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By Nicholas Kyriazi

A performance comparison of 14 rescue breathing apparatus was undertaken as an assessment of past and present worldwide technology. Rescue breathing apparatus are self-contained, closed-circuit breathing apparatus used for entry into areas having atmospheres that are immediately dangerous to life and health. Apparatus tested were the Biomarine BioPak 45, 60, and 240; the Draeger BG 4 and BG-174A; the Litton LITPAC II; the MSA Custom 4500 II Air Mask, Chemox, and McCaa; the Sabre Selected Elevated Flow Apparatus (SEFA); the Scott Rescue-Pak; the Siebe Gorman Aerorlox and Proto Ten; and the SurvivairLP-120. The apparatus were tested on an automated breathing and metabolic simulator at the Pittsburgh Research Laboratory of the National Institute for Occupational Safety and Health (NIOSH). Physiological parameters monitored during the testing were average inhaled CO2 and O2, minimum inhaled CO2, inhaled wet-and dry-bulb temperatures, and peak inhalation and exhalation breathing pressures. The metabolic demand on the apparatus was 1.35 L O2/min for most of the test with two 5-min intervals of 2.5 L O2/min, the first beginning 10 min into the test and the second beginning 10 min before the end of the rated duration of the apparatus. Results presented include apparatus service life and test-averages of monitored physiological parameters. Schematic drawings and photographs of the apparatus are also included. The Draeger BG-174A and BG 4 had the highest durations—295 and 291 min, respectively. The BG 4 hadthe lowest minimum inhaled CO2 levels, attesting to the effectiveness of its CO2-absorbent canister. The SabreSEFA had the lowest average inhaled CO2 levels, due mostly to the apparently low-dead-space mask andnose cup. The MSA Air Mask, the only open-circuit device tested, had the lowest wet- and dry-bulb temperatures, with the liquid-O2-source Siebe Gorman Aerorlox having the next lowest temperatures. The only apparatus with a CO2-absorbent not in a packed-bed design, the BG-174A, had the lowest exhalation pressures. The BG 4 had the highest level of positive pressure, with the Biomarine BioPak 60, Survivair LP-120, SEFA, and Litton LITPAC II nearly as high. The BG 4 had the lowest pressure swing, with the SEFA a close second.

Jan 10, 1999

By Julian Baldry

"This paper begins by describing the parameters of high-speed underground tracked drifting operations and the mucking cycle, and-discusses their variations. The railbound Haggloader 8HR is then described, as well as the Hagglund shuttle trains, outlining the advantages and disadvantages of the system. A comparison with conventional mucking operations is then given.IntroductionDrift and tunnel rounds are mucked by one of two major methods, either trackless with diesel loaders and trucks or with track loaders and cars. The aim of this paper is to compare the conventional tracked mucking cycle with the Hagglunds system alternatives.For many years, the traditional method of attacking the mucking cycle has been with a rocker shovel loader, a string of cars and a locomotive. The rocker shovel loads the first car and when full it is taken to a side drift, cherry picker or some other transfer point and is exchanged for an empty one. This empty car is then pushed back into the face and the loading continues until all cars are full . Then the whole train is taken out to a dump point where the cars are emptied. Normally the train's capacity is less than the volume of the muckpile, so the cars go back to the face. This loading cycle is repeated until all muck is removed from the heading. The last train out from the heading usually takes the loader and parks it in a side drift before continuing on to the dump where the cars are emptied.This simplified version of the traditional loading cycle excludes the ancillary work elements such as track laying and so on, but does point out that the cars must be transferred when full and that the whole train normally makes at least two round trips from the dump to the face. The prime aim in drifting is to cycle a round whithin a working shift which necessitates that no one part absorbs so much time as to prevent the completion of the cycle.Mucking, drilling and loading and the ancillary work such as rock bolting, vent and service pipe hanging typically each take one-third of the available work time. Drilling or mucking can and often do, because of various factors, take nearly half of the total cycle time. Once any segment of the operation starts to take this portion of the shift, the likelihood of missed rounds becomes increasingly greater."

