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In some parts of the world explosions in mines are more commonplace events. Even in the more developed nations explosions have occurred in recent times. In some cases these mines have significantly lowered their lost time injury frequency rates and as such considered themselves to be ?safe? operations. The consequence of these explosions has been devastating. It is not just the loss of those underground, but also to the families that have lost people in them and the broader community that has lost friends, income, business opportunities, and in many cases support and sponsorships. It begs so many questions, ?What went wrong??; ?Why were the circumstances not foreseen??; ?What could have been done to prevent it??; ?How is it that no one stopped it, reported it, fixed it etc.??. Being held accountable for negligence, should not be overlooked, however this may not aid the processes required to find all the answers. If the nature of explosions is understood, the circumstances and tools to prevent them are readily available then the occurrence and reoccurrence are due to weaknesses or inadequacies of the systems of work or in the implementation of them. This paper restates the fundamentals of gas explosions, examines the options to control the underground environment through ventilation, gas management and monitoring, education and training. Case studies have been used to examine systematic failures in ventilation management. In some of these cases it is not so much, the system that had been developed, but the interpretation and implementation of the system that resulted in the failure and subsequent loss.

Aug 1, 2013

By G. M. Steed, A. O. Oyinloye

The deposit lies in amphibolite-facies biotite granite-gneisses of Proterozoic age, 4 km east of a major crustal break known as the Ifewara-Zungeru fault. Gold mineralisation is contained in a zone of quartz-carbonate (mainly calcite) veins and hydrothermally altered gneisses with associated small-scale intrusions along a steeply dipping second-order fault zone that roughly parallels the main fault. Gold is present mainly at grain boundaries between quartz and carbonate as discrete particles up to 100 micrometres in size, in general association with sparse, scattered pyrite, pyrrhotite and some base metal sulphides. Adjacent to the gold-bearing zone the granite-gneisses have been hydrothermally altered with minor volume increase, most notable enrichment being in K2O, S, Pb and Cu (and Au). High concentrations of Th, Y, Ce and Zr in the granite-gneisses imply derivation from S-type granites. It is likely that their high heat production capacity maintained the high temperatures needed to transport gold in the hydrothermal systems. It is most probable that the mineralisation dates from the Pan-African orogeny (650-450 Ma)

Apr 1, 1996

By Alicia Esther Ares

The family of Zn-Al (ZA), Zn-Ag and Zn-Al-Ag (ZINAG) alloys offers high mechanical properties, ease of finishing, the advantage of low energy and pollution-free melting. Also, present alternatives to the well developed alloy systems, such as brasses, bronzes, cast irons and aluminum alloys. The main objective of this paper is to measure thermal parameters in ZA, Zn-Ag and ZINAG hypoeutectic and hypereutectic alloys directionally solidified which present columnar, equiaxed and columnar to equiaxed transition structures (CET). Also, estimate the mechanical properties, such as microhardness and tensile parameters and correlate these with the thermal (cooling rate) and structural parameters (grain size and dendritic spacing). The different types of structures were analyzed with optical and scanning electron microscopy. The results showed the influence of concentration, microstructural arrangement and thermal conditions during solidification on the mechanical properties. The results of the present investigation are compared with previous experiments.

Jan 1, 2010

By O. Naiker

Xstrata has modified and improved on the solid state reduction of chromite (SRC) process that was pioneered by Showa Denko in Japan and previously used at Xstrata?s Lydenburg plant. The history of pre-reduction is a long one, with the SRC process being the first and only commercially successful one prior to Premus. Krupp developed the chromite direct reduction and the rotary hearth for the pre-reduction of chromite. However, these processes have not been successfully commercialised. The are a number of significant differences between Xstrata?s patented Premus and the original Consolidated Metallurgical Industries (CMI) process. These include, the use of lower cost anthracite fines as the reductant source together with the use of unactivated bentonite as the binding agent. In addition the use of oxygen as an energy source is used in Premus but not in the original CMI process. There was a fundamental change in the operational philosophy of the process in that while the original CMI process aimed to maximize metallization of the pellets, the Premus process seeks to maximize the energy output from the kiln while still achieving the required efficiencies and hence increasing furnace output. Premus is designed to reduce electrical energy consumption, provide high recoveries of metallic oxides, and utilizes low cost reductant and energy sources such as anthracite and oxygen respectively. Due to the fact that Premus is one of the most technologically advanced and competitive processes used today in the production of ferrochrome it was the process selected for Xstrata?s new ferrochrome facility, Project Lion, that is due to be commissioned in the third quarter of 2006 in the Steelpoort Valley in Mpumulanga, South Africa.

