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By S McQuitty

The metallurgical treatment of Falconbridge ore re-quires an extremely fine grind of the pyrrhotite portion ? for maximum liberation and recovery of nickel. This paper describes the operation of a grate discharge pebble mill in closed circuit with cyclones regrinding 25 to 30 tons per hour of sulphides to approximately 90 per cent minus 325 mesh. The consumption of the screened ore grinding medium is of the order of 2.5 pounds per kw.-hour, or less than 1 per cent of the tonnage of re-ground sulphides. IN RECENT years, there has been a considerable interest in pebble grinding in the uranium industry, and a number of installations have appeared. However, the use of sulphide ore pebbles as a grinding medium has so far been relatively limited. There has also been some question as to their suitability for the very fine grinding of sulphide concentrates. This paper describes the preliminary results of a commercial operation involving the

Jan 1, 1964

By Seth Robertson, Alex Ciccone, Lawrence Johnsen, James Davis

The most sustainable foundation is the one that is already in place. That doesn’t mean it is the easiest to RE-use. This was realized when replacing a multi‑story hospital parking garage in Hartford, Connecticut. The previous garage was constructed in the 1980s and was supported on grade beams with 3.0‑ and 4.5‑foot‑diameter drilled shafts that extended through 100+ feet of problematic soft Connecticut Valley varved clay, glacial till, and end‑bearing in or on arkose bedrock. The proposed garage consisted of relatively high loading demands, was susceptible to excessive settlements, and was also located adjacent to an active hospital with sensitivity to vibrations, thereby limiting many foundation design options. An innovative solution was proposed to re-use many of the existing drilled shafts, supplemented with a combination of micropiles and composite piles. Composite piles were comprised of 3.0‑foot‑diameter drilled shafts with a central, 11.75‑inch‑diameter micropile to transfer structural loads to bedrock. Installation information on the existing drilled shafts was heavily limited. Dynamic load tests were therefore performed on 12 existing shafts spread across the garage footprint, confirming geotechnical resistances up to 4,500 kips and significantly improving quality assurance of the existing foundation construction. Additionally, two micropile load tests were performed in tension to 900 kips, confirming design loads could be resisted by the bond strength in rock. The foundation re‑use methodology saved over 3 million dollars in installation costs, weeks of construction schedule, and resulted in a significant carbon reduction compared to conventional, and inherently more substantial, foundation replacement strategies. Critical aspects of the foundation re‑use design strategy and limitations of alternatives will be discussed. Implementation of innovative load testing techniques are addressed, focusing on optimizing quality assurance for utilizing existing deep foundations in modern construction. Lessons learned and recommendations for future practice are presented.

Oct 7, 2024

By B. Vorobyov

A Canadian company, Sight Power Inc., has developed a revolutionary, fully automated shaft monitoring and inspection system for deep-level mining based on a 3D mobile laser scanning technology. It consists of a special hardware unit and proprietary software with Artificial Intelligence elements for object recognition, features identification and change detection. The main purpose of this system is to significantly improve the safety of underground mining as well as to increase the available hoisting time thus generating substantial additional revenue for the mine.

Aug 1, 2013

By M. Fornaro, E. Lovera, I. Sacerdote

"The analysis of the mining laws of the different European States reveals big diversities and, often, remarkable complications. This is due, firstly, to the history and culture of different States and, secondly, to the assets of natural resources. It is therefore difficult to find a common line between the text of the various mining laws, but it can be observed that nearly all the States distinguish the mining resources in two categories: the materials whose importance is considered strategic belong to the first category; in the second category the materials whose importance (above all economic) is mainly local are comprised. The paper, analysing in detail the planning procedures in some important Countries, investigates the possible perspectives of arranging some guide-lines for a correct approach to the sustainable exploitation planning/programming of mineral resources (with particular reference to "quarry" materials, such as aggregates), considering both the environment protection and the production needs."

