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By G. C. Smith

Process control applications in the minerals processing industry have grown extensively over the years due to the many benefits derived, both financially and operation-wise. Although control theory is widely known, conventional control does not address many of the problems posed by minerals processing circuits. Some of these problems include interactions between process units, large dead times, few and sometimes unreliable measurements of process variables, non-linear systems, disturbances entering the process, processes with varying dynamics, and many more. For the past 26 years Mintek has been active in the field of providing specific, customized solutions for the minerals processing industry. This paper focuses specifically on applications in milling and flotation control. The underlying control principles have been applied to a wide range of other processes including submerged arc furnaces. Advanced control application is implemented on a common advanced control platform, called the PlantStar. Much innovation has been brought to the fore in terms of adapting theoretical techniques and developing new methods to resolve specific milling and flotation circuit issues by using process control. The general approach adopted by Mintek when implementing advanced control is to enforce stabilization of the process as a solid foundation for optimization strategies to be built upon. In doing this, a robust control solution is presented, and the end result is flow and level stability, better product quality regulation, increase in throughput and a stable feed to downstream processes. The MillStar advanced milling control system ensures that the milling circuit is at all times easy to operate while achieving optimum product quality and throughput. Stabilization should include product size regulation and sump level control. Another important aspect of stabilization control is the handling of milling circuit constraints without introducing cyclical behaviour and with minimal disturbance to the final product quality. Process constraints may include factors such as process water limitations, cyclone feed density, and cyclone underflow angle. An optimization strategy should be developed to increase the mill throughput without affecting product quality or introducing mill overloads. The relationship between solids feed and power draw is examined and manipulated for optimization. The interactive nature of flotation circuits and the presence of recirculating streams tend to make satisfactory control almost impossible with conventional techniques such as PI (Proportional plus Integral) or PID (Proportional, Integral and Derivative) control. The FloatStar Level Stabilizer simultaneously considers all the flotation levels in a circuit by use of a multivariable algorithm, and takes compensating action in anticipation of disturbances. The FloatStar Flow Optimizer is an optimization strategy that controls the concentrate flow rate between various sections of a flotation circuit (e.g. Rougher ?Cleaner; Cleaner ?Recleaner). By controlling the concentrate flow between the sections, the mass pull, circulating loads and residence times of the circuit can be stabilized, thereby enhancing the performance of the flotation process. Optimizing grade and recovery are coupled with the dynamic optimization of circulating flows. The FloatStar Grade-Recovery Optimizer will maximize the grade in a section, whilst ensuring that recovery or any other process specific limits remain within predefined limits. With successful applications on several milling and flotation plants worldwide processing amongst others PGMs, gold, copper and nickel, Mintek has shown that it is a leader in milling and flotation control innovation

Jan 1, 2004

Sulfidisation Promotion Effect of Ammonium Sulfate on the Flotation of Copper Oxide Ores

Sep 13, 2010

By T. W. Beck

Drilling into coal mine roof strata to install roof bolts has the potential to release substantial quantities of respirable dust. Due to the proximity of drill holes to the breathing zone of roof bolting personnel, dusts escaping the holes and avoiding capture by the dust collection system pose a potential respiratory health risk. Controls are available to complement the typical dry vacuum collection system and minimize harmful exposures during the initial phase of drilling. This paper examines the use of a bit sleeve in combination with a dust hog-type bit to improve dust extraction during the critical initial phase of drilling. A twisted-body drill bit is also evaluated to determine the quantity of dust liberated by comparison to the dust hog-type bit. Based on these laboratory tests, the bit sleeve may reduce dust emissions by one-half during the initial phase of drilling before the drill bit is fully enclosed by the drill hole. Because collaring is responsible for the largest dust liberations, overall dust emissions can also be substantially reduced. The use of a twisted-body bit has minimal improvement on dust capture compared to the commonly used dust hog-type bit.

Feb 27, 2013

By A P. Brown

Access arrangements required with landowners and occupiers; relaxation on size of permits; lack of confidentiality for information provided; one-stop shop under RMA, removal of standing; plethora of controls; post-closure; funding for long-term security.