Jan 1, 1981

At present, discounted cash flow analysis is generally accepted as the preferred method of valuation of mine projects in surface coal mining. One of the major reasons for that is its ability to take into account the time effect of money over the life-of-mine. As a looking-forward method, it requires forecasts be made with regard to economic, technological, technical, geological, and environmental conditions that will prevail in the future. All forecasts are inherently uncertain as they are a function of the uncertainty of all variables of the discounted cash flows analysis. This indicates that the only alternative for modeling the uncertainty in the project evaluation is the development of adequate risk models to serve the decision-making process. The paper offers a model of project cash flows at risk in strip coal mining, which is an integrated corporate risk management solution. The model is based on a stochastic approach that provides the answer to the question about how large is the deviation between stochastic cash flows and the planned value (or that used in the budget) at each time step of the discounted cash flows analysis, that is due to the uncertainty in the model variables. A solution is worked out for the allowable range of the stochastic cash flows variation that determines the probability of achieving the planned value of a mine project. The Monte Carlo simulation technique is used to model the stochastic behavior of the project variables. A case study of an opencast coal mine project is provided to illustrate the applicability of the risk model. The results obtained provide a new understanding of the mine project risk as a component of the project discount rate.

Jan 1, 2006

By Evangelia Ieronymaki

Civil engineering education relies on traditional methods of teaching, comprising mostly of deductive instruction by presenting the theory, followed by applications. This paper presents an alternative way for teaching foundation design to senior students, introducing a more inductive approach. The students are presented with a project/game adapted to a real-world problem at the beginning of the course. They are asked to identify the correct steps and methods throughout the lectures, for reaching an appropriate solution that they document and present at the end of the course. To make the project more enjoyable, the project has the format of a game where the students compete with each other on completing the task faster and more efficiently. The goal of this teaching approach is to enhance the ability of the students for problem solving and the engineering way of thinking, while playing a game. In the end, the students were requested to respond to a survey regarding the efficiency of this project/game on their learning. The results showed that students prefer more inductive methods for a design course, feeling that this project made them think more like engineers, while the ‘game’ twist of it made the whole course more fun. INTRODUCTION Most civil engineering courses are taught in 2019 the same way they used to be taught 40 years ago in the 80s. Classic methods are mostly deductive, where the instructor presents the theory first and then applications of that theory follow. Little or no attention is given initially to why the theory was developed or what are the problems it solves. The applications of the theory, and thus the purpose it serves, come only in the end and after methodologies have been presented, leading the students to show low interest in the material. As Albanese and Mitchell (1993) have reported, people are mostly motivated to learn things when they clearly see the need to know them. This explains the students’ little motivation that, if any, comes mainly from the fact that the material may be useful either in other courses or in their future careers. On the other hand, there are alternative inductive methods of teaching, where the role of the students is more active. A lot of researchers have highlighted the benefits of inductive methods in teaching engineering courses. Based on Prince and Fedler (2006), inductive methods include; a) inquiry-learning, b) discovery learning, c) problem and project-based learning (PBL), d) case-based teaching, and e) just-in-time learning. Inquiry-learning is maybe the most traditional method of inductive teaching/learning methods, where the instructor asks questions or poses problems and the students need to apply the theory they have learnt in order to find the solution (Bateman 1990; Prince 2004; Lee 2012). Similar to inquire-learning is the discovery learning, where students have to find the solution to a problem or answer to a question, while the instructor simply provides feedback to the students without guiding their efforts (Leonard 1988; Westbrook and Rogers 1994).

Jan 1, 2019

By Sabry A. Sabour

Metal price is among the key variables in the open pit optimization and the economic evaluation process. The common approach to handle metal prices in open pit mine planning is to optimize the ultimate pit and the mining schedules using a fixed, flat metal price. In this respect, metal price is assumed to be known with certainty throughout the entire mine life, which in some cases can reach 80 years. Given the volatile nature of metal markets and the metal price cycles, the conventional pit optimization approach may generate sub-optimized mining plans, which results in either extra costs to mining projects or lost revenues, or both. Long-term operational decisions such as future pit expansions, beyond the currently optimized limits, cannot be handled using the pit optimization approach based on flat metal prices. There is a growing interest in the mining industry to model and manage such market uncertainty in the pit optimization as well as the project evaluations in order to maximize the value of mining operations. In this paper, the author will outline a practical approach to optimize the ultimate pit taking into account the volatility and variability of metal markets. The proposed approach uses a combination of the pit optimizer of the commercial MineSight package and a simulation-based economic modeling real options valuation approach employing the least-squares Monte Carlo algorithm. MineSight pit optimizer employs both the floating cone for initial optimization exercise and the standard Lerchs-Grossmann algorithm. In this paper, Lerchs-Grossmann option will be used to generate the pit shells. Real options valuation approach used in this study is based on generating multiple metal price scenarios. Metal price uncertainty is modeled using stochastic metal price models such as the geometric Brownian motion and the mean-reverting process. The mining operational scenarios corresponding to the simulated metal prices are then optimized using the least-square Monte Carlo approach. The objective is to maximize the discounted cash flow value of the project while incorporating the operating flexibility to revise the operating status of the open pit with time as new information becomes available. The focus of this work is on the practicality of the proposed uncertainty-based pit optimization approach and the difference it can make to the technical operational mining decisions. A case study of a hypothetical open pit poly-metallic mine is provided. The deposit contains copper, gold and silver grades. The run of mine ore is sent to the process plant that produces copper concentrate with copper as the main metal and gold and silver as the by-products. Four uncertain variables are considered here include copper, gold, silver prices and foreign exchange rate. Both copper price and exchange rate are modeled using the mean-reverting process while gold and silver prices are modeled using the geometric Brownian motion. The case study is performed using both the conventional flat price pit optimization approach and the proposed uncertainty-based one to compare results.

