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By Raimison Bezerra de Assis

Ferritas são compostos cerâmicos que apresentam propriedades magnéticas, sendo estas dependentes da sua estrutura cristalina. Dentre as ferritas, a do tipo Ni-Zn é uma das mais promissoras, em virtude da microestrutura e das propriedades magnéticas deste material. Este trabalho teve como objetivo analisar a microestrutura e propriedades magnéticas da ferrita Ni-Zn dopada com íons (TR) Sm3+. A ferrita Ni-Zn-Sm foi obtida pelo método dos precursores poliméricos, a partir da dissolução do ácido cítrico, nitrato de ferro, nitrato de níquel, nitrato de zinco, nitrato de samário e etilenoglicol. A resina obtida foi pirolisada a 350°C/2h com taxa de aquecimento de 10°C/min. e calcinada a 650°C/3h com taxa de aquecimento de 10°C/min. O material resultante foi caracterizado por difração de raios-X (DRX), microscopia eletrônica de varredura (MEV) e magnetometria de amostra vibrante (VSM). Os resultados mostraram a formação de fase única da ferrita Ni-Zn-Sm, com partículas aglomeradas e características magnéticas de material macio "soft", com baixa coercividade (0,189 kOe) e magnetização de saturação (47,699 emu/g).

Aug 8, 2018

CONCLUSIONS: Our experience, where we attempt to remove a continuous face with a seam more than six and one-half feet in thickness, has been rather disastrous. We have been able to open up faces and retreat for a distance of 75 to 100 feet in this particular character of coal, but after the supporting timbers have been removed for the second or third time we have found that the roof has a tendency to ride over the props and break off at the face of the coal support. However, where our coal is below six and one-half feet in thickness we have had very good success in handling the roof and breaking it off systematically with our break line of props, which are usually from 12 to 15 feet from the face of the coal. Our practice in to set props at about two-foot intervals in each direction and remove two sets of props, at the same time two additional sets are placed along the coal face. At all times we are very insistent that all props be removed from the gob and that entire coal be taken out so that it is possible for the roof to break freely and all come down, not having any supporting pillars in the gob. The conclusion of our experiments to date is that if all the supports beyond the breaking line of props are taken out as the long face retreats and the closer the breaking line of props is kept to the face the better the results will be. We have never felt it safe to have less than four rows of props to protect the machinery and men working along the coal face. We feel that the time element as to the rapidity with which these faces are advanced or retreated, is of vast importance, but to date have not had an opportunity or experience enough to reach any conclusion as to what time element is proper here with us. We feel that the crushing strength of the overlying strata will govern this to a large extent, and as ours varies considerably, it is going to he very difficult for us to reach a definite conclusion. My opinion is, that conditions may be such that rather than have a short face and retreat rapidly, it will be found that a longer facewith a slower retreat will work out better. Such a condition might give the overlying strata time to settle. The caving of roof or of control of this caving, with us, is quite closely associated with the ventilation. We find that the places where the strata overlying the coal are composed of shale, that the effect of the air on this is of very great importance. It is our endeavor, at all times, not only to remove obnoxious gases, but to keep a current of fresh air moving along the face and, of course, a portion of this travels over the gob and comes in contact with the roof and overlying strata. This, we find, helps materially in keeping the roof on the move. As it is our practice to keep water for sprinkling purposes at all places, we also use this water to keep the roof wet back in the gob. The effect of water on the shale is to cause it to disintegrate quite rapidly. MR. HOLMAN (Concluding): If there are any questions about this, I will be glad to answer them. I have some prints and also some pictures of the scraper loader and the shaking conveyors (distributes.) PRESIDENT MOSES: We wish to thank you, Mr. Holman, and also wish that you would thank Mr. Dupont for his paper. Are there any discussions? Do you wish to ask Mr. Holman any questions? (No response.) PRESIDENT MOSES: Mr. Holman, I would like to ask you how you take the photographs; were they taken by flashlight, or were globes used, or how were they taken, in a gaseous mine? MR. HOLMAN: Our mines being gaseous, we do not use the ordinary flashlight for taking underground photographs, but we have had considerable success, I think, with underground pictures by means of electric lights. We use 250 watt globes. We have six of these rigged up with tin reflectors, 18 inches in diameter, on standards, and we carry two or three hundred feet of wire connected with these six lights, and they can be turned on in the room where we take the photograph, by means of this extra wire. We set the kodak up and arrange the picture and take a long time-exposure. We usually take an

