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By J M. Nel

Richards Bay Minerals (RBM) has recently undertaken an exercise to improve orebody basal definition. Older mineralisation bases (pre-2005) were highly erratic using old protocol reverse circulation (RC) (also known as Air Core) drilling and interpretation. The erratic nature of this estimation tool, resulted in impractical modelling towards the end of mineralisation (EOM), difficult mine design and unrealistic reconciliation. New protocol drilling, which includes high definition SONIC drilling data, has provided more data to be available for more accurate interpretation. Recent advances in drilling methodology has allowed for better penetration depths for SONIC drilling, which was previously restricted to approximately 80 m (pre-2005). This improved data set allows more emphasis on geological interpretation with factors such as lithology, slime content, and mineral species distribution. These factors when in unique combination enable more accurate and realistic EOM data points. These points are modelled into surfaces using conditional simulation techniques and result in a vastly smoother surface and consequently, a more representative resource.

Oct 5, 2011

By D. H. Carlson

The U.S. Bureau of Mines and Michigan Technological University (MTU) are collaborating to develop an apparatus for measuring diesel tailpipe emissions in underground mines. A tailpipe emissions measurement apparatus (EMA) is described that dilutes diesel exhaust and measures the concentrations of diesel particulate matter (DPM), CO, CO2, NO, and NO2 at known dilution ratios. The EMA was evaluated by side-by-side comparison of its measurements with those by laboratory-grade instruments. Concentrations of CO and CO2 measured by the EMA were usually within ± 14 pct of concentrations measured by laboratory-grade instruments for CO and ±20 pct for CO2. EMA¬ measured DPM concentrations were, on average, 29 pct lower than those measured by laboratory-grade instruments, with the differences being fairly consistent. The CO and CO2 differences are related to the method of calculating the EMA's dilution ratio, and the difference in DPM concentrations is attributed to thermophoretic and/or condensation losses related to the EMA's unheated sample probe. Preliminary analysis of data obtained with a heated sampling probe and modifications in the sampling procedures have indicated reduced variability in DPM and in the calculated dilution ratios.

Jan 1, 2010

By A. R. Mumford

During the course of an investigation conducted by the fuel section of the Bureau of Mines , on a hand- fired return- tubular boiler plant , it was found that the average cost of fuel to produce 1,000 lb. of steam was $ 0.5287 . Tests conducted after some simple changes had been made in the furnaces , and in the control of the boilers , showed the previous high cost of fuel per 1,000 lb. of steam could be reduced to the reasonable figure of $0.3540 , or the equivalent of over 30 per cent saving in fuel cost .

Mar 1, 1923

By Sang-Beom Lee

For thermodynamic prediction, the deoxidation equilibrium of aluminum in Fe-36%Ni alloy was studied with a cold crucible under Ar gas atmosphere at 1773K. For the deoxidation reaction in liquid Fe-36%Ni base, i.e.: A12O3(s) = 2A1 + 3O The interaction parameters and the equilibrium constant between aluminum and oxygen in liquid Fe-36%Ni alloy are given in the following: eo Al = -3.8, roA1=0.9, eAl° = -6.4, rAl° = 700, rAIAIo = 3.1, roA1o = 833 within the composition range of [Al] < 1 mass%. The temperature dependence of the equilibrium constant was obtained in the range of 1773K < T < 1973K by using reported data and present results as follows: log KAo= 0.58-24463.7 / T The deoxidation equilibrium with aluminum in Fe-36%Ni could be thermodynamically well described in the range up [Al] < 1 mass% with first and second order interaction parameters as well as the equilibrium constant determined in this study.

