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By J. Dubinski

Passive tomographic imaging is a technique that can be used for assessing seismic hazard in mines with high seismic activity induced by mining. The basic idea behind such an interpretation is the assumption that there is a relationship between high values of seismic velocity and high values of stresses in the rock mass. This relationship was confirmed both in laboratory an in situ measurements and constitutes the basis for rock-burst and seismic hazard evaluation. In the paper seismic tomographic images of seismic P wave velocity and seismic P wave velocity gradient have been presented to determine rock-burst hazard and locate high seismic activity zones in the area of coal panel 306 in Zabrze Bielszowice mine. The zones of high values of P-wave seismic velocity gradient have been found to correlate with the distribution of high energetic seismic events. Passive tomographic images were obtained using modern bent ray tracing tomography algorithm with regularization. Recent study give rise to observation that not only high values of P-wave seismic velocity is important in seismic hazard assessment as many strong seismic events are located in the areas of high values of velocity gradient. This recent observation permits to increase quality of passive tomographic imaging for detection of seismogenic zones and is presented in the paper. The advantage of passive tomography comparing to active tomography is that sources of seismic waves are generated by rock mass itself and costly controlled explosions are not necessary. Another practical limitation of active tomography is the size of study area which usually do not exceed 200 m by 200 m. This restriction does not appear in passive tomography where area of study is usually 2000 m by 2000 m and the tomographic images delineate velocity structure in locations where mining induced seismicity take place. Passive tomography method utilized in analysis of rock mass state in underground mines encounter some difficulties concerned with the numerical stability of optimization algorithms. In the presented approach to passive tomography problem numerical stability is obtained by special regularization method and by application of very stable numerical seismic ray-tracing algorithm.

Aug 1, 2013

By George Burgess

1. Introduction.-THE methods of production of sound steel ingots have been described in several papers read recently before this Institute. It was thought by Director Stratton, of the U. S. Bureau of Standards, and ourselves to be of sufficient interest and importance, as a further contribution to this subject, to compare the properties of rails rolled in one mill according to American practice, from several types of ingot of substantially the same chemical composition but cast by several processes; and to compare in considerable detail the properties of an ingot manufactured by the Hadfield special-feeding process and one of the usual type for rolling into rails. We are greatly indebted to F. W. Wood, President of the Maryland Steel Co., who made this investigation possible by kindly placing his rail mill at our disposal; and he and his associates at Sparrows Point did everything possible to make the experiment a success, including the execution of several of the tests and analyses. 2. Classification and Manufacture of Ingots.-Of the ingots used in this investigation, eight were made in Sheffield and furnished by Sir Robert Hadfield, and one was made and furnished by an American steel company,' as a comparison ingot, and was supposed to represent the usual type of ingot from which rails are made.

Jan 2, 1915

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

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

By W. H. Emmons

A DIATREME is a hole blown through a rock by gases, presumably of volcanic origin. Not all pipes of ore have formed by deposition of metals in such openings, but a considerable number have so formed. In this paper are described peculiar vein patterns that are believed to be incipient diatremes or "perforation pipes." Grading into this group is one showing thorough brecciation without rotation of fragments and into the latter a group of ore pipes showing brecciation with rotation of fragments-many of them with well rounded fragments. Also, volcanic craters or volcanic necks are explosion pipes. These may carry ore deposits along their contacts and some carry deposits within the pipes. As new data appear this group becomes increasingly impressive. It has been shown that veins are concentrated in and about cupolas of batholiths and that such veins commonly lienearly parallel to the long axes of cupolas as these are shown on plans. It would be supposed that conical cupolas would have radial patterns, but few vein systems exhibit radial patterns. It is probable that explosion pipes also may form in areas above conical cupolas. From these the gases escape more violently because the loads above them are less than above an elongated cupola where larger bodies of rock are affected. The hypothesis is sound mechanically, for the top of an elongated cupola is a line whereas the top of a conical cupola is a point.

Jan 1, 1938

By D. T. G. Emmett

[ ] South Africa, by a wide margin, is blessed with the largest known world reserves of platinum. The optimal use of these resources for the benefit of all South Africa?s peoples is a goal shared by industry, government and all South Africans. The challenge is to grow the industry in accordance with demand growth and thus to manage a long-term stable growth pattern for PGM usage world-wide. Amplats and the South African producers are well able to meet this challenge [ ] South Africa plays a critical role in the provision of Platinum Group Metals for world markets. In fact no other country has the productive capacity nor the geological reserves to supply the world?s platinum needs. The world required some 5 million ounces of Platinum in 1997 and South Africa supplied a full 74% thereof. The second major supplier, Russia provided a further 18% of world demand, primarily as a by-product of nickel production. Amplats satisfied about 37% of total world demand followed by Impala 20% and Lonhro at 12%.