Jan 1, 1926

THE CHAIRMAN (J. W. MALCOLMSON, Kansas City, Mo.).-Gentlemen, Dr. P. J. O'Gara, Director of the Department of Agricultural and Smelter By-Products Investigations, of the American Smelting & Refining Co., Salt Lake City, Utah, is here; he has given attention to this very important problem and I will call on him for a few remarks. P. J. O'GARA, Salt Lake City, Utah.-I have had charge of the investigations of the smoke problem for the American Smelting and Refining Co. during the past four years. One of the worst burns we ever experienced occurred shortly after we began putting all our gases through the baghouse. This occurred in 1914 after the roaster gases had been put through the baghouse; previous to that time, the greater portion of the roaster gases was not put through the baghouse. Everyone knows why a hot-air balloon goes up. The hotter the air in the balloon, the higher it will go. In other words, hot gases will rise higher than cold gases. When gases are handled in any way that will tend to lower their temperatures, these gases are going to travel less high in the atmosphere and will therefore tend to reach the ground sooner than if they are turned out very hot. So far as injury to vegetable life is concerned, sulphur dioxide is the only element in smelter smoke that need be considered. The solution of the problem will be either to take out the excess of sulphur dioxide or to diffuse it high enough, up in the atmosphere so that the concentra-

Jan 1, 1918

By R Young

Independent Project Analysis (IPA) is in the business of benchmarking the performance of capital projects on behalf of owner companies in the processing and extractive industries. IPA assesses between 500 and 800 projects each year. The projects are conducted globally and range in size from about $1 million to over $30 billion. Using the data, we are able to benchmark the performance of Australian projects against the rest of the world. This presentation focuses on the performance of large (greater than $100 million) projects that involve chemical or thermal processing.We regularly hear calls for more processing to be conducted in Australia but on the other hand industry leaders are concerned about the poor success rate of solids processing projects in Australia. Data gathered by IPA indicates that there is good reason for the concern. Out of a sample of 31 large processing projects conducted in Australia, 23, or 74 percent, were classed as failures. This performance is significantly worse than for similar projects in other parts of the world. The presentation analyses why the performance is worse and demonstrates that many of the reasons behind the failures were manageable by the project team. Furthermore, members of the project teams on at least half the failed projects knew that the projects were on a path to failure before the project was authorised but could not be heard.The presentation discusses how large projects that involve process design and, in many cases, new technology, need to be set up for success. *This is an abstract only. No paper was prepared for this abstract.*

