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By E. Bratygin, G. Gazaleeva

"The article describes the analysis of methods and criteria of selectivity of optimal ore preparation schemes. Traditional criteria based on study of physical-mechanical properties of ores give the comparison of obtained figures with a reference standard or statistic data and the selection of the ore preparation scheme is made on a basis of a similar scheme from the existing production.Further on estimation is given to different criteria of the crushing selectivity, the new criterion is offered, the definition of which is done by results of an optical microscopy– the estimation of a valuable mineral release degree. The final estimation of the crushing process selectivity L? is performed by a calculation of a ratio of the valuable mineral free grains quantity in a desired size grade of a crushed material to the valuable mineral free grains quantity in a desired size grade of a raw product.Besides, optimality assessment of using the specific crushing unit is preliminary performed according to a maximum crushing degree of a material, however, the priority is given to the maximum selectivity criterion. The check of the proposed criterion was done when comparing the crushing of niobium bearing ores of the Vishnevogorskoye deposit in cone inertial and centrifugal impact crushers. Different operation modes of crushers were also checked. The comparison of the crushing selectivity criterion L? showed that the centrifugal impact crusher (CC) has the value of the selectivity criterion lower than the cone inertial crusher (CIC). For the crusher CC the criterion L? is 0.82 – 0.87 units, whereas for the crusher CIC – 1.43-1.755 units. The best crushing mode for the crusher CIC-300? is as follows: turning angle of debalance is 120°, peripheral speed of a cone rotation – 1,420 rpm, specific energy consumption – 6.99 kW?t/t.INTRODUCTIONIt has been shown that the most efficient techniques of mineral raw materials disintegration are selective crushing enabling to produce liberation of particles at maximum use of destruction techniques by borders of minerals cleavage. These days, such techniques are usually called special disintegration techniques (Jankovic & Valery, 2015), and they are used in 3–4 crushing stages. The selection of optimal crushing techniques is performed based on the following main parameters:- degrees of crushing, calculated based on granulometric features of raw and crushed products;- selectivity criterion defined according to useful mineral release and slime ability degree; and- traditional criteria (Gazaleeva et al., 2015) of choice based on study of physical-mechanicalproperties of ores gives the comparison of obtained figures with a reference standard or statistic data and the selection of the ore preparation scheme is made on a basis of a similar scheme from the existing production."

Jan 1, 2016

By Leonard Harris, Rosa H. Harris

"It is becoming increasingly obvious to mining and exploration companies, when operating overseas, that their activities should not conflict with the interests of the surrounding communities. The world's largest gold heap leach operation, Minera Yanacocha has installed community assistance programs in Peru concerned with health, education, electrification, agricultural assistance and road improvements. The programs are conducted with joint venture partners consisting of the company, the community people, the contractors and suppliers, government and non government organizations and a ladies association composed of the employees' wives and female staff.IntroductionThe l 990's is the decade highlighted by enormous investments in mineral exploration and mining projects in the developing world. Figure 1 illustrates the distribution of exploration spending by mining companies in 1995."

Jan 1, 1997

By B. J. Moore

This publication contains routine analyses and related source data for 374 natural gas samples from 24 States and 5 foreign countries. These samples were collected during calendar year 1966 as a part of a continuous survey of the free world for occurrences of helium in natural gas. This survey has been conducted in the United States by the Bureau of Mines since 1917. In late 1965, the survey was expanded to include foreign gasfields. The analyses of foreign samples contained in this publication represent the first samples received as a result of this expanded program. The analyses published herein were made by mass spectrometer and a special helium analysis apparatus, which are described in Bureau of Mines Bulletins 4863 and 576.4 Eight previous publications have been made to present the results of 7,244 gas analyses made? by the Bureau of Mines Helium Activity prior to 1966. The first three, Bulletins 486,5 576,6 and 617,7 contain analyses and related

