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By T. A. O’Hara

SINCE May 1948, when tungsten carbide bits were introduced at the Flin Flon mine, they have been popular with the miners because of their fast drilling speed and low gage loss. The high cost of commercial carbide bits and tipped drill steel, however, prevented their use except for the hardest rock. In an effort to extend the use of tungsten carbide on a basis economically competitive with detachable steel bits, experimental work was begun in 1950 to test the feasibility of making tungsten carbide tipped drill steel in the mine drill steel shop. This work showed that tipped drill steel could be made locally at less than half the cost of the commercial product. The performance of the local tipped drill steel was comparable to that obtained with commercial carbide bits and tipped drill steel and the cost per foot drilled was much lower. Local tipped drill steel was adopted for all mine drilling in November 1951. Since then drilling costs per foot have been sharply reduced and footage drilled per manshift has increased markedly. Experience at Flin Flon has shown that production of satisfactory carbide tipped drill steel is not difficult and that highly skilled labor and costly equipment are not required. As long as wise selection of brazing materials is made and certain simple precautions are rigidly maintained, there is no reason why small mines with relatively unskilled labor cannot produce a satisfactory product. The following description outlines the technique used at Flin Flon for making carbide tipped drill steel and discusses characteristics of the brazing process that make special precautions necessary. Drill steel is forged to four-wing shape in a conventional steel sharpening forge. Standard steel dies are modified to minimize forging cracks around the central waterhole and to forge a blunt bithead on the steel. The steel is preheated to 1500°F and held at this temperature for at least 2 min. When the temperature has equalized throughout the steel section, the drill steel is transferred to the forging furnace and heated rapidly with a reducing flame up to 2000°F. This two-stage method of heating minimizes the grain growth and decarburization of the steel while ensuring that the steel temperature does not vary greatly throughout the forging zone. After forging the steel is allowed to cool in air to about 1600°F before being annealed in a bath of vermiculite. Despite the high hardenability of the 3 pct Ni-Cr-Mo drill steel used, this simple treatment anneals the drill steel sufficiently for milling. The forged and annealed drill steel is slotted on a plain horizontal milling machine that is equipped with a quick opening chuck and a slot depth stop. The full depth of the slot is milled in a single pass of the 3-in. milling cutter which is fed at 33/4 in. per min across the crown of each bit wing. The slots are cut to a width of 0.342 to 0.344 in. Maintenance of this slot width is necessary to ensure that the optimum brazing clearance of 0.002 in. will result after assembling of shims and carbide in the slot. Prior to March 1953, when the milling machine was installed, drill steel was slotted on a small manually fed ¾ hp milling attachment mounted on the bed of a lathe. Over 16,000 drill steels were slotted on this unit, and in view of its small size and low cost it gave excellent service. Brazing of Tipped Steel Drill steel that has been milled and cleaned in carbon tetrachloride is mounted in a rotating cradle holding six drill steels, the length of which may be from 2 to 12 ft. The slots in the drill steel, the shims, and the tungsten carbide inserts are thoroughly fluxed with a fluoride flux and assembled as shown in Fig. 1. Fig. 2 shows the brazing equipment in use. As the ring burner is lowered over the bithead a spring valve opens the gas lines, and the gas mixture, preset to give a slightly reducing flame, is fed to the ring burner where it is lit from a pilot flame. The ring burner heats the drill steel over a zone about 1 to 2 in. below the bithead, which becomes heated by conduction through the steel. By this means the bithead is heated rapidly and evenly, and contamination of the brazing joint with soot from the flame is avoided. The bithead is heated to the melting temperature of the brazing alloy within 1 min. This rapid heating minimizes the disadvantage of a non-eutectic brazing alloy. The brazing alloy, a nickel-bearing quaternary alloy, is placed at the bottom of the slot below the carbide insert, as shown in Fig. 1. As the brazing alloy melts it is drawn by displacement by the carbide and by capillary action into all parts of the joint to displace liquid flux from metal surfaces. As soon as the brazing alloy melts, each insert in turn is wiped by being moved back and forth along the slot. This action assists wetting of the carbide by the brazing alloy and assists in displacing molten flux from the joint. After continuous heating for about 75 sec, when the bithead has reached a temperature of about 1500°F, the ring burner is raised and the gas supply is shut off automatically by the spring valve. As soon as heating is stopped a hand press is placed on the bithead and the inserts are squeezed down firmly. This action minimizes the clearance between the bottom of the insert and the slot. Correctly brazed steel should maintain a clearance at the bottom of the slot of 0.001 to 0.002 in. After six steels have been brazed they are removed from the cradle and allowed to cool in air. As soon as each drill steel is cool it is dressed on a grinding wheel to remove excess flux and braze and is ground to the gage appropriate to the length of the drill steel.

