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By D. -P. Georgescu

A native sulfur mineralization associated with pyrite and iron oxides containing 20% iron was quarried in Calimani Mountains, Oriental Carpathians. The wastes resulted from this exploitation activity, with a volume of 33 millions m3, were stored in four dumps while the wastes resulted from ore processing using flotation and the subsequent treatment were disposed of in a settlement pond. The investigations carried out in the entire mining area focused on the following two aspects: - the impact of environmental factors mainly on the hydrographic network; - the characterization all pollution sources, namely dumps, leachates and the quarry and the assessment of acidic drainage generation. The total sulfur content of the excavated ore was 3.8% while in dumps it ranged between 0.54 and 1.06%. Analysis of samples collected along the hydrographic network, draining the mining area, resulted in pH values ranging between 1.7 and 2.83 as well as high concentration of iron (4525 mg/L to 1.6 mg/L), copper (1.87 mg/L to 0.04 mg/L) and zinc (13.13 mg/L to 0.3 mg/L). Samples from the quarry, the dumps and the tailings pond were analyzed in order to deter-mine: - the maximum potential acidity (MPA); - the acid neutralization capacity (ANC); - the net acid production potential (NAPP).

Jan 1, 2005

By G. H. Thomas, M. C. O’Riordan, S. Rae

INTRODUCTION The British mining industry comprises 228 nationalised coal mines producing some 113 million tonnes of deep mined coal and employing 242,600 persons of whom 185,200 work below ground. There are in addition 174 small privately-owned coal mines employing 1500 persons underground and producing 1.5 million tones of coal. In the non-coal sector there are 108 privately-owned mines producing some 20 different minerals, including tin, wolfram, fluorspar, potash, rock salt, china clay, ball clay and fireclay. There are currently 4,100 persons employed in these mines of whom 2,346 are employed below ground. One mine producing potash employs 870 persons of whom 465 are underground workers. Since 1975 all matters relating to the health and safety of persons at work and of the general public who are exposed to industrial hazards have come within the purview of the Health and Safety Executive (HSE). The Mines and Quarries Inspectorate (MQI) is a constituent part of that organisation and enforces legislation applicable to all mines and quarries. The National Radiological Protection Board (NRPB) is a statutory and independent body that was established in 1970 to advance knowledge and provide information and advice about radiological protection. The NRPB has an agency agreement with HSE to provide advice and services relating to radiological protection, including protection against exposure to radon decay products in mines. There are as yet no statutory requirements controlling exposure to this source of radiation in Britain. Investigations for the presence of radon in British coal mines were begun in the mid 195Os. The concentrations found were of low order, typically around 2 pCi/l (Duggan et al., 1968). This was not surprising since big coal mines are well ventilated to control methane and airborne dust and it is now generally accepted that radon and its decay products do not constitute a health hazard in British coal mines. At one or two small mines where igneous intrusions were in close proximity to the coal workings, slightly higher readings were recorded. The MQI then commissioned NRPB to make investigations at metalliferous mines and in these much higher readings were detected. It was later agreed that the NRPB should undertake a survey of all non-coal mines in Britain to establish the extent of the problem. This survey was conducted mainly in 1973, (Strong et al.,1975) and follow-up surveys at various mines have since been conducted by the MQI. These investigations revealed that radon readings could be high in mines producing tin, haematite and fluorspar. Some small mines working other minerals in the same localities also showed high readings. More recently owners have been making environmental measurements at mines where there was some cause for concern. LEGISLATION AND RADIATION STANDARDS As a result of the survey work for MQI the Board in 1975 recommended an occupational limit of 4 WLM in a year from radon decay products together with a programme of radiological supervision (Strong et al., 1975). The MQI accepted the recommendation and advised employers and employees that the 4 WLM limit should be adopted. The limit corresponds to a concentration of 0.3 WL, equivalent to 30 pCi/l of radon-222 in equilibrium with the decay products, and was in accordance with an earlier recommendation by the International Commission on Radiological Protection (ICRP, 1959). ICRP recommendations are usually endorsed in Britain. It was recognised that some years would probably be required to achieve general compliance, and this is proving to be the case. Some confusion has been caused in Britain by the conflict between successive Euratom Directives on this and other standards of safety. As relative newcomers to the Community, we were not affected by earlier Directives, but the situation is different now. In 1959, the Commission of the European Communities provisionally recommended a maximum permissible concentration for radon-222 and daughters of 3 x 10-7 µCi/cm3, which corresponds to 300 pCi/l (CEC, 1959). The number was affirmed in 1962 (CEC, 1962) and reaffirmed in 1966 (CEC, 1966) but without any specification of the state of equilibrium. In 1976, the CEC again recommended 300 pCi/l, but on this occasion, the daughters were assumed to be present in the same quantities as in unfiltered air (CEC, 1976): this implies 40 WLM a year in poorly-ventilated mines. The formula is repeated in the latest version of the Directive (CEC, 1980), but the value is now said to be an interim one because it might not ensure compliance with the appropriate dose limit. ICRP (1955) recommended 300 pCi/l in equilibrium with the decay products in 1955 and reduced this value to 30 pCi/l for unfiltered air in 1959 (ICRP, 1959), endorsed it for equilibrium conditions in 1976 (ICRP, 1976), and held to it until 1980: these values imply respectively 40 and 4 WLM a year. The recent recommendation of the ICRP (1980) that the annual limit of intake for the decay products should be 0.02 J in a year corresponds to 5 WLM, but there is a rider that reduction should be made for exposure to gamma rays and long-lived airborne radioactivity. An allowance of 20% is suggested for uranium mines, implying 4 WLM a year for the decay products. The

