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By Robert B. Parsons

"It is quite likely that more than a handful of junior exploration companies will receive some startling news from their accountants this spring when 1989s tax returns are being prepared. These companies could find that, for the first time ever, they will have to write a cheque to the federal government on account of a new federal capital tax which became effective July I, 1989IntroductionMany mining companies which have never had to pay a federal income tax are finding that they indeed have a tax liability for 1989. Non-producing companies, such as junior exploration companies, almost automatically assume that they don't have to pay any income tax. But, a new tax imposed under the Income Tax Act catches especially the company that has no income.Effective July 1, 1989, a federal capital tax (some call it a corporation minimum tax) is payable by any corporate carrying on business in Canada where the corporation has ""taxable capital"" exceeding $10 million. As a result of the popular use of flow-through shares in recent years, many junior exploration companies are finding that their balance sheets show capital stock over $10 million, and that they are consequently caught by the new tax. Further, some of these companies are finding, to their chagrin, that the tax for 1989 was due on February 28, 1990, and that non-deductible interest is ticking away on unpaid taxes."

Jan 1, 1990

By David R. Jessey

An ore microscopic and geochemical study of the nickel-cobalt occurrence in the Southeast Missouri Mining District has revealed several significant relationships. Siegenite [(Ni,Co)3S4] is the dominant nickel-cobalt sulfide in the ores, however, several other members of the thiospi¬nel group (AB2S4) occur: carrollite (CuCo2S4), nickelean carrollite [(Co,Ni,Cu)3S4], fletcherite [Cu(Ni,Co)2S4] and polydymite (Ni 3S4). In addition, bravoite [(Ni,Fe)S2], vaesite (NiS2), millerite (NiS) and gersdorffite (NiAsS) are present. Thiospinel composition varies dramatically within the district. Paragenetic studies have shown that the stratigraphically lowest thiospinels are the most cobalt-copper-rich. A detailed ore microscopic study of the AMAX Lead Company Buick mine has also shown significant changes in mineralogical association, texture and grain size of the siegenite along the north-south strike of the major ore bodies. Siegenite commonly replaces and veins chalcopyrite and is subsequently replaced by sphalerite. Less commonly, siegenite and chalcopyrite are observed to form linear replacements along sphalerite grain boundaries. The grain size of the siegenite varies markedly, but in a predictable north-south pattern that is consistent with a genetic model formulated from geo-chemical data. The compositional as well as textural trends displayed by the thiospinel mineralization could be exploited to optimize recovery of any or all thiospinels of economic interest.

Jan 1, 1981

By E. G. Spilsbury

AS a fresh illustration of the advances made in the past few years in the overhead-cable systems of hoisting and conveying, as applied to open-cut deep mining and quarrying, I would call attention to a further development of this system of transportation which has lately been put in operation i-n this State. A few miles above Williamsport, on the Susquehanna river, the Glen Union Lumber Company, of Pottsville, Pa., owned a large tract of valuable timber-