Jan 1, 1967

By Abdolreza (Reza) Osouli

Determining the roof stability condition of underground coal mines in the Illinois Basin is sometimes very challenging. The roof stability of Illinois mines are mostly affected by two types of shales: Anna Shale and Energy Shale. These layers are sometimes weak in parallel bedding when subjected to horizontal stresses. The roof rocks in the Illinois Basin are are typically more moisture sensitive than the other coal fields. In this study the geotechnical properties of some of the shale units as well as limestone layers will be presented. The geotechnical properties of roof rocks consist of moisture content, unconfined compressive strength, slake durability, indirect tensile strength, and point load tests (Axial and Diametral). The correlation of indirect tensile strength versus diametral point load strength as well as unconfined compressive strength versus axial point load test are explored and discussed. These correlations will help in identifying the roof mass index and planning appropriate roof control measures. This study also evaluates the application of CMRR rock mass index for an Illinois coal mine.

Jan 1, 2014

By Brenda B. Mehnert

Long-term monitoring is necessary to successfully develop predictive subsidence models. Thus, the Illinois Mine Subsidence Research Program (IMSRP) has documented subsidence over a longwall mine in south-central Illinois throughout a 3-year period. No monitoring results of this duration have been previously published for longwall mining in Illinois. To study the effects of surface subsidence and the rate of subsidence, precise level surveying was conducted at the site two or three times a year. Level surveys were conducted at approximately the same time of year using a WILD NA-2 level with a micrometer. Second-Order, Class-I, closure errors were obtained in each survey. Survey monuments were designed to minimize frost-induced movements. At this particular site in southern Illinois, subsidence over chain pillars and the panel centerline continued 3 years after mining. Since mining in 1988, 6.67 feet of subsidence or 70% of the extracted height has been documented over the panel centerline, with almost 0.30 feet in the last two years. The area over the chain pillars has subsided a total of 1.44 feet, with 0.44 feet occurring within the last 2 years. Monitoring over the chain pillars and at the centerline shows the residual subsidence to be fairly uniform. Monitoring is continuing over both of these areas. Long-term monitoring data can be used to improve evaluation of subsidence effects on land and structures. However, coal companies will time mitigation activities depending on other factors including the threshold of structural damage, type of repair needed, and maintenance of landowner and company relations.

Jan 1, 1986

By Son Hoang Nguyen, Thanh‑Hai Pham, Duoc Van Tran, Dung Thi Kim Nhu

Apatite ore plays a critical and irreplaceable role in fertilizer production and is being upgraded in quality in northern Vietnam. The old tailing of the Bac Nhac Son beneficiation plant (Vietnam) has a significant apatite amount in the form of coarse unliberated particles. These millions of tons of tailing ore have 6–8% of P2O5 content, which can be retreated to provide fertilizer factories. This work presents the results obtained in recovering apatite from this tailing in the laboratory. By the combination of the processes of classification, preliminary flotation, re-grinding and cleaning in the flotation column, an apatite concentrate of 30.68% P2O5 can be obtained with the total recovery of 37.76% from the initial tailing of 7.56% P2O5. The study results not only provide an appropriate technical solution for recovering the apatite ore from tailing ore but also have outstanding meaningness in the recovery of mineral resources under the context of a circular economy.

May 24, 2023

By Joseph D. D?Annunzio

The report on cut and cover construction as assembled by the Organization for Economic Co-operation and Development covered 17 different countries based on a total of just under 400 individual replies of which about 30% of the replies were from the United States. In general, the review of replies to the questionnaire led to the impression that cut and cover construction is a very inadequate method which needed many improvements. This unfavorable judgment, however, resulted in the manner in which questions were asked rather than in the method itself, It can be concluded that the cut and cover construction method has proved itself as an excellent method and that it has reached a high degree of technological perfection, but inadequacies stilt exist, and for purposes of complete evaluation, comparisons with mining tunnelling methods should be carried out. The principal problem in cut and cover construction is the interaction during construction between the construction process and the existing use of the surface and subsurface environment. The greatest impediments to cut and cover construction have little to do with the physical works being constructed, but more with creating a site in which the work can be performed. Utility relocation, traffice maintenance, and protection of existing structures are major problems for the construction contractor, and too frequently receive insufficient attention in preconstruction planning and design. Some OF the areas in which research should be undertaken are as follows:

Jan 1, 1970

By L. Mukka

The existing main mine ventilation system for the production areas to the Kiruna mine was constructed during the period 1994-1996 and extended with 4 more shafts, including fan stations, for the sea ore body from 1997-1998. Two papers have been presented from this project at the 7th US Mine Ventilation Symposium in 1995 and at The 6th International Mine Ventilation Congress in 1997. In 2005, pre-studies for a new main haulage level were started. Different haulage methods were presented including ventilation systems. The pre-studies showed that the existing ventilation system, including fan stations for the new main haulage system in the production area, could be extended. However, a new system for the central area, which was not in the initial investment program, would be needed. In the beginning of 2006, a decision was made to design a new ventilation system for the central area to make it possible to get air to the development work for the new main haulage system, independent of the haulage method. This system was operational by the beginning of 2007. The new main haulage system, which was decided to be located on level 1365 m, only requires that the existing shafts be extended - the capacity of the fans seems to be enough. However, the problem is that the existing level 1045 m and the new level 1365 m have to be in operation in parallel for approximately 7-8 years. The challenge for the moment is determining how to ventilate the production areas for both levels at the same time, while simultaneously constructing the extended system. This paper will describe the problems associated with the extension of the main ventilation system and some potential solutions for these problems. The new main haulage level 1365 m must be in operation by 2012.

Jan 1, 2009

By J. N. Van der Merwe

"A major shortcoming of the statistical back-analysis of coal pillar strength is that it relies on the asmined pillar dimensions, not taking into account time-related pillar scaling with subsequent reduction in pillar width. The paper describes an investigation where the rate of pillar scaling was applied to the pillars in the databases of failed and unfailed pillars. The reduced pillar sizes were then used in the same method of statistical back-analysis that had been used in the past.This resulted in an equation for pillar strength which predicts significantly greater pillar strength than the previous statistical analyses, which is similar to the strength which had been found earlier by direct testing of large specimens underground. It is concluded that the lower strengths found by previous statistical analyses are due to the incorrect pillar width being used and that by adjusting the pillar sizes to compensate for scaling, more credible results that correspond to direct testing are obtained.As the newly derived strength equation is time-dependent, it follows that the safety factors and probabilities of failure are likewise time-dependant. An equation to determine the probability of failure is also developed. This results in the safety factor being close to unity at a probability of failure of 50%, which is aligned to statistical expectation. IntroductionSince the Coalbrook mine disaster, caused by massive pillar collapse, there have been several attempts at analysing and defining the strength of coal pillars. It was realized in the very beginning that performing classical strength tests on coal specimens in a laboratory and then transferring those results to real coal pillars could not be successful. The scatter of results is one problem and, more seriously, there is a size effect (at least for small specimens) that made the transformation of strength results to much larger coal pillars underground all but impossible.Two main schools of thought then emerged. Some researchers preferred to perform direct strength tests on large specimens underground, while others relied on statistical back-analysis of failed and unfailed pillar cases, attempting to determine the differences in size and shape that would allow a satisfactory prediction of failure. The latter school of thought resulted in the method that was widely accepted by the South African coal mining industry, described in the landmark publication by Salamon and Munro (1967)."

Jul 1, 2019

By J. A. Bakken, T. Magnusson, G. Sævarsdottir

This paper describes measurements of current and voltage done on submerged arc furnaces. The aim is to focus on and attempt to separate from one another the electric arc current and the current bypassing the arc through the charge. Toward this goal two measurement campaigns on AC-arcs burning on coke beds are ac-counted for. The first was done on a 150 kVA pilot furnace at SINTEF in Trondheim Norway, and the second one was done on a 48 MVA industrial furnace run by Icelandic Alloys in Iceland. The measurements were compared to simulations employing an advanced MFD arc model. The main results are that a coke bed does not necessarily behave as a linear conductor. However the warmer chemically transformed part of an operating Si-metal or FeSi furnace probably exhibits linear behaviour and passes a considerable fraction of the phase current, perhaps on some operating conditions, as much as 50%.