Jan 1, 2002

By O Arafat, R E. McIvor, K M. Bartholomew

It is well known in the industry that good classification system performance is vital to efficient ball mill circuit performance. But what is meant by good classification system performance, and how is it achieved? Unfortunately, the perceived complexity of classification systems, and the day-to-day time demands placed on the practicing metallurgist, all too often put the low-hanging fruit of classification system optimisation in the ‘I’ll do it later’ basket. Additionally, lingering confusion about basics such as the relationship between circulating load ratio and ball mill circuit performance makes the optimisation objective seem nebulous to both plant metallurgist and designer.Some of this confusion can be eliminated by taking a step back from the overwhelming detail of sharpness of separation curves, circulating load ratios, cyclone feed pressures, vortex and apex sizes etc, and focusing on the purpose or function of the classification system – to apply mill power to particles that need further size reduction! The classification system efficiency (CSE) metric captures how well the classification system is performing its function with two quick and easy size distribution analyses.Once CSE is understood and implemented as the optimisation criteria, well-tested methods for improving CSE can be implemented and quickly measured at plant scale to give confidence that the right change was made.This paper will describe CSE and how it is measured in practice. Next, a guide for improving CSE through manipulation of design and operating variables in the classification system will be provided. Examples and case studies will illustrate the gains in ball mill circuit efficiency that can be achieved by maximising CSE.CITATION:Bartholomew, K M, McIvor, R E and Arafat, O, 2018. A guide to maximising ball mill circuit classification system efficiency (CSE), for operators and equipment designers, in Proceedings 14th AusIMM Mill Operators&#39; Conference 2018, pp 553–558 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By C. C. Hustwit

The U.S. Bureau of Mines formulated a new mathematical model to characterize iron oxidation. The new model is intended to replace the currently used model, recommended by the Environmental Protection Agency (EPA), for most acid mine drainage (AMD) treatment applications. The new model was evaluated in a series of five synthetic AMD treatment tests. The initial ferrous iron concentrations ranged from 800 to 2,400 mg/L. In these tests, the Bureau model underpredicted the measured ferrous iron oxidation rates by a range of 3.24 to 8.03 pct. This was considered to be in close agreement after consideration of experimental error. The model proposed by the EPA was also evaluated using the same test data. The EPA model underpredicted the oxidation rates by a range of 77 to 100 pct in four tests and overpredicted by 19 pct in one test. The model formulated by the Bureau expresses that the rate of iron oxidation is a function of a treatment system&apos;s oxygen transfer efficiency and is independent of ferrous iron concentration when treated to near-neutral pH&apos;s. A new test method for evaluating oxygen transfer in flow-through reactors is also described.

Jan 1, 2010

By M. A. Grossman

The harden ability of most steels can be predicted within 10 to I5 per cent provided the complete chemical composition is known, including "incidental" elements; and provided the as-quenched grain size is known; and provided, finally, that the composition and heating temperatures for hardening are such as to result in austenite free from carbide particles. In the method proposed herein, the steel is considered as having a base hardenability due to its carbon content alone (the hardenability of a "pure steel" of the given carbon content, without any other elements), and this base hardenability is multiplied by a multiplying factor for each chemical element present. After multiplying all these together, the final product is the hardenability. Grain size may be taken into account either in the base hardenability or after the multiplication. Hardenability is stated in terms of "ideal critical diameter", namely, the diameter of bar, in inches, that will just harden all the way through (absence of un-hardened core) in an "ideal" quench, and the calculation may also be related to the Jominy test The data bring out certain features rather clearly. For example: i. It is quite useless to attempt to predict hardenability unless all elements, including "incidentals,,) are known; thus an "incidental,, chromium content of 0.20 per cent increases the hardenability by about 50 per cent. 2. The relative effectiveness of different alloys is sometimes unexpected; thus molybdenum when calculated in this way appears to be of the same order of effectiveness as manganese, rather than much more powerful, as would appear to be common experience; observe that an increase from no manaanese to 0.20 per cent Mn provides a multiplying factor of 1.67, and an increase from no molybdenum to 0.20 per cent Mo provides a factor of 1.63, an increase of over 60 per cent in each case. However, if to a steel containing 0.50 per cent Mn there is added "20 points of manganese," the factor is raised from 2.65 (for 0.50 per cent Mn) to 3.35 (for 0.70 per cent Mn), or an increase of only 26 per cent. Thus the fist small addition of an element has a much more powerful percentage effect than an equal further addition when some is already present, and in most cases the effect of molybdenum is considered in relation to a steel in which molybdenum is absent. 3. If two elements are equally effective, a greater hardenability will be obtained by using, for example, 0.5 per cent of each than by using I.o per cent of either of them alone. 4. A knowledge of the as-quenched grain size is essential for precise work, since a difference of only one grain size number (say No. 7 instead of No. 6) makes a difference of almost 10 per cent in hardenability (in these units); this, however, does not apply to certain It should be emphasized that in chromium steels (over 0.30 per cent Cr) and chrome-molybdenum and chrome-vanadium steels, un-dissolved carbides are likely to be present in the steel as quenched, and that in such cases the charts can indicate only a maximum Passible hardenability, whereas the extent of hardening actually obtained may be much less. Thus tests on a number of chrome-molybdenum steels have indicated a degree of hardening amounting to only 50 to 65 per cent of the maximum possible, and in chromium steels from full hardening (100 per cent) down to as low as 70 per cent. On the other hand, when the amount of such elements is small (Cr under 0.30 per cent, Mo up to 0.25 per cent in the