Aug 1, 2013

By Don Dodds

Iron Mountain Mine Rehabilitation Project Brief History The Iron Mountain Mine is located in Northern California approximately 12 miles northwest of Redding. The ore body is a large, very dense pyritic intrusion, discovered in 1860, and initially minded for silver. The silver played out quickly and 34 years later the ore body was mined for copper. Copper mining was scheduled to cease in 1939, but WWII prolonged the closure. After WWII, the pyrite ore was mined to produce sulfuric acid. In 1949 when large amounts of cheaper sulfuric acid were being generated by the oil industry the underground mine shut down. Little was done with the site conditions until 1970 when the U.S. Bureau of Mines began monitoring the pH of the effluent. The initial pH readings were around 2.0. This was lower than anyone expected. Laboratory tests showed that when the ore from the mine was submerged in water, the pH of the water would stabilize above 4.0. However, the pH got steadily lower. In 1989 with the ph around 0.7, the U.S. Environmental Protection Agency (EPA) required the owner to build a capture and treatment facility. During this initial period, the majority of the effort was spent on the treatment plant. The relationship between the mine owner and the EPA was basically an adversarial relationship from the very beginning. In 2001, North Pacific Research was hired to upgrade the capture and control facility. The general layout of the Iron Mountain Mine is shown in Figure 1.

By Gabriel S. Esterhuizen, Anthony T. Iannacchione

The roof rock in underground limestone mines in Northern Appalachia can be subject to high horizontal stresses in spite of the shallow depth of the workings. The high stresses can cause roof stability problems. The National Institute for Occupational Safety and Health staff have observed a distinctive asymmetrical mode of failure at the pillar-roof contact in two underground stone mines, which is different from the typical failure mode observed in response to excessive levels of horizontal stress. A dip of greater than 5° was identified as a possible cause of this mode of failure. Numerical model experiments showed that an increase in the dip of the workings can cause an increase in the stress at the up-dip corner of the roof beam. A case study is presented in which failure at the pillar-roof contact was observed where the dip of the workings was 7° in a high horizontal stress field. Numerical modeling showed that the relatively small dip of the workings could have induced stresses changes in the immediate roof that would explain the failures. The model results also showed that this mode of failure is less likely to occur if the limestone pillars contain weak bedding planes that provide stress relief. In addition, the results showed that for the conditions at the case study site, the stress in the roof beam was not sensitive to the thickness of the roof beam or to the excavation span. The high horizontal stresses at this site are an important contributing factor to the observed failures.