Jan 1, 1998

By Irving Palmer

DURING the past 15 years in this country there have been few additions to the literature of lead smelting. After the consolidation of the principal smelting companies at the beginning of this period it became the policy of the larger company not to permit the publication of details in regard to many of its metallurgical processes. This led to similar action on the part of some of its competitors. The result, of course, was a decided decrease in the number of papers written upon the practical side of the subject. Even upon the theoretical side the contributions have been few. It would seem that the industry as a whole has suffered from this general unwillingness to exchange information. While it is true that a given company is entitled to the reward that may come from its own development of a new process, it is also true that a still greater reward might be the result of a free exchange of ideas with other plants. This principle has been recognized for a long time in other metallurgical industries, notably those engaged in the smelting of iron and the manufacture of steel. It is pleasing to note that the leading lead-smelting company has very recently decided to modify its policy of secrecy, and to permit the publication of papers covering some details of its metallurgical methods. The statement has been made that there has been but little progress in lead smelting during the period just referred to. This impression is due in part to the scarcity of papers written on the subject. There has been, it is true, some discussion of the new sintering processes, but with little reference to detail, and to the effect that blast roasting has had upon the industry as a whole. In spite of the prevailing belief, it may be said that lead smelting, in common with other metallurgical industries, has made steady and substantial progress. It is the purpose of the present paper to discuss briefly some features of the industry as it is now conducted, with special reference to the blast-furnace practice in the larger silver-lead smelting plants of the United States and Mexico.

Jan 7, 1914

By Robert W. Row

The run-of-mine coals in the U.S. being fed to coal cleaning plants have become increasingly finer due to the adoption of modern, mechanized mining methods, and economics are dictating that finer fractions, as well as the coarser fractions, be cleaned. Traditional dewatering methods (vacuum disc filters and centrifuges) are proving inadequate for preparing fine refuse for disposal or fine coal product for downstream handling. The Electric Power Research Institute has been supporting research into the nature and limitations of mechanical dewatering, and novel ways of improving it. The efficiency of various silicone-based chemicals at aiding dewatering was investigated, and it was shown that the final moisture content of hydrophilic coals after vacuum Filtration could be reduced by 1 to 5% with up to 40% increase in filtration rate by addition of one of several of these silicones. EPRI also has an on-going program at its Coal Cleaning Test Facility to investigate the performance of the Larox pressure filter in comparison with that of the disc filter and solid bowl centrifuge. While improvements to the dewatering performance currently being obtained in existing operating plants by such means are achieveable, a fundamental study has shown that, regardless of changes in operating conditions, there may be a limit to the lowest cake moisture content attainable by mechanical dewatering of a given sample of coal fines. Determination of this limit for a specific coal would be useful in designing facilities to dry and handle it.

Jan 1, 1984

By M. J. Ferrante

Low-temperature heat capacities and high-temperature enthalpies were determined calorimetrically for Na2Cr04, A reversible transition was found at 700° K with an isothermal heat of transition of 2,263 cal/mole. High-temperature X-ray diffraction analysis showed that the transition was a crystal change from the low-temperature orthorhombic form to the high-temperature hexagonal form. At 795° K the hexagonal form had the parameters a, = 5.691 and co = 7.478 A. Powder diffraction data for the high-temperature form are given and possible structures discussed. The melting point was determined to be 1,065° K with a heat of fusion of 5,810 cal/mole. Enthalpy and entropy increments from this investigation were combined with other data from the literature to calculate enthalpies of formation, free energies of formation, and equilibrium constants of formation from 298° to 1,400° K. Equations that represent enthalpies above 298° K are given.

Jan 1, 1972

By R. A. Busch, R. R. Backer, R. W. McKibbin

The Bureau of Mines conducted laboratory tests to determine the slur- ry transport and deposition properties of coal waste. A 188.5-ft pipe test loop was constructed with instruments to measure density, flow, and power and discharge, suction, and differential pressures. A 20-ton sample of Western coal waste was crushed and screened to 100 pct minus 1/2 in. Eight loop tests using this graded waste were run at slurry densities ranging from 1.15 (25 wt-pct solids) to 1.46 (60 wt-pct sol- ids). The resultant friction-pressure gradients ranged from 0.06 to 0.24 ft of water per foot of 4-in standard steel pipe. Particle-size degradation was significant. In a typical loop test lasting 25 min, the minus 200-mesh fraction increased from 19 pct to 34 pct. Deposi- tion tests showed that without the use of a flocculant, the slurries would not dewater when left to stand for 1 week. With flocculant, ad- ditional water was released; however, the settled slurries remained thixotropic. Further testing is necessary to determine their suitabil- ity for use as backfill material in active mines.

