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By J S. J van Deventer

Geopolymerisation is a new technology that processes and transforms aluminosilicate minerals into mechanically strong and chemically stable products. In this investigation, minerals such as fly ash and kaolinite were used as the solid raw materials for geopolymerisation. Inorganic salts were added to the geopolymeric pastes to simulate ionic contamination during processing for building and construction applications. The resultant products were examined and characterised physically. It was found that the early strength development of geopolymeric pastes could be changed according to the nature as well as the extent of the contamination. At a small dosage, the effect of the contamination on the early geopolymeric system depended heavily on the nature of the artificial surfaces that were created in the geopolymeric solution through the addition of the chemical additives. For example, the samples containing MgCl2 and Mg(NO3)2 had shown the same yield stress developments due to the similar surfaces that were formed. The cations of these salts were expected to hydrolyse and form magnesium hydroxide precipitates immediately after the addition to the early geopolymeric systems. The calcium cations of CaCl2 also precipitated upon addition, but the nature of the surfaces of these hydroxides was quite different to those of the magnesium hydroxides and gave rise to different yield stress development. The samples that contained the insoluble salts (in water) demonstrated different yield stress development when compared to each other because of the different nature of the additive surfaces that were exposed to the geopolymeric solution. These results suggested that heterogeneous nucleation could be responsible for the change of the rheological behaviour upon inorganic salt addition. When the extent of the ionic contamination was large, the anions released from the insoluble salts (in water) changed the nature as well as the concentration of the geopolymeric solution and the setting processes were retarded. Oxides, in general, behaved very differently from other salts, indicating the uniqueness of the oxygen anions in a system rich in hydroxyl and silicate anions. The inorganic salt contamination in geopolymers did not show any significant signs of structural change according to the XRD diffractograms collected. The seven-day compressive strength was not lowered by the inorganic salt contamination and it could sometimes be beneficial to the product performance, for example, an increase of 20 MPa was induced by the addition of calcium oxide. Inorganic salts as chemical additives have demonstrated the potential to modify setting times of geopolymers without sacrificing much of the productsÆ strength and integrity.

Jan 1, 2000

By Douglas D. Reynolds

Intense vibration can be transmitted to the hands and arms of workers who use hand-held percussive or vibrating devices, tools, and work pieces. Continued habitual exposure to vibration directed to the hands can cause patterns of various symptoms associated with hand-arm vibration syndrome (HVAS). The International Organization for Standardization (ISO) first published ISO 5349 in 1986.3 This standard specified methods for measuring and evaluating vibration directed into the hands from hand-held vibrating devices, tools, and work pieces. The American National Standards Institute (ANSI) published ANSI S3.34 the same year.1 This standard was modeled after ISO 5349-1986 and specified methods for assessing exposure to hand-arm vibration. The Parliament of the European Union has issued the European Union Human Vibration Directive-2002/44/EC, which specifies vibration daily exposure action values (DEAV) of 2.5 m/s2 and daily exposure limit values (DELV) of 5.0 m/s2. These values have generally been accepted by medical experts, scientists, and engineers in governmental agencies, research institutions, and industry in the USA and other countries.2 When they are achieved, they will reduce the potential for the development of symptoms related to HAVS among workers exposed to hand-arm vibration. Significant improvements in measurement and analysis instrumentation, miniature and subminiature accelerometers, and medical diagnostic and assessment protocols have being introduced since 1986 when ANSI S3.34 was first published. In response to these improvements and the introduction of the EU Human Vibration Directive, ANSI Working Group S2.39 developed the revision to ANSI S3.34, which has now been published as ANSI S2.70-2006.2 Method ANSI S2.70 specifies the use of the hand-arm vibration measurement procedures outlined in ISO 5349, Parts 1 and 2.2,4,5 It requires the measurement of ISO frequency-weighted acceleration values in three mutually orthogonal axes of vibration. These values are then vectorially added to obtain the vibration total value, ahv: [ ] where ahwx, ahwy, and ahwz are the measured r.m.s. ISO frequency-weighted acceleration values in the x, y, and z directions, respectively. If multiple vibration exposure events are experienced during a work day, the overall vibration total value is obtained from: [ ] where ahvi is the vibration total value of the ith operation, Ti is time duration in hours of the ith operation, n is the total number of operations, and T is total time in hours associated with the n

Jan 6, 2006

By E. R. Darby

When bronze is melted in open-flame furnaces a considerable amount of slag is formed during the melting operation. This slag may be incidental to the melting practice or it may be formed intentionally by slag-forming constituents added to the furnace charge to provide a covering for the melted metal When this slag is removed from the furnace, it contains, in addition to the free metal carried with it by the skimming operation, a considerable amount of metal as oxide either combined or mechanically mixed with it. By the usual crushing and washing methods this metallic oxide generally goes to the tailings pile, so diluted by other foundry waste that its recovery is economically impossible. If this furnace slag is not inixed with other foundry waste, practically all of the metal oxide as well as the free metal may be recovered by some form of smelting operation. The first side of the problem to be considered is the amount of slag produced. It may not be true that the majority of foundries have no record of the amount of slag made, but it is certainly true that a very great many either have no record or pay little attention to it. Usually the collection of furnace slag for one normal week's run will be sufficient to determine the advisability of its being handled in a different manner from the crushing and concentrating generally employed. If a large representative sample of the accumulation is crushed and concentrated and the recovery compared with the results of a careful assay, it will be found that a large percentage of the total metal content was lost in the washing process. The percentage of loss will be greater than the metallic oxide content, as some free metal in extremely fine particles always is carried away in grains of slag. From such a comparison, it may be determined whether or not the slag should be melted, sold or treated as before. An example of an actual test of this kind is given below. In a foundry melting 50,000 to 80,000 Ib. of bronze per day, a week's accumulation of furnace slag and ladle skimmings amounted to 16,500 1b.—over three times as much as the foundry superintendent estimated before the accumulation was made. Two tons of this material was taken as representative of the lot. This was ground through a thoroughly cleaned ball-mill and concentrated over a Wilfley table. The