Jan 1, 1942

By M. A. Grossman

The harden ability of most steels can be predicted within 10 to I5 per cent provided the complete chemical composition is known, including "incidental" elements; and provided the as-quenched grain size is known; and provided, finally, that the composition and heating temperatures for hardening are such as to result in austenite free from carbide particles. In the method proposed herein, the steel is considered as having a base hardenability due to its carbon content alone (the hardenability of a "pure steel" of the given carbon content, without any other elements), and this base hardenability is multiplied by a multiplying factor for each chemical element present. After multiplying all these together, the final product is the hardenability. Grain size may be taken into account either in the base hardenability or after the multiplication. Hardenability is stated in terms of "ideal critical diameter", namely, the diameter of bar, in inches, that will just harden all the way through (absence of un-hardened core) in an "ideal" quench, and the calculation may also be related to the Jominy test The data bring out certain features rather clearly. For example: i. It is quite useless to attempt to predict hardenability unless all elements, including "incidentals,,) are known; thus an "incidental,, chromium content of 0.20 per cent increases the hardenability by about 50 per cent. 2. The relative effectiveness of different alloys is sometimes unexpected; thus molybdenum when calculated in this way appears to be of the same order of effectiveness as manganese, rather than much more powerful, as would appear to be common experience; observe that an increase from no manaanese to 0.20 per cent Mn provides a multiplying factor of 1.67, and an increase from no molybdenum to 0.20 per cent Mo provides a factor of 1.63, an increase of over 60 per cent in each case. However, if to a steel containing 0.50 per cent Mn there is added "20 points of manganese," the factor is raised from 2.65 (for 0.50 per cent Mn) to 3.35 (for 0.70 per cent Mn), or an increase of only 26 per cent. Thus the fist small addition of an element has a much more powerful percentage effect than an equal further addition when some is already present, and in most cases the effect of molybdenum is considered in relation to a steel in which molybdenum is absent. 3. If two elements are equally effective, a greater hardenability will be obtained by using, for example, 0.5 per cent of each than by using I.o per cent of either of them alone. 4. A knowledge of the as-quenched grain size is essential for precise work, since a difference of only one grain size number (say No. 7 instead of No. 6) makes a difference of almost 10 per cent in hardenability (in these units); this, however, does not apply to certain It should be emphasized that in chromium steels (over 0.30 per cent Cr) and chrome-molybdenum and chrome-vanadium steels, un-dissolved carbides are likely to be present in the steel as quenched, and that in such cases the charts can indicate only a maximum Passible hardenability, whereas the extent of hardening actually obtained may be much less. Thus tests on a number of chrome-molybdenum steels have indicated a degree of hardening amounting to only 50 to 65 per cent of the maximum possible, and in chromium steels from full hardening (100 per cent) down to as low as 70 per cent. On the other hand, when the amount of such elements is small (Cr under 0.30 per cent, Mo up to 0.25 per cent in the

Jan 1, 1942

By W. L. Josey

The Stibnite Project is an open pit/cyanide heap leach gold property located in the Yellow Pine mining district of central Idaho. The gold ore is in an oxidized zone of an epithermal deposit which is associated with a fault system related to Tertiary igneous activity of the region. The faults transect a roof pendant of metamorphosed sediments within crystalline rocks of the Idaho Batholith. Seasonal mining is at a rate of 3900 mt/d (4300 st/d) of ore with a 2:l waste to ore ratio. Gold and silver are recovered through a process of cyanidation, carbon adsorption and desorption, and electrowinning. Disposal of the leached ore requires chlorination of the residual cyanide.