Jan 1, 1997

By B L. Sainsbury, M F. Lee, A Penney

For most practical rock engineering purposes, stress versus strength conditions around underground openings for which rock masses fail and their various modes of failure, are reasonably well-understood and predictable. Some rock masses and particular rock types fail slowly and relatively quietly. However, others can fail rapidly, violently and maybe with the ejection of significant broken rock. In some countries and in some mines, the latter style of violent fracturing and failure of the rock mass is referred to as ‘strain bursting’. It can be alarming and extremely hazardous to nearby miners and machines. For such situations, the design and installation of appropriate ground support, to provide a safe work place, can be challenging. At any particular mine site, it’s often generally understood which local rock masses (rock types) are prone to strain bursting. Competent and strong siliceous rocks and some massive sulphides are often the main offenders. While some neighbouring, equally-competent and strong rocks might also fail, they do so more quietly and without ejection. To understand this contrast, a better understanding of the fracturing process is required other than simple and standard strength versus stress considerations. Suites of standard intact rock properties from numerous Australian mines have been used to define and help understand different high-stress failure styles for component intact rock; especially those prone to violent fracturing (strain bursting). Simple considerations of micro-fracturing (defects in the intact material), plus the available energy at failure versus the energy consumed during fracturing, distinguishes between over-stressed and competent rock types that fracture violently and those that don’t. CITATION:Lee, M F, Penney, A and Sainsbury, B L, 2018. Proneness of competent over-stressed intact rock to violent fracturing, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 170–184 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

The University of New South Wales (UNSW Australia) has been a world leader in the development of innovative virtual reality technologies over the last 15 years. AVIE (Advanced Visualisation and Interactive Environment) was developed by iCinema as a collaborative venture between UNSW?s Faculties of Engineering and the College of Fine Arts. This is the world?s first 360°-surround, virtual reality (VR) stereo projection theatre system. The School of Mining Engineering at UNSW Australia has developed 18 different virtual reality modules aimed at mine safety training and mining engineering education. These modules are being regularly used in both the mining industry and the university. The School of Mining Engineering is continuously involved in the development of different modules. Research is also currently being conducted on the implementation of other technologies into this environment. Artificial intelligence (AI) and fuzzy logic are tools that the authors would like to consider implementing in future module development. This paper will review current research in both these areas and consider options for applying these technologies.

Jan 1, 2014

By E. N. Gilje

"IntroductionDURING the past five years, over two miles of shaft sinking have been completed at Falconbridge. As sinking has continued, we have improved our equipment and technique with a resulting increase in efficiency. Throughout this period, all sinking has been done by contracting firms, under our direction and to our equipment specifications, with highly satisfactory results. Since 1949, Riddell shaft mucking machines have been successfully used in conjunction with the •benching method of sinking. Our most recent practice can best be illustrated by describing the method used in sinking No. 1 shaft at the Fecunis Lake mine.No. 1 Shaft, Fecunis Lake MineNo. 1 shaft, consisting of six compartments and measuring 13 ft. 11 l/2 in. by 19 ft. 9 3/8 in., outside timber (Figure 1), will be sunk to a depth of 4,170 feet. It was designed to accommodate two 10-ft.diam. Koepe hoists."

Jan 1, 1955

By Duah Philip, Awuah-Offei Kwame

Regulatory pressure to mitigate the cement industry’s environmental impact has intensified recently, spurring innovation in low-carbon cement technologies. In the United States, cement production, consumption, and carbon emissions are accelerating. Despite this trend, the adoption of environmentally friendly dry clinker production technology with combined preheater-precalciner systems has surged from 53.3% in 2002 to 88.5% in 2019. Concurrently, reliance on wet technologies, known for their higher environmental impact, has dropped from 16.5% to 0.6% within the same period. This paper presents the results of our comprehensive study, utilizing over 100 years of industry data, to analyze these technological shifts and their environmental implications.