Jan 1, 2007

By Jiwei Duan

The energy transferred from a hammer to the pile top is considered as the effective energy in this paper. The effective energy is related to the pile dimensions, soil and pile-soil interface properties. This relation may be expressed as an equation based on the dimensionless theory. The relationship between pile capacity and effective energy was studied by performing a series of wave equation analyses using a commercial software called GRLWEAP (Wave Equation Analysis Program) by GRL Engineers, Inc. The results show that the pile capacity increases with the blow count when the effective energy is constant, however, it has a little change as the blow count is over 100 blows per foot. The pile capacity decreases as the ratio of pile penetration depth to pile diameter increases if the effective energy is constant. The pile capacity will increase as the pile diameter and wall thickness increases if the effective energy is constant. The pile stickout (pile length above the ground surface) does not have a significant effect on the effective energy and may be neglected. Curves for pile capacity versus effective energy are presented to be used for preliminary hammer selection according to effective energy and hammer efficiency. Finally, a case from PDA testing was studied. Measurements of the effective energy and pile capacity by CAPWAP were presented and compared with the results by GRLWEAP.

Jan 1, 2006

A poured slag bottom is universally considered to be the best type bottom for copper reverberatory furnaces. It is impervious to molten matte and ready for immediate use.When molten slag is not available a monolithic silica bottom is the second best; such a bottom has been adopted at Mount Morgan. The laying of such a bottom is a phase of smelter construction on which so little real information is obtainable that a day to day record of such a recent example as Mt. Morgan's 19 ft. x 40 ft. reverberatory may be of interest. The work was done under the direction of E. Pesout, of McGill Smelter, Nevada. PREPARATION OF THE BOTTOMA 26-inch bed of closely packed firebrick with sand-filled joints was first laid on the clean, dry concrete foundation and the whole dried out by small fires maintained on the brickwork for several days. Ninety tons of silica sand was meantime thoroughly dried and stored in clean drums. This sand must be low in FeO and CaO and contain about 95% Si02 , the majority of which should be present as tridymite. This form of silica shows little expansion at high temperatures, for it is the form stable between 870° and 1470° C.Its presence thus minimises disturbance to the bottom by the expansion which, during the burning in, would mark the transformation of quartz to tridymite.

Jan 1, 1939

By J. Muller

Keywords: Pyrometallurgy, blast furnace, slag viscosity, slag liquidus temperature, phase equilibria The viscosity of molten slag in blast furnace operations is an important process variable, influencing the rate and location of reactions inside the furnace, as well as the ease with which the final slag could be removed from the furnace. To optimize these processes, it is necessary to estimate the slag viscosity and liquidus temperatures as accurately as possible, and know the target values. Models to predict viscosities and liquidus temperatures, incorporating phase equilibrium calculations, have been developed and validated with measured data. Various diagrams have been developed to illustrate how the properties vary with composition and temperature, including pseudo-ternary diagrams of SiO2-MgO-CaO-Al2O3 for viscosity and liquidus temperature. Models were also applied and methodologies developed to analyse historical blast furnace operations and determine target values.