Jan 1, 1984

By Milton H. Fies

The dedication of thin building, establishing a station of the U.S. Bureau of vines at the University of Alabama, bristles with significance and challenges the wisdom and courage of the people of this State. Here is united, in common aim, the Engineering School of our State University, recognized for its competency and usefulness, and the Federal Bureau of Mines. The Bureau is one of the most effective arms of our National Government. It outranks in accomplishment, measured by practical scientific contribution, any like governmental organization in all the other nations. It possesses none of the characteristics of many other bureaus of government which have created popular bias. Here is one Bureau devoid of political suggestion or aim with a personnel zealously devoted to labor in scientific re¬search - an organization marked by practical accomplishment, alleviating many of life&apos;s burdens, raking easier life&apos;s pathway, and unerringly pointing to life&apos;s possibilities.

Jan 1, 1936

By Zhang Shixun, Wang Xuguang, Zhong Yinting

The Mongol-Russian Joint Mining and Processing Cu-Mo Corporation Erdenet signed a contract with BGRIMM for building a plant to produce 10,000 t/y multitype emulsion explosives by using BGRIMM emulsion explosives manufacturing technology. In order to ensure the realization this contract, we have done a series of studies on the design of technological process and the selection, layout of the production equipment, and made a success.

Jan 1, 1997

By Henry O. Ash

On June 16, 1982, Rio Blanco submitted a request to the US Minerals Management Service for a suspension of operations under its federal oil shale lease. The company also wanted a five-year delay in the requirement for further minimum royalty payments and a five-year extension of the primary lease terms of 20 years. The request was granted in late July in the midst of yet another slump in oil shale activity. Rio Blanco holds one of four Federal oil shale leases that were granted under the Interior Department&apos;s Prototype Oil Shale Leasing Program in 1974. That year the DOI offered for lease six federal oil shale tracts, two each in Colorado, Utah, and Wyoming. The leases in Colorado and Utah were granted for substantial bo¬nus bids as shown in Table 1, but the Wyoming tracts had no takers. In the eight years hence, no oil shale has been produced commercially. Indeed, the prospects for successful oil shale development have fluctuated widely, continuing a 60-year- old cyclical pattern in the fledgling industry. The prototype program was developed and initiated while this infant industry struggled to begin anew, without the benefit of any federal oil shale leasing regulations, or any history of federal oil shale leasing at all. Permanent regulations were not required because the effort was limited and designed to provide the foundation for any future leasing program. Section 21 of the Mineral Leasing Act of 1920, a Nov. 30,1973 Federal Register notice, and the leases themselves constitute the legal and regulatory framework for leasing and developing prototype oil shale tracts. The prototype leases contain some unusual provisions including installment payments of the bonus bids, credit of early development costs against payments due the government, adjustment of royalty rates according to regional crude oil price fluctuations, and unusually extensive and detailed environmental stipulations. Several of the provisions were intended to provide incentives for early development and financial penalties for failure to develop. The requirement for minimum royalty payments in lieu of actual production royalties beginning in the sixth year of the lease falls in the latter category. After the sixth year, the required minimum payment is increased east year through the 15th year by an increment equal to that due in the sixth year. The minimum royalty then remains the same through the final five years of the primary lease term. Minimum royalty is calculates on a different production level for each prototype lease, based on the quantity of 125 L/t (30 gal/st) shale estimated to be recoverable from. each tract by room-and-pillar mining. That mining system was selected for reference as the only proven extraction method for oil shale at that time (early 1970s) The minimum royalty (production) rate was calculated for each tract by direct comparison of its estimated recoverable 125-L (30-gal) resource to a hypothetical tract containing 1.9 Gt (2.1 billion st) of recoverable 125-L (30-gal) shale. For the hypothetical lease tract, the minimum production for the sixth year was calculated to be 3.2 kt/d (3,500 stpd), increasing by a like amount each lease year through the 15th to a daily production rate of 31.7 kt (35,000 st). Under this formula, the hypothetical lease tract would yield $16 million in minimum royalty payments over the 20-year primary lease term if there were no production, no development credits taken, and no adjustment of the royalty rate due to fluctuation of regional crude oil prices. The production rate of 31.7 kt/d (35,000 stpd) was a somewhat arbitrary figure that was considered, in the early 1970s, to represent a minimum size for a commercial oil shale operation. A production rate of 7.9 dam3/d (50,000 bbls/day) soon supplanted that smaller mine production figure as a rule of thumb for the minimum size for a commercial facility. The proportional relation-