Feb 1, 2025

By Carl Rampacek, W. A. McKinney

The feasibility of leaching calcareous and siliceous oxidized copper ores by downward percolation with dilute acid was investigated in the laboratory and in a small continuous plant operated by the Federal Bureau of Mines . The crushed ore was leached in deep columns with solutions containing 1 or 5 percent sulfuric acid to dissolve the oxidized copper . Bench - scale laboratory tests on two calcareous ores from the same deposit indicated that a maximum copper extraction , about 72 percent , could be expected from heap leaching a feed as coarse as 1 inch . Laboratory tests on a siliceous ore showed that almost 70 percent of the copper could be extracted by leaching a 1 / 4 - inch feed . The bench - scale tests were supplemented by larger continuous runs wherein the copper was recovered from the leach liquors by precipitation on detinned scrap iron . Leaching the calcareous ore for approximately 10 days recovered 62 percent of the total copper . The sulfuric acid consumption was 9.6 pounds per pound of copper recovered . Treatment of the siliceous ore recovered 62.5 percent of the total copper in a 2.5 - day leach and the acid consumption was 3.7 pounds per pound of copper recovered . Cement copper products that assayed over 71 percent copper were obtained from the leach solutions by precipitation on scrap iron . The bench - scale and continuous tests demonstrated that downward percolation leaching of the calcareous ores was technically feasible but not economically practicable because of high acid consumption by the gangue . Leaching of the siliceous ore was rapid and the acid and water requirements were nominal . The downward percolation leach procedure is simple and requires a minimum of plant equipment .

Jul 1, 1960

By B. Salt, R. Sullivan, E. Roos, M. de Vries, K. McKenna, L. R. Nelson, M. J. N. Uys, P. Jacobs, E. Munnik, P. Lewarne

A review is presented of the furnace cooling system design and operating developments since the initial furnace start up in January 2001 of the Chambishi DC-arc furnace to produce ferrocobalt alloy. The furnace operating conditions present a number of challenges to the furnace designers and operators in particular to provide a sustainable lining that can withstand the highly corrosive siliceous slag (roughly 50% SiO2). This is exacerbated by the dual requirement to operate this slag in a relatively highly superheated state (?T ? 400°C when tapped at 1550°C) to ensure that the ferroalloy can be tapped (estimated 1370°C alloy liquidus temperature) prior to atomisation. The core design challenge occurs at the alloy-slag interface, where a single refractory type cannot simultaneously meet the dual requirement of both adequate alloy and slag corrosion compatibility. Progression by the fourth furnace campaign to involve a retrofit of the bath sidewall to Hatch high-intensity, water-cooled, copper Waffle coolers and tapblocks is described. This involves operation of the entire slag bath, the slag-alloy interface and a portion of the alloy within a more chemically compatible refractory/slag (and possibly at times even an alloy) freeze-lining. Due to the high furnace power density employed (up to 500 kW/m2 hearth area) coupled with the large slag superheat needed, the resulting imposed peak bath sidewall heat fluxes are substantial, especially in the alloy (typically averaging 170 kW/m2), and on occasion have exceeded 1000 kW/m2. The performance of the Hatch Copper Waffle coolers under such aggressive process pyrometallurgical conditions is described. Aspects of the design and especially operating strategies to effect sustained production at up to 38 MW power input are also discussed.