Jan 1, 2005

By Leonard B. Gulbransen, John R. Lewis, W. Martin Fassell, J. Hugh Hamilton

T. E. Leontis (The Dow Chemical Co., Midland, Mich.)—This paper is of particular interest to me because of my own work with F. N. Rhines on the oxidation of magnesium and magnesium alloys a few years ago. The authors are to be complimented on their development of an accurate and reproducible technique for measuring ignition temperatures and on their comprehensive study of the many variables that affect the ignition temperature of magnesium. It is indeed gratifying to see that they have obtained a good correlation between ignition temperatures and the oxidation rates reported by us. The correlation is valid not only with composition within one alloy system but also between alloy systems; that is, alloying elements which effect the greatest increase in oxidation rate also produce the greatest decrease in ignition temperature. There are a few points upon which I would like to comment. In attempting to correlate ignition-temperature data, one must be sure that the same definition of this quantity is used by all investigators. It does not appear to me that such is the case in the authors' comparison of their data with the theoretically calculated values of Eyring and Zwolinski. The equation derived by these investigators defines the ignition temperature, To, as the temperature at the gadoxide interface, whereas the present authors use the metal temperature as the criterion for ignition. The contradiction in the effect of oxide-scale thickness on ignition temperature between the predictions of the Eyring-Zwolinski equation and the observations reported in this paper indicate that some variable has not been taken into consideration. Could that be the geometry and size of the specimen? There is a marked difference in the type of specimen used in this investigation and that used in our work which formed the basis of Eyring and Zwol-inski's theoretical treatment. Another factor which plays an important role in ignition is the vapor pressure or the rate of vaporization. Combustion can safely be assumed to take place in the vapor phase by the reaction between vaporized magnesium and oxygen. Thus, a more accurate theoretical analysis may be made on the basis of the rate of vaporization which may be the controlling rate of the process. The effect of a large number of alloying elements on the ignition temperature has been reported in this paper, but beryllium was not included. Practical experience dictates that beryllium markedly decreases the burning tendency of magnesium. I was wondering if the authors plan to study the effect of beryllium in their future work. The authors predict that concentrations of sulphur dioxide in the furnace atmosphere greater than 5.8 pct would be expected to increase the ignition temperature to values still higher than those they measured. I would like to mention that large concentrations of sulphur dioxide markedly increase the rate of combustion of magnesium once ignition has started. Although it has been shown in the paper that the ignition temperature of magnesium in oxygen increases with increasing sulphur dioxide content up to about 1 to 2 pct, in practice relatively low-melting commercial cast alloys (AZ63A and AZ92A) are being continuously heat treated at temperatures just below the melting point in air containing 0.5 to 0.75 pct SO*. In regard to the change in color of the oxide scale observed on magnesium and magnesium alloys just prior to ignition, I would like to mention that in our work alloying elements were found to color the usually white magnesium oxide even though ignition did not occur. For example, the oxide formed on Mg-A1 alloys was gray, increasing in intensity with aluminum content in the alloy. Finally, I might suggest that the authors indicate their source of the value of 0.8 g per cc for the density of MgO as it is formed on magnesium upon oxidation at elevated temperatures. W. M. Fassell, Jr. (authors' reply)—The comments by Dr. Leontis are very excellent ones and I will attempt to answer them in order. First, the problem of ignition of magnesium is a rather difficult one since many factors are involved. Concerning the comparison of the To in the Eyring-Zwolinski equation, eq 4, with the experimentally determined values, it will be noted that the calculated and experimental values of the ignition temperature in Table I are not self-consistent. In the case of the 1.78 pct A1-Mg alloy the calculated value is 49°C below the experimental value; for the 3.81 pct A1-Mg alloy, 122°C below the experimental value; for Mg with 5x10-' cm film, 19°C above the experimental value; for Mg with 2x10-I cm film, 28 °C below the experimental value. Thus, if it were merely a matter of difference of location of temperature measurement the calculated ignition temperature would always be below the experimental value, the difference being due to the thermal gradient through the oxide film. The possibility of a thermal gradient in the magnesium metal must be considered. From Carslaw and Jaeger,'Y t can be shown that the maximum temperature gradient that could exist between the oxide-metal interface and the center of the sample is of the order of O.Ol°C. The geometry and size of the specimen could certainly have some effect on the ignition temperature. The equation for ignition that has been proposed in reference 14 is of the following type containing terms to account for this and other factors: M dT AHv(T) =Cp--------------\-J(.T—TB) + ZAHl-M A dx where AH is the heat of reaction, v(T) is the velocity of the reaction at temperature, Cp is the heat capacity of sample, M is the mass of sample, A is the area of sample, t is time, J is the total coefficient of heat transfer outward from the reaction zone, TR is the temperature of the bath or furnace, and AH,, is the heat associated with any phase change involved. Prior to the instant of ignition, the vapor pressure of magnesium is of no special significance. After ignition, neither eq 4 nor the above equation is applicable. The actual combination of magnesium cannot safely be assumed to take place in the vapor phase. While experimental data is lacking to support a hypothesis that ignition does or does not occur in the vapor phase, some observation on the pressure ignition experiments may be of interest. At high oxygen pressures, once ignition has occurred, the reaction of magnesium with oxygen approaches near explosive violence, the entire sample being consumed in probably less than 1 sec. At atmospheric pressure it usually requires 15 to 20 sec. Thus it appears that the oxygen concentration becomes the rate determining factor. Further, if burning magnesium is observed through darkened glass (Lincoln Super-visibility Shade No. 12) the magnesium sample is very much hotter than the "smoke" and the outline of the sample is retained perfectly. No "flame" is visible above the metal. No work was done on Mg-Be alloys. We do, however, intend to study this problem in the near future.