Jan 11, 1982

By D. C. Sharpstone

THE Polaris-Taku mine is a new gold property, situated in the far north-west corner of British Columbia. Production was started in November, 1937, at the rate of 150 tons a day, and recently has been increased to 200 tons. The property, which promises to become one of British Columbia&apos;s important gold mines, was developed and brought into production in twenty-nine months after work was started on it as a virtual prospect in June, 1935, by Edward C. Congdon and associates. Inaccessibility, extraordinarily complicated geological problems, and difficult metallurgy retarded development of the property before acquisition by the present interests, an::l, in fact, discouraged two companies that had spent substantial sums upon it from continuing development. This paper is concerned with reviewing the development of the property and describing the principal features of the geology, the mining and metallurgical practices, and the equipment employed. Acknowledgments The writer is indebted to Mr. Edward C. Congdon, President of the Polaris-Taku Mining Company, and to various members of the mine staff for their assistance in assembling the data for this paper. Special thanks are due to Mr. George G. Griswold, Jr., of Butte, Montana, who generously and painstakingly prepared the section dealing with milling and power. Microscopic studies of ores were made by Professors Frank Grout an::l G. M. Schwartz, of the University of Minnesota, and by Mr. Griswold.

Jan 1, 1938

By V Lishchuk, C Lund

"Geometallurgy aims to improve resource efficiency by creating a spatial model of an orebody in order to forecast ore processing behaviour. This enables calculation of the economic value of each ore block and the use of this information for production optimisation and hence an improved economic return. Modelling uses different approaches to link geological information with mineral processing: traditional approach uses chemical elements; proxy approach – small-scale geometallurgical tests; and mineralogical approach – minerals.A geometallurgical model is obtained during execution of a geometallurgical program encompassing the entire value chain of ore processing. Available methods for establishing a geometallurgical model are in many cases limited and not yet comprehensively validated. Some of the limitations include: sampling density, assaying methods and quality, forecast accuracy block model content, geostatistical solutions for non-additive components, blending and assessment of positive impact of geometallurgy on the project. Synthetic ore models combined with process simulations and production scheduling, based on strong case studies, present a viable way to address these shortcomings. Such synthetic ore models will have realistic elemental and mineral distribution, textural attributes and 3D variability for input into the design of an appropriate geometallurgical program.A synthetic orebody was created for an apatite-bearing iron deposit. Variability in modal mineralogy was described by geostatistical parameters combined by a set of linear and trigonometric functions. The generated synthetic deposit model was used to develop a mine production plan to generate a simulated process feed. The process feed had detailed information required for beneficiation process simulation: modal mineralogy, chemical composition of minerals, grain size and specific gravity. The synthetic system was used to measure benefits that a geometallurgical program developed for real cases could provide, compared to production where only head grade is optimised. These include a reliable sampling strategy and confidence in the feed-quality forecast for economic evaluation.CITATION:Lishchuk, V, Lund, C and Lamberg, P, 2016. Development of a Synthetic Ore Deposit Model for Geometallurgy, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 275–286 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

By M. M. Kirk

Forced-air cooling of cold-mold arc furnaces for melting zirconium and similar metals has been investigated and found to be practical from the stand-point of heat transfer. It is believed that this technique offers a significant reduction in explosion hazard from that inherent in water cooling, the current industrial practice. A copper crucible, with longitudinal, integral-type fins and with an outer jacket to confine the airstream, appears to be the best general design for air-cooled operation. Cutting the fins into short segments and bending the segments to produce a herringbone effect increased the airstream&apos;s turbulence and was found, on test, to increase the heat transfer up to 35 percent.

