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By M. de Paiva Bueno, G. Barakos, S. Luukkanen, K. Luolavirta

Lifelong development of skills and knowledge of mining professionals has been recognised worldwide as a challenge for the development of a sustainable mining sector. Practical training and learning over the mine value chain are among the key components of this knowledge building. However, the number of available mining sites for continuous education in real-life conditions is limited; also the specifications for practical training are not clearly defined. The project MINETRAIN was developed to (i) evaluate the potential of the shortly-closing Pyhäsalmi Mine, in Finland, as an educational and training underground facility, and thus, prolong the utility of the existing modern mine site, and to (ii) develop and test practical training programs for mining and mining industry related professionals. Pyhäsalmi facilities enable training among all disciplines related to the overall mine value chain and hence, the project has potential in providing education for sustainable management of any phase of mining operations. The philosophy underpinning the MINETRAIN approach to mining activities is to assist the clients in minimising the technical and financial risks by providing a competitive and skilled personnel asset while optimising the value drivers in their operations. As a first major step of the project, a pilot training course, involving basic-level online theory and hands-on training at the Pyhäsalmi mine site, was undertaken in summer 2019. Feedback from the pilot test group of mining professionals has proved to be very positive. The comments from the participants strongly emphasised the added value of the practical onsite training for improved understanding in the large scale picture of mining on the entire value chain. As a second phase of the project, a new test course was launched in June 2020. The second MINETRAIN course focuses on new technologies and digitisation in mining. Keywords: Lifelong learning, mining education, mine value chain, research and education mine, educational research project

Jan 1, 2020

By Xiaojie Yang, Gang Yang, Zhenli Hao, Manchao He, Ruifeng Huang, Zheng Wang, Chaowen Hu

Preventing the spontaneous combustion of remnant coal, which is indicated by carbon monoxide (CO), is a significant issue for coal mining safety. The N00 longwall mining method with Z-type ventilation is advantageous for the prevention and control of CO; however, research regarding this topic is limited. To investigate the distribution of CO in a goaf and working face, a sealing material and technique for blocking the goaf were introduced after proposing a forming method for a Z-type ventilation system; furthermore, a monitoring method was designed and applied in the field. The results demonstrated that the CO concentration during mining in the upper corner and return airway ranged between 4–6 ppm and 3–7 ppm, respectively. More importantly, no CO was generated in the goaf. The results indicate that the advantages and mechanisms regarding the prevention of spontaneous ignition using the N00 longwall mining method with Z-type ventilation are as follows: (1) The adoption of the downward ventilation technique is not only beneficial for reducing the air pressure difference between both ends of the working face, but also prevents CO accumulation in the upper corner. (2) The basis of a spontaneous combustion disaster is eliminated owing to the absence of coal pillars reserved in the goaf. This study proposes a superior method for mitigating spontaneous combustion disasters related to remnant coal in the goaf in a coal mining field and contributes to the better understanding of the N00 longwall mining method for coal mining safety.

Nov 28, 2022

By C. R. Dodson, W. T. Cardwell

This paper presents the results of a theoretical and experimental study of the effect of preperforated liners on well productivity. The analysis concerns the rectangular type of slot, either machine or torch cut, which is common in California. Muskatla treated this problem recently in a mathematical analysis of the flow of oil or gas from natural reservoirs into both gun-perforated and slotted liners. This paper presents additional and simplified analytical considerations and conclusions for the slotted liner, the solutions being applicable in the laboratory as well as in the field. An application of the analytical methods to permeability measurements on mounted core samples is given, to show the effects of sealing wax or other obstructions on the end faces of the cores. Results Fig. I 'Ompares the results of some theoretical calculations with the results of electrical analogy experiments performed in the Standard Oil Company of California's Production Technology Laboratory at La Habra, California. Fig. I indicates that the flux into short, "three-dimensional slots" of the relative dimensions encountered in practice is approximately equal to that into full-length, "two-dimensional slots" having the same open fraction of pipe. (The same numbers of slot rows, and the same radius ratio are also implied.) It can be seen from physical considerations that the closeness of this approximation must improve as the ratio of the external radius to the internal radius increases. Therefore, it is possible to calculate the flow into a slotted liner using a simple modification of the Darcy equation: where N is .the number of slot rows and O is the open fraction of the pipe. Fig. 2 shows the results of calculations . using Eq. 28. The most remarkable features of Fig. 2 are the high flux values for very small open fractions of pipe. For instance, with six slot rows and a radius ratio of 500, 74 per cent of the open-hole flux may be obtained with only o.I per cent of pipe Open. Fig. 2 shows that for a given open frac tion of pipe, the flux is greater, the greater