Jan 1, 1892

By M. Li, W. R. Reed, J. P. Rider, J. D. Potts

"The National Institute for Occupational Safety and Health (NIOSH) has recently developed a series of validated models utilizing computational fluid dynamics (CFD) to study the effects of air-blocking shelves on airflows and respirable dust distribution associated with medium-sized surface blasthole drill shrouds as part of a dry dust collector system. Using validated CFD models, three different air-blocking shelves were included in this study: a 15.2-cm (6-in.) wide shelf; a 7.6-cm (3-in.) wide shelf; and a 7.6-cm (3-in.) wide shelf at four different shelf heights. In addition, the dust-collector-to-bailing airflow ratios of 1.75:1, 1.5:1, 1.25:1, and 1:1 were evaluated for the 15.2-cm (6-in.) wide air-blocking shelf. This paper describes the methodology used to develop the CFD models. The effects of air-blocking shelves and dust-collector-to-bailing airflow ratios were identified by the study, and problem regions were revealed under certain conditions. The simulation results could be used for future development of dust control methods, particularly to address respirable silica dust, for a surface mine blasthole drill shroud.INTRODUCTION Surface blasthole drilling can generate considerable amounts of respirable silica dust. These high dust concentrations can be exacerbated by the high silica content as drilling cuts through silica-bearing materials such as sandstone and shale, causing overexposures for miners to respirable silica dust. These overexposures can lead to silicosis, an occupational lung disease that has no cure and is often fatal. A review of the Mine Safety and Health Administration (MSHA) respirable silica dust sample database from 2010 to 2016 for metal/nonmetal mining shows the following overexposure rates for occupations related to blasthole drilling: • 14.1% of rotary air drillers were overexposed to respirable silica dust (22 out of 156 samples) • 5.8% of rotary drillers were overexposed to respirable silica dust (9 out of 155 samples) • 7.3% of drill helpers were overexposed to respirable silica dust (3 out of 41 samples) • 7.5% of blaster/shotfirer were overexposed to respirable silica dust (3 out of 40 samples) Of significant interest is an x-ray surveillance study recently completed under the NIOSH Enhanced Coal Workers’ Health Surveillance Program (ECWHSP). This program offered free chest radiographs to surface coal miners in 16 states. Pneumoconiosis was found among 2.0% of the 2,328 screened miners with at least 1 year of mining tenure. Twelve miners had radiographic changes consistent with progressive massive fibrosis, and 9 of these 12 miners reported no underground mining tenure. All but 1 of the 9 miners worked at occupations that were in the vicinity of blasthole drilling, either as a driller, blaster, or blast crew member [1]."

Jan 1, 2018

By M. Li, W. R. Reed, J. P. Rider, J. D. Potts

The National Institute for Occupational Safety and Health (NIOSH) has recently developed a series of validated models utilizing computational fluid dynamics (CFD) to study the effects of air-blocking shelves on airflows and respirable dust distribution associated with medium-sized surface blasthole drill shrouds as part of a dry dust collector system. Using validated CFD models, three different air-blocking shelves were included in this study: a 15.2-cm (6-in.) wide shelf; a 7.6-cm (3-in.) wide shelf; and a 7.6-cm (3-in.) wide shelf at four different shelf heights. In addition, the dust-collector-to-bailing airflow ratios of 1.75:1, 1.5:1, 1.25:1, and 1:1 were evaluated for the 15.2-cm (6-in.) wide air-blocking shelf. This paper describes the methodology used to develop the CFD models. The effects of air-blocking shelves and dust-collector-to-bailing airflow ratios were identified by the study, and problem regions were revealed under certain conditions. The simulation results could be used for future development of dust control methods, particularly to address respirable silica dust, for a surface mine blasthole drill shroud. INTRODUCTION Surface blasthole drilling can generate considerable amounts of respirable silica dust. These high dust concentrations can be exacerbated by the high silica content as drilling cuts through silica-bearing materials such as sandstone and shale, causing overexposures for miners to respirable silica dust. These overexposures can lead to silicosis, an occupational lung disease that has no cure and is often fatal. A review of the Mine Safety and Health Administration (MSHA) respirable silica dust sample database from 2010 to 2016 for metal/nonmetal mining shows the following overexposure rates for occupations related to blasthole drilling: • 14.1% of rotary air drillers were overexposed to respirable silica dust (22 out of 156 samples) • 5.8% of rotary drillers were overexposed to respirable silica dust (9 out of 155 samples) • 7.3% of drill helpers were overexposed to respirable silica dust (3 out of 41 samples) • 7.5% of blaster/shotfirer were overexposed to respirable silica dust (3 out of 40 samples) Of significant interest is an x-ray surveillance study recently completed under the NIOSH Enhanced Coal Workers’ Health Surveillance Program (ECWHSP). This program offered free chest radiographs to surface coal miners in 16 states. Pneumoconiosis was found among 2.0% of the 2,328 screened miners with at least 1 year of mining tenure. Twelve miners had radiographic changes consistent with progressive massive fibrosis, and 9 of these 12 miners reported no underground mining tenure. All but 1 of the 9 miners worked at occupations that were in the vicinity of blasthole drilling, either as a driller, blaster, or blast crew member [1].