Jan 1, 2007

By J. P. Harrison, M. Grenon

"This paper summarizes the results of numerical studies on drift stability in moderately jointed rock masses. It investigates the relative pertinence of geomechanical factors such as joint orientation, spacing and stress regime. This is undertaken using a distinct element model and a series of parametric studies. IntroductionIn evaluating the stability of drifts in jointed rock, it is often necessary to compromise by employing models that capture the complexity of the rock mass, but which are still simple enough to allow for their routine implementation. The distinct element method can arguably provide an eloquent method to accomplish this. This paper explores the applicability of the method by means of a parametric study aiming to identify the dominant factors that influence the stability of drifts in moderately jointed rock. To this end a commercially available 2D distinct element numerical model was used. A series of techniques are presented in an effort to quantify the results of any instability predicted by the model.This paper also demonstrates a case study where a 3D joint network model was integrated into the distinct analysis software package. The advantages, as well as the current limitations of this approach, are discussed."

Jan 1, 2001

By J Stipcevich, N Patel, J Faithful5, S Narendranathan, L Tatnell4

ENGIE, the current owner and operator of the Hazelwood Open Cut coalmine, are in the process of rehabilitating the mine site in preparation for mine closure. Bucket wheel excavators (dredgers) used during coal mining operations are in the process of being demolished, which involves the controlled felling of dredger components using explosives. With the mine located in close proximity to the Morwell township and other critical public infrastructure, it is crucial that the use of explosives and the possible implications for mine stability is well understood. The felling of Large Mining Equipment (LME) will apply an impact load immediately to the underlying coal benches with the adjacent coal batters potentially at risk also of adverse slope stability impacts. This paper presents the 3D stability modelling methods employed to successfully assess the potentially adverse stability implications associated with LME collapse.

Jun 21, 2022

By Liping Gu, Gordon A. Irons

"A 1/3 scale ""thin-slice"" model of an electric arc furnace was developed so that gas could be injected to simulate foaming gas evolution. The influence of arc jets, oxygen lancing and bottom bubbling was also simulated. Liquid velocities were measured by a white light particle image velocimetry system. Numerical simulations of some of these effects were made using commercial CFD packages. Comparisons with measured results show that considerable care is required to select the appropriate model and to interpret the computational results.INTRODUCTIONDuring recent years electric arc furnaces (EAF) have captured a significant share of steel production, partially due to increased productivity through the introduction of advanced technology such as oxygen injection, slag foaming and long-arc operation, process automation and post-combustion. However, systematic analysis of the process and these effects on heat transfer and fluid flow in the furnace is still in its infancy. Most investigationsl1 6l have been restricted to a specific application such as oxygen lancing or bottom stirring. In addition, it is very difficult to make comparisons between these computed results and experimental/industrial results. In this work a 1/3-scale ""thin-slice"" model of Dofasco's 190 ton EAF was built. The fluid flow in the bath due to the arc jets, bottom stirring, oxygen lancing and CO evolution was simulated. The velocity distributions in the bath were measured using Particle Image Velocimetry (PIV) System. In addition, numerical simulations of bottom stirring and CO evolution were carried out using two commercial CFD packages (Fluent and Phoenics). The objective of this work is to validate the mathematical models so that they can be applied to the full-scale furnaces."