Jan 1, 1945

By G. W. Grove, H. R. Burdelsky, F. E. Griffith

"INTRODUCTION The work of using a system of air locks in unsealing a mine in Ohio following a fire probably was one of the outstanding jobs of its kind in a medium-thick coal bed.The fire in this mine, thought to have been of electrical origin, was discovered at about 1:30 a.m. on December 11, 1941. Forty-four men were working in the affected section at the time; 5 escaped unassisted, and 37, who had barricaded themselves, were rescued. The bodies of two men, who had succumbed after breathing fumes given off by the fire, were brought part way out of the mine before it became necessary to seal the mine.Direct attempts to extinguish the fire with water were abandoned when hazardous amounts of 'carbon monoxide, methane, and other inflammable gases were detected in return air passing through the main haulageway and escape-way of the mine.The mine was completely sealed on the surface at about 8 p.m. on December 12, 1941, with the two bodies remaining in the mine. Sixteen days after the mine had been sealed, the two bodies were recovered. This was accomplished by air-locking through a partitioned-off stairs compartment of a ventilation shaft. This work was done under adverse barometric and atmospheric conditions.A method for collecting representative air samples from the immediate fire area was used successfully by utilizing a vacuum pump on about 10,000 feet of existing water line that had been broken near the fire area.The immediate fire region was sealed, and confined to a relatively small area; the remainder of the mine was recovered, and operations were resumed February 9, 1942, or 59, days from the time of sealing the mine on the surface. This recovery was made by air-locking the main entries inby from the shaft and No. 2 right-face entries.The distance between air locks averaged about 500 feet, and as each new set of seal stoppings and air locks was completed, the old or outby seals were taken out and the area between was ventilated. In ventilating this area, it was often necessary to use line curtains to direct the air into dead-end places."

Oct 1, 1947

By Edward C. J. Urban

The nature of certain pernicious dusts commonly encountered in the removal of iron ore from the underground mines of the Lake Superior district is recognized, and appropriate measures for the protection of individuals subjected to these dusts have been applied in most mines for more than 10 years. Undoubtedly the principal concern has been demanded by dusts that contain as a constituent silicon dioxide in its free or uncombined state, the mineral we all know as quartz or chert. Investigators1 agree that among the factors that determine injurious exposure are: the composition of the dust, the concentration suspended at the breathing level and the duration of the exposure. We have on hand data that describe the experience of several underground iron-ore mines in the Upper Peninsula of Michigan and Wisconsin in regard to these factors, which in a large measure also reflect the general experience of the other mines of the area. We believe that a discussion of these data will indicate the effectiveness of the dust-control methods that have been considered good practice and that have been employed over a long period of time. A silicosis-prevention program was ini- tiated by the iron mines of the Upper Peninsula in 1934 and records of mine dustiness are available from that date to the present time. For dust conditions previous to 1934 the results of observations conducted in other industries can be utilized together with local information pertaining to the development and initial application of rock-drilling equipment. Development of Rock DriI,I.s Thc earliest percussion rock drill that later had any industrial use was patented in this country in 1851.2 This machine was later modified and improved in many respects but its Piston-Type design remained standard for drilling equipment in most mines for the next 63 years.3 The piston drill was operated either by air or by steam. It employed a solid drill steel, fastened directly to the reciprocating piston. The drilling was done dry except in holes directed at an angle below the horizontal. For these water was manually injected into the hole and the pumping action of the rod expelled the sludge. The dry cuttings from drillings inclined above the horizontal fell from the hole by their own weight. A development of major importance from the standpoint of dust generation and control occurred with the invention by J. G. Leyner of a hammer-type rock drill that embodied the essential feature of a drill steel with a hollow core. His patent was secured in 1897. The machine was a radical departure in drilling method. The steel was held loosely in a chuck while

Jan 1, 1949

By M. Holuzko, Y. R. Zhang, J. Alexander

The use of dual-energy X-ray transmission (DE-XRT) technology in the mining industry has been developing rapidly in the past decade. In many projects, DE-XRT images are analysed to classify ore and waste in sulphide deposits where the sulphides and non-sulphides have distinct characteristics in density. In the current practice of DE-XRT image processing, every pixel in the image gets classified into two categories, depending on the intensity of the pixel from both low and high energy X-ray response. If the pixel’s intensity is above the threshold curve generated in the high-energy and lowenergy scatter plot, it is classified as “higher density” material, and the remaining pixels are classified as “lower density” material. This paper elaborates on an alternative approach to DE-XRT image processing for in-depth characterisation of samples from a lead-zinc sulphide deposit. The detailed characterisation is achieved by providing density distribution of individual rock using DE-XRT laboratory sorting unit (COMEX) located at the University of British Columbia. Instead of classifying the X-ray images into two different classes from one threshold curve, multiple density classes are created between two threshold curves to generate the density distribution. By manipulating the two threshold curves, we can analyse images of samples obtained at different resolutions of density distribution. If the two threshold curves become infinitely close to each other, it will create a single-threshold scenario similar to the current methods being used in the industry. This method provides the ability to change the threshold curves, hence providing more flexibility on material characterisation based on the needs of the operation. Furthermore, the DE-XRT images from individual rock specimens are compared to the test results from hand-held X-ray fluorescence (XRF) analyser as well as assays of various elements for the same set of samples. The data analysis will provide information regarding ore characteristics and also the theoretical performance of the sensor-based sorting. Keywords: Dual-energy X-ray Transmission (DE-XRT), X-ray fluorescence (XRF), ore characterisation, density distribution

Jan 1, 2020

By Charles E. Van Barneveld

RBH Note: the following item is a little different than the usual, in that it shows how explosives were used in the past to assist in putting down holes for iron ore prospecting on the early Mesabi Range of Minnesota. The way explosives were used would not even occur to the modern explorationist, given the equipment available today. The reason for the great difficulty in putting down a hole was that the area was covered with debris from the great glaciers of the past (the same ones that carved out the Great lakes). This debris consisted of boulders, clay, sand, and any other imaginable mixture of materials.