Jan 1, 2019

By Claude M. Moreland

It is common practice when thermally drying coal from wet coal preparation facilities that the coarse +0.5mm fractions and fine fractions (-0.5mm) are combined and fed to a thermal dryer. There are very good reasons for this concept. The coal feed is much easier to handle in this condition for it is not as likely to stick inside the dryer or hang up in the material handling equipment, and is easier to feed and discharge. However, it is generally only the -0.5mm fines that really need to be dried. The +9.5m fraction of the cleaned coal will leave the dewatering screens, or drain and rinse screens, at an acceptable surface moisture, probably 5 or 5 ½ percent. The 9.5m x 0.5mm mesh fraction can be centrifuged to about 8 percent. The -0.5m fines can range from 18 to 22 percent or more, depending upon the type of equipment employed to dewater, and upon the size distribution within this size range.

Jan 1, 1982

By John W. Pack

At present the manufacture of metallic zinc, or spelter, in Missouri, is carried on only at the establishments located at Carondelet, or South St. Louis. Although the development of the industry has not yet reached that fulness which it has attained elsewhere in the United States, still the commanding advantages of position of these establishments are so evident that the prospective importance and magnitude of spelter manufacture in this State, and especially at this locality, can hardly be doubted. A study of the chemical and technical aspects of the business of these points cannot be without interest to the metallurgist as well as to the capitalist. I have endeavored to investigate, as far as was practicable, within the limits of a paper of this character, some of the problems connected with this industry, at the same time making a few comparisons of the results attained at some of the Carondelet works with those of other establishments with regard to which information could be obtained. In the prosecution of these studies I have been kindly aided by the managers of Carondelet establishments, and by Prof. Charles P. Williams, Director of Rolla School of Mines. To the latter I am especially indebted for assistance in regard to methods of analysis. To all these gentlemen I return my thanks for their valuable aid. Other sources of information will be duly credited in their proper places.

By Sam S. Taylor

[The discovery of many oil fields in Illinois since 1936 and their rapid development have brought tine problem of brine disposal to the foreground in that State. The ultimate magnitude of the is not yet known, but already it is great enough to warrant technical investigation. The Bureau of Mines, therefore, has made a general survey of the methods of handling oil-field brines in Illinois and has studied in detail three subsurfaces disposal systems.]

Jan 1, 1940

By S. E. Bretherton

Mr. AllaW Gibb, of Mount Perry, Queensland, Australia, in an interesting and instructive paper,* describes fully the great difficulties metallurgists encounter in seeking to produce marketable copper by smelting refractory ores containing arsenic, antimony and bismuth. After reading this description, I prepared a paper for the Institute, which was, however, through a misunderstanding on my part, published in the San Francisco Mining and S'cie~ztiJic Press of April 4, 1903, after its acceptance by the Council, and without credit to the Institute. Consequently, it cannot be reprinted for the Transactions of the Institute; and I can only here summarize its contents and conclusions, referring the reader to the article above named for further particulars. In that article I omitted the col~sideration of bismuth, which we do not encounter in our smelting at Val Verde, and added lead and zinc, which Mr. Gibb had not discussed; and I undertook to shorn, by the record of a year's practice of the Val Verde Copper Co., Arizona, that copper-ores containing arsenic, antimony, lead and zinc may be successfully smelted, without preliminary roasting, in an ordinary blast-furnace with hotblast, making a normal slag, such as one would aim to make under ordinary conditions with cold-blast, so as to produce a marketable matte, containing not less than 45 per cent. of copper, free from arsenic and antimony, and below the limit in lead and zinc penalized by the refineries.