Jan 1, 1952

H. H. Stoek, Scranton, Pa. (communication to the Secretary): Mr. Rothwell's condemnation of my paper on the International Correspondence Schools as not giving an impartial view of the whole field of such enterprises, because it was confined to the description of a single one, might be legitimate if I had undertaken or professed to do anything else than that which I was requested by the Secretary of the Institute to do—namely, to give an account of this particular institution. This is a sufficient answer to all his direct or implied criticisms upon my omission to mention either persons or organizations outside of the special limits set for me. Recognizing fully that personalities have no place whatever in the Transactions, I regret the necessity of making the following statement in answer to Mr. Rothwell's charge that I omitted to give due credit to Mr. Ewald, a former principal of the International Correspondence Schools. The Colliery Engineer School of Mines was started in 1891, and its first student was enrolled October 16, 1891. During the fall of the same year the mine-mechanical course was laid out along the same general line that had already been followed in the mining course. This mine-mechanical course was the foundation from which sprung the later developments of the several mechanical courses. The latest circulars of information of the International Correspondence Schools give the following as their distinctive features: (1) courses of instruction for particular occupations,

Jan 1, 1900

By R. S. Olson

An analysis of the wash thickener-two stage filtration circuit was undertaken to quantitatively describe the circuit performance. On the basis of information developed in this study, the factors of major importance in reducing soluble loss are: filter wash, wash zone size, filter capacity, filter feed pulp dilution, blow back, filter cloth, water to repulp, water to primary wash, water to feed pulp and thickener underflow. The wash thickener-two stage filtration circuit has been used in many hydrometallurgical plants to separate solute from leach solids. As many as 20 to 30 plant operating variables determine the performance of the wash circuit and hence the soluble loss. The complex relationships of these variables, with each other and with soluble loss, make the evaluation of their effect on mill performance difficult. An analysis of this type of circuit was undertaken to quantitatively describe the circuit performance. THE CIRCUIT The diagram in Fig. 1 represents the operation of the wash thickener-two stage filter circuit. This circuit provides for counter current, cross current or for a combination of countercurrent-cross current operation. The feed pulp stream, J, is intended to include any solution or water, other than primary and secondary filtrate, which enters the wash thickener. Except for filter washing, the operations in the circuit are simple mixing operations which can be described by material balance. The concentration of solute in the filtrate and cake solution after application of wash liquor cannot be described by simple material balance. The displacement washing equation, derived by Rhodes1 and extended by Choundhury and Dahlstrom,2 appears to best describe this case of miscible displacement in porous media. In a useful form, it is