By Yang Tian-zu

Sodium pyroantimonate which is widely used as the clarifier in the glass industry and as the synergist in anti-flaming industry respectively has a production history of about 30 years in China. At present there is the production capacity of about 20,000 tons per year in China and China is chief producer of sodium pyroantimonate. In this work the production processes adopted by the producers of China, the operation conditions and the product qualities of sodium pyroantimonate prepared by different preparation methods have been summarized, and on the basis of comparing the advantages and disadvantages for each process the most advanced method - the air-oxidation method was introduced in detail. The production processes of sodium pyroantimonate mainly include the method of oxidation by sodium nitrate, the oxidation method of hydrogen peroxide, the potash salt method, the chlorination method and the air-oxidation method etc. The method of oxidation by sodium nitrate in which the crude antimony oxide, sodium carbonate and sodium nitrate are separately used as the raw material, the flux agent and the oxidant, has the advantage of short and simple technological process, but low grade quality of sodium pyroantimonate and bringing about serious environmental pollution are the disadvantages and this method has been substituted by other methods. Sodium pyroantimonate can be directly prepared by the hydrogen peroxide oxidation method with the advantages of simple technological process, high grade quality of the product and little environmental pollution. High grade quality of sodium pyroantimonate can be also synthesized by the potash salt method, in which industrial antimony white, the hydrogen peroxide, potassium hydroxide and sodium hydroxide are applied as the starting material. The disadvantages of these two preparation methods are high production cost and high requirements on the raw materials. The chlorination method with the crude antimony oxide, hydrochloric acid and chlorine as the raw material has the advantages of low production cost and high grade quality of the product. However complex process, high requirement of anti-corrosion for the equipments and producing plenty of waste water limit its application. The air-oxidation method with stibnite/jamesonite concentrate, sodium sulfide, and sodium hydroxide as the raw material and atmosphere as the oxidant is a newly developed method and its advantages are low production cost and the complex raw material can be treated by this method and its disadvantage is that the quality of sodium pyroantimonate is lower than that prepared by the oxidation method of hydrogen peroxide and the potash salt method.

Jan 1, 2006

The Diamond mechanical soot blower system, manufactured by the Power Specialty Co., Detroit, Mich., consists of a permanent installation of units arranged in groups to fit the design of the various boilers to which they are attached. The units are so arranged as to effect the cleaning of the fire surfaces of the boiler, particularly the tubes, by means of steam jets. When this steam is released among the tubes at a velocity of approximately 2700 ft. per sec. (assuming a pressure of 150 lb.), a perfect whirlwind of currents is created, reaching all the nooks and crannies of the boiler, and thoroughly removing the particles of soot and fine-ash accumulation. This operation requires but a moment's time-from 20 to 50 sec. per unit¬and is performed from two to four times a day, while the boiler is under full load, using a greatly decreased amount of steam and labor as compared with the hand hose or any other system of blowing. Moreover, this cleaning is effected without admitting air, with its cooling effects, to the boiler setting. The Diamond model "G" blower consists of five principal elements as follows: 1. A head through which the steam passes from the main supply line when the valve is opened. The head is the foundation element to which the other elements are attached. 2. A wall box which is bricked into the setting. The box supports the unit, and is so constructed that the blower can be turned freely without admitting air to the inside of the setting. 3. A blower element running cross-wise of the bank of tubes, and consisting of a small header with nozzles so placed as to discharge steam between the boiler tubes. 4. A sheave or hand wheel by means of which the blower element is rotated. 5. Bearings attached to the boiler tubes which support the blower element and permit it to revolve freely. The nozzles from which the steam issues are designed on the principle of the Venturi tube, and vary in design for different pressures. The expanding walls of the Venturi nozzle increase the velocity of the steam over 200 per cent., as compared with a nozzle having a straight orifice. Thus, the velocity of the steam issuing from a straight nozzle at 150 lb. pressure is 912 ft. per sec., whereas, at this pressure the velocity of the steam issuing from a Diamond Venturi nozzle is 2795 ft. per second.

Jan 5, 1918

By Eugene F. Poncelet

This paper attempts to analyze the phenomena involved in the fracture of brittle solids by simple compression. Glass squares standing on edge, and compressed between two parallel steel jaws, developed jagged fractures, roughly parallel to the compression. Polarized light disclosed the presence of high local stress concentrations indicating poor contact all along the contact surfaces with the steel jaws. Teszar has shown that partial contact between the jaws and such a compressed solid causes tensile stresses in the free surfaces of the solid normal to the compression. Furthermore, Scoble has shown that in brittle solids fractures are caused by, and develop normally to, a sufficient tensile stress. It follows that such "partial-contact" cracks must be expected. With optically flat jaws and specimens, no "partial-contact" cracks developed. Instead, very smooth cracks vertical to the compression appeared 'upon release of the pressure. Investigation showed that the friction with the rigid steel opposed the return of the glass to its original width, after creeping on the steel as compression built up, thus setting up tensile stresses in the glass surfaces. It was not found possible to eliminate these "release" cracks. Compressed to destruction, such specimens flew to dust. A Microflash photograph taken during the explosive disintegration revealed: (I) oblique parallel fragmentation cracks, 30" off the direction of the compression; (2) transverse forking fragmentation cracks, normal to the former; (3) release cracks parallel to the compression; (4) a disintegration cloud. Given a distribution of submicroscopic cracks of random orientation throughout the glass, Griffith has shown the highest tensile stresses to be located at the tip of the cracks 30" off the direction of the compression. Although the preexistence of such cracks or flaws in unstressed glass is doubtful, they can be generated by the distorting forces set up by the compression itself. Atoms, in a structure held together by Morse-type forces or bonds, have under tension both a stable and an unstable state-of equilibrium. The difference in energy level between these two states is a function of the tension, and this difference must be borrowed from the thermal energy level of the bonds to reach the unstable state and result in a broken bond. As the fraction of the bonds having at any instant a sufficient thermal energy level is a known function of that level, it is possible to determine, in function of the prevailing tension, the fraction of the bonds that are broken. Upon breaking of the bond uniting two atoms at the tip of an advancing crack, the two atoms accelerate away from each other, causing in the solid propagation of a