Aug 8, 2011

By A. Simkovich, R. W. Lindsay

The possibility of using sodium sulfide slags to remove copper from ferrous alloys has been investigated by Jordan1 and by Langenberg.2, 3 In these studies, such slags were determined to be capable of removing copper and sulfur from the melt. The present work represents additional effort to clarify the effects of temperature on copper removal. The experiments were performed in a 17-lb induction furnace. Graphite crucibles contained the melts and kept the baths saturated with carbon. Temperatures were measured with a calibrated optical pyrometer and were controlled by manipulation of power input to the furnace. Estimated accuracy of temperatures in this investigation is ± 10°C (18°F) for measurements prior to slag additions, and + 20°C (36°F) after slag formation. The procedure consisted of melting 800 g of electrolytic iron. During this step, powdered graphite covered the exposed iron surface. After a predetermined temperature was reached, copper shot was added. A sample of the molten alloy for chemical analysis was then aspirated into a silica sheath. Next, a slag-forming mixture of sodium sulfite and graphite was added instantaneously to the melt. The sodium sulfite amounted to one-tenth the charge weight of iron; sufficient graphite was added to combine with oxygen in the sodium sulfite, assuming formation of carbon monoxide and reduction of the sulfite to sulfide. Subsequent to the slag addition, the molten alloy was sampled periodically, with the exception of heat A in which no intervening samples were taken between the slag addition and the end of the run. The iron was poured into a graphite mold, and the ingots sectioned and drilled for samples. Results of selected heats are presented in Table I. Analyses of samples drawn from the iron prior to slag addition are listed under zero time. Two samples from heat D were reported with copper contents greater than the initial concentration in the bath. Owing to the gradual but complete disappearance of slag during this heat, it is believed copper momentarily became more concentrated in the upper portion of the bath while reverting from the slag. This is the region from which samples were drawn. It should be noted that analysis of the ingot was equal to the copper content at the time of slag addition. The terminal temperatures of heats D and E, and the initial sulfur content of heat A are also to be noted. Because of the large temperature drop which occurred when slag was formed in heat D, power input to the furnace was increased in heat E after the slag addition, causing a higher terminal temperature. In heat A, the initial sulfur concentration was relatively high as compared to heats B through E owing to contamination by some slag remaining in the crucible from a previous heat. It is evident from Table I that copper was removed at the onset of slag formation. Roughly 30 pct of the copper was taken into the slag, with the exception of heat D, which had approximately 50 pct removed. For a comparatively short time of slag-metal contact, it appears that no gain is to be made in copper removal through use of high or low temperatures. If the slag initially formed remains in contact with the iron for an extended period, temperature has a marked effect upon copper removal, as can be seen by studying results for the two extremes in temperature. At about 1425°C, the copper level remained relatively constant after the initial removal by the slag. However, in the region of 1670°C, a definite reversion of copper occurred. Reversion was incomplete in heat D, and complete in heat E. The final temperatures of heats D and E differed by about 75°C. This temperature difference is thought to be the reason for only partial copper reversion in heat D. It is believed the effects of temperature noted above are related to the evolution of a white fume, which appeared in every run except heat A. (In the case of heat A, the fume was practically indiscernible.) After each slag addition, a yellow flame formed for about 5 sec. When the flame subsided, a white fume appeared. Upon contact with surrounding cooler surfaces, this fume deposited as a white solid. In the experiments made at 1425°C, evolution of fume continued unchanged to the end of the runs. However, heats D and E exhibited a different behavior. A very noticeable decrease in fume evolution from heat D was observed. Furthermore, this heat had much less slag remaining than did runs A through C when the experiments were terminated. No slag remained at the end of heat E; evolution of fume from this heat ceased prior to pouring. Spec-trographic analysis of the white deposit indicated sodium to be the major metallic element, with the maximum concentration of iron and copper as 0.1 and 0.01 pct, respectively. It is supposed the white fume observed in these experiments is principally sodium oxide (Na2O), formed by oxidation of sodium in the slag and subsequent sublimation. (Sodium oxide is a white to gray substance in the solid state; at 1275oC, it sublimes.4) According to this mechanism, elevated temperatures would accelerate removal of sodium from the slag, sulfur pickup by the

Jan 1, 1961

By Parker JB

The Channar Iron Ore Mining Project (the Project) is a joint venture involving the Hamersley Holdings group (Hamersley) and the China Metallurgical Import and Export Corporation (CMIEC). The Project represents the first new iron ore mine development in Australia since 1971/72 and is the first major overseas mining investment by the People's Republic of China (PRC). The Channar mine is being constructed in the Pilbara region of Western Australia and will operate for more than 20 years supplying the PRC with 200 M tonne of high grade iron ore. The Project is configured to take advantage of Hamersley's existing production and support facilities. The Joint Venturers will invest A$280 million to achieve an initial production rate of 3 M tonne/yr and an additional A$120 million over the next 8 years to achieve full capacity of 10 M tonne/yr. Financing the Project to meet the requirements of both Hamersley and CMIEC, with the dependance of the Project on Hamersley's infrastructure, and in a depressed iron ore market, required an innovative financial structure to be arranged in a limited period. This paper describes both the financial structure and its implementation, and draws a number of conclusions regarding the financing of resource projects.