Jan 1, 1904

By Hal D. Peterson

In research sponsored by the Bureau of Mines, a technique was developed for simulation of heap leaching on uranium ores. The equipment used for this simulation consisted of 2-ft-diam columns, constructed in sections, such that ores as coarse as minus 4-in could be treated in bed depths ranging from 8 to 18 ft with test condition variations. The technique consisted of the separate steps of ore preparation, column loading, charge wetting, leaching, washing, and residue recovery. For the ores studied, sulfuric acid was found to be generally superior to alkaline lixiviants. The rate of uranium extraction by acid leaching proved" to be primarily a function of percolation rate, alkalinity of the ores and oxidant requirements of the ores. Column leaching was closely comparable to field-scale heap leaching in terms of uranium extraction rate and recovery results. Acid and oxidant requirements were lower for the field-scale operation. All column tests, as well as the field-scaly heap, achieved in excess of 96-pct uranium extraction on a Bear Creek, Wyoming ore.

Jan 1, 1985

By Hongbo Zhu, Jinhui Peng, Linqing Dai

"To optimize the process variables for making direct reduced iron (DRI) by microwave heating, the reduction temperature, reduction time and the ratio of coal to material were studied with the central composite design (CCD) and their interactions on the Metallization rate were also investigated. The predictive polynomial quadratic equations model was analyzed by ANOVA. Optimal conditions of making DRI can be concluded as follows: 28 min at 1139?, the ratio of coal to material is 20.95%. Under these conditions, the metallization rate is up to 97.06%.1. IntroductionThere are some technical problems in making direct reduced iron (DRI) or sponge iron using iron ore concentrates containing coal and lime. In reduction of the iron ore concentrates containing coal with microwave heating, the strength of pellets needn’t to be considered. It reduces the costs and improve quality of materials, but also eliminates the need for pellet preparation process, simplifying the process and achieving a cleaner production.Microwave energy is a clean energy, and has characteristic of selective heating. The speed of heating is also fast. It researched by Chen Jin that when the mole mass of C: O is 0.627 and the microwave power is 15KW, it only spends 15min from 15? to 1021.6? with 1 kg iron ore containing coal with microwave heating [1]. There are many factors affecting metallization rate. In order to find a suitable condition for reduction process, it is necessary to design a series of suitable experiments.Response surface methodology (RSM) is one of the relevant multivariate techniques which can deal with multivariant experimental design strategy, statistical modeling and process optimization [2–5]. It is used to examine the relationship between one or more response variables and a set of quantitative experimental variables or factors. This method is often employed after the identification of the vital controllable factors to find the factor settings that optimize the response. Designs of this type are usually chosen when a curvature in the response surface is suspected. The process optimization of the microwave reduced iron ore was seldom reported in literature. Hence the present work intends to assess the effects of variables such as reduction time, reduction temperature and the ratio of coal to identify the optimum conditions using a central composite design (CCD). The characteristic of product was assessed using scanning electron microscope (SEM)."

Jan 1, 2011

By Larissa Koroznikova, Steve Durkin, Patrick Bouwmeester, Michael Tuck

In currently used mining methods, blasting techniques often causes dilution of valuable ore and results in costly processing requirements. In the context of narrow vein mining of thin and highly concentrated orebodies there is a need of a mining method that can reduce dilution in order to remain economically viable. This research project explored the viability of a new mining technology, thermal fragmentation, in narrow vein mining. Thermal fragmentation technology uses a flame jet to produce extreme heat that spalls the surrounding rock to a strategically located drill hole, as an alternative to traditional blasting. This paper creates a net present value (NPV) model of a mining method using thermal fragmentation, as well as for an existing method used for narrow vein mining; comparisons and evaluations were made regarding the feasibility of the new technology. It was found that while overall costs for thermal fragmentation were relatively high, reductions in wages, haulage and processing costs, as well as increased productivity and ore recovery, meant that the new method would improve the financial performance of any operation. These results identify that there is an opportunity to introduce the thermal fragmentation technology to narrow vein mines within Australia, in order to lower costs and increase profit.