Jan 1, 2011

The rapid advance in the adaptation of scientific knowledge to the treatment of ores, resulting in improved chemical and metallurgical processes, has enormously increased the possibility of working lodes profitably, which a few years ago were passed over as unpayable. By the discovery of chlorination, and later of "Cyanidation," by the highly improved methods of concentration, and by the rapid advance in our knowledge of smelting (for which we are greatly indebted to Americanmetallurgists), and by the gradual replacement of scientifing principles and enlightened management, for rule-of-thumb methods formerly in vogue, many large low grade bodies have been opened, and are now paying dividends, which were,twenty years ago, looked upon as hopelessly unpayable. It has been more especially in these large low grade ore bodies, that modern methods have been benefical; for small lodes must be rich to pay at all, as they cannot be worked on the same gigantic scale, but even in these a percentage of metal is saved which formerly was absolutely lost.To make these large deposits remunerative it is, however necessary to try and find the point of greatest economic advantage, to handle the ore in vast quantities, to erect huge plats to treat it, to substitute automatic appliances for hand labour as far as possible, and to combine several processes supplementary to one another, so that ore which is not amenable to one process, shall be saved, by another; and by doing...

Jan 1, 1901

By Jean-Sébastien Lambert, Joseph Kabuya, Richard Simon

Managing a highwall failure risk in an open pit mine by controlling and mitigating the impact of vibrations produced by blasting operations is key to achieving safe and cost-effective operations. The impetus behind this study was the risk of multiple bench failures in an open pit mine that was detected by Slope Stability Radar (SSR) following blasting at the foot of a highwall. The unstable high wall was over 140 meters(459.3 ft)high and 200 meters(656.2 ft)wide. The quantity of associated material was estimated at 3,000,000 metric tons(3,306,934 US tons). The overall slope angle was 44 degrees. In order to continue mining without causing the highwall to fail, this study was initiated as part of a risk management plan to mitigate this major risk by identifying mitigation measures related to blasting operations and assessing the response of the rock mass to each blast. Identifying blasting mitigation measures consisted of optimizing blasting parameters to avoid failure of the highwall. First, a signature hole operation was carried out to obtain seismic records that were georeferenced and associated with known single charges. The method used made it possible to determine the propagation velocity of the P-waves associated with the various lithologies studied and to calculate, from the geomechanical properties of the rock masses, the maximum peak particle velocities to be respected. Subsequently, the targeted area of operation was modelled with the I-Blast software in order to optimize the firing sequences to be used according to the different lithologies present. The model also made it possible to optimally predict the particle velocities associated with each blast and ensure compliance with the previously determined peak particle velocity limits. The assessment of the rock mass’s response using SSR consisted of defining alarm thresholds and an intervention plan for various possible alarms as well as quantifying the damage to the highwall from blasting. Controlling the impact of blast vibrations on the highwall’s stability was successful, as the mitigation measures used during mining did not accelerate the highwall’s movement. No upward trend in the stabilization time of the rock mass was observed by the SSR after each blast. Post-blast geotechnical inspections and analysis of post-blast deformations in the highwall also did not reveal significant damage to the highwall as a direct consequence of the blasting.

By V. Petr, G. G. W. Mustoe, T. G. Rozgonyi

We introduced a numerical method that is applicable for the analysis of static and dynamic deformations of elastic media. In this numerical study, each elastic body is modeled with a system of several hundred rigid particles that are joined together elastically. Maximum shear stress fields due to external loading are obtained and compared with those obtained by photoelastic experiment. Preliminary results are presented for the stresses generated in circular disk shaped bodies idealized with connected systems of particles. Extension of this simulation method to fragmentation of brittle materials is also illustrated in an example.