Jan 1, 2004

By L. J. E. Hofer

This Bureau of Mines study, part of a broad program of research in air pollution related to minerals, was conducted to determine the nature and extent of the poisoning of catalysts when used to remove certain air pollutants from exhaust gases. The poisoning examined in this investigation occurs when leaded gasoline is used as a fuel for automotive engines. This research was conducted in cooperation with the Public Health Service. The smog in the Los Angeles area is attributed to the effect of radiant energy on the mixture of hydrocarbons and nitrogen oxides found in the atmosphere there. A large portion of the hydrocarbons is believed to originate in the exhaust gases of automobiles. To prevent the escape of these hydrocarbons into the atmosphere catalytic mufflers have been proposed. In principle, these are simple and nonmechanical, consisting of a vessel charged with a catalyst capable of destroying the exhaust hydrocarbons by catalyzing their oxidation. However, these catalysts are severely affected by the particulate matter, mainly in the form of lead sulfate (PbS04), lead oxysulfate (PbS04.Pb0), and Lead chlorobromide (Pb(Cl,Br)2), contained in the exhaust gas. Deposition of lead on the catalyst as PbS04, PbS04, PbS0, and Pb(Cl,Br)2 restricts catalytic activity. As a result,after 20,000 miles of operation the temperature necessary to achieve 50-percent conversion of isopentane in a mixture of 0.2 percent isopentane, 2 percent carbon monoxide and 4 percent oxygen, and the balance nitrogen at 10,000 hourly space velocity has risen from 400° C. to 500° C. even with a catalyst taken from the exit portion of the catalyst bed. The rest of the catalyst was more severely affected. Therefore, it would appear that operation for 20,000 miles is near the limit for a catalyst.

Jan 1, 1963

By T. L. McCall

AUTHENTIC records show that coal mining in Cape Breton dates from the year 1720, when the French military authorities obtained coal from what is now known as the Blockhouse seam, in the Morien district, for use at Louisburg, where the fortress, now of such historic interest, was being built. Thereafter, coal was mined more or less continuously, and although, for many years, in a very haphazard way, there was, on the whole, a steady growth towards a better regulation of the industry, culminating in the year 1825 with the arrival of the General Mining Association in Nova Scotia. This Company, more generally known as the G.M.A., had acquired extensive tracts in Brazil and Colombia and had also purchased the mining rights for the whole of Nova Scotia-excepting such areas as had been previously leased-the attraction being the rich veins of copper currently reported as occurring in the Province. These fabulous copper veins, upon examination, failed to materialize, but the engineer sent out to report on them returned with such good accounts of the coalfields he had seen that steps were taken immediately to investigate them. From that rime onward, the history of coal mining in Nova Scotia is one of progress and achievement. One of the first acts of the G.M.A. was to acquire the areas already leased to other concerns and develop these in a masterly fashion, employing the best machinery and talent available at that time. Such dominant control over the minerals of the Province, although it led to orderly development in the several coalfields, also, later on, resulted in restricting operations in the various coal centres to such an extent that a determined effort was put forward by the people of Nova Scotia to end such a state of affairs. Consequently, in 1857, the original lease of the G.M.A. was abrogated, the Company, however, being allowed to retain certain areas in the districts it had already developed. This opened the way for the formation of other coal-mining companies, and numbers of these immediately sprang into existence in the various coal-mining centres. In 1893, the Dominion Coal Company, Limited, was formed by the amalgamation of seven companies operating in the neighbourhood of Glace Bay, in Cape Breton. Later, in 1910, the Dominion Iron & Steel Company, Limited, obtained control of the Cumberland Railway & Coal Company, which had been formed in 1884 to acquire the rights of the two companies working in the Cumberland County coalfield.

Jan 1, 1936

By F. T. Cawood, C. C. Birc

The South African mineral resource royalty system, initiated in 2010, employs a hybrid ad valorem approach tied to profitability ratios. This rate is then applied to the total revenue from mineral sales. In this article we explore how the profitability ratio in the mineral resource royalty formulae interacts with the gold tax formula as applied to South African gold mines, devising a strategy for cut-off grade optimization. The study focuses on nine narrow, tabular-reef (Witwatersrand) gold mines. The approach involves a mixed-integer cut-off grade optimizer utilizing Excel Solver, incorporating relevant royalty and income tax formulae. Different mineral resource royalty cost scenarios were examined. Ignoring these costs in the break-even analysis lowers the cut-off grade but slightly boosts profits compared to considering expected costs. The developed optimizer model, recognizing the profitability aspect of the mineral resource royalty and income tax formulae, yielded a 6% lower cut-off grade than the base case. This is significant for increasing mine life. While the profit differences are marginal (0.2% compared to the break-even base case), substantial improvements were noted: a 6% increase in stope tons above cut-off grade, and a 4% rise in total revenue. Enhanced resource utilization (3% extra gold kilograms) reduces necessary development and overall costs, allowing for a reduction in the cut-off grade and ultimately, more profit.