Jan 1, 1965

By Luciano Ramos, Daniel Santos

Igarape Bahia Mine, which is currently the largest gold producer in Brazil, belongs to Companhia Vale do Rio Doce - CVRD and is situated 250 miles south of Belem, state of Para, Brazil. It came into operation in 1992 with the startup of the CIP plant, fed with high grade ore, presently at 3.4gAu/t. The ore is totally oxidized, containing extremely fine particles of free gold. The ore is soft and low cyanide consumer, with a maximum work index of 12kwh/sht. In 1993 a Heap Leaching Plant was set into operation, handling low grade ore of · 1.s g Ault. The Heap Leaching Plant has undergone two major expansions. In 1994 its capacity was raised from lMtpy to 2,2Mtpy and in 1998, with a thorough remodeling of crushing operations, it was raised again to 3.5Mtpy. The CIP plant, which was originally designed with a capacity for l. lMtpy of ore, was expanded to a 2.2Mtpy capacity in 1994. Presently, the CIP plant accounts for most of Igarape Bahia's total gold output, with approximately 220,000oz per year. Igarape Bahia's annual output is around 360.000 oz. Over the last three years, work has been intensified to improve performance of both processes, aiming at maximizing productivity of the two industrial plants. Reduced costs and increased revenues have been given a special emphasis in the unified improvement program. Cost · reduction has involved developing methodologies to rationalize consumption of materials and inputs, while increased productivity has been given top priority as a means of generating increasing revenues. Thus, all of the improvement and development efforts have had a common goal: to increase operating income.

Jan 1, 2000

By Ian Wark

IT has been shown1 that in the absence of heavy metal salts, the nature of the alkali used to promote differential flotation-whether caustic soda, lime or sodium carbonate-is unimportant. The hydroxyl ion is the active depressant upon which differentiation depends, and the pH value determines whether a film of collector forms on any given mineral surface. Even when sodium cyanide is used in conjunction with alkalies, it is the pH value and not the nature of the alkali or its concentration that is important, for upon the pH value alone depends the fraction of the added sodium cyanide that is present as cyanide ion, which is the true depressant. When heavy metal salts become dissolved from partly oxidized ores, or are added intentionally as activators, conditions are different, and metallurgists have found that lime and sodium carbonate do not give identical results, even when added in amounts sufficient to produce the same pH value. In the few instances in which we had attempted to confirm this view, we had failed to do so, for, with nothing to guide us, we had inadvertently chosen solutions that gave contact curves of a "stand-ard type," which correspond to conditions of ready floatability. Now, however, we find that, in the presence of copper salts, and within a restricted range of reagent concentrations, the nature of the anions present is important, and that sodium carbonate does influence the results, apart from the usual effect due to its influence on pH value. With the recognition of this anion effect, and with the control of temperature already introduced,' it appears that the factors of major importance in connection with this set of reagents have been taken into consideration, and that the results obtained and the conclusions drawn from contact tests will be of more immediate value to the operator.

Jan 1, 1939

By Earl Sweet

This overview of economic and financial market prospects examines the current and future state of the global and Canadian economies with a focus on commodities. Housing prices and global conflicts all have effects on the overall economy and metal prices and mining developments in particular. This presentation outlines where we are today and where the author and BMO see the economy moving over the next few years.

Jan 1, 2016

Blasthole cone samples are used for grade control in most of the iron ore mines of the Pilbara of Western Australia. This method of sampling is straightforward when the target sampling interval matches the mining bench height but is problematic when the blasted bench is 'flitch' mined in two or more passes. Crucially, the sampler needs to decide which part of the blast cone represents the upper and lower flitches to be mined. This sampling selection decision is usually made by assuming a certain proportion represents the subdrill component of a sampling face in a blasthole cone and the two flitches are split equally from the remaining portion. In this study of a West Australian channel iron deposit, sampling experiments were carried out to determine the actual proportions of a sampling face in a blasthole cone by spray painting the blast cones at flitch and subdrill boundaries during the drilling process. Measurements were then taken to assess the differences between assumed sampling face proportions and actual proportions. A set of paired samples was also collected from a number of blastholes with one set using the assumed proportions of each flitch for each sample, while the other set sampled the actual proportions, which were identified from the cone painting process.