Jan 1, 1944

By Eugene F. Poncelet

This paper attempts to analyze the phenomena involved in the fracture of brittle solids by simple compression. Glass squares standing on edge, and compressed between two parallel steel jaws, developed jagged fractures, roughly parallel to the compression. Polarized light disclosed the presence of high local stress concentrations indicating poor contact all along the contact surfaces with the steel jaws. Teszar has shown that partial contact between the jaws and such a compressed solid causes tensile stresses in the free surfaces of the solid normal to the compression. Furthermore, Scoble has shown that in brittle solids fractures are caused by, and develop normally to, a sufficient tensile stress. It follows that such "partial-contact" cracks must be expected. With optically flat jaws and specimens, no "partial-contact" cracks developed. Instead, very smooth cracks vertical to the compression appeared 'upon release of the pressure. Investigation showed that the friction with the rigid steel opposed the return of the glass to its original width, after creeping on the steel as compression built up, thus setting up tensile stresses in the glass surfaces. It was not found possible to eliminate these "release" cracks. Compressed to destruction, such specimens flew to dust. A Microflash photograph taken during the explosive disintegration revealed: (I) oblique parallel fragmentation cracks, 30" off the direction of the compression; (2) transverse forking fragmentation cracks, normal to the former; (3) release cracks parallel to the compression; (4) a disintegration cloud. Given a distribution of submicroscopic cracks of random orientation throughout the glass, Griffith has shown the highest tensile stresses to be located at the tip of the cracks 30" off the direction of the compression. Although the preexistence of such cracks or flaws in unstressed glass is doubtful, they can be generated by the distorting forces set up by the compression itself. Atoms, in a structure held together by Morse-type forces or bonds, have under tension both a stable and an unstable state-of equilibrium. The difference in energy level between these two states is a function of the tension, and this difference must be borrowed from the thermal energy level of the bonds to reach the unstable state and result in a broken bond. As the fraction of the bonds having at any instant a sufficient thermal energy level is a known function of that level, it is possible to determine, in function of the prevailing tension, the fraction of the bonds that are broken. Upon breaking of the bond uniting two atoms at the tip of an advancing crack, the two atoms accelerate away from each other, causing in the solid propagation of a

Jan 1, 1944

By John A. Church

In the year 1874, when the price of pig-iron was still high, that staple product became the subject of discussion in the newspapers and among those philosophers who are determined to know the "reason why" for all things. The object of the inquiry was to find out Why iron did not fall in price as much as some other things had fallen. Indeed, there seemed to be an impression that iron and gold ought to move on nearly parallel lines, for the fact was constantly brought forward that iron then cost twice as much as it did before the war, though gold had returned to within twelve or fifteen per cent. of its former value. After some time the papers, by what seemed to be nearly unanimous consent, reached the conclusion that the maintenance of high prices was the work of the furnace proprietors, who, as the result of their shrewdness, were pocketing enormous profits. No proof was given to support this opinion, and the figures of cost and sale that were produced were of the most general kind, and used in the most vague manner. Visiting the president of the Thomas Iron Company, whose great works at Hokendaqua and Catasauqua, in Pennsylvania, are so well known as the largest and among the best conducted in America, I found him quite willing to meet the vague suspicions of the newspaper wiseacres, by a publication of the figures contained in the furnace-books of his company. These were accordingly given to the world through the Engineering and Mining Journal, and attracted the general attention they so well deserve. Though my own name became connected with these statistics, from the fact that I used them as the basis for the first attempt ever made to calcalate the furnace economy of an American works, I can, without the suspicion of egotism, say that the statistics referred to are the most valuable contribution the literature of American blast-furnace work has yet received, for their interest is independent of personal considerations, and due entirely to the fact that they represent faithfully the details of practice in one of the most prominent American works. They are an exact transcript of the furnace-books, and form a body of accurate information of the greatest value to the metullurgical student. Their importance is shown by the fact that though the theoretical conclusions which I based upon them have been several times attacked, not one of my critics has made the slightest contribution to