Aug 8, 2024

By C. H. Johnson

The Big Jim mine of the United Eastern Mining Co. is at Oatman, Mohave County, Ariz. Oatman is in the foothills on the western side of the Black Mountains, at an elevation of about 3,000 feet. It is on a paved highway about 30 miles west of Kingman, the principal distributing; center for mining supplies in the district, and 25 miles from Topock, en the Colorado River. Both latter towns are stations on the Santa Fe Railroad; Oatman has no roil connection. Mining conditions are typical of western Arizona; water is scarce, no mine timber is available locally, and the winters arc mild and the summers hot. The supply of a killed labor is usually adequate, but in 1933-34 it w not more than sufficient for the restricted operations. Wages have been $4.00 to $4.50 for miners. Electric power was supplied from Kingman and cost about 2.6 cents per kilowatt-hour.

Jan 1, 1935

By M P. Cipold, J F. Bachmann, C C. Bachmann, J Ha, H B. Wurst

In washability, the objective is to establish the set of maximum possible separation performance criteria for a given mineral feed. The method of choice for this has been the conventional sink–?float analysis. Using traditional sink–float method it is not possible to sample and automatically measure. This fact, as well as time, cost and safety aspects, as well as required and manageable sample size of the sink–float method have prompted the search for a new approach for establishing separation performance criteria. This is especially important as performance levels of separator plants can be maximised if reliable information based on an analysis of washability in conjunction with particle size distribution, elemental constituency becomes available in time. The new method to establish washability and associated performance information utilises at least two measurement technologies concurrently: single energy X-ray transmission and image processing. For porous or very inhomogeneous materials the results can be refined further by using dual energy X-ray transmission and image processing.If used as a standard laboratory device the continuous build-up of the washability curve allows to evaluate the statistical significance of the sample size during measurement.The measuring method can even be improved if it is combined with an automatic sampling system that feeds a continuous and representative particle stream to the analyser. Then the value of the collected data increases considerably since the washability data are additionally available in time to be considered in the control of the process plant. The washability curve can be derived and is continuously updated based on the particles measured during the past time interval. Older curves and results are stored in a database and can be retrieved for later reference. This technology now enables automated sampling and integrated automated measurement, which was previously not possible.This incorporation of an automated washability monitor at a representative sample gives near real-time data that would impinge positively on the profitability of the plant operation. The method was extensively tested using a range of samples with different densities, varying from 1 g/cm3 (coal) to 5 g/cm3 (iron ore).CITATION:Bachmann, C C, Bachmann, J F, Cipold, M P, Wurst, H B and Ha, J, 2017. Inline analysis of washability parameters for process control, in Proceedings Eighth World Conference on Sampling and Blending , pp 213–218 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By M. R. Hosseini

Among all gravity concentration methods, Shaking-table is the most effective one from the metallurgical point of view. In this study, Shaking-table was used for concentrating a manganese ore sample from Ghasem Abad area in Kerman, Iran. Experiments were designed by using L8Taguchi design with five variables and each variable in two levels. The influence of each variable and their interactions on the operation of the device was studied. The variables under investigation were: table slope, table frequency, water flowrate, feed rate, and particle size. The manganese concentrate grade and recovery were used as response variables. It was shown that water flowrate, table slope, feed rate, and particle size are the significant variable on concentrate grade; while, all the variables influence manganese recovery.

Jan 1, 2014

By T. H. Etsell, P. Kerr

Over the past 50 years, there have been a number of geological, mineralogical and metallurgical studies carried out on the Clear Hills ironstone deposits in northern Alberta, Canada. Due to the low grade and complex mineralogy, previous metallurgical studies have achieved limited success. After rejecting some of the more ‘conventional’ techniques for iron ore beneficiation, the authors chose to investigate the grain enlargement/iron segregation process for the beneficiation of the Clear Hills ironstone. The segregation reactions for iron have not been explored in as much detail as they have for copper and nickel. Iron segregation roasting involves the addition of a carbonaceous reductant and a chloride transport agent to an iron oxide ore, and heating the mixture to an appropriate temperature. The benefit of this process is that it allows ores in which the value minerals are intimately associated with gangue minerals to be processed in circumstances where conventional iron ore processing techniques are not suitable or capable of upgrading the material. As iron is one of the few elements that can be transported through this process, a clean high purity metallic concentrate can be achieved.Through the use of crucible testing followed by Davis tube magnetic separation, scanning electron microscopy, X-ray diffraction and atomic absorption spectroscopy analysis, characterization of the reaction products has been made with an improved understanding of the reaction mechanism.