Jan 1, 1963

By G. R. St. Pierre, S. W. Gilby

The equilibrium vapors established over Fe-Cr-Ni alloys containing up to 50 pct Cr at 1600°C were determined by collecting the effusate from thoria Knud-sen cells. From these data, the activities of iron and nickel in the ternary alloys have been calculated. In a separate set of measurements, a mass spectrometer was used to determine the activities of the three components in a series of Fe-Cr-Ni alloys. In binary Ni-Cr melts, the activity coefficient of chromium referred to pure solid chromium as standard state is less than one at low chromium contents but is greater than one for melts containing more than 30 at. pct Cr. For Ni-Cr melts containing less than 10 at. pct Cr, In ?Cr = 0.15-0.92 N2. Ternary activity diagrams for Fe-Cr-Ni alloys are presented. In iron-rich alloys, the interactions between the three components are negligible; however, in nickel-rich alloys, the three principal interactions have significant positive values. THE activities of components in binary Fe-Ni alloys have been measured both in the solid solution1-3 and liquid solution4-6 temperature ranges. The determination for Fe-Ni melts have been made by three different methods: transpiration,4 Knudsen-cell effusion with analysis of an accumulated condensate,' and Knudsen-cell effusion with direct measurements of relative component intensities by mass spectrometric methods.6 The results obtained by the three different methods in three different laboratories are in excellent agreement and considerable confidence is placed in the activity coefficients for Fe-Ni melts. At 1600°C, the activity coefficients in dilute solution are ?°Ni = 0.66 and ?oFe = 0.40 for the pure liquid components as standard states. Measurements made on solid solution Fe-Cr alloys7-10 indicate that the component activities are slightly greater than ideal. The measured activities in iron-rich Fe-Cr liquid alloy11-16 indicate almost ideal behavior for this system. The recommended activity coefficient of chromium in dilute solution in liquid iron at 1600o C, ?oCr, is 1.18 when referred to pure solid chromium as standard state and 1.04 when referred to pure undercooled liquid chromium as standard state." The conversion factor of 1.14 is based on a melting point of 1857°C and an enthalpy of melting of 4047 cal per g at. for chromium.21,37 The activity coefficient of iron in dilute solution in liquid chromium has not been measured. Activity measurements on solid Ni-Cr alloys have been made by several investigators.17-20 During the review of this manuscript, Fruehan12 published the results of a study on liquid Ni-Cr alloys at 1600°C. He used galvanic cells22 with ZrO2(CaO) as the electrolyte and Cr/Cr2O3 for the reference electrode. His results indicate that ?oCr in nickel at 1600°C is about 1.01 when referred to pure solid chromium and about 0.89 when referred to pure undercooled liquid chromium as standard state. The activity coefficient of nickel in liquid chromium is unknown. The compositions of vapors freely evaporating from some molten Fe-Cr-Ni alloys have been measured by Lyubimov, Granovskaya, and Berenstein.23 Their results, combined with known vapor pressures, indicate strong positive deviation for chromium in iron-rich alloys and strong negative deviation for chromium in nickel-rich alloys. EXPERIMENTAL PROCEDURES This research was undertaken to determine the activities of components in ternary Fe-Cr-Ni melts containing up to 50 at. pct Cr. At the initiation of the research, we decided to use a Knudsen-cell method wherein a portion of the effusing vapor is condensed on a cold plate.' The composition of the equilibrium vapor for each alloy composition is determined in this method by holding the system under constant conditions for a period of time sufficient to collect enough condensate on the cold plate for chemical analysis. The activities of components in ternary alloys can be calculated from such vapor composition determinations by Eq. [1].24