Jan 1, 1959

By Shu Xinqian, Zhou Weining, Shan Zhiqiang, Feng Jifu

"Modified calcination conditions on the extraction of rubidium resources from muscovite (KAlz[AlSi301o](OH)2) ore are systematically studied. The impact of various factors on the rubidium leaching rate is also analyzed, including the agent types, roasting time, the mass ratio, roasting temperature, etc. The results indicate that the best agent is a combination of sodium chloride and calcium chloride, its mass ratio ofmuscovite/NaCl/CaClz is 1:0.25:0.25. Calcined at 8500 for 30 minutes, the rubidium leaching rate is up to 90%. The reaction of muscovite ore with the chlorinating agent CaC12 has been studied by TG/DSC, the calcined intermediates are identified by XRD, and the surface morphology before and after leaching is characterized by SEM. Rubidium chloride products can be obtained with a purity of 92% by using t-BAMBP extraction, hydrochloric acid reextraction and purification.IntroductionRubidium is one ofrare alkali metals, which is found in 1981 by German R.W. Bunsen and G.R. Kirchhof. Rubidium includes the following features: atomic number is 37, each atom weight 85, the body-centred cubic lattice, each unit cell has two metal atoms, and the density is 1.532g/cm3. The properties of low melting point (312.46 K) and malleability make rubidium have a strong chemical activity and photoemission. Rubidium provides a main material for manufacturing phototube with a wide range oflight waves, which possess high sensitivity and stability. Efforts are being made to carry out research on rubidium ion gngine, magnetic flux generators and thermoelectric transudcer, etc. Rubidium exists in muscovite ore(KA12[AlSh01o](OH)2 (monoclinic system and hexagonal layered structure) as isomorphism and replace the large radius of the potassium atoms, as well as part of rubidium locate between structure layers. The complex components bring great difficulty in comprehensive recycling."

Jan 1, 2013

By D. Duffy, V. Papangelakis

"MgCl2 brines present a number of potential advantages for the processing of saprolite ores for nickel production. With a chloride system it is economically requisite to recover hydrochloric acid, which can be performed in a more energy efficient manner by precipitation and subsequent decomposition, as compared with conventional pyrohydrolysis. In addition, concentrated MgCl2 solutions enhance the activity of acid used, allow atmospheric leaching at elevated temperature and inhibit magnesium dissolution, which reduces acid consumption and increases metal selectivity. Leaching and iron control experiments have been conducted in concentrated MgCl2 brines, up to 4.5 m, to determine the most amenable process conditions. INTRODUCTIONChloride metallurgy has received an increasing amount of attention in recent years, especially with respect to the treatment of nickel laterites. Although the limonitic fraction can sometimes be economically treated by High Pressure Acid Leaching (HPAL), acid-consumption problems are encountered with the higher magnesium content of the saprolite fraction. As magnesium does not hydrolyze under HPAL conditions, as iron and aluminum do, any sulphuric acid used to leach magnesium is subsequently neutralized and wasted. In addition, this magnesium must later be disposed of in some fashion, further complicating the process. The most common method used to process saprolite is the Rotary Kiln-Electric Furnace (RKEF). This uses a kiln to dry and preferentially reduce nickel, before sending the calcine to an electric furnace. Although nickel recoveries are high, and the problems of acid consumption are avoided, the capital and energy costs associated with this process are high (de Bakker, 2011).The problems outlined above have led to the investigation of an atmospheric chloride process for the treatment of saprolite ores. The use of a chloride system presents a number of advantages, but also a few limitations. Metal chlorides tend to be very soluble, with a few notable exceptions (silver, mercury and lead), allowing for smaller processing equipment through more concentrated solutions. This solubility also serves to reduce the potential for scale formation. In addition, a hydrochloric acid leach is more aggressive than a comparable sulphuric acid system, and the activity of the hydrogen ion can be augmented by further chloride addition (Peek, 2011). This activity augmentation in concentrated chloride systems, first reported by Jansz, has repeatedly been referenced as a motivating factor behind chloride hydrometallurgy (Jansz, 1983; Harris and Magee, 2003; Harris et al., 2004; Demopoulos et al., 2008; McDonald and Whittington, 2008; Demopoulos, 2011; Peek, 2011; Verhulst and Lakshmanan, 2011). This activity increase allows a similar amount of acid to perform more in a concentrated solution than in a dilute solution, decreasing overall acid requirements. As such, much focus has been placed on leaching in concentrated chloride systems, or brines, using calcium, magnesium or sodium as the cation (McDonald and Whittington, 2008)."