Keywords: Pyrometallurgy, furnace, Lonmin, platinum, PGM, UG2 Lonmin, earlier than any other primary platinum producer, had to deal with the concentrating and smelting of UG2-rich ores and concentrates respectively. Smelting was performed at a fairly modest scale compared to its industry peers in the Platinum Group Metals (PGMs) industry, and was focused on smelting concentrates obtained through low mass pulls while still maintaining high recoveries at its concentrators. Lonmin gradually smelted larger quantities of UG2 concentrates. Initially two 2.3 MW Infurnco furnaces were commissioned in 1982, followed by three 5 MW circular 3-electrode Pyromet furnaces in 1991. Deep electrode immersions and moderately high hearth power densities were used in all the designs. Lonmin decided to continue with circular furnace technology when it planned its new high-intensity No. 1 Furnace. Neither Lonmin nor the technology supplier and EPCM company foresaw the challenges that scale-up would bring when applied to the smelting of UG2-rich concentrate blends. Superimposed on the high chromite content, was the low base-metal loading, mineralogical difficulties, and fine particle size that came with milling of all concentrates (especially UG2 and recycle materials within the smelter). Through a process of fundamental diagnoses of furnace run-outs and wear patterns observed during repairs, the main challenges operating this high-intensity furnace were identified as sulphur vapour corrosion of the copper coolers, uneven and unpredicted movements of refractory bricks with associated copper cooler lift and matte tap-block movement, the formation of three-phase ?mushy? zones, high refractory and tap-hole wear rates, uncertainty in matte level and associated insufficient matte buffer height, and high furnace operating temperatures. These factors seldom worked in isolation and will be explored in more detail in the paper. The variability in furnace feed characteristics led Lonmin to re-design Furnace No. 1, to invest in back-up furnace capacity, and to invest in improved monitoring and control. These improvements consisted of online and high-frequency off-line feed chemistry and mineralogy monitoring, online pressure monitoring of water-cooled circuits, an alternative matte liquid level measurement in the furnace, and electrode immersion estimation. Improvements were also made at the converters by installing and utilising optical spectral analysis of the converter flame to characterise converter behaviour and achieve the desired iron end-point for white matte. This paper also briefly discusses the current and future expansion plans, as well as ancillary operations at the smelter such as flue-gas handling and materials handling.

Jan 1, 2011

By Matteo Montesi

Deep foundation elements are often used as discrete elements in externally supported earth retaining structures such as soldier piles and lagging walls, secant/tangent pile walls, anchored walls, etc. These structures can be designed using different approaches such as conventional limit equilibrium methods, finite element/difference analyses, and soil spring (p-y) analyses which are contained in various commercially available computer programs used by engineers worldwide. Although these programs usually undergo various degree of verification and validation, users must be aware of each program’s limitations and conditions in which each design approach should or should not be used. These programs may incur difficulties when non-continuous deep foundation elements are used as part of the retaining system and in cases of non-uniform geometry where closed-form solutions are not available. Design input parameters such as passive arching width, wall-to-soil friction, and p-multipliers (pm) are not always given enough attention which may result in potential design deficiencies. This paper presents the main features of various earth retaining structure design approaches and discusses their advantages/disadvantages. The paper addresses some of the limitations and difficulties that may be faced with when utilizing non-continuous discrete deep foundation elements and identifies key elements that are sources of potential deficiencies and errors. An example comparing the design outcomes using different design approaches is presented. INTRODUCTION Earth retaining structures are used to retain soil and create or maintain a difference in ground elevation between grades in front and behind the retaining structure. Deep foundation elements have been used for decades to support earth retaining structures. Earth retaining structures that are supported by deep foundation elements are typically referred to as externally stabilized system (O’Rourke and Jones, 1990). An externally stabilized system uses an external structural element, in contrast to internally stabilized systems which rely on reinforcements installed within and extending beyond the potential failure mass. The design of externally supported retaining structures can be accomplished using different approaches, which can be challenging under certain conditions. Each approach has been discussed extensively in available literature, and is well summarized in the and Federal Highway Administration (FHWA) guidelines. The objective of this paper is to present the most widely used design approaches and discuss their limitations and difficulties in one concise and brief document, giving the reader a good source for comprehensive understanding of the various design approaches, supported by a technical example. This paper is not intended to be a comprehensive review of all available design methods and formulations available in the literature.