Jan 1, 1981

By Guohua Bai, Lin Zhang, Shan He, Xuling Chen, Min Gan, Xiaohui Fan, Yanbin Tang

"For reducing the dust content in flue gas during copper smelting in the Ausmelt, technologies and mechanism of strengthening the pelletizing of copper concentrate were studied. The results show that adding bentonite, organic binder or composite binder can effectively enhance the strength of green pellet and reduce the rate of shedding powder in drying process, and composite binder works best. When adding 1 % of composite binder, the drop strength of green pellet increased from 1.9 times/P to 4.5 times/P, the rate of damage dropping from 2 m reduced from 50.12% to 13.08% and the rate of shedding powder reduced from 60.28% to 25.13% at drying temperature of 400111. In addition, the mechanism of binder absorbed in the surface of mineral was investigated by infrared spectroscopy. The results showed that composite binder absorbed in the surface of copper concentrate was in the form of chemisorptions, which made hydrophilic enhanced, while bentonite absorbed in the surface of copper concentrate was physisorption.IntroductionWith the technologies of copper smelting developing, the equipment of copper smelting steps into a large-scale and efficient level in recent years. As Ausmelt smelting is characterized by being efficient, environmentally-friendly and energy-saving, it is generally recognized by the enterprises of copper smelting [l]. The process of Ausmelt smelting starts with materials mixing, and then pelletizing. From this point onwards, pelletized balls are charged to the top of the furnace, undergoing drying and smelting reaction[2]. Since copper concentrate is fine-grained, directly adding into smelting furnace can make it blown out of the furnace together with the airflow upwards, rendering flue gas high in flue dust, which can cause the loss of concentrate greatly, and make smelting environment poor as well. The concentrate transported into the furnace and smelted after disc balling is conducive to improve the capacity of melting furnace during copper concentrate smelting, reduce the powder shedding rate and improve enterprise benefits."

Jan 1, 2013

By S. Sener

In this investigation, the changes in the grindability of both ulexite and colemanite related with the different structural modifications created by the heat treatment were examined with respect to work index (Wi). The effects of the heat treatment temperature and time on grindability index were also investigated. It was observed that, heat treatment significantly affected on the grindability of ulexite. The work index of original ulexite was found 7.11 kWh/shton and decreased to 3.5 kWh/shton level within 200-480°C for 60 minutes. On the other hand, the work index of original colemanite was found as 6.48 kWh/shton and no appreciable change was observed when heated up to 320°C. Thermally assisted comminution can be used to improve the grindability characteristics of ulexite. Besides, the different behaviour of ulexite and colemanite under heat treatment can be used for their separation. The artificial mixture of ulexite and colemanite could be separated with over 90% recovery by utilizing their different grindability characteristics by thermal treatment, semi-autogeneous grinding and screening.