Jan 1, 1989

By W. J. Carlson, S. M. Parmerter, I. Iwasaki

This article discusses the effect of physical and chemical modifications of starches on the anionic and cationic flotation of silica from oxidized iron ores and magnetite-taconite concentrates. It also deals with the results of the interaction of starch, pH, and calcium ions on the flocculation, clarification, and filtration of iron ore slimes and magnetite-taconite tailings. Starches, particularly when an-ionically modified, were found to be effective depressants in anionic silica flotation. British gums and dextrins were beneficial for oxidized iron ores, but none of the starches or starch derivatives appeared to have any effect on magnetite-taconite concentrates. The flocculation and filtration of iron ore slimes were affected most strongly by the level of starch, to a lesser degree by the pH, and virtually not at all by the level of calcium chloride. The clarification of iron ore slimes was complexly dependent on these three factors. The flocculation of magnetite-taconite tailings depended mainly on the level of starch, whereas their clarification depended on the level of lime. As previously reported in the literature, selective flocculation and partial upgrading of a semitaconite could be readily attained with starch and calcium chloride. A previous article discussed some of the parameters that are important when various corn starches are applied to the soap flotation of activated silica from iron ores and to the flocculation of iron ore slimes.' Methods of solubilizing the starches and the molecular size after mechanical and thermal treatments, as inferred from the viscosity measurements, were shown to influence the adsorption characteristics on iron ores, and the residual concentrations of both calcium ion and starch in the pulp liquor were in direct correlation with the flotation and flocculation behavior.2-4 The present article summarizes the results of more recent tests conducted to ascertain the role of starches and starch derivatives in various phases of iron ore beneficiation. The scope of this investigation has been extended to include not only anionic and cationic flotation, but also the flocculation and filtration of iron oxide slimes, the selective flocculation and partial upgrading of semitaconites, and the flocculation and clarification of magnetite-taconite tailings —all problems that are of vital importance to the iron ore industry of Minnesota. By combining a practical approach with adsorption studies and by using single-mineral systems5 in these research efforts, it is believed that the ultimate goal of developing a suitable modification of starch structure may be realized. The topics covered in this article are limited to the empirical aspects of the research program on the depressant activity of starches and products in the cationic and anionic flotation of silica and of their role in flocculation, water clarification and filtration. ANIONIC SILICA FLOTATION In the soap flotation of activated silica from iron

Jan 1, 1970

In dense medium processes for the beneficiation of minerals the consumption of medium (generally ferrosilicon) represents a major component of the operating cost. This provides a strong incentive for improvements in heavy medium recovery circuit design and operation. A major thrust in heavy medium plant design over the past decade has therefore been towards circuits which reduce medium losses. In this paper recent trends in plant design are examined in relation to the factors infl uencing circuit performance and medium losses. Operational details ' which have a significant effect on losses are discussed and illustrated by means of plant and pilot plant test data, with particular emphasis on the performance of wet drum magnetic separators. While the emphasis is on reducing the consumption of the high-cost ferrosilicon medium used in minerals beneficiation, most of the principles described in this paper are equally relevant to the design and operation of magnetite recovery circuits for coal preparation plants.

Jan 1, 1987

By F. K. Mulenga, M. Lusambo

"Geometallurgical models are a strategic tool used primarily in planning, diagnosis, and process optimisation. This paper discusses the development of a predictive model for flotation recovery at Kansanshi concentrator in Zambia as a function of key operational parameters. Daily data for milling and flotation from the circuit processing the copper sulphide ore was analysed. This data, recorded over a period of one year of production, was collected from diverse sensors measuring semi-autogenous milling and flotation parameters. Statistical analysis showed that copper head grade, acid-insoluble, and acid-soluble copper, SAG mill throughput, ball mill load, and tonnage of feed material to the SAG mill in the -125 +63 mm size fraction contributed to the prediction power of the proposed empirical model by as much as 90%. Finally, a proposal was made for the development of a simulation model under MODSIM®, a flexible tool for mineral processing applications. This was to test the limitations of the empirical model for future refinement. IntroductionKansanshi is the largest copper mine in Africa and the world’s eighth largest copper producer, with a total of 253 272 t of copper produced in 2016 (Newall, 2017). It is located approximately 10 km north of Solwezi, the capital and administrative headquarters of the North Western Province, Republic of Zambia. The ore deposit consists predominantly of chalcopyrite and malachite with variable amounts of chalcocite, bornite, chrysocolla, and azurite. Because of the varying proportions of sulphide and oxide copper minerals, the ore is classified into three distinct categories: sulphide, mixed, and oxide ores. As such, three separate processing circuits are used to mill the ore down to a standard targeted grind size of 80% passing 150 µm before froth flotation and leaching.A lot of research and development has been done to improve the performance of the sulphide ore circuit at Kansanshi. Despite the research efforts and subsequent operation, the sulphide ore circuit did not attain the target recovery and grade for 2016, primarily because of the complexity of the orebody.The constant drive to achieve the set production target of recovery and concentration grade after froth flotation motivated the development of a tool for the prediction of recovery. The anticipated benefit is not only to enable the plant to apply corrective actions when needed, but also to make a tool available for the quick assessment of the plant performance of the. For that purpose, data spanning June 2016 to May 2017 was analysed, cleaned up and curvefitted to an empirical model. In doing so, key parameters were determined and calibrated to represent the behaviour of the plant."