Jan 1, 2015

By L. F. Pett

ALTHOUGH Kennecott's orebody has long been outlined, it is still necessary to define further its limits. This mine, long an advocate of churn drill methods, recently supplemented its practice by using a diamond drill. So far, the diamond drill holes have been along the main canyon floor which traverses the mine area. This zone, highly fractured, is cut by numerous friable quartz veinlets encountered within the porphyry. These veinlets, formed by filling contracted porphyry fractures with hot, siliceous, mineral-bearing solutions, are often more highly mineralized than adjacent porphyry. They pulverize on contact, yielding only slight amounts of core. For this reason neither analysis of core recovered (amounting to 37.5 pct of approximately 10,000 ft of hole drilled) nor sludge analysis considered separately would give correct assays. The following equation has been adopted: actual weight theoretical actual weight core — theoretical weight sludge Drilling is done un.der contract. Collection of core and sludge samples together with double-checked assaying is done by the company. Microscopic, specto-graphic, and X-ray examination of rock specimen saved is carried on at the University of Utah. A fund, set up by Kennecott Copper Corp. at the University, has encouraged cooperation in solution of mining research problems. The drill used is a Longyear hydraulic powered by a Waukesha 26-hp gasoline motor. A pump, Bean Royal type, is powered by a 7-hp Briggs-Stratton gasoline motor. Weight of unit is 4000 lb and is placed under tripod of 35-ft wood poles within a 25-ft sq area. The following results have been attained: Avg depth of hole drilled, ft 1215 Avg drilling depth per 8-hr shift, ft 10.11 Core Size, In. Avg depth of NC casing, ft 21 2 11/16 Avg depth of NX casing and size core, ft 178 2Ye Avg depth of BX casing and size core. ft 427 1% Avg depth of AX casing and size core, ft 585 1 5/32 Through the use of metal sludge collection boxes, 12 ft long, 18 in. wide, and varying from 12 to 24 in. in depth, sludge recovery has been maintained above 95 pct. Samples, both core and sludge, are assembled for each sample interval and averaged in 50-ft bench increments. Determinations for mineral values are made at both the mine and mill. Assay results varying in excess of 0.05 of copper assay values are rerun to determine correct assay. The organization chart for the Kennecott Utah copper mine is shown in Fig. 1. The Utah mine requires approximately four men per shovel shift to drill solid rock, properly slope banks, and trim loose material from the top of bench sections in making them safe for men and equipment working below. During recent years, while the mine has been worked day and night, powder crews have worked day shift only. Series of toe holes drilled back of advancing electric shovels comprise approximately 70 pct of all mine drilling. These holes, spaced about 21 ft apart, have a depth of 24 ft. Of the footage drilled, 25 pct is for vertical holes either 24 or 28 ft deep, spaced about 28 ft centers along the top rim of banks. This combination of toe holes, using an Ingersol-Rand DA 35 drifter mounted on a tripod and the vertical holes using an Ingersol-Rand Type X71 wagon drill is standard practice above the pit elevation. These benches are 68 ft high, with a few exceptions, while benches below the pit elevation are 50 ft high and require very little drilling other than toe holes. Five benches, varying in height from 80 to 100 ft, require a combination of toe holes plus vertical holes and bank holes or electric churn drill holes. Bank holes using the DA 35 drifter and electric churn drill using Bucyrus-Erie Model 29T, account for 5 pct of feet drilled. Until the past year, drilling of fill material on upper levels was accomplished by churn drill method. Wagon drills are now used for fills because they can be set up and drill a hole an hour. It is necessary to run an air drill 7 hr for each 8-hr shovel shift. Average drilling rates during the past year were as follows: DA 35 drifter using 2 and 3-man crews — 4.36 ft per man hr X 71 wagon drill using 2-man crews — 6.45 ft per man hr DA 35 drifter bank holes, 3-man crews — 2.53 ft per man hr Bucyrus-Erie churn drill, 2-man crews —1.9 ft per man hr Total mine avg — 4.57 ft per man hr Mine practice assigns powdermen to job locations where they perform all drilling and blasting work within the area. Almost as much labor is consumed in springing and loading holes as in drilling them. Average powder consumption is approximately as follows: Per Electric Shovel Shift, Lb To spring and chamber holes — 50 Blasting — 500 Trim banks — 50 One fifth of all powder used is Hercules hercomite bag powder while four fifths is Hercules gelamite 65 pct strength in 1 Vi in. x 8 in. sticks. Frozen material, overhanging from top banks, creates a winter hazard to shovel operators and equipment. Experiments are being conducted with a high-speed Hercules powder in an effort to eliminate this condition. Banks are trimmed by powdermen who lodge 8 or 10 bundles of powder, containing 15 or 20 sticks each, across the upper surfaces of each working face. To each bundle of powder is attached a primer with 6-ft fuse which is lighted as the powdermen, who have lowered themselves down the bank on ropes, ascend to safety. Three angle dozers, D8 caterpillars, make grades and move supplies for the average 31 drilling machines working daily. Compressed air is furnished by three Ingersol-Rand compressors, each with rated capacity of 3750 cu ft of air per min at 100 lb reservoir pressure.