Jan 1, 1955

By Karl J. C. Harries

"7.1. INTRODUCTIONThis chapter addresses the joint venture and examines a number of general concepts of the joint venture relationship. Chapter 8 looks at some specific concepts related to the provisions of a joint venture agreement, which are equally applicable to a shareholders’ agreement or other contractual relationship. The term “joint venture” is used because the author is a common law solicitor (a matter of convenience for which apologies are offered to civil law readers), but it should be noted that, in many aspects, the legal structure of an entity is irrelevant. It is the relationship of the parties, be they venturers or shareholders, that is important, and the areas of concern are surprisingly similar (except for terminology).The details of the joint venture relationship and agreement could fill several books. Accordingly, this chapter and the next focus on selected topics that appear to have caused problems or confusion or to be the basis for disputes. And remember that, although the following discussion is in the context of a joint venture, many of the comments apply, either directly or indirectly, to virtually any common undertaking and the agreement defining it.For interest’s sake, this chapter starts with a summary of some of the earlier types of agreements and arrangements that have evolved into the joint venture agreement of today in Canada, the philosophy of their format, and the reasons for their demise.7.2. EVOLUTION OF THE CANADIAN JOINT VENTUREThe Early DaysMuch of the opening up of Canada’s north is due to exploration for and development of mines. The early discoverers were relatively free to roam and explore at will and, if they were fortunate enough to find something, they dealt with the appropriate government to acquire exploitation rights. The early interests lay in gold or silver. Workings were small, usually worked by the owner alone or with only a couple of helpers to extract the native gold or silver. If the owner wanted to move on, he could sell his property outright and use his stake to look elsewhere.As logging and mining camps were opened, the development of better roads and railways made access easier. Entrepreneurs saw the opportunity to make money by assembling large parcels of land and working them as a single operation. In order to do this, some owners who were not ready to sell out had to be convinced to do so through more attractive offers"

Jan 1, 2003

By C Beal

Proper understanding and management of production drilling in sublevel cave (SLC) mines is essential for meeting production targets and for avoiding inefficiencies that could lead to costly bottlenecks. In many SLC mine management models drilling is considered a derivative activity of firing and bogging. Its scheduling is considered to be implied in a firing schedule and given an arbitrary target completion, ‘several weeks’ or ‘a month’, ahead of production firing. Like development, drilling activities are most efficient when conducted separately from other production activities. Unlike development, however, it is not apparent during planning and feasibility how far in advance of firing drilling activities ought to be conducted. This paper presents a method of defining a minimum threshold of drilled stocks above which drilling, firing and bogging operations occur most efficiently. As an example, the drilling activities, procedures, and maintenance at Glencore Xstrata’s Ernest Henry Mine are presented, and the larger scale consequences of improvements to these areas are presented to illustrate their importance in the operation of an SLC mine.CITATION:Beal, C, 2014. Understanding the role of production drilling and drilled stocks in efficient operation of a sublevel cave mine, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 235–242 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By Raíza da Silva Juvenal, Suelen Silva Rocha, Artur Ângelo Alcântara de Assis, Jorge Luiz Valença Mariz