By A J. Moore

Blasting operations in open cut mines results in airblast, ground vibration and rock movement with the potential for wild flyrock. In some mines there is a trend towards blasting closer to established infrastructure as extraction boundaries alter with increased resource prices and changed economic circumstances. Mine infrastructure in close proximity to blasting has the potential to be damaged by ground vibration and wild flyrock especially. To provide guidance for the determination of appropriate buffer distances and blast specifications, this paper describes methodologies that may be used to determine the potential for damage to infrastructure by predicting:dynamic blast;near field ground vibration levels;dynamic blast strains induced in infrastructure by consideration of ground vibration,ground strain and response amplification;potential for damage by comparing induced strains with allowable working strainsfrom structure design codes and serviceability requirements; andflyrock throw distances based on models developed from flyrock investigations.This report demonstrates the use of strain prediction methodologies developed during Australian Coal Association Research Programme (ACARP) Project C14057 for infrastructure such as conveyor belts, transmission poles and towers, buildings and bridges.

Nov 8, 2011

By Ralph S. Mattson, Frederick T. Hancock

GENERAL DESCRIPTION The Granduc mine is situated on the west side of Granduc Mountain in the Leduc River area of north¬west British Columbia about 51 km (32 miles) north¬east of the town of Stewart. Entry is by means of a 16.6-km (10.3-mile) long haulage adit from the concentrator site at Tide Lake. The principal ore mineral is chalcopyrite. Ore mineralization varies from dis¬seminated to irregular semimassive stringers generally associated with highly folded and sheared biotite-rich sections of metasediments. The ore bodies vary in width from 3.6 m (12 ft) to over 18 m (60 ft) in a zone some 800 m (2600 ft) in length dipping at an average of 75° to the west and plunging steeply to the south. Capping varies in thickness from 0 to over 300 m (1000 ft). SELECTION OF MINING METHOD In the early stages of planning the primary require¬ments to be met by the mining method were: (1) suit¬ability to high output and productivity, (2) adaptability to mechanization in order to minimize operating costs, and (3) minimization of the costs and the development period for preproduction. Considering these, the selection of a mining method was narrowed to a choice be¬tween sublevel long-hole stoping or sublevel caving. A method utilizing backfill was ruled out as uneconomical, especially since mill tailings would not be readily available. At the time, the following facts about the ore body were recognized: (1) the average width of the ore body was 12 m (40 ft), (2) 30% of the stoping blocks were in areas of 6 m (20 ft) or less in width; (3) in many cases, ore zones were located in parallel lenses or were irregular in shape; (4) major faults were present in the hanging wall close to the ore, with minor faults within the different ore zones; and (5) the ore itself was lami¬nated with fractures and jointing crossing such bedding planes. Considering these factors, the two stoping meth¬ods were compared. Sublevel Open Stoping The characteristics of sublevel open stoping are: 1) It is generally not used for widths below 6 m (20 ft) since the ratio of waste development increases as level intervals are reduced to insure good control of long-hole drilling. 2) This method is used where ground is relatively competent. Otherwise more pillar support is required, increasing ore losses or mining costs with pillar recovery. 3) The method is not easily adaptable to parallel ore zones separated by relatively narrow waste bands. 4) Close control of pillar and stope design will per¬mit some degree of dilution control. Sublevel Caving Sublevel caving is characterized by the following: 1) The method can be used for a variety of ore widths, recognizing that 3 m (10 ft) in width may be an economical minimum. 2) It is particularly favorable where wall rock is weak. 3) The method has flexibility so that irregular ore bodies or parallel lenses can be mined. 4) The need to leave pillars for support is eliminated 5) Dilution and tonnage recovery are interrelated. With concern about the stability of the hanging wall and the geometry of the ore bodies, there was a growing favor to choose the sublevel caving method. However, this method was relatively new to Canadian mining practice. Teams were sent to Sweden to tour mines using sublevel caving and to study their methods and any extraction and dilution problems. The objective was to develop cost and performance data and ore recovery grades and tonnages for feasibility comparisons. Next, detailed mine plans were developed covering a three-year period using sublevel caving and sublevel open stoping. Extensive analyses of the development and production costs for each method were made. With sublevel caving, ore losses of 10% with a 20% dilution rate were used. The same ore loss was estimated for sublevel open stoping with dilution varying from 11 to 15%. As a result of these studies, the decision was made to use sublevel caving partly because slightly lower op¬erating costs, together with higher productivity, offset the adverse effects of higher dilution; and partly be¬cause, at the start, there would be less risk in being able to handle the ground, and the method offered more flexi¬bility. Open stoping was not ruled out for testing at a later date. It was recognized that some other method of mining would be required for mining the ore extensions below the glacier, where surface subsidence must be avoided. During the operating life of the mine, open stoping was tested in narrow width ore zones and found un¬satisfactory due to wall failure. A cut-and-fill stope, utilizing run-of-mine waste from development, was also operated on a trial basis. A sequence was finally estab¬lished that permitted ground control; however, produc¬tivity needed considerable improvement in order for open stoping to be considered as a sole alternative for mining below the main haulage level. At the termination of operations in June 1978 over 13 000 000 t (14,500,000 st) had been extracted from above the main haulage level at production rates varying from 3270 to 7250 t/d (4000 to 8000 stpd). Sublevel Caving Methods Two basic methods were used, transverse and longi¬tudinal. The transverse method was used mainly in the upper C ore body in widths greater than 18 m (60 ft), and occasionally as an alternative to the longitudinal method in localized areas where strike faulting inter¬fered with normal development. The longitudinal method was used in ore bodies varying in width from 3.6 to 18 m (12 to 60 ft), with a multilongitudinal varia