Jan 1, 1970

By T. A. Lang

For permanent structure underground, where rock is not competent, support usually consists of concrete or reinforced concrete. However, temporary supports in the form of timber or steel are often needed during construc-tion. Although the history of rock bolting is relatively short—50 years— their use has become widespread in general engineering construction as well as mining. This profusely illustrated and detailed study covers a broad range in the field of rock belting: from the behavior of rock and bolts including rock properties, masses, and structure; mathematical formulae for rock bolt applications; through analysis of various operations and descriptions of the bolts themselves. Rock construction is one of the oldest of the engineering arts and its origin is lost in antiquity. From the days when early man decided that he wanted to improve the natural caves which he was using for shelter and protection or made a river crossing by placing rocks to form a ford or causeway, we have had structures made of rock. It is not too much to say that rock in situ as a structural material forms part of every major engineering undertaking. Rock bolting is one means whereby the inherently good characteristics of rock in situ are preserved and used to the best advantage and the bad characteristics ameliorated. In many cases the latter are accentuated by construction processes used. Rock bolting, as with other rock construction techniques, is only just beginning to emerge from being an art. Consequently, its theory and practice is still more descriptive than mathematical. ROCK BOLTING In underground excavations, where the rock is not competent, support is provided. For the permanent structure, this generally consists of concrete or reinforced concrete. However, support may be needed during construction before the concrete can be placed, and conventionally this consists of timber or steel support in the form of ribs, struts, and lagging. Alternatively, rock bolts may be used. Although their use dates back over 50 years, it is only in recent years that rock bolts have become widely used, not only in mining but in general engineering construction. A rock bolt is a steel bar which is inserted in a hole drilled in rock. The end away from the rock face has a device which permits it to be firmly an- chored in the hole and the projecting end is fitted with a plate which bears against the rock surface. The bolt is placed in tension between the anchor and the plate, thereby exerting a compressive force on the rock. The essential feature of a rock bolt is that it is placed in tension. This distinguishes it from anchor bars which are grouted into holes in rock, but which are not prestressed. The difference between an anchor bar and a rock bolt when the rock bolt is grouted in is analogous to the difference between the reinforcement in ordinary reinforced concrete and in prestressed reinforced concrete. The view that rock bolts only pin or nail blocks or slabs of rock which are loose to the sounder rock behind them is erroneous. Rock bolts are useful for this purpose and have been so used for a long time. However, the term rock bolting, as used here, means the designed use of rock bolts to reinforce and develop the rock around an excavation into a structural entity which can competently play its part in a structure such as a powerhouse or a mine installation. Rock bolts behave quite differently than steel ribs. They can be installed at the working face directly after blasting and within a short space of time can be exerting a stabilizing pressure on the loosened rock surface. This early installation not only partially restores loosened blocks of rock to their original unloosened positions, but also it prevents the gradual relaxation or loosening of the decompression zone behind the new rock face. In contrast, steel ribs generally use timber blocks, wedges, and lagging between the ribs and the irregular rock surface. The timber is relatively compressible, and loosened blocks of rock must move outwards an appreciable distance before any load builds up on the steel ribs. Also, the ribs may settle as the foot blocks and foundation become compressed. Hence, it may take several days or weeks before the rock has moved sufficiently to make the steel ribs carry an effective load.

Jan 1, 1961

By B. R. Richards, B. G. W. Robinson

"This paper shows how metal values can be profitably won from a comparatively small and remotely located but high-grade orebody. To do so is not uractical for a large mining company, with its necessarily high overhead of services and organization essential to a large producer.To evaluate, develop and profitably exploit a limited but defined ore deposit becomes only a problem of determining the dollars required to win the dollars available. This paper will record how this was carried out during 1965 and at the same time describe the history, development and production plant of this property."

Jan 1, 1966

By Mohammad Noaparast, Hamed Haghi

"In this research work, it was attempted to reduce the iron contents (Fe2O3) in relatively low grade silica sample which was obtained from Shenin silica mine, using magnetic separation. The prepared sample contained 93.75%SiO2, 0.44%Fe2O3, 2.78%Al2O3, and 0.08%TiO2 and the Fe2O3 had to be reduced to less than 0.08%. Mineralogical studies and XRD analysis indicated the presence of Hematite, Pyrite, and Goethite as the major iron impurity phases. Mineralogical studies of polished and thin sections demonstrated that the degree of freedom of silica is more than 80% in -300+212 ?m size fraction. After primary stages including crushing, milling, scrubbing and de-sliming, various methods such as low intensity magnetic separation (LIMS), high gradient magnetic separation (WHGMS) and dry high intensity magnetic separation (DHIMS) were applied to find out that which process would be appropriate for the elimination of iron impurities from this sample. Different experiments were carried out on the product of wet rod mill and scrubber. However the magnetic field intensity of WHGMS was increased down to 2Tesla, and there was not any acceptable result in reduction of Fe2O3 content lower than 0.08%. Thereafter the product of wet rod mill and scrubber were dried and reported to DHIMS. In these experiments, maximum magnetic field intensity was 1.4Tesla. In this case very significant results were achieved, and Fe2O3 content of final product were reduced to less than 0.05% which meets glass manufacturing standards."