Um estudo de viabilidade econômica deve considerar parâmetros e incertezas inerentes a um projeto para produzir resultados econômicos consistentes, fornecendo ao investidor uma ferramenta de tomada de decisão. O presente estudo objetivou efetuar inicialmente a avaliação econômica convencional do cenário realista de uma mineração de agregados em fase de implantação localizada em Pernambuco, cujo Valor Presente Líquido (VPL) obtido foi R$ 24.006.309,20, o que representa um resultado positivo. Através da metodologia de Monte Carlo foi efetuada uma análise de risco considerando incertezas em 100.000 cenários, nos quais nove variáveis assumiram valores estocásticos dentro dos domínios definidos, produzindo assim uma distribuição estatística do VPL ao invés de um único número, sendo ainda efetuada uma análise de sensibilidade para identificar e quantificar as variáveis que mais impactam no resultado da avaliação. Por fim, os dois fatores mais influentes identificados na análise de sensibilidade inicial foram isolados e submetidos a uma nova análise de risco em outros 100.000 cenários. Foi concluído que o projeto é economicamente viável e que a probabilidade de gerar lucro aos investidores é de 99,77%, embora a probabilidade do VPL ser igual ou maior que o previsto pelo método convencional seja de 49,98%.

Oct 3, 2019

By Tristan H. Jones

Rib-related accidents in underground coal mines continue to cause injuries and take lives at a rate of 1.3 fatalities per year for the last 17 years. Prevention of these accidents cannot simply be a matter of installing more rib support, though the eventual final solution will undoubtedly include that component. Ultimately, rib safety will be brought about through an understanding of the causes and mechanisms that cause rib fatalities, and by fine-tuning mining methods and procedures that are currently in place. This paper investigates the factors likely to be most influential in the occurrence of rib fatalities. A careful analysis has been conducted of the rib fatality reports documented by the Mine Safety and Health Administration between 1996 and 2013. The primary outcome is the determination of the relative impact of each type of rib fatality factor and a better understanding of the impact each factor has. Among the rib fatality factors considered, the influence of rib profileis considered for the first time. Mining height, rib stratigraphy, mining cycle, and the standing position of a miner next to a pillar are also considered. Questions are also raised about the current interpretation of rib fatalities due to mining depth, with the indication being that mines operating at the greatest depths may actually have the most respectable rib fatality record among all mining depths. Additionally, a new method of considering rib exposure is developed. Based on the results, recommendations are made for future work. By gaining a better understanding of how particular factors contribute to rib fatalities as opposed to rib failure or stability, it is hoped that future efforts can be more precisely directed towards the desired outcome of fatality reduction.

Jan 1, 2014

By D. H. Desy

The Bureau of Mines studied the properties of copper that had been oxide dispersion-strengthened with 2 vol pct of either yttria or alumina. Compounds of copper and aluminum or yttrium, coprecipitated from aqueous solution, were converted to oxides; the copper was reduced, compacted, sintered, and extruded. For the resulting copper-dispersoid composites, room temperature tensile properties after annealing in hydrogen, elevated temperature tensile and stress-rupture properties, and electrical conductivity were determined. Compared with oxygen-free copper, copper-2 vol pct alumina or yttria had improved strength after annealing between 400° and 1,000° C, and improved elevated temperature strength, with little loss in electrical conductivity.

Jan 1, 1969

By A Dal-Corobbo, R Zick

The coal mining industry is now faced with new regulatory requirements that limit the discharge of chlorides into receiving streams. To meet these new regulatory standards for treated mine water, a zero liquid waste approach was developed using advanced treatment technologies. The approach described in this paper will generate clean water for reuse in various energy endeavours or discharge to the environment, with no substantial liquid or solid waste from the treatment process leaving the site.CITATION:Dal-Corobbo, A and Zick, R, 2013. Zero liquid discharge in coal mining operations, in Proceedings Water in Mining 2013 , pp 221-226 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