Jan 1, 1952

By Lixun Kang

Bedded structure is the characteristics of coal measures strata. Bedding planes, coal streaks and soft rock bands existing in roof strata destroyed its continuity in the direction perpendicular to the coal seams. The statistics from a large number of long wall faces in which serious roof fall accidents occurred in Datong coal mining area suggest that beddings planes, coal streaks, soft rock bands, especially the first layer thickness of immediate roof have a great influence on its stability. As the first layer thickness increases, the probability of serious roof fall accidents decreases. Based on the investigations finite element method is used to analyze stress and displacement in the first layer of immediate roof. The results indicate that the thicker the first layer is, the less the roof convergence in the unsupported area is. Stress and strain features in stratified immediate roof are quite different f7om those in non- bedded immediate roof. In stratified immediate roof, every stratum has each own tensile zone in the upper portion and compressive zone in the lower portion, while tensile zone in the first layer is less, the tensile stress is the greatest.

Jan 1, 2005

By M. B. Shongwe, G. Freemantle, M. Botha, S. Brill, J. D. Steenkamp

During steelmaking, one important factor for production efficiency is the free opening of the ladle nozzle during continuous casting. Free opening occurs when the well filler sand and the steel flow freely when the slide gate is opened. Well filler sands are used for the slide gate of a ladle to separate the slide gate refractory and the molten steel. The well filler sand is partially sintered by the heat of the molten metal, which can result in the ladle nozzle failing to open if the sand is not sintered to an appropriate extent. It is therefore of great importance that the sintering behaviour of well filler sand is understood. We studied the effect of the well filler sand chemistry on its sintering properties. Samples of well filler sand with various chemical compositions were sintered in a high-temperature chamber furnace. Quantitative X-ray diffraction (XRD), inductively coupled plasma optical emission spectrometry (ICPOES), and wavelength dispersive spectrometry (WDS) by electron microprobe analysis (EMPA) were used to analyse the various well filler sand mixes. Points of analysis were indicated on backscattered electron (BSE) micrographs. Quantitative element mapping was undertaken by energy dispersive spectrometry (EDS). Thermodynamic calculations in FactSage predicted the percentage liquid phase present for the various well filler sand compositions at equilibrium conditions. The results of this work show that the degree of sintering of chromite-silica well filler sand depends strongly on the silica content. The amount of liquid phase formed increases with increasing silica content.

Feb 2, 2022