Jan 1, 2000

By F Gervilla, J Laukkanen

The Kevitsa Ni-Cu-PGE deposit in northern Finland contains disseminated Ni-Cu-Fe sulfides, and platinum-group minerals (PGM). The 2.05 Ga-old ultramafic-mafic, funnel-shaped intrusion is intruded into a volcanosedimentary sequence. The major ore minerals are pentlandite, hexagonal pyrrhotite and troilite, chalcopyrite, cubanite and magnetite. Pyrrhotite and pentlandite are replaced by pyrite, millerite and heazelwoodite near the surface. Nickeline, maucherite and gersdorffite are locally abundant. A pilot feed sample of 575 t of the main orebody at Kevitsa had average grades 0.45 per cent Cu, 0.36 per cent Ni, 1.85 per cent S, 0.33 g/t Pt, 0.34 g/t Pd and 0.19 g/t Au, with 0.06 per cent Ni contained in silicates. Higher PGE contents (2 - 27 ppm) occur in a vertical N-S bearing fault zone. The PGM (n = 6022) mapped in specimens of four drill holes, consist of sperrylite 34.7 vol per cent, braggite and cooperite 18.5 vol per cent, geversite 17.2 vol per cent, moncheite 14.7 vol per cent, merenskyite 6.7 vol per cent, stibiopalladinite 4.3 vol per cent, sobolevskite 1.9 vol per cent, kotulskite 0.7 vol per cent, michenerite 0.6 vol per cent, keithconnite 0.3 vol per cent, rustenburgite 0.1 vol per cent, melonite, froodite, laurite together 0.1 vol per cent and only a few grains of sudburyite, irarsite, vysotskite, erlichmanite, hongishiite, Pt-Cu-alloys, three undefined Pd and Pt bismuthinides and antimonides and a Pd-Cu-bearing Pt-Ni-Fe sulfide. Most of the PGM, 62 per cent, are enclosed in silicates, 29 per cent are intergrown with sulfides, nine per cent are within sulfides. A portion of the PGE occurs in solid solution in sulfides and arsenides. Ni flotation yielded smelter-grade concentrates averaging 12.2 per cent nickel. The average sulfide Ni recovery is 78.8 per cent and total Ni recovery is 65.3 per cent. The average Pt is 10 g/t at 55 per cent recovery, Pd 7.7 g/t at 38 per cent recovery and Au 1.6 g/t at ten per cent recovery. Cu concentrates averaged 28.6 per cent Cu with an average recovery of 78 per cent and further 11 per cent in the nickel concentrate. The copper concentrate contains 4 g/t Au at 24 per cent recovery. The average Fe to MgO ratio 17.9 is well above the targeted minimum of 4.5. The pilot results show that the previously used TETA (triethylene tetramine) can be avoided and good quality, smelter-grade concentrates can be produced from the Kevitsa ore at economical levels of recovery by conventional flotation methods.

Jan 1, 2008

By John G. Henning

The Hoyle Pond underground mine is located in the Porcupine Gold Camp, 18 km east of downtown Timmins. Since the initiation of production in 1985, the operation has produced in excess of 1.9 million ounces of gold at a grade of 15 grams per tonne. Currently, Hoyle Pond ranks 8th in total ounce production for all past and present producing mines in the camp, with the highest head grade of any of the significant past, or present producing mines in Timmins. Cut and fill, shrinkage, longhole and panel mining methods are employed to excavate the narrow gold bearing structures. Using illustrative examples, this paper provides an overview of factors, such as geologic setting, rock mechanics challenges and mine infrastructure, which influence mine design at Hoyle Pond.

Oct 1, 2004

By Hirst H, McAfee J

In December, 1934, Mr. R. Sussex extruded some silver lead alloy wires in order to determine their electrical resistance. After the determinations had been made each wire was threaded through a sheet of paper on which the composition was noted. The wires were placed in a more or less tangled mass in a cupboard in March, 1935. It was not until September, 1936, that they were re-examined, and it was found that during the eighteen months some of the wires had corroded very badly. It was not until March, 1938, that the wires and the corrosion product could be examined, and the results are given in the following pages. It is of interest to note that none of the numerous silver-lead alloys which have been used for creep testing in the rolled condition have shown a similarphenomenon. The silver content of the rolled alloys was, however, never greater than 0·1%.GENERAL APPEARANCE OF THE CORRODED WIRESThe wires contained 0·1% up to 28% of silver, and a general examination showed that the corrosion had progressed farther the higher the silver content. This can be seen in Fig. 1, which is a photograph taken in March, 1938, of typical wires at a 'magnification of five diameters. In the early stages the corrosive action seems to be localized and an area of grey, yellow or reddish brown powder appears. The grey corrosion product is by far the one, most frequently seen, and in the later stages the only one visible from the outside. The reddish coloured product is sometimes seen on the residual metallic centre when the grey product has been broken away.The irregular extent of the corrosion can be seen in Fig. 2, which shows (natural size) the 5% silver wire after standing two years. The corroded parts are seen to be much thicker than the slightly corroded ones. The corrosion product, being non-metallic, is quite brittle and falls away from the wire on bending. In the extreme case the' whole wire has corroded and presents a cracked appearance as seen in Fig. 1 (281% silver).An approximate idea of the extent of corrosion can be obtained by measuring the diameter of the corroded part. The original diameter in each case was ·050/ ·051 inches and in Table I are given the diameters of the corroded parts after Hand 3 years respectively. When the diameter...