Nov 1, 2018

By Michael Perch, Charles C. Russell

The fact that coal is a complex organic material and heterogeneous in composition has made its study extremely difficult, particularly in regard to obtaining a fundamental concept of the processes involved in the formation of coke. It has been stated1 recently that coals are natural polymers and that in a complex reaction such as occurs during carbonization the differences between coals are quantitative and not qualitative. Such being the case, it would be expected, therefore, that a series of coals subjected to exactly the same conditions of coking would produce results that should be related quantitatively to the chemical analyses of the individual coals. In the past, however, this fact was not generally recognized and consequently many empirical test methods were devised for testing coals. The interpretation of the results obtained depended upon comparing a set of data with other data on coals of known characteristics. The testing procedure had to be conducted in precisely the same manner each time and not the least requirement in making the interpretation was experience. The test methods varied from those utilizing a gram of coal to the plant-scale tests in which one or more ovens were charged with the test coal. The aim had been principally to find some simple, inexpensive means by which the coke-making and other properties of coal could be predicted. It is not necessary to emphasize the importance of these endeavors, for anyone acquainted with the problems involved in the selection of coals for making coke is well aware of the time and expense involved in making a series of definitive tests. Several attempts have been made to derive formulas for such predictions. The relations of both ultimate and proximate analyses to coke physical properties and byproduct yields have been studied. Lowry and co-workers1 correlated statistically the U. S. Bureau of Mines-American Gas Association assay test results with coal analyses, using 90 coals and coal blends. Parry,² Spooner,3 and Gabinskii and Krym4 made similar correlations. Lowry concluded that there were definite relationships between the proximate analyses of the coals and the yields and properties of the products. Although this statistical study was made on test results from carbonizing coal in a cylindrical steel retort, the conclusions are important and provided one of the first steps in the simplification of this complex subject. Studies should be made of other references on this subject just cited. Studies of the coking process reported in the literature frequently were confined to a small number of coals, and very few have included coals from various ranks. It would seem logical, however, that a program for studying factors involved in the formation of coke should be based on a number of coals selected from each of the various classes or types of coking coal. The possible use of the scheme of classifi-

Jan 1, 1950

By Thomas P. Mucho, David C. Oyler, Jim Pablic, Dennis R. Dolinar

A room and pillar coal operation in central Pennsylvania was experiencing roof cutters and long running roof falls caused by high horizontal stresses. The roof conditions created hazards for the miners, and caused several production panels to be abandoned prematurely. The mine requested assistance from the National Institute for Occupational Safety and Health (NIOSH) in applying the "advance and relieve" mining to reduce the affects of the horizontal stress during panel development. The "advance and relieve" plan that was developed called for removing a pillar on one side of the panel as it was being advanced, thus creating a cave. The caving then relieved a portion of the horizontal stress across the panel. Because the horizontal stress direction is key to the success of this method, stress mapping as well as mining experience was used to establish the direction of the horizontal stress. Subsequent field measurements provided an estimate of the stress magnitude. Three panels were mined using this technique, and good roof conditions were achieved in all these panels. Instrumentation was used to monitor the stress changes in the roof created by the cave. Stress relief of over 1,000 psi was measured 120 ft from the cave and to depths of 20 ft into the roof. The lateral extent of the stress relief zone across the panels appears to be about 400 to 500 ft. This case history has provided much insight into the practical application of the "advance and relieve" method discussed in this paper.