Jan 1, 1999

By G. N. Goldberg, Robert A. Rapp

The steady-state kinetics and microstructures of simultaneous internal oxidation and external scale formation were investigated for the oxidation of Cb-Zr and Ch-Zr-Re alloys in pure oxygen at 1000°C. For each binary alloy, a constant rate of scale formation was observed at steady state; the internal oxidatton zone reached a steady-state, time-independent thickness. The dependence of the scaling rate on zirconium content exhibited a maximum at a low zirconium colttent; more concentrated alloys oxidized at a slower rate than pure columbium. This composition dependence of the scaling rate may be partially attributed to the effect of the internal oxidation process as an intenla1 sink for oxygen, with the formation of voluminous, and relatively impermeable ZrO2 precipitates. However, the morphologies of the ZrO2 internal oxide precipitates also affected the scaling rate. The internal oxide precipitates were distributed uniformly in the external scale. The steady-state thicknesses of the internal oxidation zones were not in agreement with those predicted theoretically from an idealized simple model. However, in this particular alloy systen1 the ideal model is not satisfied experitmentally. For ternary Cb- Zr - Re alloys with Nr10e = 0.02 or 0.05, the steady-state thicknesses of the internal oxidation zones were less than those for the corresponding binary Cb-Zr alloys. For ternary alloys with NRe(0) - 0.05, a more adherent and much less porous external scale was formed, and a reduction in the kinetics of scale formation was observed. A number of authors'-7 have reported linear oxidation kinetics for the reaction of pure columbium at 1000°C in air or in oxygen of about 1 atm pressure. A thin and tightly adherent oxide layer is found at the metal/oxide interface beneath a porous external layer of scale. Although the suboxides CbO and CbO2 represent thermodynamically stable phases at 100O°C, generally only Cb2O5* is identified as a reaction product in quenched specimens. Studies at 1000°C at lower oxygen pressures8,10 (Po2 - 10-4 Torr) show that both CbO and CbO2 do form and that CbO2 is a protective oxide. The growth of CbO2 results in parabolic kinetics until Cb2O5 is nucleated and grows. Apparently because of the large volume change associated with its formation, Cb2O5 is porous and does not serve as a diffusion barrier; linear kinetics prevail after the surface is covered with Cb2o5. On the basis of these observations, several authors5,9,10 have suggested that the linear oxidation of pure columbium at 1000°C in air or oxygen of 1 atm is limited by a Loriers-type mechanism,11 whereby the total rate or reaction is controlled by ionic diffusion through a thin layer of CbO2 which is maintained at a constant thickness throughout the linear oxidation. However, the concept of a Loriers model for a three-phase CbOlCbO2 Cb2O5 scale, with the only oxygen activity gradient across the NbO2 phase, is not self-consistent. Further, the suboxides CbO and CbO2 were not observed in the high-temperature X-ray diffraction study of columbium oxidation by Goldschmidt.8 From investigations in which the large PO2, dependence of the linear scaling rate was demonstrated, the importance of an oxygen adsorption or dissolution step has been suggested.4,5,7,10,12 Thus the mechanism for the linear oxidation of pure columbium remains quite controversial. Since further insight into the oxidation mechanism of pure columbium is not provided by this investigation, the authors wish to emphasize at the outset that their experimental results are not interpreted in terms of a particular rate controlling step for the oxidation of pure columbium. Previous investigations13-18 of the oxidation of Cb-Zr solid-solution alloys in air or oxygen at 1000°C suggest a remarkable dependence upon zirconium content. For an alloy of bulk zirconium mole fraction, N2r(0) equal to 0.05 or 0.10, the oxygen uptake is reported to be as high as four times greater than that for pure columbium;13,14 for alloys with NZr(0) 0.05 or 0.10 the oxygen uptake is reported to decrease with increasing NZr(0); to a minimum uptake of about one quarter that for pure columbium at NZr(0): = 0.50.13-15Since zirconium forms a more stable oxide (ZrO2) than the lowest columbium oxide (CbO), and since columbium exhibits a high solubility19,20 and diffusivity21,23 for oxygen, the internal oxidation of the zirconium component to ZrO2 is expected at a reaction front ahead of the advancing metal/scale interface. The external scale is then formed at the metal/scale interface by the inward migration of oxygen through the scale, probably through a series combination of molecular and ionic diffusion. Internal oxidation in conjunction with external scale formation has been investigated by Maak24 for the oxidation of dilute Cu-Be alloys in pure oxygen at 850°C. For cu-Be and many other binary alloys,24-27 very small, essentially uniaxial internal oxide particles are formed in the most dilute compositions; for somewhat more concentrated alloys. the internal oxide particles precipitate as platelets or needles. From pertinent solutions to the diffusion equation, Maak28