By E. C. Tarpley, H. M. Cooper, R. F. Abernethy

"INTRODUCTION The determination of moisture in coal is an empirical test, because the time, temperature, and equipment are specified. The essential features of the equipment, and details of the procedure are given in A.S.T.M. Designation: D271-44 5/. According to the specifications, the following features should be incorporated in the equipment used:1. Oven must have uniform temperature in all parts.2. Oven must be constructed with minimum of air space.3. Oven must have means of changing the air two to four times per minute.4. Air must be dried over sulfuric acid (sp. gr. 1.84).The analytical procedures governing most of the nation's coal purchases are those given in A.S.T.M. Standards referred to above. These standards are general and of such scope as to cover all ranks and types of coal. In cases where only the moisture is required, a faster method than the standard procedure is desired.The semi-automatic moisture tester manufactured by the Brabender Corp. has proved itself in many industrial processes where the rapid estimation of moisture is of vital importance. With the approval of A. C. Fieldner, chairman, A.S.T.M. Committee D-5, and through the courtesy of Lowell Malan, chief chemist, United Electric Coal Companies, who loaned the tester, tests were made to determine the feasibility of using this instrument to determine the moisture in coal."

Jun 1, 1948

By Cleland N. Conwell

Intermittent coal mining has been conducted in Alaska for over a century. The first report of coal in Alaska was by the Veechy expedition of 1826 and 1827. Whaling ships used coal from Corwin Bluff near Cape Beaufort during the early and latter parts of the 19th century. The first coal mine in Alaska -- and probably the oldest in the Pacific coast of North America -- was opened by the Russians at Port Graham on the southwest tip of the Kenai Peninsula. The mine was operated from 1855 until 1867 when the United States took possession of Alaska. From 1880 to 1915 coal mining was carried on intermittently at several localities including Unga Island, Herendeen Bay, Chignik Bay, Kachemak Bay, and several points along the Yukon River. Small coal mines along the Yukon River provided fuel for the river steamers traveling from St. Michaels to Whitehorse. The area of sustained coal production began with the completion of the Alaskan Railroad from Anchorage to the Matanuska coal field in 1916 and to the Nenana field in 1918. Coal production continued to increase until 1953, at which time the oil and gas fields in Cook Inlet were developed. Oil and gas eventually replaced- the Matanuska coal as a means of providing heat and power. The military in the Anchorage area shifted from coal to oil in 1967 causing a decrease in production. The increased demand for electric power in the Fairbanks area offset the decline and coal production started to increase. In 1969-1970 the Golden Valley Electric Association constructed a mine-mouth power plant 70 miles southwest of Fairbanks near Healy and a power line to Fairbanks. LOCATION OF COALS IN ALASKA Coals are found in all geographical subdivisions of Alaska. Coal in Alaska ranges in age from Carboniferous to Tertiary, and grade from lignite through anthracite. Figure 1 shows the locations of Alaskan coal fields. The only fields that have been developed extensively are the Matanuska and Nenana. Figure 2

Jan 1, 1972

By R. E. Aune

In the present work, a critical assessment of the thermodynamic information available for the ternary system Ni-W-0 has been carried out; This was complemented by experimental measurements of the standard Gibbs energy of formation of the ternary compound NiWO, by solid state galvanic studies involving Zr02-CaO solid electrolyte. The assessed thermodynamic data, together with the phase diagram information available in literature, were employed to generate the ternary phase diagram for the system Ni-W-0 in the temperature range 1073-1273 K. The calculation of the phase diagram was carried out by computer optimization according to the "CALPHAD" method. The computed phase diagram has been confirmed at isothermal sections by identification of the phases present in samples equilibrated at defined temperatures.

Jan 1, 1994

By I. B. Lambert

The guide has been developed by officials from resources and environment agencies in the Australian Government and the South Australian and Northern Territory Governments. It elaborates on the Australian Government?s policy that new uranium mines are to be approved subject to world best practice environmental and safety standards. The guide outlines the best practice principles and approaches that apply generally to mining in Australia, before giving more detailed consideration to best practice environmental protection and regulation for ISR mining.