Jan 1, 2013

By E. V. Koporulina, I. Zh. Bunin, M. V. Ryazantseva

"We studied the mechanism of structural and chemical transformations of columbite, tantalite, zircon and feldspar surface during the treatment by high – power electromagnetic pulses (HPEMP) using the methods of X - ray photoelectron spectroscopy (XPS), analytical electron microscopy (SEM - EDX), Kelvin probe force microscopy and flotation experiments. According to XPS and SEM - EDX data, major changes in the atoms chemical state of investigated minerals surface are related with formation and modification of functional cover and changing in concentration (% at.) of hydroxyl groups on the mineral surface, leading to increase in the contrast of technological properties of the rare metals minerals. Research has identified the optimal parameters of HPEMP - treatment of rare - earth minerals Npuls ??104, Ttreatment ??100 sec. improving the selectivity of the flotation separation of columbite and zircon. The difference in mineral recovery into the froth (? e) before and after HPEMP - treatment was about 11 % and 24.5 %, respectively. No significant increase was observed in the flotation activity of feldspar. This study's findings demonstrate the effectiveness of HPEMP - treatment in inducing a controlled change in the structural and chemical properties of rare - earth minerals and improving the process efficiency for the processing of rougher tantalum - niobium gravity concentrates.INTRODUCTIONThe importance of improving the processing methods of tantalum-niobium ores and concentrates is explained by the mineralogical properties of rare - earth ores. These are properties such as the variable mineral, chemical, and grain size composition, complex morphometric properties, and the dissemination of ore mineral grains of micro - and nanometric size. Iron oxide and hydroxide films on the mineral surfaces neutralize the surface physicochemical and absorption properties of the minerals. Both in Russia and globally, nonconventional (non - mechanical) methods are being developed based on physicochemical, electrochemical, and pulsed energy impacts to improve the efficiency of disintegration and liberation processes of disseminated mineral complexes and to enhance the contrast of the structural, chemical, and technological (floatability) properties of refractory ores. This study is devoted to the investigation of the applicability of high – power electromagnetic pulses (HPEMP – treatment) for the change in the structural and chemical properties of rare - earth minerals and improving of the flotation properties contrast."

Jan 1, 2018

By A. Boehm

Mineral selective flash roasting (Steiner process) in oxidizing atmosphere was tested successfully for the efficient low intensity magnetic separation of siderite from ankerite and other gangue minerals within the ore of the Steirische Erzberg, Austria. The calcium, magnesium and manganese containing siderite turns selectively into a ferromagnetic oxide phase while ankerite remains paramagnetic. The expulsion of CO2 increases the volume specific surface depending on the selected temperature. Unlike pure iron oxides absorbing SO2 below 530°C, the calcined siderite (enriched in the magnetic product of the Steiner process) acts as a sorbent for SO2 between 500 and 820°C. The effect was studied by thermogravimetric analysis and particle bed tests complemented with neutron thermo diffractometry (D20/ ILL). The energy input was simulated with the help of a tube furnace. The composition of the artificial combustion gas permeating the bed was close to that measured in pilot scale tests (6 % CO2, 12 % O2, 700 vpm SO2). The evaluation of data relating the SO2 variation in the off gas monitored by electrochemical cells to the actual phase composition revealed an intermediate ferrite phase as absorber. During the rapid decomposition in oxidizing atmosphere starting at 500°C the occurring siderite alters to a (CaO, MgO, MnO)Fe2O3 ferrite as an intermediate phase accompanied with sudden SO2 absorption. SO2 disintegrates the ferrite structure forming MgSO4 and MnSO4, taking Fe2O3 as a catalyst. Ankerite (enriched in the non magnetics of the Steiner process) starts to absorb at temperatures above 1000°C forming gypsum. The findings support the SO2 controlled operation of the Steiner process by ?autodesulfurisation?. It explains aspects of the SO2 implementation reactions within off gas recycling in Austrian sinterstrands. Keywords: magnetising flash roasting, neutron thermo diffractometry, SO2 absorption, siderite, ankerite