Jan 1, 1982

By C. E. Drennan

Your Chairman and the Program Committee have limited the time allotted to me, and so, with your permission, I will omit the preliminaries, and wade right into my subject. I believe I am safe in taking for granted that you are all agreed that electricity is the most feasible means of transmitting the power used in coal mining. If you are not, then the papers read before the Association in previous years have failed of their object. If you are so convinced, the only question is as to the choice between alternating and direct current, and the voltage. Fundamentally, this is primarily a matter of dollars and cents. With the sole exception of your trolley locomotives, there is no part of your equipment that cannot be operated as well on alternating as on direct current. With your pay rolls at their present high level and the efficiency of the labor you employ at the stage it now occupies, I believe you will agree, first, that everything possible must be done to increase the tons of coal produced per man employed; and, second, that ample power at all times must be available at every machine so that every one of your men can really get out a day's work without any avoidable delay. Speaking Electrically, all this means is a machine instead of a man wherever it is possible for a machine to do a man's work, and full voltage at the machine always, so that there can be no excuse for the work not being done promptly and efficiently-and dependably. The real reason for the selection of one kind of current or the other can best be understood by considering the formula which determines the size of wire required for transmitting electrical energy. The size of the copper required, and, consequently; its weight and cost, depends upon four factors: 1. The amount of energy or power to be transmitted. 2. The distance to be covered. 3. The percentage of loss allowed. 4. The voltage at the receiving end of the line. These factors enter into the problem to different extents, and, briefly, the size of wire depends directly upon the amount of power and the distance transmitted, and inversely upon the percentage of loss allowed and the square of the voltage. From this it will be seen that, with the amount of power and the distance fixed, if the loss is to be cut in two, the wire must be doubled with a given voltage, while if the loss remains constant and the voltage is doubled, the wire may be cut in four; that is, reduced 75 per cent in weight and cost. Now, standard motors are designed and recommended by their builders to run within a range of 10 per cent under voltage to 10 per cent above voltage - 440-volt motor, then, should preferably not be operated on less than 400 volts