Jan 1, 1999

By Bruno Nordberg

THE high cost of coal in Butte and the development of large amounts of cheap electric power from the Missouri river caused the Anaconda Copper Mining Co. in 1908 to make an investigation as to the possibility of adapting this cheap electric power for the operation of their many large hoisting plants. At this time electric power had already been adopted for practically all other power purposes around the mines, such as pumping, compressing air, etc. The map (Fig. 1) shows the location of the mines of the Butte mining district. These mines are located within a circle of approximately a 1-mile

Jan 9, 1913

By Carl E. Reistle

ENGINEERS as a group are often criticized because they have been responsible for the development of many technical improvements only to allow the administration of them to pass into the hands of others who may have little understanding of how they operate and little competence in their use. These criticisms are not always justified. There are many engineers who have shown themselves to be great administrators in addition to being great creators or builders, and more and more one sees the engineer demonstrating unusual administrative ability as he moves into the executive ranks. The last five or ten years have seen a new movement developing in this country-a movement brought about through the recognition by business statesmen that successful business administrators and executives just cannot be expected to develop like Topsy, that the management of a company must deliberately take steps to cultivate the proper climate or environment for growth if it wishes to provide an adequate and capable succession of executives in the future. This movement has taken on a number of names-Management Development, Executive Development, Personnel Development-but are these not actually misnomers? Can we develop anybody? In the strictest sense of the word, the author doubts that anyone can change any other person, at least, not much. The change must come from within. Perhaps this movement should be looked at not as a program of developing people but as one of giving people an opportunity to develop themselves. In the Humble Oil & Refining Co. there is an Executive Development Program; maybe it might better be called a Program of Development for Executives.

Jan 8, 1954

By D. Lelinski

Air-sparged hydrocyclone (ASH) flotation is one of the new promising technologies. Numerous applications have been successfully tested since its conception; nevertheless, research and development efforts have continued to provide for further improvements with respect to operating conditions and design considerations. Results from experimental studies on bubbles generation in the ASH system are presented in this paper. Bubble size distributions generated during ASH operation have been determined using a high-speed photographic technique. The influence of many factors including surfactant concentration, water flowrate, and porous tube pore size has been studied. Test results indicate that these process variables have a profound effect on the bubble size distribution. With an increase in surfactant concentration or water flowrate (shear field) the bubble size distributions become narrower and shift toward smaller average bubble diameters. However, the influence of pore size is more complicated. Depending on the experimental conditions the average bubble size was found to range from about 100 to 300 microns in diameter.

Jan 1, 1995

By Arthur F. Wendt

The hot-blast oven of which the accompanying plate gives complete details, was designed by the writer for the spiegel-furnace of the Lehigh Zinc and Iron Co., at Bethlehem, Pa. Members of the Institute who visit these works will have occasion to notice the large amount of zinc-dust in the gases, and the consequent dirtiness of both the hot-blast and the boilers. The zinc collects in a crust on the iron hot-blast pipes and leads to the frequent burning of them. A large vertical standing-pipe stove, a modification of the "Pleyer," thus gave considerable trouble by the burning of pipes and, on one occasion, the toppling over of a whole row. These inconveniences are avoided in the hanging-pipe stove under consideration, and less heating-surface has been found to be required to heat the blast to a given temperature. The construction of the stove is so simple and so plainly shown in the plate that it is needless to enlarge upon it. Each pipe rests in a box and is connected by U-branches with the next adjoining pipes. An iron sleeve is placed around the joint and tightened, as usual, with iron filings. In case of a pipe burning out, that particular section is cut off by the valves and the sleeves are broken. The pipe is then readily lifted out of the box and a new one is substituted. The whole operation requires only a short time and the section can then be at once put into use. The boxes supporting the pipes cover the whole surface of the hot-blast and are closed by iron covers rammed with fire-clay. Owing to the air-space around the U-pipes there is little loss of heat in the upper part of the stove. Each section as built has 280 square feet of heating-surface. The furnace is blown by a horizontal engine with 46-inch stroke and 48-inch diameter of blowing-tube. At 28 to 30 revolutions, the engine delivers 3000 cubic feet of air per minute, at a pressure of 5 pounds. When five sections of stove are in use, the blast enters the furnace at a temperature of 850" F. That is, there are over two cubic feet of air per square foot of heating-surface heated to that temperature per minute.