By Onton W

During the latter part of 1936, conditions encountered in sinking the Main Ore shaft made it impractical and uneconomical to continue deepening it to serve haulage levels below the present No. 5 level. Plans were therefore made for an interior shaft, to serve new lower levels, that terminated above No. 5 level so that ore hoisted could be drawn off on this level and transported to the Main Ore shaft for final conveyance to the mill crude ore bin. The No. 2 interior ore shaft was only developed below No. 4 level and borehole ropeways with a winder located on the surface were installed in preference to an underground winder installation. No. 2 ore shaft with borehole ropeways commenced lead ore hoisting operations in September 1938. Apart from the copper ore hoisted in the war years of 1943-46, lead ore was hoisted to No. 5 level until September, 1951.

Jan 1, 1988

By R. K. McGeary, B. Lustman

Orientation relationships and rates of annealing of 97 pct cold rolled zirconium have been studied by X-ray techniques, metallography, and by hardness measurements. The process of annealing occurring by the formation of domains, and their subsequent growth without recrystallization, is demonstrated. NUMEROUS investigators1-' have revealed sub-grain formation after annealing cold worked metal single crystals and have generally described the results in terms of a dislocation mechanism of polygonization.5,6 in heavily cold rolled polycrystal-line aluminum the first stages of heating cause an apparently spontaneous subdivision of regions of the deformed matrix into a great number of slightly disoriented domains followed or accompanied by rapid growth of certain domains to form a recrystallization texture.7,8 In the present investigation the isothermal rate of annealing of cold worked polycrystalline zirconium has been studied by X-ray techniques, metallography, and by hardness measurement. It will be shown that, rather than by the usual nucleation and growth mechanism,"12 the process of annealing can best be explained by the formation of domains early in the annealing treatment and the subsequent growth of these domains without recrystallization. Experimental Procedure All the experimental work was performed on an arc-melted ingot of high purity zirconium crystal bar made by Westinghouse in a production scale deBoer unit. The weight percentage of impurities in the zirconium were as follows: C, 0.04; Fe, 0.05; Hf, 0.014; W, 0.01; Ti, Ni, and Al, 0.003; Si and N, 0.002; Cr and Cu, 0.001. Metallographically the zirconium showed only a single phase. The 4 in. diarn ingot was forged at 760°C to a 1 1/2x3 in. plate and was then rolled at 760°C to 0.150 in. thickness and annealed at this temperature for 20 min to produce an equiaxed grain structure of about 0.1 mm average grain diameter. The hot rolled plate was roller-leveled, surface ground to 0.135 in., cold rolled 83 pct reduction in thickness on a two-high mill, and then to a total reduction of 97 pct on a Sendzimir mill. The 0.004 in. strip material so produced received no subsequent mechanical treatment. Specimens % in. square were sheared from the rolled strip and isothermally annealed at 104 mm Hg in an all glass vacuum system equipped with a Vycor furnace tube. The temperature data presented are the average temperatures during the runs and did not vary more than about k3"C. X-ray Diffraction Methods: When zirconium is drastically cold rolled, a high degree of preferred orientation results in which prism poles, {1010}, are perpendicular to the rolling direction. After fully annealing this material an entirely new texture develops which corresponds to a 20" or 40" rotation of the deformation texture about hexagonal axes of

Jan 1, 1954

By D. R. Zuchelli

In order for MSHA to require the use of a noise control, the control must be proven ?technologically feasible?. Technologically feasible controls must reduce a miner?s noise overexposure to the permissible exposure level (PEL) or achieve at least a 3 dBA reduction in the miner?s noise exposure. The miner?s total noise exposure should be examined from an occupational viewpoint and not solely on a machine or equipment basis. This examination requires a field evaluation of the occupation as opposed to an equipment test conducted in a laboratory. The operator of a Hilti TE MD20 LS electro-pneumatic drill realized a full shift noise exposure (TWA8) reduction exceeding 8 dBA as compared to when operating a Gardner Denver S83F pneumatic jackleg drill while installing an equal number of drill holes into the Idaho Springs Gneiss rock formation. However, the Hilti drill?s penetration rate was only 27% of that of the jackleg drill. Although the Hilti drill was proven to be technologically feasible as a noise control, its slower penetration rate would require its economic feasibility to be evaluated on a case-by-case basis.