Jan 1, 2018

By D. G. Westlake

Levine has concluded that prismatic slip in titanium is controlled by two distinct thermally activated processes in the temperature ranges O° to 220°K and 220" to 300°K. We feel that his evidence is not convincing. He has drawn a curve through the plot of activation volume u* vs temperature T with a break at 22WK, Fig. 2. We believe that considering the possible experimental error apd the fact that there are only four points between 4° and 196°K, one could be justified to represent these points by the smooth curve we show as a solid line in Fig. 12. Levine believes that (8rf/aT)+ is discontinuous at 220°K, but he hints that he recognizes the possible doubt. We doubt seriously that the results shown in Fig. 1 indicate a discontinuity at 220°K. Certainly if he were justified in drawing a straight line through only four points on Fig. 2, we are even more justified in drawing straight lines through the seven points between 77" and 297°K on Figs. 1 and 3, because a straight line touches all seven of his circles. If his measurements of t were so precise that his decision to draw two curves in Fig. 3 is valid, then the experimental technique is deserving of further mention. We have drawn a smooth curve through the results in Fig. 3 and used this and the curve in Fig. 12 to obtain the values of shown in Fig. 13. Finding continuous is a contradiction of Fig. 4, which deserves comment itself. Levine obtained the solid circles in Fig. 4 using Fig. 2 and a smooth curve through the points shown in Fig. 3. We have replotted these solid circles in Fig. 14. We submit that there is no basis for his two straight lines (dashed in Fig. 14), and therefore have superimposed the curve that we believe is indicated. Certainly, there is reasonable doubt that a discontinuity in dH/dT exists at 220°K, even when one accepts Fig, 2. Furthermore, if one draws the straight lines indicated by least-squares treatments of the results presented as open circles in Fig. 4, their intersection is at about 196°K, not at 220°K as shown. Thus the alleged agreement between Figs. 3 and 4 with respect to "break" temperatures is not apparent. In Fig. 9, Levine has drawn a straight line very much different from the one indicated by a least-squares treatment. Obviously, he has done this because he recognized that H = 0 when T = 0. He is consistent in drawing the low-temperature curves in

Jan 1, 1968

By George S. Rice

"""To prevent the propagation of mine explosions, the Bureau of Mines, Department of Commerce, recommends rock-dusting all coal mines, except anthracite mines, in every part, whether in damp or dry condition. It also recom¬mends that rock-dust barriers be used to sectionalize the mine as additional defense; but these should not be regarded as a substitute for generalized rock-dusting."" (Mine safety decision, No. 5, April 7, 1927.)The purposes of this circular are to present concisely the requirements for effective rock-dusting of coal mines in order to prevent propagation of mine explosions, and to point out the principal related facts on explosibility of coal dusts and mixed coal-mine dusts, as determined by extensive testing in the Experimental Mine of the Bureau of Mines. study of these principal facts on coal-dust explosibility in conjunction with the recommendations for rock-dusting, should help to make clear to mine operators and others concerned the necessity for systematic rock-dusting in all coal mines except anthracite mines, and the futility of inadequate or sporadic rock dusting.The descriptions of nearly 1000 explosion tests in the Experimental Mine are available in a forth-coming bulletin2 and in earlier publications.3For detailed specifications as to the kind of rock-dust, amount to use, where to be applied, and method of sampling and testing, the reader is referred to Bureau of Mines Serial 26064, and American Engineering Standard Committee ""Recommended American Practice for Rock-Dusting Coal Mines"" quoted in Bureau of Mines Information Circular No. 60305.Summary of Specifications for Rock-Dusting1. Mines to be rock-dustedAll coal mines, except anthracite mines, should be rock-dusted.2. Extent of rock-dustingAll open accessible parts of a coal mine should be rock-dusted, including slopes, entries, cross,-cuts, and rooms, headings and pillar workings to within at least 50 feet of the face.3 Amount of rock-dustingAfter cleaning up the coal dust as thoroughly as practicable, sufficient rock-dust should be used to have the mixture of remaining coal-dust plus the rock-dust in any zone contain 65 per cent or more of inert or noncombustible matter. First dusting may require 3 pounds to 5 pounds per linear foot of passageway.4. When to rock-dust In any zone or section (made by dividing the mine for rock-dusting purposes) when the incombustible content falls to 55 per cent, that section should be re-rock-dusted so as to bring the incombustible content well above 65 per cent."

Jun 1, 1927

By V. J. Sikora, W. J. McGuire

Several years ago, we used an electric analogue computer to study the effect of vertical fractures on the productivity of wells in expanding fluid-drive reservoirs. The results of this work were used both as a guide to research and as a basis for the design of actual fracturing operations. In 1958, Dyes, Kemp and Caudle' published portions of the data from this study. The results were released in 1959 to service companies for use in fracturing guides. The purpose of this note is to discuss the experimental conditions and assumptions which were made to derive the information which is already partially a matter of public record. ASSUMPTIONS This study applies to fractures which extend in a vertical plane symmetrically from the well. For simplicity, we chose to study the case of a well in a square drainage area with its fractures extending toward two of its nearest neighbors. We have assumed no migration of fluid across the boundary of this area. The quadrant shown in Fig. 1 models this system. We made the usual assumptions that the reservoir is homogeneous and isotropic, that the fluid is homogeneous and that production is effected by fluid-expansion drive. We also assumed that the fracture extends from top to bottom of the reservoir. Of more practical interest is our assumption that the fracture does not change the size of the drainage area. This premise is appropriate if all the wells in the reservoir are fractured in the same way. But where a single well is fractured, its drainage area may be enlarged at the expense of its neighbors. This could be of economic importance. The assumption that the fracture will not extend beyond the established drainage area of the well also limits the application of the results. Such long frac- tures will yield benefits in excess of those predicted here. When we assume homogeneity, we disregard the frequently observed zone of damage in the vicinity of the well. Where damage is severe, the benefits resulting from by-passing it will far exceed those predicted by this study. All comparisons were made of systems producing in the semi-steady state. By this we mean that equal volumes contributed equally to flow (pressure was dropping at a uniform rate throughout the system) when the rate of flow was observed. EXPERIMENTAL SET-UP We used the pool unit of an electric analyzer' in this study. The squase drainage system of a well was rep-