Aug 1, 2010

By John Blackshaw

NEW sources of supply of tungsten concentrates became necessary when the flow from the Far East was blocked in 1942. In Canada, one source has been noted in the workings of the Hollinger mine at Timmins, Ontario. For many years, a small amount of scheelite had been removed with the gold ore, and an appreciable tonnage of scheelite still remained in the mine. At the request of the Dominion Government, a plant was built to recover this scheelite in the form of a concentrate that could be used by the alloy-steel industry without further preparation. The plant began operating early in March, 1942, and a small tonnage of ore, carrying scheelite and gold values, has been treated daily since then. The possibility of recovering the scheelite had been considered on several previous occasions. Each time the proposal had been given up because the total profits from the enterprise were estimated to be insufficient to repay the capital cost of building the recovery plant. At the time the plant was built, additional information had been gathered about the tonnage and grade of the scheelite ore in the mine, and several thousand tons of ore grading higher than 0. 25 per cent WO3 (tungstic trioxide) were known to be available.

Jan 1, 1943

By Douglas G. Baker, Jason K. Braund-Read, Megan R. Sheffer

"Mica Dam, owned by British Columbia Hydro, is an 820 ft high earthfill embankment dam that was completed in 1973. The dam consists of a thin central impervious till core supported by upstream and downstream compacted sand, gravel and cobble shell zones, founded on bedrock and valley bottom alluvium deposits. In 2012, a project was started to install new piezometric instrumentation within the dam’s foundation and downstream shell zone.Deep drilling in large embankment dams is very difficult and requires careful planning to select a methodology capable of reaching the target locations, while ensuring that no damage is done to the surrounding dam fill zones. At Mica Dam, the drilling could only be carried out from the Dam Crest and drillholes had to be carefully oriented to avoid intersecting any existing dam instrumentation. The drilling work therefore required angled drillholes, at depths of up to 855 ft, through heavily compacted sands, gravels and cobbles, and to significant depths below the standing water table.During the planning stages for the Project, in early 2013, the preferred drilling method was selected and a drilling contractor was engaged for the drilling work. Critical dam safety concerns were identified and addressed during the design stage of the Project with the drilling contractor, so that risk mitigation measures and response plans could be effectively developed.This paper explains how background information and prior drilling experience influenced the final design and implementation of the drilling techniques used in 2013 to successfully drill the deep holes and carry out the piezometer installations in Mica Dam. Drilling procedures, instrumentation design and installation methods, as well as risk mitigation measures, are discussed."

Jan 1, 2017

By J. S. Dunning

Recent developments in energy conversion, mining and minerals processing, pollution control, and chemical equipment have placed in-creased demands on heat-resistant alloys. Higher corrosion rates from oxidation, hot corrosion, and sulfidation have increased consumption of materials that contain critical and strategic minerals and in many cases have limited productivity. The Bureau of Mines is conducting re-search on substitute materials having greater resistance to high-temperature corrosive atmospheres above 6000 C compared to austenitic stainless steels such as AISI 304, 316, and 310, with the goal of conserving critical metals such as Cr and Ni. The substitute materials are based on Fe-Cr-Ni-Al compositions. Aluminum additions to Fe-Cr-Ni alloys significantly improve high-temperature corrosion resistance, particularly sulfidation resistance. In the alloy, devised by Bureau researchers, aluminum additions are held in the form of a second-phase precipitate, 8NiAl. Minor additions of Ti, Mn, and Si improve workability.