Sep 1, 2012

By T. Mützea, L. Schützenmeister, G. Kacheb

High-Pressure Grinding Rolls (HPGR) are employed in cement plants worldwide. Although the technology has been mastered in the majority of applications, design-, scale-up- and operationalchallenges still exist today. In addition to modelling throughput, energy requirement and comminution success, the prediction and avoidance of malfunctions is of great importance in current research. The performance of the mill is highly influenced by the roll speed and the fineness of the feed material. Another parameter influencing the composition of mill feed is the circulating load, which is determined by the grinding circuit configuration, the target fineness and the separation efficiency of the classifier. In order to meet this circumstance and simultaneously maintain a reasonable level of material input and testing effort, a complete grinding circuit on a small technical scale was set up in the research centre of Thyssenkrupp Industrial Solutions AG. The present study varied the roll speed (0.5-2 m/s) and the proportion of fines in the feed (0-75 %) in order to investigate the influence of both parameters on the finish grinding of a Portland cement with a target fineness of 4000 cm²/g according to Blain. The experimental results are compared with model concepts and critically examined with regard to their transferability to industrial plants. Keywords: Cement grinding, grinding circuit, HPGR

Jan 1, 2020

By R. F. Shurtz

R. Duval (Mining Engineer, Ancien eleve de PEcole Polytechnique, Paris, France) I do not agree with the Eq. 3, reading: m =1/100- [(0.214x30.4) + (0.7B6 x0.00)] =6.5pct CaO If 0.214 and0.786 were proportions by weight, the equation would represent the well known mixtures law of the conventional arithmetics and 6. 5 pct CaO would be the correct average content. But it is not the case as the author states: "In samples consisting of single grains of mineral, those grains must, as already mentioned, be either of dolomite or magnesite. Since 78.6 pct of the deposit consists of magnesite and 21.4 pct of dolomite (excluding for present purpose the presence of other minerals), for any single grains picked at random the probability will be 0.214 that is it dolomite and0.786 that is it magnesite. In 1000 such samples the expected numbers of dolomite and mapesite grains will be 214 and 786 respectively." 0.214 and 0.786 would be proportions by weigbt under the necessary condition that all grains of dole mite and magnesite should have an identical weight. Obviously it is not the case, as the specific gravities are not the same for mapesite and dolomite and the volumes of the grains are different. Furthermore, because of these differences the conditions for a random sampling are not fulfilled and we are not authorized to state that the probabilities are, respectively, 0.214 and 0.786. The author however makes a simple application of Eq. 1: M = 1/n— ? fi x i . n Should we deduce that this relation is wrohg? Not at all, but when applying Eq. 1 you must not overlook what it actually. means. Eq. 1 gives a definition of the arithmetic mean of a total of n observed values Xi and nothing else. But the average conteht of a deposit has not the same significance. It is the ratio between the weight of concerned mineral in the deposit and the total weight of the deposit. As from 1000 particles the 214 of dolomite and the 786 of magnesite have not the same weight, the two definitions do not concur, and when applying Eq. 1 the result is an arithmetic mean of figures which has no connection with what is named average contentof a deposit. The situation is similar to the calculation of an average velocity. If a car travels a first mile over at 30 miles per hr and a second mile over at 60 miles per hr, when applying formula 1 you find as average velocity for the 2 miles: 30+60 ------- - 45 miles per hour. Many people calculate in this way and they do not realize that a mistake is involved. In fact the definition of he average velocity for the 2 miles is the quotient of the distance of 2 miles by the time (in hours) necessary for 2 miles travel, i.e.: 2 ---------- = 40 miles per hr. 1 + 1 30 60 In other words, the average volocity wanted is not the arithmetic but the harmonic average of the two velocities. The above mentioned bias in the calculation of the average contents of deposits is frequent, even in the assessments made by experienced engineers and is independant of what is named the sampling error. In order to supress the bias and to be able to use Eq. 1, you must apply a correction. An example on the subject can be found in an article by Duval et al. in the January 1955 issue of the ''Annales des Mines" (French), page 19. R. F. Schurtz (Author's Reply) Mr. Duval's position is quite correct. The proportions shown for dolomite and magnesite., respectively, of 0.214 and 0.786 are, in fact, proportions by weight uncorrected for specific gravity. In our day to day operation of producing magnesite from these mines at a very substantial rate, we do not normally make corrections for the difference between the specific gravity of dolomite and that of magnesite. If these corrections are made in Eq. 3 as shown in my article, then the numbers of grains turn out to be in proportions of 0.226 dolomite and 0.774 mapesite instead of the values actually shown in the equation. For the purposes of our work, and in view of the inherently low accuracy of the data, this correction was not deemed worthwhile making.