Jan 1, 2011

By E Philippa, B Miller

Material characterisation, ore sequencing and process capability are commonly regarded as the key parameters to consider in designing and operating an effective pit to plant supply chain. Therefore, an understanding of rock mass properties is the starting point for controlling the presentation of problematic material to the ore handling plant and optimising process utilisation. There are different methodologies for characterising problematic material and ore handleability which exist across a spectrum of sophistication. At one end of this spectrum are simple indices built using first principles geology and at the other are complex algorithms developed using advanced analytical methods, machine learning and ore traceability. Regardless of the method, without an integrated process for careful management of problematic ore, material handling delays (ie blocked, bridged, blinded and buried events) will interrupt the stable flow of production. The ability of the ore control geologist to mitigate these delays is central to the ore handleability challenge. This paper describes a quick and easy solution that links the presentation of ore to the operating window at the crusher. This operating window is based on material characterisation, using the physical, textural and mechanical properties of the rock mass, and crusher performance. It also describes how a basic problematic ore index (POI) can be developed using ore tracking to the production face, petrographic studies of ore texture and mineralogy, and existing variables in the geological model. Finally, this paper will demonstrate that by coding the POI into the geological model for mine planning, together with performance monitoring through compliance to plan routines, material handling delays at the crusher can be reduced, unlocking significant resource value.

Mar 22, 2022

By Hugh S. Spence

In the winter of 1941-42, muscovite mica was discovered by a young prospector, Justin Purdy, in the township of Mat-tawan, Nipissing District, Ontario, a few miles north of the small settlement of Eau Claire, which lies on the main Transcontinental Highway (No. 17) between Mattawa and North Bay. Muscovite had been known for many years to occur in the general Mattawa region, and various small efforts at production had previously been made by individuals working scattered deposits in the area. These minor activities, however, had never led to production of any consequence or the discovery of any important mica bodies. The Purdy finds were made in a heavily timbered, mountainous area lying about three miles north of the low-lying settled land along the highway, and in a section not hitherto prospected. The mica taken by Purdy and two partners from surface outcrops proved so remarkable in point of size of trimmed sheets that the find attracted considerable attention, and the area surrounding the three claims taken up by the prospectors was quickly blanket-staked by various interests. Though a number of other mica-bearing pegmatite dikes were found later on these stakings, none proved as rich as those on the original Purdy ground; very little commercial mica was obtained from them and attempts to develop them were short-lived. Purdy and his partners sold their initial production to dealers in Ottawa and Hull, where the Canadian mica industry long has centered. Young Purdy, unfortunately was killed in a motor accident in the early summer of 1942. His father continued mining operations for a few months and produced and sold a considerable amount of mica. In October 1942 the Purdy holdings were acquired by lnspiration Mining and Development Co. of Amos, Quebec, which had been active in staking in the area. This company formed a subsidiary, Purdy Mica Mines, Limited, to develop the property. Under the energetic management of Joseph R. Norrie, rapid progress was made, and production soon attained important proportions. Realizing from the outset the unusual character and potential value in the war effort of the Purdy discovery, the Metals Controller, Department of Munitions and Supply, at once arranged with the Department of Mines and Resources for an official geological report on the area, and also drew the attention of the War Production Board in Washington to what seemed likely to prove a highly important source of strategic-quality mica. Officials of WPB were sent up in the Summer of 1942 to inspect the property, and so impressed were they with the possibility that it might become of high