By Creelman C. A

The removal of a horizontal piIlar between Nos. 9 and 11 levels (1,116 feet from the collar of main shaft) was successfully carried out by blast hole drilling, and, to the best of the authors' knowledge, this mass blast of 40,000 tons of ore (requiring the use of 7,150 Ibs. of explosive and 5,385 feet of primacord) was the first occasion on which a task of this magnitude has been carried out underground in an Australian mine.Blast hole drilling was introduced by this company approximately fifteen months ago at New Cobar Mine in sub-level stoping.This experimental drilling and blasting gave very valuable information as to the correct technique to be adopted when the final plans were being prepared for the removal of the horizontal pillar at New Occidental Mine.The dimensions of this pillar were approximately 15 feet long, and from 40 to 60 feet wide, and 65 feet in depth.The block of ore extended to 35 feet above the level and 30 feet below, the upper section being partially excavated to form the bell-hole floor and grizzly chambers used for drawing off the broken ore in the shrink stope above the No. 10 level.Description of Ore-bodyThe main ore-body is situated almost vertically in a well-defined shear zone at the junction of sandstone and slate, the sandstone forming the eastern and the slate the western wall.The ore-body consists of highly silicified slate (chert) sparsely impregnated with sulphides, containing a small percentage of copper and bismuth.The total length of the ore-body is approximately 450 feet, with a maximum width of 60 feet in the main section, which is about 200 feet long.This is the section shown in longitudinal and cross-section (Fig. 1).The ore-body is divided by a horse of mullock in the northern end into two legs, the eastern and western, which are approximately 250 feet long and averaging 25 feet in width.This ore-body is strong, hard and abrasive with extremely strong walls, an ideal proposition for open stoping and mass blasting.MiningIn all levels above No. 9 the cut and fill method of ore extraction was adopted. On Nos. 10 and 11 levels the shrinkage method was introduced with economical results, and no difficulty had ever been experienced with shabby or bad walls.Shrink stoping had been carried upwards from No. 11 level to within 10 feet of No. 10 level, when mining hazards developed, and instead of...

Jan 1, 1944

By Ken Dyck Sr.

"Construction of the Cigar Lake uranium mine in northern Saskatchewan is well under way. Production is anticipated to begin in the first half of 2007. When in full production, the mine is projected to produce a net 18,000,000 pounds of U3O8 per annum from a high grade ore averaging approximately 20% U3O8.Water management was a major issue that required substantial effort in the Cigar Lake design. Not only is effluent quality controlled, a tight water balance to minimize effluent volume is also designed into processes in order to minimize environmental impact. This paper focuses on the water treatment process and water handling equipment, as well as the operating philosophies that will reduce fresh water consumption to a minimum. An overview of the mining and ore handling systems, as well as aspects of controlling water production from the ore formation and the mine development areas will also be discussed.INTRODUCTIONThe Cigar Lake ore body has proven and probable reserves that total 232 million pounds U3O8. Cigar Lake is the world’s largest undeveloped high-grade deposit. It is the second largest highgrade deposit surpassed only by the McArthur River deposit another Cameco - operated mine in the Saskatchewan Athabasca Basin. Cigar Lake is owned 50% by Cameco Corporation, 37% by AREVA (COGEMA Resources Inc.), 8% by Idemitsu Uranium Exploration Canada Ltd. and 5% by TEPCO Resources Inc. When production starts in 2007, ground ore slurry from Cigar Lake will be shipped to the McClean Lake mill (owned by AREVA) some 80 km away. All of the Cigar Lake uranium will be leached from the ore at McClean Lake. During initial production ramp up, all uranium will be processed at McClean Lake to yellow cake (uranium oxide). After the average production has exceeded a 10 million pound per year rate, Cigar Lake uranium will begin to flow from the McClean Lake Mill to the Rabbit Lake Mill. At this point, subject to regulatory approval, 57% of the Cigar Lake uranium, in a uranium rich solution, will be transported to Rabbit Lake mill (owned by Cameco) for purification to yellow cake. A complete summary of the flow of uranium from the Cigar Lake ore body through to product at both McClean Lake and Rabbit Lake can be found in a recent paper titled “The Cigar Lake Project – Mining, Ore Handling and Milling” in the March 2004 issue of the CIM Bulletin. The paper was written by C.R. Edwards."