Jan 1, 1961

By E. Eichen, G. S. Ansell. L. R. Sefton

A Technique for the Preparation of Thin Films of Two-Phase Alloys Suitable for Use in Transmission Electron Microscopy In order to obtain foil sufficiently thin to permit transmission electron microscopy of two-phase alloys, particularly alloys in which the second phase is present as a very finely dispersed second phase, a technique has been developed to permit the thinning of these materials without either the large degree of pitting present in the electrochemical or chemical thinning process or the slicing of the second phase particles accompanying the mi-crotoming process. The metal or alloy to be thinned is first ground to initial sheet thickness of approximately 0.002 in. by hand grinding on 4/0 emery paper. This starting sheet of material is then reduced in cross section by an air-abrasion technique using an S. S. White dental abrasive unit. This unit consists of a supply of compressed CO, which provides a stream of high velocity gas impinging on the metal surface. Incorporated in the machine is a hopper-type of device in which fine abrasive particles of alumina are introduced into this high velocity gas stream. The abrasive particles are irregular in shape, approximately 30 across. If any particles are introduced into the sample during thinning, they should be readily observable. Impingement of this gas abrasive mixture on the metal surface causes rapid removal of the metal in a rather uniform manner. The rates of metal removal may be adjusted by both regulating the gas pressure and amount of abrasive material introduced into the gas stream. By moving the gas stream abrasive mixture slowly about the specimen or sheet surface, a large area of the sheet may be uniformly reduced to a thickness which is then suitable for use in transmission electron microscopy. Attendant with this metal removal is a result in cold working of the metal surface due to the impingement of the abrasive particles. The resultant foils, therefore, if made from materials which are desirable to view in the annealed condition, have to be subsequently heat treated. If the alloy from which these foils have been made has been initially heavily worked or deformed during manufacture, then no subsequent heat treatment is, necessary as in the case of the SAP-type alloys. Following this initial abrasive thinning technique, the material is then given a final electropolish which is just sufficient to remove any oxide which has formed on the metal surface during the abrasion technique. Care has to be exercised in this technique to make sure that none of the attendant problems associated with the chemical or electrochemical removal process are encountered. An electrolyte solution which has been found suitable for aluminum-aluminum oxide SAP materials is one containing 78 ml of perchloric acid, 120 ml of distilled water, 700 ml of ethanol, and 100 ml of butylcellosolve. The current density used is 1 amp per sq cm and a polishing time of 3 sec on each side of the foil is used. Extreme care must be used in choosing the current density and time to ensure that just the oxide film on the specimen is removed. Times or current density which are greater than this will lead to perforation of the film due to the removal of the A1,0, particles or a complete loss of the film due to complete solution of the matrix. Once the oxide film is removed the specimens must be kept in alcohol until placed in the microscope to prevent any further oxidation. Fig. 1 shows -an electron micrograph at X28,000 of a thin foil of an aluminum-aluminum oxide SAP-type alloy prepared using this technique. The alloy consists of a matrix of commercial purity aluminum containing a very finely dispersed second phase of aluminum oxide flakes. In the electron micrograph, the dark patches are the aluminum oxide particles. The lighter portions of the micrograph are the aluminum matrix. Within the aluminum matrix, the dislocation structure of the alloy is clearly visible. This structure consists of intersecting twin boundaries and single dislocations. No abrasive particles were observed in the samples. In addition, thin films of commercial purity aluminum were prepared using this method and observed in the electron microscope. Here also, there was no evidence of abrasive particles introduced in the samples. The technique presented provides a satisfactory method for the production of thin films of aluminum-aluminum oxide SAP-type alloys for use in transmission electron microscopy where none of the usual thinning techniques appears to be satisfactory.

Jan 1, 1962

By Jean Howard