Jan 1, 1967

Local conditions at Hatches Creek are described, together with the mineralogy, mining and treatment of tungsten ores, with particular reference to procedures adopted by Pimer Mines, N.L.The Hatches Creek tungsten field covers an area of about 20 square miles in: Central Australia; 285 miles' N.E. of Alice Springs. Access is by rail to Alice Springs, then by 200 miles of bitumen highway to Barlo Creek and 8,5, miles of bush track to the field. Passengers can also reach the mines by... A semi-desert climate prevails Rain; generally in the form of thunderstorms falls in the summer months: December to March. Average annual rainfall is approximately 15 inches but there is great variation from year to year, occasional droughts being followed by torrential downpours. Vegetation is sparse except in the wide creek beds. Surface water occurs in rock holes some of which, spaced miles apart, carry permanent water...

Jan 1, 1955

By W. I. Duvall, D. E. Stephenson. Discussion by R. G. K. Morrison

R.G.K. Morrison (Chairman, Dept. of Mining Engineering and Applied Geophysics, McGill University, Montreal, Que. Canada) - The authors have introduced a subject, the intensive study of which has awaited scientific developments. These are now such as to justify a thorough review of our thinking on the causes and control of rockbursts. The formulae developed in the paper, the comparison between rockbursts and earthquakes in figure 1 and the energy data relating to explosives serve as an effective foundation. In the paper this writer is quoted as suggesting that energy of the order of 10 x 10 11 ft. lb. may have been released in some of the heaviest rockbursts in the Kolar Gold Field. He now hastens to add that such an estimate has no scientific validity. It arises only from a casual comparison of surface and underground effects of the medium heavy bursts, referred to in the paper and estimated by Jones at 3.0 x 1010 ft.lbs., with those of obviously greater magnitude. In any case, we have the estimate by Jones above, that by a team of seismologists 1 on the Witwatersrand of 10 x 1010 ft.lbs., that by Coates of 3.8 x 10" ft.lbs., also referred to in the paper, that by Blan-chard2 for the catastrophic rockburst at Springhill, N.S., of only 7.4 X 107 ft. lbs., and finally the range covered by Cook, referred to in the paper. It is probable also that data for the heavy rockburst at Merkers3 in the WERRA potash area of Germany are available as it was recorded on seismographs from Moscow to Spain. It is the writer's impression that some standardization is present in the estimates above for 10 x 10 10, 3.8 x 1010 and 7.4 x 107 ft.lbs., respectively. The latter, for Springhill, represents a relatively low order of magnitude. It is obvious that any energy estimate must be related to the local geometry of openings at the time of the event. It is also obvious that any explanation of the causes of rockburst must be capable of providing a wide range of energy. However, we must remember that the energy release is an effect indicating magnitude. The cause lies with the mechanism initiating the primary failure, generating the seismic activity, to which much of the damage is probably due.