Mar 1, 2010

By Mohana Rao Andavarapu

Sandur Schist Belt of Karnataka Craton constitutes a hub of iron ore mining activity. The iron ore deposits of the region are closely associated with Banded Hematite Quartzite (BHQ), Shale, phyllites, & Graywackes. The gradual depletion of high-grade iron ores due to extensive utilization has necessitated the exploitation of low-grade iron ore including BHQ, which are available abundantly in different parts of India. In the present investigation, BHQ samples from Kammaturu mines of Sandur schist belt was studied to assess the possibility of upgrading the iron content to the maximum possible extent, so that it may find the better use in iron and steel industries. The sample collected from the study area analyzing 39.76% Fe, 39.8% SiO2, 0.66% Al2O3, and 1.92% LOI was subjected to beneficiation studies. The mineralogical studies reveals that the samples are predominately consists of hematite as an iron mineral and quartz as gangue mineral. Liberation studies and chemical analysis were carried out prior to the beneficiation process, which is flotation. Cationic Reverse flotation studies were carried out and discussed by varying different process variables such as feed size, depressant, collector cum frother dosage & pH. The results obtained are encouraging with an appreciable improvement in the iron content up to 62.8% Fe with Fe recovery of 50.43%. The concentrate obtained from the flotation can be used for pellet making process.

Sep 1, 2012

It is well known that the electrochemical precipitation of metals from aqueous solutions is conducted by the electrochemical standard potential scale. This scale indicates the order of nobility of different metals. Consequently, the standard potential scale shows the nocivity of metals when they are present as impurities in aqueous solutions containing metals. In the hydrometallurgy of cobalt, nickel and zinc are co-deposited during cobalt electrodeposition. Nickel has almost the same chemical, physical and electro-chemical properties as cobalt. Furthermore, the electrochemical potential of nickel is relatively higher than that of cobalt. For this reason nickel is preferentially reduced at the cathode in the presence of cobalt. The presence of small amounts of zinc in solution leads to contamination of cobalt electrowon. In spite of its lower electrochemical potential in comparison to cobalt, zinc is reduced on the cathode because of its higher hydrogen-overvoltage. But, in the hydrometallurgy of zinc, small amounts of cobalt present solution will contaminate the zinc reduced at the cathode. This is easily explained by the fact that cobalt is more electropositive than zinc. The co-deposition mechanism of sulphur at the cathode is not yet well understood. Great attention has to be paid to the mechanism of co-deposition of sulphur during cobalt electrolysis. This work aimed to investigate the influence of some parameters on the co-deposition of nickel, zinc and sulphur during cobalt electrowinning. The following parameters have been investigated: Co2+concentration in the electrolyte, current density and ammonium sulphate (NH4)2SO4additions. It has been established that important nickel, zinc and sulphur co-deposition takes place at relatively low Co2+ concentration in solution. The behaviour of these elements with regard to the cathodic phenomena has been established.

Jan 1, 2005

By Dennis C. Drehmel

Because of their relatively low cost, carbonate rocks are desirable in abatement of air pollution by acid gases. Although many processes and types of equipment have been proposed to control or prevent the emission of acid gases, this paper considers only three: dry-limestone injection, wet-limestone scrubbing, and chemically active fluid beds. For the dry-limestone-injection process, surface area-especially from pores in the 0. 2- to 0.3-micron range-is one of the important parameters which will discriminate between less and more reactive carbonate rocks. Marls, chalks, oolitic calcites, and other fine- grained limestones show larger-than-average surface areas and better- than-average removal efficiencies. Inert chemical impurities such as silica detract from reactivity. A dolomite is not necessarily less re- active than a calcite. For the wet-limestone-scrubbing process, dissolution rate of the carbonate rock controls the differences in stone reactivity. In this process the dolomites are less reactive than the calcites. In addition, more than 90 percent of the dolomitic impurity in Fredonia white lime- stone is inert during limestone scrubbing. Dissolution rate is also a function of particle size for large-grained carbonate rocks. Fine grinding of such rocks will increase the dissolution rate to that of a fine-grained rock. Hence the requirement of high dissolution rates may be met by some grinding of fine-grained material (e. g., marl, chalk, or other limestone) or by fine grinding of coarser grained material. For chemically active fluid beds, chalks have the best capacity to absorb both SO2 and H2S. However, most chalks are probably unsuitable for fluid beds because of high attrition rates. Because dolomites are generally more reactive than calcites with both SO2 and H2S, a dolomite would be the feed material of choice for a chemically active fluid bed.