ThE formation of the (100) [001) texture in 3-1/4 pct Si-Fe strip was first reported by Assmus ef a1.l in 1957. Since then much experimental work has been carried out with a view to establishing the mechanism involved. The papers cited above state that the (100) [001] texture was developed in strip rolled from material melted and cast in vacuum. (The impurity content of the ingot is reported as 0.005 pct.) The present note records that similar results can be obtained in material processed by powder metallurgy. A processing schedule is described.which enables the texture to be formed in strip up to 0.010 in. thick, but there seems no reason why this should not be achieved in thicker strip, provided that large grains are developed after sintering. The materials were prepared from Carbonyl Iron Powder Grade MCP (particle size 5 to 30 p) of the International Nickel Co. (Mond) Ltd. The powder contains about 0.15 pct 0, 0.01 pct C, 0.004 pct N, <0.002 pct S, $0.005 pct Mg and Si, and 0.4 pct Ni— that is, it is substantially free from metallic impurities other than nickel, which is thought to be unimportant in the present work. The silicon powder was 99.9 pct purity, or material of transistor quality (ground in pestle and mortar). The mixed powders (3-1/4 pct Si to 96-3/4 pct Fe) are heated in hydrogen at 350" and 650°C to deoxidize the iron before sintering loose at temperatures between 1350" and 1460°C (depending upon the ultimate thickness of strip required) for up to 24 hr. The object of the high-temperature sinter is to develop a large grain size at this stage. Alternatively, the loose sintering can be done at a lower temperature followed by rolling or pressing and then annealing at temperatures between 1350" and 1460°C. Both methods produce large grains, which remain large throughout the process. The compact is then hot-rolled to approximately 1/8 in. with high-temperature interstage anneals if necessary. This step is taken to avoid intercrystalline cracking which would occur if the material of such large grain size were cold-worked. The bar is then annealed at 1050°C and reduced to its final thickness by approximately 50-pct reductions and 1050°C interstage anneals. Throughout the process the dew point of the hydrogen furnace atmosphere is maintained at about -40°C. Samples were annealed in hydrogen at various temperatures and times. Secondary recrystalliza-tion to (100) [001] was developed on the thinner material (i.e., up to 0.002 in.) by annealing in hydrogen at 1050" to 1200°C with a dew point of - 40°C or in vacuum (10-5 Torr) at 1050°C. With the thicker materials (i.e., up to 0.010 in.) the best results were obtained by annealing in hydrogen at 1200°C with a dew point of - 55°C. Complete secondary recrystal-lization to (100) [001] textures was obtained. Above these temperatures secondary recrystallization to (110) [001] tended to develop. The final annealing of samples was normally carried out with the samples placed between recrystal-lized alumina plates, but some experiments were performed with the samples suspended so that their surfaces were not in contact with anything except hydrogen, and these were equally successful in developing secondary crystals. An approximate determination of the proportion of material (before secondary recrystallization took place) having crystals with the (100) or (110) planes in or near the rolling plane showed that 11 pct of the sample had (100) and 16 pct (110). The method used for the determination is described below. A sample was annealed at a temperature just below the secondary recrystallization temperature and etched to reveal the (100) planes. The approximate area covered by crystals having (100) or (110) in or very near the surface was measured on the screen of a Vickers projection microscope. This was repeated for twenty positions chosen at random and a mean of the results calculated. The main hindrance to developing the secondary crystals in the thicker materials was the difficulty of obtaining a large enough initial primary grain size before secondary recrystallization. This was overcome by increasing the particle size of the silicon powder used. During the course of the work, it had been observed that the larger the grain size after sintering the more likely it was that the material would be successful in developing secondary crystals at a later stage. An experiment was therefore carried out to determine whether the material with the larger grain was more successful in developing secondary crystals due to the large grain produced at the sintering state per se or whether it was due to the greater reduction of silica brought about when the sintering temperature was raised in order to increase the grain size. A comparison was made between two compacts, one of which was made with silicon powder of 60 to 100 mesh, the other with silicon powder which was finer than 200 mesh. F?r this experiment, use was made of a phenomenon previously observed that the larger the particle size of the silicon powder employed in making a compact, the larger is the grain size of the compact. The silicon powder was ground