Jan 1, 1967

By S. Klingmann, G. Blackwell

"Terrestrial photogrammetry has been used for many years to build three-dimensional pictures for analytical purposes. Photographs can include images from open pit mine walls, ground topography, mine buildings infrastructure or accident scenes. Analysis and co-ordinate estimation in early photogrammetric systems demanded the use of specialized and costly optical-mechanical stereocomparators. Few of these instruments could be interfaced with computers to provide databases, and all required trained personnel with good perspective vision to collect data.Digital photogrammetric software has replaced these older systems, and the digital video plotter (DVP) is a simple computerized system utilizing readily available PCs to quickly produce a stereoscopic model of any scene where an area of overlap exists in photographs taken from two different locations. Specialized cameras providing reference marks or grids on the photographs are still necessary for detailed analysis in three dimensions.This paper describes the application of DVP to the three-dimensional analysis of strike, dip and location of dominant fracture planes visible on an open pit wall slope, and the intersection of planar features which might reduce wall stability. Where reference survey points cannot be provided because access by personnel to the photographic scene is impossible for safety reasons, normal case terrestrial photogrammetry can provide the necessary control points for more complex aerial photogrammetry geometries.DVP recreates the geometry existing at the moment photographs are taken, providing the user with the ability to safely recreate, compare and analyze the structures determining the stability of open pit wall slopes."

Jan 1, 2003

By John Meuth, Brandon Fryman, P. E. Hoffman, Joshua Ph. D., P. E. Calnan, Josh Ph. D.

Quantitative Risk Assessment (QRA) can be specifically tailored for use by the commercial explosives industry to calculate risk to employees and the public from commercial explosives facilities and operations. This methodology has several advantages over the American Table of Distances (ATD) which is solely a Quantity Distance (Q/D) analysis. While the ATD has remained essentially the same for 100 years, the safety of commercial explosives in storage has increased immensely. However, the concepts behind QRA calculations are not as intuitive and even at the end of the calculations the values of risk must be compared to something. A few regimes for assessing an acceptable level of risk are discussed in this paper. Risk levels for various United States Federal entities and international agencies are highlighted and suggestions for acceptable risk criteria for individuals and groups are provided. Referencing risk levels from normal daily activities is also done in this paper to make relative the accepted risk criteria. This paper presents some relevant individual risks for everyday activities and some scenarios for group risk for comparison. Individual risk is a straightforward concept however group risk is more complicated. This paper aims to clarify these concepts and provide means to communicate them to the public.

Jan 1, 2019

By D. Tepavac, S. Simmonds

"The high virgin horizontal in situ stress field in the Sydney Basin and its impact on civil engineering projects is a well-known and accepted phenomenon found in significant literature (e.g., Pells, 2013). The prevailing high-stress effects depend on many factors — rock quality, tunnel orientation, proximity of geological features to tunnel crown, size and shape of opening, depth of excavation and stress magnitude. The behavior of the rock mass under high in situ stress conditions can cause stress fracturing and consequent dilation of the tunnel periphery, resulting in rock spalling at the tunnel crown/ invert or raveling of rock blocks on the tunnel sidewall. This type of failure is of a brittle nature and may create construction and safety risks during tunnel excavation, if it occurs behind the excavation face where ground support has already been installed. Therefore, the associated risks need to be managed during construction. This paper presents the design strategy adopted to mitigate these adverse tunneling conditions for the WestConnex Project to-date. Overview of the WestConnex M4 East and new M5 tunnel projectsWestConnex is one of the New South Wales government’s key infrastructure projects. This 33 km (20.5 mile) project aims to ease congestion and connect communities and is the largest integrated transport and urban revitalization project in Australia. It was a key recommendation of the State Infrastructure Strategy released in October 2012. It brings together a number of important road projects, which, together, form a vital link in Sydney’s Orbital Network. These road projects include a widening of the M4 east of Parramatta, a duplication of the M5 East, and new sections of motorway to provide a connection between these two key corridors. The WestConnex project includes a number of stages: Stage 1a: M4 Widening; Stage 1b: M4 East; Stage 2: New M5; and Stage 3: M4–M5 Link. The tunnel design referenced in this paper relates to Stage 1b and Stage 2 of the WestConnex Project (Fig. 1)."