Jan 1, 2014

By E. Larsen

Marine tailings disposal may in some cases be the best solution to one of mining’s major environmental problems. The public standing of this solution is poor, partly because of past bad practices. Among the tailings problems spreading of super-fines and toxic substances stand out. The two are often interrelated as heavy metals, insoluble to sea water, tend to adsorb on fines and spread piggyback on them. Only one paper seen, has addressed the mineralogy of fine particles in suspension. The topic was a secondary tailings deposit. Studies on the sedimentation of fines in large river estuaries show that different minerals sediment sequentially as the salinity of the river increase. Mineralogy in the particle size range where particle separation takes place, has so far, naturally been in focus for the mineral processors. Next to nothing is known about the fines which may cause environmental problems. Tailings disposal has, due to its complexity, to a large extent been taken over by specialists in other fields, biologists, engineering geologists, chemists. The basics of the subject ought also to be given more input by the mineral processing community. A drawn out battle about tailings discharge into some fjords has taken place in Norway. The detractors focus at fine particles (could be dangerous nanoparticles) spreading and emphasize the effects. Calculation is simple: 5 million tonnes tailings annually discharged, containing 1 wt. % < 1µm determined by laser diffraction, results in 50 000 tonnes of nanoparticles let loose in the environment. The industry`s difficulty to answer, is lack of basic research on this topic. Test work to develop a methodology to address the problem is described. One thing the mineral processors know is that final tailings may be comprised of tailings from separate separation processes, with different properties, some of which may require special treatment. One tailings stream from a Norwegian mine is investigated in such a context.

Jan 1, 2016

By C. L. Wilsey, K. M. Oakes, W. J. Reich, M. J. Protani

The time frame required to complete the National Environmental Policy Act (NEPA) process for a new mine or the expansion of an existing mine can range from nine to 24 months, or even longer. The time frame will vary depending on whether an environmental assessment (EA) or an environmental impact statement (EIS) will be completed Completing properly designed baseline studies to collect natural resource data to support the EA or EIS prior to the start of the NEPA process can shorten this time frame. In this paper, a two-year natural resource baseline study by Homestake Mining for the Ruby Hill Mine, Eureka County, NV, is used as a case study to describe how a properly designed and executed natural resource baseline study streamlined the NEPA process.

Jan 1, 2003

To develop more accurate means of measuring low ventilating air velocities in underground mines to facilitate compliance with ventilation standards. The Approach: A commercially available vane anemometer which produces an electrical, rather than a mechanical, output was coupled to electronic circuitry to produce the Bureau&apos;s Digital Low Velocity Anemometer (see photo). [ ] How It Works: The experimental instrument has a vane anemometer head which produces an electrical output signal with a frequency directly proportional to the rotating speed of the vane anemometer shaft. The commercially available ABBIRKO FLOWMASTER anemometer head was used in this instrument. In the signal conditioning circuits the electrical signal from the anemometer is converted into pulses, which are counted during either 3-second or 1-minute intervals, depending on the position of the operator&apos;s control switch. (Readout is either "instantaneous" or averaged for a period of one minute.) The number of pulses counted is then translated (integrated and divided by a constant) in the logic circuits to get average speed. Average air velocity in feet per minute is displayed in illuminated digits. (See simplified diagram of circuitry.)

Jan 1, 1975

By Scott W. Hazen

This report describes the design, analysis, and results of a special experimental mine-sampling project conducted as part of the Bureau of Mines continuing program on sample-methods research. The experimental project was conducted at the Bureau?s Denver Mining Research Center and at the Climax molybdenum mine within the framework of a cooperative research agreement between the Bureau and Climax Molybdenum Co., Division of American Metals-Climax, Inc. The primary objective of the project was to investigate the interrelation-ship between sample volume, number of samples, the standard deviation of an assay distribution, and the precision required for an estimate of grade of ore, and to develop this interrelationship into a relatively simple method of application for use in mine sampling.