Jan 1, 1964

By Yun-tao Xin, Yu Yi, Kang Yan, Gang Li, Xiao-dong Lv

This study investigated the selective removal of arsenic from high-arsenic dust in alkaline systems and the effects of different leaching conditions. The results indicated that the liquid–solid ratio, NaOH concentration, and sulfur dosage had a significant influence on the process. The leaching efficiency of arsenic reached 99.57%, while that of lead was as low as 0.03% under appropriate conditions. In particular, the addition of sulfur can effectively promote the leaching of arsenic and reduce the dissolution of lead in the solutions. An artificial neural network was used to model the leaching process. It consisted of a back-propagation artificial neural network model with a “6–10–2” structure that could effectively simulate and predict the value with more than 99% accuracy. Based on the difference in weights of the different parameters in the neural network model, the relative importance of the parameters related to arsenic and lead leaching efficiency was obtained, which followed the order of NaOH concentration, liquid–solid ratio, sulfur dosage, temperature, time, and stirring speed.

Jul 29, 2021

By N. D. Policow, J. S. Orphanos

D.A. Dahlstrom In the paper, a comparison was made between flowsheets using the screen bowl centrifuge and disc filters, respectively, as the primary dewatering device. The authors very effectively show that by raising feed temperature, mechanical dewatering is enhanced through reduced viscosity of the water. When a final coal moisture of 10% is required, the net effect of heating the feed is a reduction in thermal requirements because of the lower evaporated load on the dryer. The flowsheet comparison on this, however, was not completely correct. Disc filters do not generally utilize enhanced mechanical dewatering through increased temperature. Horizontal belt filters, through the optimal use of steam heating of the filter cake, have shown significant cost reductions over disc filters where steam is not used. In addition, the authors forgot to mention that there is a certain fraction of the feed that is so fine that it cannot be recovered by the centrifuge economically. That is the function of the filtration device shown in Fig. 4. In subbituminous pipeline coals, this fraction represents more than 10% of the feed. Even after flocculation and belt pressing, the moisture content in the fine fraction is 30%-40. Recent tests on a subbituminous pipeline coal slurry, compared an Eimco-Extractor horizontal belt filter with the screen bowl centrifuge flowsheet shown in Fig. 4. Mechanical dewatering was enhanced through the use of a steam dry zone after the initial dewatering zone. Vacuum pump horsepower was reduced by more than 50% through the use of condensors, and heat recovery was utilized to minimize thermal costs. Thermal drying to 10% moisture by weight was required in both flowsheets. The 0.5 wt% fines in the filtrate were flocculated and recycled to the filter, thus eliminating the belt press and solids mixers. When the three major operating costs (which normally account for 50%-75% of the coal operating costs) of steam, electricity, and flocculants at current prices were calculated, the flowsheets utilizing the Eimco-Extractor had cost savings vs the centrifuge flowsheet of 89¢/t (81o; per st) of surface dry coal. This advantage increases with increasing thermal and electricity prices. I congratulate the authors for the strides made with the screen bowl centrifuge. However, filtration flowsheets continue to be improved and their simplicity and low thermal and electrical requirements should not be so quickly dismissed. ? Reply by N.D. Policow From a technical standpoint, a number of extensive test programs have demonstrated that by raising the feed temperature, mechanical dewatering is enhanced whether it be centrifuges or filters performing the separation. Economic constraints usually become the over-riding issue. If an inexpensive source of thermal energy is available, i.e., waste steam, etc., the approach becomes commercially acceptable. For the specific case of coal slurry pipelines, since the dewatering plants will be located close to power generation stations, the preheating approach becomes especially attractive. The cost comparison shown in Fig. 7, which was prepared by a major West Coast engineering firm, contrasted screen bowl centrifuges and disc filters. In both alternatives, a thermal dryer was specified to produce a contract moisture of the final product. Sizing of the thermal dryer on the screen bowl alternative was substantially smaller than that of the filter, since the screen bowl was credited with a four to six percentage point decrease in product moisture compared to the disc. This comparison was made at ambient temperatures and reflected solely the mechanical dewatering capabilities of the two devices. No preheating of the feed was included in this particular comparison. Early generation screen bowl centrifuges operated under the principle of maximum bulk tonnage throughput. Second and third generation machines have followed the trends dictated by the coal industry calling for higher one-pass recoveries than previously demanded. The current generation of high-speed, long-bowl centrifuges permit recoveries of ultrafine solids directly comparable to commercial filter operation. On typical coal preparation plant feeds with nominal 600 E. an X 0 (28 mesh X 0) clean coal, recovery levels in excess of 99% are now obtained. The material that does escape the centrifuge and reports to the main effluent is typically -10 µm, high ash material. We have documented substantial beneficiation effects in screen bowl centrifuge performance with the cake product containing lower ash levels than the feed. On the referenced pipeline, coal tests performed on Western subbituminous coal, we found that the percentage of -10 µm material in the feed was six to eight times greater than normally encountered on eastern Appalachian seams. Our 45-t/h (50-stph) production centrifuge was never set up to maximize recovery. It was used as part of the two-stage dewatering circuit that had been chosen by designers of the proposed pipeline. Recoveries up to 95% were recorded. With field adjustments, that performance could most certainly have been improved. In contrast, manufacturers of the horizontal belt filter provided a small pilot-scale unit that operated on a side stream at the test facility. Numerous combinations of chemical additions were evaluated to determine those that had the best impact on filter performance. Whereas the screen bowl centrifuges operated without chemical addition, producing a dry, fluffy cake, the belt filter required two stages of chemical addition to attain satisfactory performance and high recovery levels. On the two-stage circuit that incorporated a belt press as the secondary

Jan 12, 1983

I do not regard the criticism by Rankin of my work as valid or even meaningful. I strongly reject the assertion by Rankin that my results are invalid. As stated in my paper, the equation to solve (in e. m.u. ) is: [ ] where n2 =4 ps? . This may be rewritten as three scalar equations in Cartesian coordinates, one of which is [ ] The equation to be solved is the diffusion equation, and it is not appropriate to consider waves in this context. In the method a point-wise solution to the problem is approximated. As discussed in my paper the approach taken at discontinuities in conductivity is to choose the normal component of [E] at the discontinuity as the average of the normal components on either side of the discontinuity. This is similar to the well-known assumption made in obtaining Fourier series representations for piecewise continuous functions (Pipes, 1958; p. 51). In the geophysical problem this means that the discontinuity is approximated by a transition zone, which is a very good approximation for such cases. By using this approach, both the values of the function and its derivatives give a good estimate for those associated with the physical situation. It is not apparent and Rankin has not shown that "all the field components and the derivatives are grossly distorted in the region about the boundary" for geophysically realistic situations. He has given no evidence to support his assertation that the results are "invalid". Although Rankin declares that one cannot calculate the electric components from the magnetic, in his discussion he has not stated why this is so. After once calculating the magnetic components the electric components may be calculated, but one must insure that the initial solution is accurate enough to prevent significant round-off errors.