Dec 1, 2018

By Harmon D. Nine

DISCONTINUITIES in the strain vs cycles to failure (S-N) fatigue curves have been reported for polycrystalline materials by Porter and Levy for copper,&apos; by Benham and Ford for mild steel,&apos; and by Williams and Stevens for 18/5 stainless steel.3 However, the considerable scatter of the points for polycrystalline materials have made a statistical analysis of the data necessary to confirm the discontinuities. By using oriented single crystals of high-purity material and low grown-in dislocation content, the reproducibility has been improved to a degree which makes statistical analysis unnecessary and clearly shows discontinuities in the S-N curves. A recent publication4 reports an abrupt change of slope in the S-N curve for ( 100) axially oriented copper single crystals fatigued in torsion. The reported discontinuities in the S-N curve, along with a different mode of failure in the two branches of the curve, have been interpreted as being due to two different amplitude-dependent mechanisms. Further work on copper single crystals of (111) axial orientation has now been completed which supports the existence of an amplitude-dependent change in the mechanism of fatigue. In the present study torsional-fatigue experiments were performed at room temperature in air on (111) axially oriented single crystals. The crystals were grown and prepared as reported in Ref. 4. Two different methods of growing the crystals also discussed in Ref. 4 produced crystals of two different perfections: Type I with dislocation density of >l06 per sq cm and Type I1 with dislocation density of <l05 per sq cm. In accordance with the accepted hypothesis that crystal flaws are primarily responsible for the scatter of fatigue lifetimes, particularly at low amplitudes, the values for the three lowest amplitudes for Type I crystal are believed to be uncertain. They are marked in braces in Fig. 1. On the other hand, at amplitudes above a surface shear strain (SSS) of *0.003, the data for both types of crystals are in agreement. Most remarkably the S-N curve for torsional fatigue of (111)-oriented single crystals is composed of three segments separated by rather large discontinuities, Fig. 1. Metallographic and X-ray measurements indicate that the lower branch corresponds to the low-amplitude (F) region of the (100) copper single-crystal results as discussed in Ref. 4, and the two upper branches correspond to the high-amplitude (H) region. The lower discontinuity in the (111) S-N curve and metallographic evidence give further support to the hypothesis that different mechanisms of failure operate for high and low strain amplitudes. Reference to the resolved shear stress diagram, Fig. 2, shows that the highest stressed systems in

Jan 1, 1965

By G. H. Babcock, D. A. Knight, V. G. Medina

"Production commenced at the Golden Sunlight Mine in 1983 and at Kidston Gold Mine in 1985. These projects were both constructed under the management of the Placer Dome Inc. (Technical Services) group of Vancouver, B.C.Golden Sunlight uses a conventional rod mill/ball mill grinding circuit while Kidston uses a SAG /ball mill, and later on, SABC circuit. Since the hardness of the ore and the fineness of the primary grind in both mines are about the same, a meaningful cost comparison can be made. The comparison is made on the basis of capital and operating costs.The result of this study indicates that a significant capital cost savings can be achieved in favour of the SAG /ball mill circuit when the ore is amenable to semi-autogenous grinding. With a good maintenance program, the operating cost for the SAG /ball mill circuit could also be lower that the conventional rod mill/ball mill circuit.HISTORYGolden Sunlight MineGolden Sunlight is a 5 530 t/d gold mine located near Whitehall, a small town 60 km east of Butte, Montana. Originally optioned by PDI U.S. in 1958, the development of the property waited until 1981 when economic conditions became favorable for the project to proceed.Semiautogenous grinding tests were conducted during the final stages of the feasibility study. The hard near surface ore was not quite amenable to semi-autogenous grinding. The best result produced from a series of tests using a 1.83 m diameter mill was a semiautogenous work index of 39. This is substantially higher than the 22 Bond work index of the sample used in the testwork. A conventional SAG/ball mill circuit for this ore will be extremely inefficient, and for this reason, this circuit was rejected in favour of a conventional rod mill/ball mill circuit."