Jan 1, 2012

"IntroductionNOTE.-All figures on the supply and consumption of coal in Canada cited in this paper were collected and reported by the Dominion Bureau of Statistics. The friendly and efficient co-operation of the officials of the Bureau. is the foundation upon which it has been my privilege to rely in the preparation of this and two earlier papers ( 1). My thanks are due also to our own staff on the Dominion Coal Board, who spared neither time nor effort in helping to prepare this paper.STATISTICS are comparatively dull and, to many people, coal statistics are duller than most. Coal is, however, the one basic mineral upon which our whole way of life depends and it may be well if attention is drawn to this fact by an occasional statistical review of the coal position in Canada. The recent furore over the oil and gas discoveries has so obscured the basic position of coal in many minds that a little basic coal information may be useful. It is for this reason that, for the third successive year, a factual survey of the coal position is being presented.In the previous papers (1) explanations have been given of the basic considerations in the preparation of these reports and some special phases of the overall picture have been examined. For these explanations, the earlier papers may be consulted."

Jan 1, 1952

By J. D. Noll

An effective technique to minimize miners? respirable dust and diesel exposure on mobile mining equipment is to place mine operators in enclosed cabs with designed filtration and pressurization systems. Many factors affect the performance of these enclosed cab systems, and one of the most significant factors is the effectiveness of the filtration system. High-efficiency particulate air (HEPA)-type filters are typically used because they are highly efficient at capturing all types and sizes of particles, including those in the submicron range such as diesel particulate matter (DPM). However, in laboratory tests, minimum efficiency reporting value (MERV) 16 filters have proven to be highly efficient for capturing DPM and respirable dust. Also, MERV 16 filters can be less restrictive to cab airflow and less expensive than HEPA filters. To verify their effectiveness in the field, MERV 16 filters were used in the enclosed cab filtration system on a face drill and roof bolting mining machine and tested at an underground limestone mine. Test results showed that DPM and respirable dust concentrations were reduced by more than 90% when the cabs were properly sealed. However, when the cab door was opened periodically throughout the shift, the reduction efficiency of the MERV 16 filters was reduced to 80% on average.

Feb 1, 2014

By Robert S. Sanford

The Broughton and Ring magnetite deposits in Essex County, N. Y., are thought to be in a mineralized belt that extends from the Lake Sanford mining district to Lake Champlain and passes through the Mineville district. The country rook is syenite-gneiss, similar to that in the Mineville area. The Bureau of Mines conducted a vertical magnetometer survey of the area. On the Broughton properties there are no indications, magnetically, of magnetite in any degree of concentration. &apos;On the ding properties a number of sharp, localized magnetic anomalies were observed. One core-drill hole was bored to a, depth of 48.3 feet, also, to test one of the strongest anomalies. The field work was in progress from December 1943 to July 1944. Although no iron ore was intersected by&apos; the drill hole, it showed that the amphibolite rook, at a depth of 6 to 100 feet below the surface, contained enough magnetite and iron sulfides to account for the anomalies recorded by the magnetometer survey.

Jan 1, 1949