Jan 1, 1948

By Hugh S. Spence

In the winter of 1941-42, muscovite mica was discovered by a young prospector, Justin Purdy, in the township of Mat-tawan, Nipissing District, Ontario, a few miles north of the small settlement of Eau Claire, which lies on the main Transcontinental Highway (No. 17) between Mattawa and North Bay. Muscovite had been known for many years to occur in the general Mattawa region, and various small efforts at production had previously been made by individuals working scattered deposits in the area. These minor activities, however, had never led to production of any consequence or the discovery of any important mica bodies. The Purdy finds were made in a heavily timbered, mountainous area lying about three miles north of the low-lying settled land along the highway, and in a section not hitherto prospected. The mica taken by Purdy and two partners from surface outcrops proved so remarkable in point of size of trimmed sheets that the find attracted considerable attention, and the area surrounding the three claims taken up by the prospectors was quickly blanket-staked by various interests. Though a number of other mica-bearing pegmatite dikes were found later on these stakings, none proved as rich as those on the original Purdy ground; very little commercial mica was obtained from them and attempts to develop them were short-lived. Purdy and his partners sold their initial production to dealers in Ottawa and Hull, where the Canadian mica industry long has centered. Young Purdy, unfortunately was killed in a motor accident in the early summer of 1942. His father continued mining operations for a few months and produced and sold a considerable amount of mica. In October 1942 the Purdy holdings were acquired by lnspiration Mining and Development Co. of Amos, Quebec, which had been active in staking in the area. This company formed a subsidiary, Purdy Mica Mines, Limited, to develop the property. Under the energetic management of Joseph R. Norrie, rapid progress was made, and production soon attained important proportions. Realizing from the outset the unusual character and potential value in the war effort of the Purdy discovery, the Metals Controller, Department of Munitions and Supply, at once arranged with the Department of Mines and Resources for an official geological report on the area, and also drew the attention of the War Production Board in Washington to what seemed likely to prove a highly important source of strategic-quality mica. Officials of WPB were sent up in the Summer of 1942 to inspect the property, and so impressed were they with the possibility that it might become of high

Jan 1, 1948

By Damon C. Gray

This report contains a recommended program of acoustic research to be conducted by the U. S. Bureau of Mines. The program is directed toward the abatement and control of machinery noise in underground coal mining operations. The study conclusions and specific problem areas which led to the definition of the recommended program are also briefly described.

Jan 1, 1971

By R. Karl Zipf, Suresh Bhatt

Highwall mining is an important coal mining method. It appears that upwards of 60 highwall miners are presently in operation, and they may account for approximately 4% of total U.S. coal production. A review of the Mines Safety and Health Administration (MSHA) data over the 20 year period from 1983 to 2002 identified 9 fatalities attributable to auger and highwall mining of which inadequate ground control accounted for 1/3. In the past 5 years, 1 fatality occurred in highwall mining. Estimates of the manpower requirements in highwall mining suggest that its fatality rate is essentially the same as for surface coal mining, thus highwall mining appears to be a very safe modern mining method. Highwall stability is the major ground control related safety concern, and operators are required to develop and follow an appropriate highwall mining ground control plan. The plans usually specify the following geotechnical parameters: hole width, maximum hole depth, maximum overburden depth, seam thickness, web pillar width, barrier pillar width and number of holes between barriers. Calculated web pillar stability factor exceeded 1.3 for most designs evaluated. This study examined records from 5,289 highwall miner holes with a total completed footage of about 2,560,000 feet to understand the reasons for early pull out. Average lost footage is typically about 20% of planned footage. Only 35% of the holes reached planned depth, and 20% were short due to rockfalls. Water and adverse geology accounted for 15% of the losses. Mechanical/electrical problems, guidance, and slope stability problems accounted for the remaining 30%. Web pillar stability factor for these holes also exceeded 1.3 in 95% of cases. Best practices to avoid trapped miners include: avoid mining in stream valleys, avoid mining near outside corners, careful alignment of each hole and the use of an onboard guidance system. Several issues in highwall mining ground control require further investigation including highwall mining through old auger workings, highwall mining near old underground mines, multiple-seam and multiple lift highwall mining and finally the size and frequency of barrier pillars.

By Robert E. Miller

Mass blasting of open stopes and pillars is a normal operating practice at Isa Mine. Three main parameters are used to classify the types of mass blasts. These are type of explosive used, blasthole diameter and system of initiation. Both electric and non-electric initiating systems are currently employed. The advantages gained by large firings are many. More efficient utilization of manpower, due to the concentrated and continuous charging, is achieved. There is no requirement for men to re-enter areas where localised ground conditions may have deteriorated and exposure of ringfirers to stope voids is reduced.

Jan 1, 1980