Jan 1, 1974

The work that has formerly been done by the Waste-reclamation Section of the War Industries Board has been tranferred to the Department of Commerce and will be continued as the "Waste-reclamation Service" of this Department. The plan contemplates the establishment of a national waste-reclamation system-comprehensively organized, effective and systematic, and animated by a sympathetic understanding of the purposes and principles of the Service. It is the expectation of the Service that the work will be taken over by the several municipalities or other local governments and made permanent through the creation of a department or bureau for the salvage of waste materials. The local Waste-reclamation Council should be composed of representatives of cooperating national organizations, selected by the mayor; or chief executive officer of the city, town, or village. It will be the duty of the Council to inaugurate an educational campaign to conserve materials so as to make less waste, together with a reclamation movement that shall reclaim the residue of waste in the community, thus conserving the resources of the country both in man power and in materials. The most important materials to reclaim at the present time are: paper, steel, copper, wool, iron, brass, cotton, zinc, aluminum, rubber, babbitt metal, tin-foil, lead, tin and leather.

Jan 8, 1919

By Joel K. Waterland

GENERAL DESCRIPTION Open cut-and-fill stoping for many years was prob¬ably the most widely used mining method in under¬ground metal mines. Then for a time this method was largely supplanted by the blasthole stope. It again be¬came popular as many mines reached depths or grades where methods requiring large open voids to remain open for extended periods of time became unsuccessful, often as a result of excessive dilution. The open cut-and-fill method is very flexible and is readily adaptable to almost any ore body. The standard application requires that a slice of ore usually 2.4 to 3 m (8 to 10 ft) thick be removed from the back of the stope, and as the ore is taken down, the back is dressed and rockbolted. After the back is secured, the broken rock is removed through rock passes to the level below. When the rock has been removed, the rock passes are extended upward the height of the ore removed, the stope is backfilled, and another cycle is mined. This method is best employed in plunging ore bod¬ies with considerable vertical extent, ore areas that re¬quire selective mining, ore areas where weak wall con¬ditions exist, and ore bodies that have an ore value that will carry this relatively expensive mining method. Blast¬hole stoping, shrinkage stoping, and other mining meth¬ods that do not employ rock passes in a stope are not efficient in plunging or flatly dipping ore bodies because the footwall makes ore removal quite difficult. Since mining is accomplished by taking down slices of the back, only small areas of the wall rock are exposed at any one time, and these only for short periods. Due to limited back height, areas of uneconomic rock may be left in place, or they may be mined and the material gobbed in the stope. Because the miners in the stope must work under freshly blasted areas, the amount of ground control is usually great. The volume of rock that is broken during one section of mining is relatively small and the amount of nonproductive work required is high. This results in limited productivity for the scope and, be¬cause of the nonproductive work that must be done on a regular basis, the production from the stope may be quite cyclical. SUITABLE ORE BODIES The open cut-and-fill method may be adapted to al¬most any type of ore body with a relatively high vertical extent. The ore body must be accessible at both top and bottom as well as at regular intervals throughout its vertical extent. Although adaptable to most ore bodies, the method is probably best employed where the ore has poor con¬tinuity and where most types of bulk mining would pro¬duce excessive dilution. In areas of poor ore continuity, the capability of continuous and extensive sampling dur¬ing the mining of each cycle makes this method very desirable. This capability also minimizes the amount of evaluation sampling that must be done before mining is started. Because of the extractive system used, the size and shape of the stope may be as readily changed as the sampling mandates. Probably the only ore characteristic demanded is that the ore has strength enough to be sup¬ported through the use of rockbolts or cable bolts dur¬ing the mining and backfilling cycles. Good planning, systematic sampling, and careful supervision will pro¬duce a product with less dilution than any other open stoping method. PLANNING Evaluation Planning Once it has been decided that the open cut-and-fill method would be the most efficient for mining a par¬ticular ore body, the next considerations would probably be the availability of an economical backfill material and the selection of an efficient transport system for this material. Although hydraulically transported mill tail¬ings are the most widely used product, this is not always practical due to mill location or the quality of the tailings. In such cases, backfilling may be used. The type of backfill and the type of equipment used will determine if a floor or capping on the backfill is required to minimize dilution during ore removal. The early selection of rock removal equipment is im¬portant since equipment usually determines the amount of development work required to bring a stope into pro¬duction and the size of the openings needed. The size and continuity of the ore body will usually determine the type of loading equipment. The use of slushers or load¬haul-dump (LHD) equipment captive in the stopes will minimize the amount of development. If the ore con¬tinuity is such that a ramp system for extraction can be used, the cost of development will be increased but the flexibility of continuous mining will minimize the cycli¬cal nature of the production. The height of the mining section usually is deter¬mined by the strength of the wall rock and the amount of back bolting required. Once this has been decided, the appropriate drilling equipment can be chosen. The number and sizes of the rock passes employed depends upon the type and size of the extractive equipment and the type of backfill that is to be used. Since the miners must enter and leave the stope each shift, the level inter¬val is usually maintained at approximately 45 m (150 ft). Access from the level above into the stope must be main¬tained at all times. The employees perform all the work in the stope and adequate ventilation must be provided. Stope Planning Due to the flexibility of the method and the vari¬ability of the ore zones, layout is usually done on a stope basis. In areas where continuity is a problem, the size of the stope is usually determined by the boundary of the ore (with all of the ore within that boundary being removed). In areas of good continuity where ramps are to be used, the length of the stope may be determined by the length of time each of the cycles (preparation, back¬fill, mining, and ore extraction) requires. The ramp work is then laid out so that access to the various parts