Jan 1, 1962

By Malcolm Ross

Fibrous minerals are common in nature but asbestiform minerals are rare. The unique mineralogical characteristic common to all the asbestos minerals is their morphologic form (or habit of crystallization) as polyfilamentous fiber bundles. The individual fibrils within the bundles have a tendency to be very long with a narrow range of diameters and grow with their long fiber axis in parallel orientation to the bundle length. The asbestiform habit imparts to the asbestos minerals sufficient flexibility and tensile strength so that most can be woven into cloth. In the past application has focused on their ability to insulate against the transfer of heat. However, these minerals possess other properties which make them useful in many industrial applications. © 2007 Elsevier Inc. All rights reserved. Keywords: Asbestos; Asbestiform; Fibrous

Jan 1, 2007

By N. Schmalbein, L. von Ketelhodt, L. Kotelo

"The mining industry continues to remain challenged. Lower commodity prices, increasing labor costs and more importantly diminishing reserves are amongst some of the primary factors contributing to the difficulties faced. Ore sorting technologies have demonstrated their potential to improve the downstream beneficiation performance. The general processing route sees all primary crushed material milled for liberation and then processed in accordance with characteristics of the gold therein, i.e.: refractory golds would require direct leaching or oxidation whilst free gold would follow the conventional cyanidation process, after gravity concentration. Ore dilution increases the associated operating costs within each stage of this process. The opportunity to process the same throughput at a higher grade can only be achieved through pre-concentration. Sensor based sorting has been examined on numerous occasions by various gold producers. Test work carried out on different bulk low grade gold ores has shown the successful applicability of ore sorting to remove waste, whilst maintaining high recovery rates of gold. The results of a combined XRT/Laser sensor sorter have shown grade improvements of over 300% from feed grades of 0,25g/t to 1,15g/t at gold recoveries of 66% and waste ejection rates of 85%. This paper presents some of the results from the work undertaken by STEINERT with respect to gold sulphide ore sorting and presents a modelled financial scenario, which can enable a company to target and reclaim low grade ores.INTRODUCTION The mining and minerals processing industry presents with many different challenges including, volatile commodity prices, increasing labor costs and more importantly diminishing reserves. Conventional processing routes will continue to play a crucial role in the beneficiation process, however innovative methods of reducing operating costs and optimizing on the overall circuit performance is where technologies such as sensor based sorting come into the fore. Sensor based sorting as a dry mechanical process not only reduces water consumption by eliminating unnecessary waste but also reduces the requirements for reagents in the downstream processes. Sensor based sorting under dry conditions presents a highly efficient alternative to wet pre-concentration methods such as dense medium separation (DMS)."

Jan 1, 2017

The Mount Kasi gold deposit in Vanua Levu represents an epithermal, high sulphur, gold-enargite system occurring within a volcanic caldera of Middle to Upper Miocene age. High level breccia pipes flare upward and form flat lying, near surface, mushroom shaped sheets parallel and near to the present land surface. The upper portions of the pipes contain explosion breccias which are capped by sandy siliceous sinter-like beds and are cut by alunite veins. Outlying intercalated scoriasceous lavas and breccia sheets show evidence of subsidence within a shallow basin. Detailed mapping of breccia types, intensity of silicification and the presence of multiple quartz veining and enargite abundance outlined blocks of more consistent gold mineralization within the breccias. This work has provided a model for mineralization which will aid future exploration and ore reserve estimations. A significant eluvial concentration of gold occurs in soils overlying the breccia sheet.

Jan 1, 1987

By Daniel J. Rider

In 1980 the Niagara Frontier Transportation Authority contracted for three miles of cut and cover box structure and station construction in Main Street, Buffalo, New York. Project depths ranged from 30 feet to 60 feet below street grade. Soils, Figure No. 1, consisted of loose to very compact sands and gravels. Ground water level was generally 20 feet below grade. Specifications required a soldier pile and timber lagging support system with external dewatering. Soldier piles were to be installed in prebored holes with bentonite slurry and lean concrete backfill. Conventional pile driving was not allowed due to concerns over excessive noise and vibrations and its effect on nearby historic structures and the surrounding population

Jan 1, 1987

By A. M. Thyer

Throughout the history of mining, underground fires or explosions have killed many hundreds or thousands of people. Fortunately, the safety record has gradually improved as a result of many people’s work on improving safety standards and work practices. However, fire and explosion still remains a threat in many situations. As a result the UK Mines Inspectorate commissioned a project to study materials used underground and to develop a fire/explosion risk assessment system. Such a system has been developed, following the study of incidents, specified fire resistance criteria, and fire safety practices. The final methodology uses a series of check sheets to identify potential fuels and ignition sources, preventative measures, fire/explosion mitigation systems, and items for inclusion in emergency plans.

Jan 1, 2003