Jan 1, 1923

By R. Lapcevic, P. Lokin

"Among other components, ore body „T“ was discovered in the Bor pit copper mine at a depth of app. 500m. The ore body is relatively small: 60-65m long, 40-45m wide and up to 50m high, but with an extremely high content of copper and accompanying metals. The rock mass containing the ore body is composed of altered andesite of relatively poor physical-mechanical characteristics. The exploitation of this ore body by way of conventional mining methods was associated with high risk, cost, and potential losses of the rich ore. Due to everything specified hereinbefore, the adopted solution involved ore exploitation with complete ensuring of excavation stability, thus forming an extremely large underground opening. The excavation and supporting were performed according to the NATM principles along with applying shotcrete, 9-25 m long cables, and horizontal and vertical lattice girders. The excavation of the ore body was completed in app. 2.5 years, all the while ensuring complete stability and security of people and equipment; construction works were constantly monitored. The proposed solution for ore excavation enabled more than 95% exploitation of ore as requested by the Investor, while depletion due to waste-rock excavation was less than the allowed 5%. INTRODUCTIONThe Bor copper deposit is located in Eastern Serbia. Copper exploitation in Bor commenced as far back as 1902. A total of 26 ore bodies varying in size and valuable mineral content have been excavated so far. One of the most recent excavated ore bodies is ore body „T“, which is one of the smallest bodies by volume, but also a body with one of the largest content of copper and accompanying metals. An important fact regarding its exploitation is that it is located at a depth of app. 500 m; that its shape is irregular: 60-65 m long; 40-45 m wide and app. 50 m high; and that it lays within a rock mass of relatively poor physical-mechanical characteristics. In addition, above it, in its immediate vicinity, ore body „F“ was previously excavated and its room was refilled with backfill and water. Excavation methods common in mining practice (excavation with caving, excavation with backfilling, excavation with protective pillars) were inadequate in this case for various reasons:• The immediate vicinity of the previous ore body „F“ room could cause expansion of the caving zone also to that room, along with penetration of groundwater and flotation sludge;• Possibility of deformation expansion caused by caving as far up as the surface of the terrain, including the urban area;• Difficulties in maintaining the stability of excavation sidewalls having steep slopes and great height;• Significant loss of ore mass with a relatively high content of metals etc."

Jan 1, 2015

By E. J. Lintner

This paper is one of a series being prepared by the United States Bureau of Mines describing cloy reining, crushing, and grinding methods and costs at various operations throughout the United States. These Toners are designed to disseminate technical information regarding methods used. The cost tabulations represent local operating expenditures only and not total production casts. Although the publication of total production cants might not be advisable in sore instances, yet a knowledge of operating costs is essential to the technical discussion end study of the methods employed. The attention of the reader is specifically coiled to this differentiation in order that no misunderstanding of the scope of the cost tabulations shall ensue. The following report deals specifically with the methods employed by the Clay City Fire Co., Uhrichsville, Tuscarawas County, Ohio.

Jan 1, 1936

By R. S. Dean

[Methods s of preparing the manganese metal that has been used in chemical and metallurgical investigations have passed through a cycle. Most of the investigations made before the turn of the century, used Moissen's pyrophoric manganese.3/ This material was made by electrolysis into a mercury cathode and distillation in vacuo. The material was essentially pure, and investigations in which it was employed (with proper precautions against oxidation) led to accurate results on tie intrinsic behavior of metallic manganese.]