Jan 1, 1982

By Paul Girodo

Government Approval For Foreign Workers 1.Service Canada Work Permit Tasks: i. Prove there are no Canadian Journeymen; ii. Use heavyweight stakeholders; iii. Understand the FWU transformation taking place today. Government Approval For Foreign Workers 2.Provincial Nominee Program (PNP) i. Build relationships; ii. Use heavyweight stakeholders; iii. Continually demonstrate benefits to the Province from foreign Journeymen. Government Approval For Foreign Workers 3.Letter from Union Supporting Temporary Foreign Workers i. Easiest path to approval ?CLAC model

May 1, 2006

By F. S. Norcross

MANGANESE has long been considered one of the United States' most important strategic raw materials. Its indispensability in steel manufacture makes it vital to the nation's industrial life. Coupled with this indispensability is the fact that domestic producers have not been able as yet to supply more than a small part of the nation's manganese needs; consequently the United States has had to depend upon such distant foreign sources as the U.S.S.R., British India, the African Gold Coast and Brazil for the major portion of its ore. The great bulk of the mineral is consumed in the form of ferromanganese, an alloy that averages 8o per cent manganese and 20 per cent mainly iron and carbon. Manganese is used as a deoxidizer and desulphurizer in steel manufacture. By combination with the residual oxygen and sulphur of the bath, it helps to produce a clean, sound metal. Approximately 14 lb. of manganese, it is estimated, is used in making one ton of steel. For standard-grade ferromanganese, the ore should have a minimum of about 48 per cent Mn and preferably more. Beneficiation methods have not yet been developed for making the immense low-grade deposits in the United States salable at prevailing prices. Dependence upon foreign sources thrust the United States into serious difficulties during the World War. At the start of hostilities in 1914, the nation was obtaining about four-fifths of its ore from the Eastern Hemisphere. Domestic production had been negligible, amounting to only 3000 to 7000 tons a year. As the war continued, shipments from the East, endangered by sea attack and blockade, dwindled rapidly. Imports from Brazil mounted. In this grave emergency, the United States, starting almost from scratch and under the stimulus of high prices, succeeded by 1918 in producing a peak of 311,000 tons, about 3 5 per cent of its consumption that year. Since the war, the United States has depended upon foreign sources for more than nine-tenths of its high-grade ore. During 1936, 1937 and 1938, imports for consumption averaged 736,000 tons a year. During the same period, domestic shipments averaged 21,000 tons a year. Of the imports, the U.S.S.R. accounted for an annual average of 280,000 tons, the Gold Coast 208,000 tons, India 74,000 tons, and Brazil 72,500 tons. But India and Brazil, which in 1936 ranked third and fourth in the list of importers, were displaced in 1937 and 1938 by a new and much nearer source-Cuba. From less than 1 per cent of United States imports in 1931, the island republic increased its shipments to more than 27 per cent in 1938. Cuba's average for this three-year period was 97,000 tons, and its total for 1938 was 131,000 tons.

Jan 1, 1940