Jan 1, 1991

By A Karrech, E Hamdi

The uncertainty of geomaterials properties is commonly encountered in resource engineering applications, especially in mining engineering where the rock mass texture, joints, fractures and damage zones can impose the mining method and operation cycles. In the present research work, we focus on the elastic properties of rock masses. We use stochastic numerical simulation of fractured rock masses to estimate their Young moduli. The procedure starts with field mapping campaigns which are conducted to evaluate the locations, dip angles, strike and trace lengths of pre-existing fractures. We develop statistical law parameters of the natural fracture network to represent the weakness patterns. In addition, we use HLA-Dissim, a recently developed and validated Matlab program, to simulate the propagation of monochromatic longitudinal waves and derive the equivalent elastic modulus of fractured rock masses. Such procedure applies within a simulation domain, which is greater than an Elementary Representative Volume.At a second stage, we conduct finite element simulations to validate the previous approach. The matrix and the actual fractures zones are reproduced in the model to mimic the natural set-up. The block is excited on particular positions and its response is evaluated on predefined spots. Real field rock fracture mapping data and monitored single-hole blast induced vibrations are used to validate the theoretical and numerical approaches discussed herein. Comparison between the simulated and measured vibration records shows a great agreement which confirms that the proposed methodology has a full potential to predict and optimise multihole/multidelay blast-induced vibrations.CITATION:Hamdi, E and Karrech, A, 2015. A methodology for rock mass characterisation to control blast-induced vibrations, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 89–96 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By J. H. Zietsman

Ilmenite smelting furnaces must be operated with a freeze lining of frozen slag to protect sidewall refractories from chemical attack by liquid slag. The process has two main control objectives: Firstly the slag TiO2 content must be controlled according to specification, and secondly a safe freeze lining thickness must be maintained. This text only concerns itself with the second objective. Freeze lining thickness cannot be measured directly due to conditions inside the furnace. It is therefore inferred from signals of thermocouples installed into the furnace sidewall. These signals are used as the &apos;controlled variables&apos; in the freeze lining thickness control loop. The dynamic characteristics of these signals are the focus of this text. During this study a model of the process was constructed to study interactions between the slag bath and the freeze lining. This model was used to conduct simulation-type experiments to study the influence of various parameters (temperature, slag composition, ilmenite feed rate, reductant feed rate, power input, etc.) on the slag bath, freeze lining, metal bath, furnace wall and other parts of the process. In the case of the furnace wall, the modelling results were first plotted in the form of time-dependent isotherms. The dynamic thermal behavior of the furnace wall became evident from these results. To illustrate the dynamics of sidewall thermocouple signals, the results were also plotted as time lags between events occurring in the slag bath, and the time when a thermocouple registered a temperature change. The events studied included sudden changes in slag bath state (composition and temperature), changes in operating setpoints (ilmenite feed rate, reductant feed rate, power input), and operational errors (complete loss of reductant or of all feed). All of the result sets showed significant lags between occurrences in the slag bath and temperature changes in the refractory wall. The delays are such that it appears impractical to rely on sidewall thermocouple signals for warning of impending damage to the refractory wall, or for purposes of actively controlling freeze lining thickness. Some measures are suggested to improve the usefulness of these signals during furnace operation.

Jan 1, 2005

By P. R. Clifford

Phosphogypsum is a byproduct of the manufacture phosphoric acid. There are approximately one billion tons of phosphogypsum stored in stacks in northern and central Florida. The Florida Institute of Phosphate Research (FIPR) is investigating potential uses of phosphogypsum. These include using phosphogypsum as a road base material, as a soil amendment in agriculture, as a daily cover in landfills, and as a construction material. The results of this research will be presented and will be discussed in light of regulatory barriers to the use of phosphogypsum.

Jan 1, 2002

By R. C. Milici

Historical production data and reserve decline rates can be used both to predict future coal production rates and to estimate when the reserve will be depleted. Reserve estimates inferred herein from historical production data and decline rates are commonly lower than estimates made from geologic data. The latter include significant amounts of lower quality (deeper, thinner, higher sub) coal which may be mined when the better deposits are depleted.