Jan 1, 1941

By P Steinier

Most operations, when considering their comminution circuit, are primarily concerned with achieving the desired particle size distribution to obtain adequate liberation for the separation process, at minimum cost. Invariably, the effect the grinding media has on the subsequent separation process (ie flotation, leaching) is not taken into account when selecting the media. This choice is driven by cost rather than metallurgical outcomes. Maximising valuable mineral recovery, with optimum selectivity against gangue minerals, is about good particle preparation. And, particle preparation is strongly related to achieving adequate mineral liberation with the correct pulp chemical conditions. However, the pulp chemistry of a system is largely ignored and relies heavily on pH control and the addition of appropriate reagents (collectors, frothers, activators, depressants). The changes effected by the addition of the various reagents are obtained through conditioning the ground pulp prior to or during the concentration stage, subsequent to grinding. Frequently, reagents are added during grinding with good effect, but the impact of the grinding media on down stream processing is usually overlooked. A number of case studies examining the effect of grinding media on pulp chemistry and subsequent flotation performance were completed on several different sulfide mineral systems (for example, lead/zinc, copper, refractory gold). The evidence suggests that a change in grinding media type from forged to high chrome (ie conditioning the pulp during grinding) resulted in a shift in pulp Eh to less reducing conditions, an increase in the dissolved oxygen content of the pulp and a reduction in the iron species present. The changes have a positive impact on flotation behaviour. The results of the case studies are discussed.

Jan 1, 2004

By Y. K. Wong

This paper describes in detail various problems concerning the construction of a pile-raft foundation supporting a 42-storey tower block. The subsurface generally consists of highly weathered sedimentary rock. A comprehensive instrumentation program is implemented to monitor the behaviour of the foundation system throughout the construction period of the superstructure. Instrumentation observations are highlighted in the second half of the paper.

Jan 1, 2010

By F. Calizaya

This paper describes the basic requirements for the utilization of belt air to ventilate working sections at remote locations in relation to the main fans. It highlights the risks involved, potential problems, and the provisions that must be followed by mine operators for the safe use of belt air. The paper also presents a brief review of the current regulations, conveyor belting tests, atmospheric monitoring system requirements, and the experience gained by mine operators to utilize the belt entries to ventilate remote working sections or areas where mechanized equipment is being installed or removed. The paper concludes with a set of recommendations with respect to the utilization of belt air and the composition of belt material in under-ground coal mines.

Jan 1, 2008

By Sandra Penttinen

There are several ways to get a positive reaction from neighbors to blasting projects. When you need to enter structures to conduct a pre-blast inspection, the impression you make on the structure owner must be positive from the very start. Public relations with neighbors is the key. The purpose of all communications with persons residing near planned blasting operations is to educate structure owners as to what will take place during the blasting operations, to explain that safe blasting practices will be followed and vibration and airblast limits will be imposed to protect surrounding structures from damage. Once access into the structure is obtained, the structure owner can be shown, by careful documentation, all existing cracks and defects in the structure prior to blasting and construction. The objective of your visit and all communications is get your foot into the door and win the confidence of the structure owner so that he or she understands blasting will not compromise the structure.