Jan 1, 1887

By B. Burger, T. Bester, S. Deonarain, L. S. Schultz

"This paper details the process followed to raise drill and equip a 6.1 metre diameter vertical service shaft at Glencore’s Mindola Deeps project in Kitwe, Copperbelt Province, Zambia. The project scope comprises a shaft collar, bulk air intake, a pre-sink to -90 metres and raise drilling the shaft from surface to a final depth of 1,892 metres underground, in four separate lifts of approximately 475 metres each. Equipping of the shaft steelwork, headframe erection and commissioning the shaft for hoisting. In addition, the scope included the development of a decline down from 5430’ level underground to the new shaft position on 6205’ level. This paper will provide some insight into the successes and challenges encountered during execution of the shaft works to date and provide some learnings for future projects of a similar nature. INTRODUCTION The Mindola Mine is situated to the west of the city of Kitwe in the Copperbelt Province of the Republic of Zambia (Figure 1). The mine is owned and operated by Mopani Copper Mines Plc (MCM), whose shareholders are Glencore AG (73.1%), First Quantum (17.9%) and ZCCM Holdings Plc (10%).Productivity at the Mindola mine has been adversely affected because of significant ventilation and ore re-handling challenges, which are caused by the dilapidated infrastructure, and design of the existing vertical and sub vertical shaft system. In the circumstances, it was apparent that it was necessary to sink a new vertical shaft that would increase the volume of fresh air down the mine, improving the underground working environment and the current hoisting restrictions, resulting in an increase in productivity and efficiency. Because there were two existing levels ideally situated and perpendicular to one another and because a third level was in close proximity to that perpendicular and because the mine required a ramp down to shaft bottom in any event, consideration was given to raise drilling the shaft as opposed to using conventional blind sinking methodologies."

Jan 1, 2019

By G. T. Lapidus, O. J. Solis-Marcial

Currently, ores with high-grades of copper oxides and secondary sulfides have become increasingly scarce, which has encouraged industry to look for alternative raw materials to obtain copper. The problem is that primary sulfides are difficult to dissolve at near ambient conditions. Proposed hydrometallurgical processes have not been implemented beyond a pilot or demonstration scale. One of the principal disadvantages is that the majority of these processes are not viable with ores that contain less than 4% copper. One of the more recently proposed techniques involves the use of organic solvents in the dissolution of copper concentrates with a chalcopyrite content of between 20 and 30%, which has shown high efficiency. The present research explores the ability of this process to leach copper from low grade chalcopyritic ores (<4% Cu) at different temperatures ranging from 20 to 50°C. It was found that depending on the temperature, high percentages of copper were extracted in 40 to 100 minutes, using different oxidants. Additionally, distinct leaching behavior was observed with each oxidizing agent. In some cases, the extraction kinetics adjusted to the diffusion controlled shrinking core model.

Jan 1, 2014

By W. A. Tiller, J. D. Harrison

HOT pressing of powder particles has gained importance recently, since it affords a method in which high densities are rapidly attained. In a recent study on hot pressing of alumina powders, Mangsen, Lam-bertson, and Best1 have shown within the limits of their experiment, the validity the rate equation for densification in hot pressing, originally proposed by Murray, Livey, and Williams.2 i.e., where D = fraction of theoretical density, t = time, P = pressure and 77 is the viscosity, all in appropriate units. This equation was derived from earlier work of Shuttleworth and Mackenzie3 which is based on the closing of spherical pores under the action of surface tension. The integrated form of this equation yields a straight-line relation between log (1 - D) and t. Such conditions were found by Felten4 to be true only after extended periods of pressing. In an endeavor to quantitatively evaluate the exactitude of this equation in the initial sintering stage, i.e., prior to sealing of the pores, spherical anti-

Jan 1, 1962