Jan 1, 2006

By George Piscsalko

"Quality control of drilled shafts and CFA piles is greatly dependent upon the practices of the site personnel. In many applications it is difficult or not possible to fully inspect the shaft prior to concreting, such as when the shaft is drilled under slurry. CFA piles are cast with no ability to inspect the shaft prior to concreting. There are numerous methods currently available to assess the integrity of drilled shafts and CFA piles. This paper will compare evaluations by three existing non-destructive test (NDT) methods with a new method of Thermal Integrity Profiling for assessing integrity in drilled shafts and CFA piles. This Thermal Integrity Profiling method determines the integrity over 100% of the shaft cross section, both inside and outside the reinforcing cage, by measuring the hydration temperature of the concrete along the length of the shaft. These temperature measurements are made typically only 12 to 48 hours after concrete placement, thus accelerating the construction process. Thermal Integrity Profiling will be described in detail and the Thermal Integrity Profile (TIP) results from several shafts with purpose built defects will be discussed along with several case histories from drilled shafts and CFA piles where Thermal Integrity Profiling was used. The case histories will compare TIP with various other NDT tests utilized on these particular piles including comparisons with Low Strain Integrity Testing and Cross Hole Sonic Logging. Test results for the various methods will be discussed and the excavation of the upper portion of one drilled shaft will be shown and discussed. IntroductionDrilled shafts can be a desirable foundation element in many applications due to the large axial and lateral capacities which are attainable for very large drilled shafts. Drilled shafts can be cast in a dry hole which allows for inspection of the hole prior to casting, but the casting process is very difficult to nearly impossible to inspect with any accuracy. Drilled shafts are frequently cast under slurry as a means to stabilize the surrounding soils during the construction process. When casting a drilled shaft under slurry, it is very difficult to accurately inspect the hole prior to casting and it is equally difficult to inspect the shaft during the casting process. CFA piles are another popular choice for their relative ease and speed of installation. The trend has been larger CFA piles carrying greater loads, which makes quality control critical for these piles. Many of the above mentioned processes are blind to inspection and therefore the chances increase for having defects present in the completed elements."

Jan 1, 2014

By M. Shapiro, L. A. le Grange, B. Henning

Effective pyrometallurgical process vessel design requires accurate assessment of the heat fluxes through the walls of the furnace. This is particularly important for freeze lining operation which is designed to protect refractory materials exposed to chemically corrosive molten contents, or facilitate high temperature operation when the refractory materials are used at conditions close to their service limits. Numerical modelling of fluid flow and heat transfer in process vessels is often used to aid in the design of process vessels. Sophisticated models are used to analyse the three dimensional flow and heat transfer predicting the effects of electrical heating, magnetic stirring, buoyancy, shear forces, various cooling effects and ultimately heat fluxes at the walls of the furnace and refractory. Traditionally these models are applied to the separate single fluid systems in a vessel such as the freeboard region including the arc, the slag region and the metal bath. Boundary conditions such as shear forces and heat fluxes between connecting regions such as the slag and metal bath are either estimated or carried over from separate solutions. Shortcomings in these traditional approaches include the estimation of sometimes critical boundary conditions leading to unreliable heat flux calculations. Also when boundary conditions are carried over between solutions, the process is difficult to set up, time-consuming and finally not fully coupled. In this paper the definition and results of a fully integrated numerical model of a complete arc furnace are presented. The most important mechanisms acting in an arc furnace were considered, including the fields of electrical potential, current, magnetism, momentum, heat transfer and radiation. Temperature dependant properties included electrical conductivity, density, viscosity, and thermal conductivity. The geometry consists of the freeboard, the arc, slag, metal baths and different refractory regions. Although the combined model of air, slag and metal would be defined as a multi-phase problem it is not solved as such. Instead the different fluids are separated by sets of special solid baffles. These baffles allow the implicit transfer of current, magnetism, heat transfer and shear forces between the different fluids and disallow mixing of the separate fluids. The strengths of the integrated model are threefold: Firstly, it provides robustness in defining the geometry and boundary conditions for the overall model. Secondly, it provides the capability to switch on and off individual mechanisms such as buoyancy, magnetic stirring and shear forces in order to observe their individual importance. Finally, it provides a useful tool in the design process through its ability to obtain results of parameter changes in short time scales.

Jan 1, 2004

By John R. Eckel

A Iaboratory drilling rig has been devised and placed in operation which permits the application of hydrostatic, terrastatic, and formation pore pressures to a rock sample for drilling under controlled conditions. Calibrated rests of this equipment in-dicate qualitative agreerment with field results on the effect of weight on bit, rotary speed, and rate of circ~~lation on drilling rate. Tests under pressure, using water and air as the drilling fluids, indicate that when drilling saturated limestone samples, a pressure differential between hydrostatic and formation pressure is the only pressure which affects drilling rate. A differential in either direction causes a reduction in rate, but the same diflerential from the wellbore into the formation causes the greater reduction. Also, in lime.stone, air favorably affected drilling rote only when compared to water at high hydrostatic pressurre, all other conditions remaining constant. Large increases in drilling rate with air could be obtained in the laboratory only while drilling a shale, which offered a limited permeability to nitrogen and which could be drilled with a differential from the formation into the wellbore. In general, the magnitude of the changes due to the pressure effect were small compared to changes in drilling rate obtainable by changing either bit weight or rotary speed. INTRODUCTION In the course of Humble's drilling research program, the desirability of investigating several factors which lend themselves to laboratory rather than field study became obvious. Principal reasons for this are the expense involved in field studies and the difficulty of exercising the required degree of control. Of primary importance are investigations of the effect of fluid pressure on rock rate, the effect of pressure on rock drillability, and a study of the fundamental mechanics of rock failure and chip generation. Several years ago some work in the laboratory on the effect of mud properties on drilling rate was re- ported to the API', but the completion of this work and the investigation of the other variables affecting drilling rate required equipment somewhat more elaborate than that used on these earlier tests. This equipment has been assembled to form Humble's Drilling Engineering laboratory. To date only the effect of pressure on rock drillability while using water and air has been investigated, and this work is not yet complete. Facilities are available to investigate the effect of fluid properties on rate of penetration, and information is being assembled on a current basis on the mechanics of drilling. The sequence of presentation in this progress report is somewhat different from that in which the data were obtained since the order of ex-