Jan 1, 2006

By W. W. Adams

"The eighth National Safety Competition, that for 1932, conducted by the United States Bureau of Mines, was participated in by 322 mines and quarries operating in 34 States. Tables 2 to 6 show the relative standing of the various plants. Bronze trophies, known as the 'Sentinels of Safety,' which were provided by The Explosives Engineer magazine to be awarded annually by the Bureau of Mines, were awarded to five companies, each company having established the best accident-prevention record within one of the five classes into which all companies were divided, - namely, anthracite-mining companies, bituminous-coal-mining companies, metal-mining companies, operators of nonmetallic-mineral mines other than coal mines, and quarries or open-cut mines. The trophies are awarded for one year and are then transferred to the winners of the succeeding year's contest.The trophy for bituminous-coal mines, which has been in New Mexico for the past year, was won by a mine in West Virginia. The metal-mine trophy goes to New York State for the first time in the history of the contest. Previously it had been won by mines in Minnesota, Michigan, Kansas, and Oklahoma. The trophy for nonmetal mines remains for another year at Retsof, N. Y.; that for quarries and open-cut mines is transferred to a quarry in Michigan. The trophy for anthracite mines, of course, remains in Pennsylvania, as all anthracite mines in the contest are in that State.In determining the relative standing of the various companies enrolled in the competition, all mines and quarries were graded according to their accident-severity rates, - that is, according to the number of days of disability of employees on account of accidents in proportion to the number of man-hours of work done by all employees at the plant. Reports of accidents and of man-hours worked were furnished by the companies to the Bureau of Mines. These reports were carefully examined and classified to insure uniform weighting of accidents according to the number of days of disability to be charged and to insure uniformity of methods of calculating the accident-severity rates for the various mines and quarries. When the relative standing of all plants had been determined in this way, a summary of the records of the companies having the lowest accident-severity rates in their respective groups was sent to the members of a committee of award for examination in accordance with the rules governing the safety contest. The members of the committee of award were W. H. Cameron, managing director of the National Safety Council; William Green, president of the American Federation of Labor; James F. Callbreath, secretary of the American Mining Congress; T. T. Read, Vinton professor of mining engineering of Columbia University; C. B. Huntress, executive secretary of the National Coal Association; A. J. R. Curtis, assistant to the general manager of the Portland Cement Association; and A. T. Goldbeck, director of the bureau of engineering of the National Crushed Stone Association. When two or more mines within the same group operated without an accident during the year, the contest rules provided that the trophy for the group should be awarded to the plant whose record showed the largest number of man-hours worked."

Jul 1, 1933

By Charles H. Jacoby

The solution mining of evaporites has a history of approximately 5,000 years. The insitu leaching of metalliferous ore will have many of the same characteristic problems which have confronted industry for years in the production of evaporite minerals. Basically we will be converting the less soluble forms of valuable metals into their more soluble salts and producing these in much the same manner as, for example, sodium chloride. There are a number of problems which we are faced with that are more complex and complicated and, therefore, will require a higher degree of sophistication than previously employed. One of our most important undertakings will be the clearcut definition of the problems involved. While a great deal of expertise can be drawn from the solution mining of evaporites and the oil industry, problems peculiar to the hydrometallurgical field will need to be solved. One of these is the introduction of permeability into an impermeable rock so as to expose the sought after values of metals which by volume are infinitesimal when compared with sodium chloride, potash or other type water soluble mineralization. After establishing this permeability which will be miniscule when compared to a solution cavity, there will be the maintenance of this permeability. In general, salt deposits are of a somewhat uniform chemical composition and thus their chemistry was relatively simple. Each metalliferous ore body is somewhat unique in its chemistry and will need to be treated accordingly.

Jan 1, 1974

By Michael S. Wieland

This report focuses on the U.S. Bureau of Mines research regarding desensitization and malfunction of delay detonators and cap-sensitive explosives. Delay blasting in underground coal generates tremendous shock waves and rifting forces that fracture and heave coal. The shock waves from charge explosions radiate outward in all directions through the coal, impacting neighboring charges with later delays. The impact phenomena have debilitating effects, causing degrees of reduced charge performance ranging from misfires or weak detonations to out-of-sequence detonations. Impaired charges react poorly, fracture less coal, and produce fumes with higher toxicity. Coal-mine, underwater, and simulator tests were conducted to more thoroughly understand the shock damage mechanism. Recently the U.S. Bureau of Mines partially resolved the malfunction problem by demonstrating the feasibility of formulating a rugged explosive. This rugged cap-sensitive candidate passed the requisite tests for certifying an underground coal-mine explosive. With natural aging the formulation rapidly lost initiation sensitivity. The charge formulation had to be rebatched to finish the testing and therefore could not be certified.

Jan 1, 1992