Jan 1, 1941

By M. S. Seltzer

In a previous paper1 it was shown that activation energies for steady-state creep in lead sulfide single crystals varied with the concentration of electronic defects. For n-type lead-excess crystals, values for the creep activation energy, Qc, decreased from 2.9 to 2.2 ev as the carrier concentration increased from 1017 to 1017 electrons per cu cm, while Q, increased from 1.6 to 2.3 ev for p-type sulfur-excess crystals as the concentration of holes increased from 6 X 1017 to 1018 per cu cm. It was also shown in Ref. 1 that the Q, values could best be correlated with the sum of the self-diffusion activation energies in PbS, i.e., for a given concentration of electrons or holes Q, = QPb + Qs, where Qpb and QS are the activation energies for diffusion of pb210 and s35 in PbS. This result suggests that creep rates are controlled by at least two atomic defects, one for lead migration and one controlling sulfur transport. It would be most desirable to interpret these creep energies in terms of an atomic defect disorder model for lead sulfide. Studies of creep rates as a function of sulfur pressure2 indicate that the defects controlling creep rates are singly ionized lead and sulfur vacancies, VPb and Vs. This hypothesis can be more firmly established by determination of Q, under conditions of constant atomic defect concentration rather than constant electronic defect concentration as has been previously done. Therefore, several measurements have been made of Q, on crystals whose con- centration of lead vacancies or sulfur vacancies has been fixed by equilibration under appropriate sulfur pressure and temperature. It was indicated in Ref. 2 that high-temperature isotherms are available which show how the concentrations of various atomic point defects in PbS vary with composition over the entire homogeneity range for the compound. Two such isotherms are shown in Fig. 1 for 900" and 1000°K. Utilizing these isotherms the concentration of a given defect, say Vbb, may be held constant over a particular temperature range by adjusting the sulfur pressure as the temperature is changed. Thus the concentration of one defect suspected of controlling creep rates may be held constant while other defect concentrations vary. The apparent activation energy for formation of these defects can be obtained from the isotherms giving their concentration as a function of sulfur pressure. Activation energies for creep obtained under conditions of constant defect concentration can then be compared with defect formation energies and certain possible atomic defects can be eliminated as creep-controlling species. We have performed creep experiments as a function of temperature at a stress of 244 g per sq mm under conditions which fixed the lead vacancy concentration, V', at 1016, 3.2 x 1016, and 3.2 x 1017 cm-3, and one were the sulfur vacancy concentration, Vs, was held at lo17 cm-3. The results are presented in Fig. 2 as creep rate, k, vs reciprocal of absolute temperature, 1/T. Specimens 34, 35, and 36 were deformed under constant lead vacancy concentration while specimen

Jan 1, 1969

By C. S. Kucukkargoz, A. Bhalla

"The solid-state reduction of ilmenite concentrate with graphite was studied under argon atmosphere at temperatures between 1050°C and 1350°C. The rate of reduction increased with increasing temperature and decreasing particle size, with the maximum extent of reduction found to be 80% at 1350oC with a particle size in the range of -106 +90 µm. The activation energy for the reduction reaction was calculated to be 257 kJ/mole. The trivalent iron oxide was reduced first to divalent iron oxide and then to metallic iron. An annular ring of a partially reduced oxide phase containing metallic nuclei was formed around the core of ilmenite, which grew in size as the original ilmenite core shrank with progressive reduction. The reduction of titanium dioxide also took place during this stage, which produced the Magneli oxides in the form of TinO2n-1. The metallic phase had very high Fe and low Ti contents throughout the reduction, whereas the partially reduced oxide phase adjacent to the metallic phase was high in Ti and depleted in Fe. The metallic phase globules formed during this stage were small and are considered suitable only for leaching and not for other concentration processes. The core of the particle generally retained the composition of the original ilmenite. The purpose of this work was to remove iron and increase the titanium oxide content of the ilmenite to produce a high-titanium containing charge for subsequent smelting.IntroductionThe titanium-bearing minerals ilmenite (FeTiO3) and rutile (TiO2) are the primary natural raw materials used in industry to produce titanium metal and titanium-based products. The common titanium-bearing minerals are ilmenite, rutile, anatase, brookite, leucoxene, arizonite, and perovskite. Although the formula of ilmenite is commonly expressed as FeTiO3, the mineral can accommodate Mn and Mg, which are chemically bound in the form of (Fe,Mn,Mg)TiO3. Ilmenite concentrate can also contain Fe2O3 and Al2O3 in addition to the physically separable oxides such as SiO2 and CaO. The crystal structure of ilmenite is an ordered derivative of the corundum structure with a trigonal-rhombohedral symmetry. Fe2+ and Ti4+ ions occupy the octahedral sites, leaving the tetrahedral sites unoccupied. In each FeTiO3 unit, there are three O2- ions creating three octahedral sites, two of which are occupied by Fe2+ and Ti4+ (Wilson et al., 2005)."

Jan 1, 2017