By A. C. Fieldner

"Gas hazards may be divided into two principal classes- namely (1) inflammation and explosions; (2) asphyxiation and poisoning. Inflammation and explosion are prevented first by adequate ventilation so that gas sufficient to propagate flame or explosion can not accumulate, and second, by avoiding sources of ignition such as open lights, flames, sparks, electrical arcs, short circuits, or sparks. Ventilation is also the best preventive measure against asphyxiation and poisoning from toxic gases. When men must enter unventilated or poorly ventilated places containing poisonous or asphyxiating gases, suitable gas masks, hose masks, or oxygen breathing apparatus should be worn. As a rule, loss of life and property from gas hazards is due to ignorance of the properties and hazards of various gases commonly used, and consequently a failure to use proper measures for eliminating and meeting these dangers. It is the purpose of this paper to describe briefly the common gases and vapors with respect to the hazards involved in their use and to discuss the necessary precautions and methods for avoiding these dangersLight, heat, and power are prime essentials of our modern industrial civilization. The combustible gases and vapors used for the production of these necessary elements of our existence comprise the largest and most common group of domestic and industrial gas hazards. Almost every individual in his daily life subject to possible danger from asphyxiation, fire, or explosion due to one or more of these gases. For evidence one need only refer to the daily newspaper where headlines such as the following are frequently seen; ""Family asphyxiated by carbon monoxide from gas heater;"" ""Schoolhouse wrecked by gas explosion in basement;"" ""Man killed by automobile exhaust fumes in his garage;"" ""Fireman overcome by smoke fumes in basement ""Woman"".""Boman severely burned by accidental ignition of gasoline used in cleaning gloves;"" and 'in recent Years we often see ""Building wrecked by explosion of still in residential district.""It will be noted that these accidents are due to inflammation, explosion, or asphyxiation. Any combustible gas or vapor, when mixed, with air or oxygen within certain limiting proportions, will explode on ignition. The violence of the explosion varies with respective proportions of air and gas, and with the degree of confinement of the mixture. For example: Mixtures of natural gas and air containing between 4.5 and 12 per cent natural gas will explode, or propagate flame throughout the mixture without the continual presence of the source of ignition that started the inflammation. The table following gives the explosive limits of the common combustible gases and air:"

Jun 1, 1926

By Fredrik Sarvell, Leo-Ville Miettinen, Veli-Matti Uotinen

"Abstract The continued concentration of building activities in cramped areas of city centres means more underground construction and facilities. Demand to minimize execution time and any kind of disturbance to business, general traffic, people and nearby technical structures is increasing. To answer these challenges foundation business has to develop schedule-wise and technically reliable deep excavation and foundation solutions in vulnerable surroundings and challenging soil conditions.Drilled pipe pile wall – RD-pile wall – is a combination of Ruukki drilled steel pipe piles and interlocking sections. Individual RD piles, typical diameters between 219 mm to 610 mm, are connected to each other with interlocks welded to the piles on the automated welding line at the workshop. The matching dimensions of the drilling system and interlocking sections allow the installation of RD piles through stones and boulders and even into the bedrock, if necessary, in a single operation. Installation is done by a down-the-hole hammer (DTH) using the centric drilling system and oversized ring bits or a drilling system consisting of a pilot bit and reamer wings.RD-pile walls have been used as permanent and temporary retaining walls for different kinds of demanding excavations and structures, such as bridge foundation, as quay wall and as sealing structure in rehabilitation of an existing dam. Rigidity and watertightness of the lower end of the RDpile wall is being researched as well as fire design issues examined. One of the most fascinating RDpile wall solution is to utilize the wall as a permanent underground basement wall. Reliable and rigid connection to bedrock decreases the bending moment and deflection of the wall and good fire resistance in addition to large variety of pile dimensions and steel grades gives possibilities to design and execute a cost-efficient basement wall and foundations without temporary retaining wall works. 1. IntroductionIn northern Scandinavian soil conditions, installation of sheet pile walls, or even secant pile walls, is often problematic due to the existence of hard and large boulders in the overburden, and especially when the retaining wall structure should penetrate to or into the hard bedrock. On the other hand, an efficient and reliable piling technique suitable for the demanding soil conditions described above, percussion drilling, has been developed since the 1960’s, and especially from the mid-1990’s, when the concentric drilling method with the ring bit and pilot was introduced. A new retaining wall application, the RD pile wall was developed by combining that drilling method with the widely used retaining wall solution based on steel pipe piles and welded interlocks (a pipe pile wall normally installed with vibratory or impact hammers)."