Jan 1, 1984

By E. Johansson, T. Siren, K. Kemppainen

"As part of the preparations for the design of an underground repository for radioactive waste at Posiva’s Olkiluoto site on the west coast of Finland, there has been considerable site investigation work. Except the extensive surface based investigations, Posiva started in 2004 the construction of an underground characterisation facility termed the ‘ONKALO’. The ONKALO has now been completely excavated to the anticipated repository depth i.e. -430–450 m. The research conducted in the ONKALO gives further information on the bedrock and groundwater conditions of the final disposal site, as well as on the impact of the construction. The ONKALO provides an excellent opportunity to investigate the rock at tunnel scale, to conduct in-situ testing in rock, to develop excavation and final disposal techniques in realistic conditions. The ONKALO has aided in collecting the data needed, supported by a Preliminary Safety Assessment, for the application of the construction license that was submitted in the end of 2012. The target is to begin disposal operations in 2022. According to current plans, the final disposal would end in 2112 and the repository would be sealed up by 2120. The testing in the ONKALO still continues several years. This paper presents the current characterisation and research activities in the ONKALO. INTRODUCTIONAfter the Finnish Government's favourable policy decision in 2001, Posiva Oy, which is an expert organisation responsible for the final disposal of spent nuclear fuel of the owners in Finland, has focused further investigations at the Olkiluoto site, western Finland (Figure 1) and has started in 2004 the construction of an underground facility termed the ‘ONKALO’. The ONKALO is not a nuclear waste repository, but more than just an underground research laboratory URL – it is an underground rock characterisation facility. The ONKALO has now been completely excavated to the anticipated repository depth i.e. -430–450 m. In addition to facilitating bedrock research, the ONKALO also provides an opportunity to develop excavation techniques and final disposal techniques in realistic conditions. Later, the ONKALO facilities can be put into use when building and using the repository. The construction methods and materials have been selected in order not to jeopardise the long-term safety of the repository site. The findings of the ONKALO and other investigations will provide the knowledge needed for an application, supported by a Preliminary Safety Assessment, to construct a final repository for spent nuclear fuel at the site. Posiva has submitted an application for the construction licence for a disposal facility in the end of 2012. The target is to begin disposal operations in 2022 and the repository would be sealed up by 2120."

Jan 1, 2015

By J. K. Whyatt

U.S. Bureau of Mines researchers reviewed an in situ stress estimate developed from measurements of overcore strain taken at test sites on the 7300 and 7500 levels of the Star Mine near Mullan, ID. Although the field measurements of over coring strain were found to be useful, significant deficiencies were found in the stress estimation procedure. A new stress estimate was developed incorporating statistical methods and an improved understanding of stress concentration factors for interpreting overcore strains recorded by doorstopper' cells. Spatial variability of overcore strains and the implications for stress field variability were explored.

Jan 1, 2010

Sampling protocols and metallurgical flow sheets should be designed to suit the style of mineralisation in question. In support, the a priori need for evaluation programs is the characterisation of the mineralisation type(s) and domain definition. The proportion and size of gold particles has a direct impact on the effectiveness of sampling protocols and metallurgical processes employed. Within the coarse- to-fi ne gold spectrum, the samplability of a deposit ranges from relatively simple for fine-grained disseminated particles, through to difficult for coarse gold mineralisation. Conversely, metallurgical processing may be relatively simple for coarse gold ore and more difficult for very fi ne gold ores. Most deposits contain a proportion of both particle size types. The nature of mineralisation, mineralogy and associated gold particle sizing can only be determined by preliminary sampling to facilitate testing. Mineralogical studies, focussing on ore mineralogy and gold particle deportment, size and distribution are required as part of wider paragenetic and geometallurgical investigations. This paper discusses key issues and presents a methodology for the determination of gold particle size in coarse gravity-gold ores to support sampling optimisation and which also contributes to metallurgical characterisation.

