. 73°45’ and 75°30’ W. It comprises an area of some 4,000 square miles and is centred 300 miles north of Montreal.The first recorded report of metal deposits in Chibougamau was that of James Richardson in 1870-71 (I). In 1903 discoveries of asbestos, copper, iron, and gold made by Peter McKenzie stimulated considerable activity in the district. By 1909 interest had become so marked that considerable pressure was brought to bear on the Quebec Government to build a rail-way into the region. It was finally decided to have an authoritative opinion on its economic possibilities, and the Chibougamau Commission was appointed for this purpose (2). They examined the district in 1910 and concluded that at that time a railway was not warranted. Nevertheless, the district continued to be prospected and from time to· time gold and copper discoveries were made. Thus the Merrill Island deposit was discovered in 1920, the Campbell Chibougamau Cedar Bay deposit in 1922, Obalski in 1928, Opemiska Copper in 1929, and Norbeau in 1930.Difficulty of transportation was a severe detriment to the development of the region. Consequently the Quebec Department of Mines completed a road into Chibougamau in 1949. Further activity was encouraged. The Chibougamau Explorers, Quebec Chibougamau Goldfields, and Kayrand deposits were found in 1950. ** The copper-siderite zones on Copper Cliff ground were found in 1952 and 1953. Opemiska Copper, Campbell Chibougamau Mines, and Chibougamau Explorers were encouraged to prepare for production following favourable exploration results."

Jan 1, 1949

By Wall J, Platell R, Knight P

The Argyle diamond mine is located in the East Kimberley area of Western Australia, some 100 km south of the town of Kununurra. The diamondiferous Argyle olivine-lamproite diatreme was discovered in August 1979. Production began in December 1985 at a planned throughput of 3 Mtpa to recover 25 million carats of diamonds. Presently, Argyle is the world's largest producer of natural diamonds and produced over 40 million carats in 1993 from seven million tonnes of lamproite ore treated. Two phases of evaluation have been carried out. The first in 1981-82 was designed to prove around 60 million tonnes of ore to justify the establishment of a mine with a minimum life of 20 years. The second phase, in 1989-90 involved drilling the orebody to greater depths to provide data on which to plan the final pit design. The final pit contained 120 million tonnes of ore at 5.3 carats per tonne (ct/t) grade. Mining of the planned pit will extend the mine life to around the year 2004, following which it is possible that a further cut-back could be mined. There is also the potential for future underground mining. A substantial amount of geotechnical work has been undertaken due to the fact that the pit will be nearly 500 m deep and is located in a mobile zone causing abnormally high residual stresses to occur in the pit walls. The occurrence of low strength rocks at depth in the pit walls further complicated the situation. Hydrogeological investigations were done to determine pit dewatering requirements and to assess the effect of dewatering on pit wall stability.

Jan 1, 1994

By D. Piper, J. Noble, M. Richardson, J. Wojcik, L. W. Cope

L. Cope Welcome. I am pleased to see so many people interested enough to stay until this last session of the last afternoon of the convention. Of the people on the panel here, everyone has done everything in placer mining, but we have tried to sectionalize it with each one of us taking a particular portion. Any one of us could probably answer any question. I am very pleased to see some well experienced people in the audience. To start the discussion, Joe, can you give us a brief comparison of pit and drill sampling? J. Wojcik They both have their place. If you expect your deposit is something that is going to have widespread fairly uniform grades, then the pits may be the way to go. If you are going to have something you have to trace for several miles down the river valley, you can't do it with pits and do a proper sampling job. C. McLain Joe, the end result of all of this is to go ahead and mine it if it is a deposit. Do you try to relate your drilling results to something in mining itself against the recovery, because the drilling in different types of ground will give you different kinds of recoveries, and mining in the end is what you are going to do? Do you try to relate this to your mining method?