Jan 1, 1973

By Hermann Schmalzried, Carl Wagner

UPON heating a metal oxide or sulfide in H2, first only oxygen or sulfur is removed from the surface. Thus the metal/nonmetal ratio in the oxide or sulfide increases and the thermodynamic activity of the metal increases accordingly. For the formation of metallic nuclei, a certain super saturation is necessary, i.e., the activity of the metal must become greater than unity. Once nuclei of the metal have been formed, the supersaturation is supposed to decrease. Even under steady-state conditions of a reduction process, however, the activity of the metal inside an oxide or sulfide phase must be greater than unity in order to have a driving force for the individual steps of the reduction process. As an example, consider the reduction of Ag2S with a high mobility of cations and electrons. As shown schematically in Fig. 1, the reduction process involves the following steps:&apos; 1) sulfur removal along the entire Ag2S surface in form of H2S, 2) migration of Ag&apos; ions and electrons from the Ag2S surface to nuclei of metallic silver, 3) transfer of Ag+ ions and electrons from the Ag2S phase to the metal. For a finite rate of step 3, the activity of silver in Ag2S next to the Ag2S-Ag interface must be greater than unity. Moreover, for a finite rate of step 2, the activity of silver at the gas-Ag2S interface must be still somewhat greater. Theoretical considerations lead to the tentative conclusion that at least under steady-state conditions the supersaturation will be rather low.&apos; To test this conclusion, the average activity aAg of silver in a Ag2S tablet about 2 mm thick with a diameter of 5 mm was followed as a function of time by measuring the electromotive force E of the cell Pt I Ag | Agl I Ag2S I Pt where AgI is used as a solid electrolyte in which only Ag+ ions are mobile. The activity aAg is calculated as aAg=exp(-EF/RT) Since the diffusion coefficient for transport of silver in Ag2S is very high (- 3 . l0-2 sq cm per sec),3 the average activity aAg is representative for both the bulk and the surface of the Ag2S tablet. Experiments were conducted at about 400°C. The cell was surrounded by an aluminum block in order to secure a uniform temperature and to eliminate thermoelectric effects. With the help of a potentiometer, first a potential of 0.16 v was applied to the above cell in argon in order to obtain a well defined initial state with a Ag/S ratio close to 2.000.~ Then the cell was disconnected from the potentiometer and a gentle stream of H2 was passed along the Ag2S surface. In view of the removal of sulfur in form of H2S, the electromotive force E decreased within 20 to 30 min to zero corresponding to a Ag/S ratio of 2.003 and became negative, indicating a value of aAg> 1. A representative plot electromotive force E vs time is shown in Fig. 2. The minimum of the electromotive force characteristic of the onset of

Jan 1, 1963

By William Carton, Cyril U. Cooledge

BecauSe the rich La Colorada orebody of the Cananea Consolidated Copper Co. does not outcrop at surface, after its discovery (by churn drill) and before mining was begun, a large amount of development work was done to learn the shape and extent of the ore and to provide for economic and effective mining. As these openings progressed, the actual shape of the orebody became known and with this knowledge came the realization that mining this rich and unusual deposit without loss of ore or danger to workmen presented several difficult problems. This paper will describe one of these problems; namely, the handling of an inverted cone-shaped core of barren rock, on the periphery of and below which the ore occurred. It mas clear that as the ore was mined out from around and under this inverted cone of rock, weighing about 3,000,000 tons, it would hang over the workings, held suspended only by its own tensile strength and threatening to crash into the stapes below with disastrous results to men and mine. The method of mining developed during the years 1928,1929 and 1930 anticipated and provided for the eventual collapse of this core of rock. In November 1936 the core settled down gradually on the edges of the stopes, descending a total of about 50 ft. vertically. The men in the stopes had ample warning and no lives were endangered. There were no air blasts in the mine. The core broke loose from its top support at the 500-ft. level along old fracture planes and evidently is crushed to the ninth level. Movement was of short duration; the following shift went to work in the stopes as usual. No ore was lost. Geological Structure The shape of the orebody is illustrated in Fig. 1. The geology of the Colorada orebody has been described by V. D. Perry1 and by W. G.

Jan 1, 1940

By John J. Collins

Page Purpose and scope..................... 521 Core sampling......................... 521 Sludge sampling....................... 527 Combining core and sludge assays ...... 533 Deviation and surveying of drill holes. . . . 537 Conclusions........................... 544 Acknowledgments...................... 545 Bibliography.......................... 545 Purpose and Scope The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described,25,26,31,32,50,65 but more informa- tion is needed on the results shown by actual mining operations. Core Sampling The accuracy of diamond-drill-hole Sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says; in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change; often indicated by a color change in the sludge water, should be a a for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

By Cyril U. Cooledge, William Carton

BecauSe the rich La Colorada orebody of the Cananea Consolidated Copper Co. does not outcrop at surface, after its discovery (by churn drill) and before mining was begun, a large amount of development work was done to learn the shape and extent of the ore and to provide for economic and effective mining. As these openings progressed, the actual shape of the orebody became known and with this knowledge came the realization that mining this rich and unusual deposit without loss of ore or danger to workmen presented several difficult problems. This paper will describe one of these problems; namely, the handling of an inverted cone-shaped core of barren rock, on the periphery of and below which the ore occurred. It mas clear that as the ore was mined out from around and under this inverted cone of rock, weighing about 3,000,000 tons, it would hang over the workings, held suspended only by its own tensile strength and threatening to crash into the stapes below with disastrous results to men and mine. The method of mining developed during the years 1928,1929 and 1930 anticipated and provided for the eventual collapse of this core of rock. In November 1936 the core settled down gradually on the edges of the stopes, descending a total of about 50 ft. vertically. The men in the stopes had ample warning and no lives were endangered. There were no air blasts in the mine. The core broke loose from its top support at the 500-ft. level along old fracture planes and evidently is crushed to the ninth level. Movement was of short duration; the following shift went to work in the stopes as usual. No ore was lost. Geological Structure The shape of the orebody is illustrated in Fig. 1. The geology of the Colorada orebody has been described by V. D. Perry1 and by W. G.

Jan 1, 1940