By T. Koulomzine

The theory of the Askania mag-netometer, as well as a complete discussion of all factors influencing magnetometer readings, is very ably described by J. Wallace Joyce.1 We will assume that the reader is thoroughly familiar with this theory and we will use whenever possible the same notations for the physical values involved in the theory. On the other hand, we believe that the principal formulas of the magnetometer can be deduced much more easily by the use of differential calculus than by the methods employed by Joyce. Consequently, we are presenting herewith a resumé of the theory of the vertical magnetometer in order to determine the correct formula of the temperature coefficient which we propose to compensate. Short Theory of the Vertical Magnetometers We shall accept the following notations: m = total mass of the magnet system, in grams. g = acceleration of gravity, in dynes. a = projection on the axis of the magnet system of the distance between the axis of support and the center of gravity of the magnet system, in centimeters. h = projection on the perpendicular to the axis of the magnet system of the distance between the axis of support and the center of gravity of the magnet system, in centimeters. M = 2pL = M, magnetic moment of the magnet system. Z = vertical component of the earth's magnetic field. ? = deflection angle of the magnet system. = focal length of objective lens in the optical system S = scale reading. In a vertical magnetometer the axis of the magnet system is kept horizontal within 2". Therefore, referring to Fig 1, we see that the magnetic torque of the magnet system influenced by Z must be counterbalanced by the gravity torque. MZ cos ? = mga cos ? + mgh sin ? or dividing by cos ? MZ = mga + mgh tan ? If we now move to another magnetic station and the temperature conditions change, the values of all the variables will change also: M, a, and h under the influence of temperature. 2, both with time and space. ? and S, as a result of all the above changes. Taking increments in Eq 2 we obtain: (Mo + ?M)(Z + ?Z) = mgao + mg?a + mg(h + ?h) tan (? +??) but tan (? +??), because ? is small, can be replaced by tan ? + tan ??; in fact tan (?+ ??) = sin ?.cos ?? + cos ?.sin ??/ cos ?.cos ?? — sin ?sin ?? but sin ? sin ???0 tan (8 + ??) ? tan ? + tan ??. Fur-thermore, neglecting all terms of the equation containing two differentials becomes: or a differential and a tan ??, Eq 3

Jan 1, 1950

By Fred E. Beasley

Metallurgists have, in the last ten years, overcome many difficulties of high zinc in lead blast-furnace slags. This problem was brought to the front at the close of the war, by the price obtained for lead and the high cost of labor, fuel, and flux. The Port Pirie plant of the Broken Hill Associated Smelters, Australia, early in 1919 had increased the zinc oxide in its blast-furnace slags from 13.5 to 20 per cent., at the same time reducing the iron oxide in the slag from 33.5 to 25.6 per cent. These results were obtained by the addition of 10 to 12 per cent. granulated slag to the roaster charge, improved conditions of the crushing, mixing, and roasting of the constituents of the sinter. By experiments and improvements, the practice at this plant has developed a normal slag that carries 22 per cent. zinc oxide, and at times runs as high as 33 per cent. zinc oxide, with only slight furnace trouble. The granulated slag is used not only as a diluent to the sulfur in the roaster charge, but on account of its value as a slag-forming material and the physical properties it imparts to the finished sinter. The ore is mixed with flux and given a double roast. All the first roasting is by sintering with the Dwight & Lloyd machines. The final roasting is performed partly by Dwight & Lloyd machines and partly by H. & H. pots. The charge to the blast furnace consists chiefly of the final sinter, slag shells, siliceous and non-bearing sulfur lead ores. High blast and ore columns are the practice. The greatest care is taken to obtain in the final sinter as low a sulfur content as possible. Attracted by the success of the Port Pirie plant, The Consolidated Mining & Smelting Co. of Canada, at Trail, B. C., in 1920 treated its crushed primary sinter with about 15 per cent. granulated slag. The Trail smelter at this time was short of ore and a large part of the charge to the roasters consisted of the zinc-plant residue having the following analysis: 12 to 16 per cent. lead, 12 to 16 per cent. zinc, and 25 to 30 per cent. iron. The lead in the final sinter was low and the physical condition was so bad that all the sinter had to be screened before it could be charged to the blast furnace. Also it disintegrated on being trammed and dumped