Jan 1, 1992

By Karin Soldenhoff

Recovery of uranium in conventional ion exchange circuits, using strong base or weak base resins containing quaternary or tertiary amine functional groups, is based on an anion exchange mechanism. In saline liquors, chloride competes with uranium for the active resin sites and reduces resin loading. This results in significant negative impacts on the IX process in terms of plant throughput and operating costs. In this paper we present the results of a study on the recovery of uranium from an acidic in situ leach liquor containing approximately 9 g/L chloride. A range of resins was tested, and a high capacity strong base resin was identified with significantly enhanced loading compared to conventional resins. Other process-relevant criteria such as loading kinetics and elution efficacy are also discussed. The process developed was successfully tested during a recent ion exchange pilot plant operated in conjunction with a field-leach trial.

By E. Asselin, G. K. Yue

"In the present study, electrochemical measurements, especially cathodic potentiodynamic polarization (CPC) and cyclic potentiodynamic polarization (CPP), have been employed to study the kinetics of ferric reduction on chalcopyrite in aqueous Fe(II)Fe( III)H2SO4 solution at 90 °C and 110 °C. Results from the calculation of speciation in the above solution reveal that, at both 90 °C and 110 °C, the real Fe3+/Fe2+ ratio is much lower than the nominal ratio, which leads to smaller VALUES (for the reversible potential of the ferric/ferrous couple. Exchange current densities of the ferric/ferrous couple were obtained and are on the order of 106 10 5 A/cm2 a t 9 0 °C and 1 07 10 6 A/cm2 at 110 °C, respectively. The Tafel slopes from the CPP method are higher than those from the CPC method at both 90 °C and 110 °C. The transfer coefficients were about 0.360.59 at 90 °C and 0.370.65 at 110 °C. Based on the exchange current densities obtained, the rate constant kc was calculated. It was found that kc at 90 °C was slightly larger than at 110 °C, especially for the results obtained by the CPC method. The average rate constants are 862.17 cm/s for CPC and 65.30 cm/s for CPP at 90 °C, and 477.69 cm/s for CPC and 63.53 cm/s for CPP at 110 °C for nominal Fe3+/Fe2+ ratios ranging from 1:1 to 1000:1. INTRODUCTIONChalcopyrite is one of the most abundant copper-bearing minerals, accounting for approximately 70% of the world’s known copper reserves (Wang, 2005). Most of the currently proposed leaching processes are based on sulfuric acid with ferric ions and/or dissolved oxygen as oxidants. Some of these processes have been piloted in attempts to implement these technologies on an industrial scale, such as Activox (U.S. 5,232,491, 1993), Nenatech (U.S. 5,993,635, 1999), Dynatec (U.S. 5,730,776, 1998), AAC/UBC (U.S. 6,503,293, 2003), Total Pressure Oxidation (U.S. 5,698,170, 1997), CESL (U.S. 5,874,055, 1999) and Galvanox (U.S. 7,846,233, 2010).To date, much effort has been devoted to understanding the kinetics and mechanisms of the oxidative processes (Biegler & Horne, 1985; Córdoba, Muñoz, Blázquez, González, & Ballester, 2008, 2009; Ghahremaninezhad, Asselin, & Dixon, 2010; Hirato, Majima, & Awakura, 1987; Hiroyoshi, Kuroiwa, Miki, Tsunekawa, & Hirajima, 2004; Warren, Wadsworth, & El-raghy, 1982) or the cathode reduction reactions of chalcopyrite itself (Arce & González, 2002; Biegler & Swift, 1976; Biegler, 1977; Nava, González, Leinen, & Ramos-Barrado, 2008), by chemical or electrochemical methods. However, as the most important cathodic processes in chalcopyrite leaching systems, the cathodic reductions of ferric ions or dissolved oxygen are worth measuring due to their effects on the overall leaching rate, as predicted by mixed potential theory."

Jan 1, 2012

By Thomas Shumka

El presente texto señala la norma internacional ASIM E2905 aplicada a la limpieza e inspección de corona en los molinos SAG, bolas, engranajes de molinos y hornos donde se busca detectar electrónicamente los defectos de ruptura en la superficie y de dimensionar con precisión, el tamaño de cualquier grieta encontrada en la superficie del material fundido y forjado de la corona y dientes del piñón en molinos SAG y de bolas.

Sep 1, 2014