Jan 1, 2014

By A. J. Monhemius

"The paper reviews the history of modern iron control processes in hydrometallurgy. Particular attention is paid to the processes developed in the 1960s by the electrolytic zinc industry for the removal of iron from zinc sulphate leach liquors - namely the Jarosite, Goethite and Hematite processes. Problems faced by the industry in the safe disposal of the iron residues produced by these processes are discussed. The lessons learnt over the past half century by the zinc industry are projected on to the current efforts to develop hydrometallurgical processes for the treatment of copper sulphide concentrates. It is argued that any attempt to create a “hydrometallurgical copper smelter” is doomed to ultimate failure, unless and until methods are developed that can recover iron from solution in a form that can be directly utilised by the steel or pigment industries.INTRODUCTIONYou suddenly realise that age is catching up with you and that technology is moving on apace when you are asked to talk about the history of a process or a technology that you’ve grown up with. Such was my reaction when, out of the blue, I received an email last year from the chairman of the CIM’s Historical Metallurgy Committee, asking me if I would give a talk on the early developments of iron control in hydrometallurgy. Once I’d got over the initial shock of realising that I had now become living history, I was of course flattered that the invitation was to give a talk at the prestigious IMPC, a conference series that has been going for as long as I have been in the business. That this particular IMPC was to include the 4th Symposium on Iron Control in Hydrometallurgy, made the invitation even more appealing and appropriate, because I was the co-chairman of the First Symposium on Iron Control, together with my old friend, John Dutrizac (known by many as Dr Jarosite), which was held 30 years ago in Toronto.first became involved with iron at the very start of my career in hydrometallurgy – the research topic for my Masters degree thesis at UBC, back in the mid 1960s, was “Reductive leaching of Goethite with SO2”. My thesis supervisor was the late Professor Ian Warren, who himself was a protégé of Professor Frank Forward, one of the founding fathers of modern hydrometallurgy. Little did I realise at that time, now 50 years ago, that the problem of iron and how to deal with it in hydrometallurgical processes would be a theme that followed me throughout much of my 40-year career in academia."

Jan 1, 2016

By S. Makni, A. Faucher, F. Robichaud, A. Bouajila

"Optimizing the ball charge in a grinding mill may be needed to maximise throughput or alternatively, to achieve proper product size in more challenging situations. Efforts to optimise the ball charge however are essentially empirical and suffer from the lack of appropriate, accessible and reliable tools. To address these limitations, a new model-based technology was developed that relies on 3D Discrete Element Method and population balance techniques: it takes on the form of a simulator that aims to significantly decrease the time required to optimise the size distribution and the grinding media load within a ball mill. This paper summarizes the research work related to the development of this simulator. The calibration of the different constituting sub-models will be highlighted. The successful use of the simulator to predict the effect of ball make-up on the grinding product size distribution in one plant is also reported.INTRODUCTION Adjusting the characteristics of the grinding media charge in grinding mills holds significant potential for optimisation of mill performance. For example, properly adjusting the ball size distribution in a ball mill can help achieving specified targets for final product size distribution and/or throughput capacity. However, several problems appear when reviewing the process and the available tools to optimise ball mill load: • The current empirical simulation tools are lacking with regard to the ability to predict the performance of grinding when changes to mill ball make-up are made; • Generally, these tools estimate the top size of the charge so it is difficult to predict the effect of mixed ball make-up; • As the quality of prediction of these tools is low, the operators try to optimise the ball charge by a ""trial and error"" method and proceeding with conservative changes. The influence of these changes on grinding performance is often less than the measurement precision and the improvements then remain imperceptible; • A change in ball make-up may require an average of about 6 months waiting period before reaching the stationary state of mill ball load. This exercise is due to be costly if applied change decreases the mill performance. In its effort towards the comminution process optimisation, COREM R&D portfolio had addressed the herein described opportunity during the last five years This paper describes the technology already developed for the optimisation of ball charge. It takes on the form of a simulator that aims to predict the evolution of the grinding media load, the grinding energy available in the ball mill and the final product size distribution. The simulator was first validated in a pilot scale ball mill and then applied in an industrial case."

Jan 1, 2012

As in other countries, gold mining inNew Zealand has been cyclic in character. Looking at these cycles may provide insights as to what may occur in the future and that is the basis for this study.   New Zealand has experienced four gold mining cycles. The first began in 1857, peaked in 1866 and terminated in 1890. This cycle was characterised by alluvial gold mining by individuals and small groups. This is referred to as the Labour Led cycle. The second cycle began in 1890 and reached its peak in 1906 and terminated in 1929. This cycle was characterised by the use of dredges in alluvial gold mining and was also the peak of hard rock mining. This cycle is referred to as the Technology Led Cycle. The third cycle peaked in 1940. This cycle was far less significant than the previous cycles, arising out of the depression when the unemployed embarked on small scale alluvial mining. This overprinted the continuing decline that occurred after 1906. This cycle is here referred to as the Depression Overprint Cycle.   The fourth cycle began in 1984 when production began to climb rapidly from a very low base. The growth in production was spurred by rising gold prices in the early 1980s resulting in relatively low grade gold-bearing alluvial ground becoming economic. But in the 1990s, production from the new hardrock mines in Coromandel and Otago has overshadowed alluvial production. This cycle is here referred to as the Price Led cycle.   What are the future prospects for gold in New Zealand? Although there are many factors involved, this will be largely determined by two factors; firstly, continued access to prospective gold-bearing properties, and secondly, the price of gold. Based on these factors, three future industry scenarios have been developed. The first of these is industry collapse, in which access to prospective property is constrained and prices mayor may not weaken. The second is industry consolidation, in which the available ground is accumulated by a small number of large companies who can use economies of scale to weather price fluctuations. The third scenario is industry growth, in which access to prospective ground is constrained in environmentally sensitive areas but is otherwise relatively unconstrained. Efficient companies are allowed to work in a policy and economic framework that allows them to make a major contribution to the economy.   While the industry cannot control these critical factors, a good knowledge of the positive role gold mining could play in the future economy may provide a framework by which the access issue, in particular, may be addressed from a balanced perspective.