Feb 24, 1997

By M. T. Ityokumbul

Experiments have been carried out to determine the efficacy of bubble column flocculation. An acid-base neutralization method was used to determine the mixing time in a 0.064 m x l.5 m bubble column equipped with a Mott porous sparger. The mixing times decreased as the tracer injection point approached the gas sparger. Preliminary flocculation experiments on two clay samples were carried out. The effects of frother addition and air flow rate are presented and discussed.

Jan 1, 1998

Uranium exploration over a wide area of the Southern Frame Embayment, South Australia, has defined a number of Lower Tertiary fluvial palaeochannels incised in older rocks. The buried channels contain similar stratigraphicsequences of interbedded sand, silt, and clay, probably derived from the adjacent uranium-rich Olarv Provinee. Uranium mineralization is pervasive within two major palaeochannels, and four small uranium deposits have been found in the basal sands of these channel sequences, at the margins of extensive tongues of limonitic sand.One of the channels is examined in detail with respect to channel morphology, stratigraphy, and rock alteration and the geology and ore controls of the contained uranium deposits are described.A geochemical zonation reflected by variation in limonite, pyrite, carbon, and humic malleI&apos; is defined across one of the deposits. A genetic model is proposed suggesting formation by a uraniferous geochemical cell which migrated down the stream gradient and concentrated uranium on its lateral margins adjacent to the channel bank.

Jan 1, 1978

By Maria I. De Rosa

Table 1 and figure 1 show the number of fires and fire in¬juries that occurred in underground coal mines by state during 1990-1999. Table 1 also shows by state the risk rates, employ¬ees&apos; working hours, lost workdays, and coal production. Over¬all, 87 fires occurred in 12 states. Twenty-seven of those fires caused 34 injuries (the yearly average was 8.7 fires and 3.4 injuries). One fire and one injury involved a contractor. The underground mine fires required 25 mine rescue team interventions and 30 mine/section evacuations followed by 13 mine/section sealing/flooding/CO2/N2, gas injections. The Ewhr value was 1,003 x 106 hr (In = 0.007), the CP value was 4,008 x 106 st (Frr = 0.022), and the LWD value was 208. Virginia had the most fires (15 fires and 7 injuries). Pennsyl¬vania had the most fire injuries (12 fires and 9 injuries), fol¬lowed by Kentucky (12 fires and 6 injuries), and Alabama (12 fires and 4 injuries). Among these states, Alabama had the highest fire risk rate value (Frr = 0.073), whereas Pennsylvania had the highest injury risk rate value (In = 0.016). Table 2, partly illustrated in figure 2, shows by time period the number of fires, fire injuries, risk rates, employees&apos; working hours, lost workdays, and coal production. The number of fires and fire injuries show a decrease followed by an increase during the five time periods (see table 2 and figure 2). This was ac¬companied by a decline in employees&apos; working hours throughout the periods and an overall small decrease in coal production. The In and Frr values follow patterns similar to those shown by the fire and injury values. By comparison, data from Pomroy and Carigiet [1995] show that during 1978-1992 a total of 11 states were involved in 164 underground coal mine fires (yearly average, 10.8) with 43 injuries (yearly average, 2.9) and 27 fatalities (yearly average, 2; however, the 27 deaths occurred during a single fire caused by an overheated air compressor [MSHA 1984]). The CP value was 5,340 x 106 st (yearly average, 356 x 106 st) (Fn = 0.031). Data on employees&apos; working hours and injury risk rates were not available. Tables 3-8 show the number of fires by ignition source, method of detection and suppression, equipment involved, location, and burning material by time period. Figure 3 shows the major variables during 1990-1999. Table 9 shows the number of fire injuries per number of fires causing injuries and total fires by year, ignition source, equipment involved, and location.

Jan 1, 2004

Size Separation Pretreatment and Alumina Recovery from Bayer Red Mud

Sep 13, 2010