Aug 1, 2010

By V. H. Gottschalk, John Grodd, R. S. Dean

"Elemantary ConsiderationsAlthough magnetiem was first discovered in and derives its name from an oxide of iron found in magnesia and called lodestone its pracical uses date from the discoveries of Faraday concerning its relation to the electric current. We may therefore start our discussion of magnetic properties with Faraday's law for relation between induced electromotive force and the change in a magnetic where e is the electromotive force induced in an electric circuit by a unit rate of change of magnetic flux N, H is the intensity of the magnetic field, S is the area enclosed by the circuit, and i an angle between H and the normal to dS. In the simplest case we could use this law to measure the strength of a magnetic field by passing it at a measured rate through a coil of copper wire and determining the,electromotive force. As a matter of practical convenience, it will be simpler to nee the total quantity of electricity produced by a change of magnetic field. For this purpose we may let i denote the strength of the electric current induced by the change of magnetic field in time in a circuit of area a and resistance R. Then the total electric quantity. Q is given. by"

Mar 1, 1934

By I. Yochev, A. A. Metelev, A. I. Stepanenko, A. P. Purgin, M. G. Viduetsky, A. M. Panshin, I. F. Garifulin

"Pneumatic flotation cells of the KFM series are intended for the concentration of various ores and technogenic materials. The key operating principle of the flotation cells of the KFM series is the preparation of pulp for flotation in blowers, offering the advantage of an accelerated flotation process. This report summarizes the results of the implementation and operation of the KFM cells at Russian and foreign projects in 2015-2017. Their implementation and operation experience confirms the high efficiency of the KFM cells, especially in the head of lead and zinc, copper and zinc, copper flotation processes:head KFM-600 flotation cell recovers from lead-zinc ore (Rudmetal AD modular concentrator) up to 52% of lead in a single stage (without cleaner flotation) in a salable lead head concentrate (grade of up to 76% lead, 2.3% zinc). The flotation cell proved to be much more efficient than originally expected in the process description in the design phase,flotation of the copper-zinc ore from Novo-Shemur deposit (PJSC Svyatogor concentrator) in two KFM- 1400M cells recovers up to 65% of copper and 70% of zinc in a high-grade bulk copper-zinc concentrate containing 12-13% of copper and 25-28% of zinc. The installation of the two cells with a combined capacity of 26 m3 allowed to decommission 4 chambers of FPM-16 pneumomechanical flotation cells with a combined capacity of 64 m3 with a simultaneous increase in the overall ore concentration performancea single KFM-1400M cell (13 m3) recovers from the copper ore from Volkov deposit (PJSC Svyatogor concentrator) 53% of copper into a high-grade copper head concentrate (33.0% of copper). For comparison, the pneumomechanical cell FPM-16 (capacity 16 m3) has almost half as low copper recovery (29.6%) into the head concentrate at a lower grade (30.5% copper)."

Jan 1, 2018

By C. B. Foraker

THE Magma mine, of the Magma Copper Co., at Superior, Pinal County, Arizona, is 68 miles east of Phoenix and 21 miles west of Miami, Arizona, on highway U. S. 180. TEMPERATURES AND UNDERGROUND WATERS Surface Temperatures.-The average surface dry-bulb temperature at Superior is 72.4°; wet-bulb, 57.4°; with an indicated relative humidity of 38 per cent. The average yearly rainfall from 1921 to 1936, inclusive, was 18.74 in. The maximum condition of the atmosphere is in June, July and August, and the minimum is in January and February. The surface temperatures, of course, influence the temperatures of the down-cast shafts; however, the variation between winter and summer causes a difference of only 2° to 3° in the stopes and working places that are not close to the intake shafts. Average atmospheric conditions are shown in Fig. 1. Rock Temperatures.-Rock temperatures on the deeper levels of tile mine are as follows: LEVEL DEGREES LEVEL DEGREES 2000 109 3200 127 2250 112.5 3400 130 2550 116 3600 134 2800 120 3800 137 3000 124 4000 140 There appears to be no difference in rock temperature for any particu-lar horizon of the various formations or in the vein. Water encountered is always the same temperature as the rock from which it flows, and for this reason it is thought that the water is heated by the rock; not the rock by the water. The country rock apparently has been heated by the various intrusions, perhaps the vein itself, and has not yet had time to cool off. The geothermal gradient (Fig. 2) as indicated by the rock tem-peratures given above is 1 1/2 ° per 100 feet.

Jan 1, 1938