Jan 1, 1994

By A. F. Neil, A. S. Bahaj, P. Watkins, P. M. Hyslop, C. E. Kirby, P. A. B. James

The results of initial work indicate that for 3 of the 5 quarry sands bioleaching reduces the iron oxide content to a commercially acceptable 0.035 wt% Fe20 3, although the time taken to achieve this is 14-22 days against a target leach time of 7 day. Subsequent work has identified strains of Aspergillus niger that are more efficient at producing the array of complexing agents for iron and is yielding excellent and rapid reductions in Fe levels. Potentially, bioleaching offers a more environmentally friendly and energy-efficient process for the removal of Fe from glass-making sands than conventional routes but the production of organic acids by fermentation must be achievable with low-cost substrates such as sugar-beet molasses and lactose permeates for bioleaching to offer a commercially viable alternative to treatment with hot mineral acids. Silica sands usually contain three groups of Fe minerals: iron oxyhydroxides, which comprise crystalline types such as goethite and limonite tending to occur as thin, discontinuous coatings and in cracks and crevices on the surface of the quartz; the assemblage of heavy trace minerals, which include ilmenite, rutile, leucoxene, chromite and haematite, commonly present at levels of 0.1 wt% or less as discrete grains or sometimes cemented to the quartz; and clay minerals cemented in pits on the surface of the quartz and in association with weathered, discrete feldspars. The response of the assemblage of iron minerals to short periods of exposure to hot inorganic acids is broadly related to which group they belong to, those that form coatings being accessible to acid and tending to react rapidly whereas the trace minerals are usually resistant and react very slowly, if at all, and the clays are probably digested. The note reports on a preliminary study of the use of microorganisms to develop an alternative clean technology that will give results comparable to those achievable with hot inorganic mineral acids but with more environmentally acceptable procedures. Various fungi and bacteria produce organic acids when grown in certain media. When these acids come into contact with Fe-stained sand grains the iron is removed as soluble complexes. Several strains of Aspergillus niger have been found to produce organic acids from sugar, predominantly citric, gluconic and oxalic acids, their relative yields depending on cultivation medium and conditions. Sand samples were prepared from five quarries in the UK: Levenseat, Lothian; Redhill, Surrey; King's Lynn, Norfolk; Chelford, Cheshire; and Oakamoor, Staffordshire; ranging in iron oxide content from 0.035 to 0.153 wt% Fe203

Jun 18, 1905

By A Kurniawan, A Sjadat and M Akhasyah

PT Freeport Indonesia’s Deep Ore Zone (DOZ) is a block cave mining operation that has been in production for 15 years. DOZ is located approximately 900 m below surface and has an average production rate of more than 5000 t/d. The footprint of the extraction level will ultimately be 900 m long and average more than 200 m wide. The undercut level is located 20 m above the extraction level.While developing Panel 1K on the northern portion of the extraction level, a major geotechnical problem was encountered when there was a change in lithology and an associated fault intersection. A large collapse of the hanging wall of the extraction level drifts occurred in the area of the fault intersection between February and March 2011. This was followed by a collapse of one of the undercut level drifts directly above the extraction level in May 2011.It was suggested that the best approach to mitigate risk during rehabilitation was to systematically open up the failed area under the protection of a steel set support system. To achieve this, approximately 60 t of grout was injected from the undercut level to fill the voids created when the rock mass between the two levels failed. A ground penetrating radar system was used to survey the condition of the rock mass between the two levels before and after the injection of grout. After the grout injection phase was completed to a satisfactory level of confidence, rehabilitation activities started on the extraction level by taking out the now consolidated material from inside the drift in 2 m advances from under steel sets initially constructed in secured ground.Controlled blasting was the key aspect to the success of the initiative to open up just enough of the drift to systematically support the rock mass before the next set was installed. The systematic support consisted of in-cycle shotcrete sprayed over mesh pinned to the exposed rock, as effectively as possible, with splitsets before spiling the next advance’s protection canopy with rebar. Once secure, a combination of steel set and pumped concrete was installed as the final support. This final support provided a safe platform to repeat the process until the failed zone was safely traversed.CITATION:Kurniawan, A, Sjadat, A and Akhasyah, M, 2016. Rehabilitation of the collapsed panel 1K north in the Deep Ore Zone block cave mine, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 763–768 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016