Jan 1, 1925

By John A. Church

IN the year 1874, when the price of pig-iron was still high, that staple product became the subject of discussion in the newspapers and among those philosophers who are determined to know the "reason why" for all things. The object of the inquiry was to find out why iron did not fall in price as much as some other things had fallen. Indeed, there seemed to be an impression that iron and gold ought to move on nearly parallel lines, for the fact was constantly brought forward that iron then cost twice as much as it did before the war, though gold had returned to within twelve or fifteen per cent. of its former value. After some time the papers, by what seemed to be nearly unanimous consent, reached the conclusion that the maintenance of high prices was the work of the furnace proprietors, who, as the result of their shrewdness, were pocketing enormous profits. No proof was given to support this opinion, and the figures of cost and sale that were produced were of the most general kind, and used in the most vague manner. Visiting the president of the Thomas Iron Company, whose great works at Hokendaqua and Catasauqua, in Pennsylvania, are so well known as the largest and among the best conducted in America, I found him quite willing to meet the vague suspicions of the newspaper wiseacres, by a publication of the figures contained in the furnace-books of his company. These were accordingly given to the world through the Engineering and Mining Journal, and attracted the general attention they so well deserve. Though my own name became connected with these statistics, from the fact that I used them as the basis for the first attempt ever made to calculate the furnace economy of an American works, I can, without the suspicion of egotism, say that the statistics referred to are the most valuable contribution the literature of American blast-furnace work has yet received, for their interest is independent of personal considerations, and due entirely to the fact that they represent faithfully the details of practice in one of the most prominent American works. They are an exact transcript of the furnace-books, and form a body of accurate information of the greatest value to the metallurgical student. Their importance is shown by the fact that though the theoretical conclusions which I based upon them have been several times attacked, not one of my critics has made the slightest contribution to

Jan 1, 1876

By M. Silva, G. Olivares

In a leach pile, the drainage layer or commonly known as the cover layer has been a matter of discussion due to the multiple phenomena or events that occur within it which impacts negatively the leaching process. All these events are directly related to the loss of draining capacity and irregular flow net due to an excessive compaction and entrapment of fines reducing the flow velocity which in turn cause more fines to get trapped. Over time the problem will continue to worsen rapidly and accelerating continuously like a geometric progression. When the cover layer gets to a certain level of deterioration, the only alternative to continue operating safely is to replace the entire cover material, a huge cost for any mining operation. TESRA has developed a solution to counteract the accelerated advance of the deterioration of the drainage layer. The solution consists of using large agricultural tractors coupled to a proprietary design of a heavy- duty subsoiler-ripper for deep tillage for loosening and breaking up the ground at various depth and with different legs spacing from 0.4 m or above. Also, if the ground still requires additional breaking, TESRA proposes the use of a rototiller that breaks any lumps to homogenize the ground to a certain particle size. With the subsoiler-ripper/rototiller combination equipment a layer of 0.5 m deep is properly treated to match closely the original conditions of the ground thus regaining the flow net as close as to the original seepage condition. Once the ripping is completed the subsoiler-ripper tool is replaced with another that is similar to a plow and thus continue with the installation of drainage pipes. The tool digs a Y-type profile trench allowing the pipe to sit precisely at the bottom of the trench, contributing additional lateral and lower support. This tool controls the installation depth of the drainage pipe at different inclination grades and also eliminating any irregularities that may block the flow of the leaching solution. Sometimes the obstructions block completely the flow of the solution. The main result of this methodology is the reduction of groundwater levels and at least 2% less humidity in the leaching gravel due to an increase in the draining capacity of cover. The methodology developed by TESRA applied during a couple of cycles of a dynamic leach pile has proven to be a real alternative to remedy and recuperate the cover close to its original conditions. The methodology requires only one large tractor and two accessories operated in different stages while replacing the more traditional and less efficient machinery such as the Bulldozer, Motor-grader, and backhoes.

Jan 1, 2019

By H. V. Winchell

It seems to me that uniformity of mining law in North America must of necessity start with uniformity in each country. I understand that the Provinces of Canada have the right to make their own individual laws and conditions, to prescribe the terms under which the prospector or the applicant for mineral titles may secure the same. That is not true of the states in the United States. In this country, the Federal Mining Law was made applicable to the public domain, which, at the time of the adoption of this Act in 1872, was chiefly west of the Missouri River; it is impossible to locate a mining claim in most of our eastern states. It is true, however, that there are certain holdover rights, mostly obsolete, arising from the fact that at the time of the formation of the Confederation of States, the land that belonged to each state became the property of the Federal Government. I am not quite correc-t in that, it was sometime after the adoption of our Constitution, I think it was 1778, but those rights were largely, by agreement, surrendered to the Federal Government; but it is a fact that the great acquisitions of territory in the West, with the single exception of Texas, were of domain that came into the possession of the Federal Government and our Federal Mining Law applies to that territory. No state has the right to (by legislative enactment) enlarge upon the rights given by the Federal Mining Law, governing the acquisition of mining claims in its territory, within its boundaries. Each state, however, may, and some of them do, limit somewhat the rights given to the prospector. For instance, the Federal Mining Law provides that a claim may be taken, staked upon the public domain, of 600 ft. in width and 1500 ft. in length. Some of the states, like South Dakota and Colorado, have at times said "that is an excessive amount of land for a single mining claim; we will therefore limit it to 300 ft. or 200 ft. in width." There is, therefore, not any direct and immediate need for uniformity of mining laws in the United States, such as there might possibly be if it were considered at all desirable upon the whole in Canada. The initiation of mining rights under the Federal Mining Act is the act of discovery of valuable mineral in place, referring now to the lode mining claims, but the same is true of the placer claim. It is true that the emphasis has never been laid upon the word valuable and that

Jan 1, 1920

By S. Yano, S. Takagi, Y. Nagaya, T. Akai, T. Ishida, Z. Bennour, T. Suzuki

"It is often reported that the concept of single-plane-fracture is not sufficient to account for the productivity from shale gas/oil wells. Many researchers introduce the concept of the stimulated reservoir volume to describe observed production behaviour. However, the characteristics of the stimulated reservoir volume are not well understood. In this study, to investigate the characteristics of fractures and surrounding region induced by the hydraulic fracturing, a devised hydraulic fracturing experiment was conducted using the cylindrical shale specimens in the laboratory.Two resins were prepared as the fracturing fluid for the experiments: cyanoacrylate, an instant glue, mixed with a fluorescent paint (resin A), and methyl metaacrylate, a thermosetting acrylic resin, mixed with a fluorescent paint (resin B). These resins were able to fix within the specimen after fracturing. This is because the present study aimed to only detect fractures induced by hydraulic fracturing. Cut sections of the specimens were observed under ultraviolet light irradiation. It is expected that the hydraulically induced fractures and the surrounding regions will be detected because the induced fractures filled with the resin should emit light, while the other parts will not.The specimens, which were collected from the Kushiro Coal Mine in Hokkaido, Japan at the depth around 275 m, were roughly 85 mm in diameter, 170 mm in length, and cored normal to the sedimentary planes. An injection hole with a 10-mm diameter was drilled onto the center of the specimen parallel to the sedimentary plane to simulate hydraulic fracturing in shale gas development. The experiment was conducted under a uniaxial loading condition of 3 MPa and the fracturing fluid was injected into the sealed injection hole at a constant flow rate using a syringe pump.After the fracturing experiment, the resins were fixed in the specimens and then the cut sections were observed under ultraviolet light irradiation. As a result, the hydraulically induced fractures, which were filled with the resin, are clearly observed. Detailed microscopic observations show that the main fractures are tortuous and are accompanied by many thinner ramified fractures. Additionally, fractured regions where the resin penetrates significantly are observed around the main fractures in the specimens fractured by the resin B. These induced fractures and fractured regions are considered to be the stimulated region in which permeability is improved. The necessity of estimation on the features of fracture tortuosity and ramified fractures is indicated for more accurate understanding of the well production and the stimulated reservoir volume."

Jan 1, 2015

By D. J. McAdam

In a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,t and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or poten- tial* fracture. An example of a rechnlcal cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stresst determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated Sz In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal eithel to S3 or 31). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 3I I to 324), attention will be confined to stress combinations with

Jan 1, 1942

By D. J. McAdam

In a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,t and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or poten- tial* fracture. An example of a rechnlcal cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stresst determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated Sz In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal eithel to S3 or 31). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 3I I to 324), attention will be confined to stress combinations with

Jan 1, 1942