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By Diego Ruben Martin

Este trabalho objetivou comparar, por meio de experimentos desenvolvidos em uma fundição, através da adição de três inoculantes diferentes (IM-22 com FeSi-Ba/Zr, G 20 com FeSi-Ba) e IMSR 75 com FeSi-Sr), para observar as características da vida util de ferramentas na usinabilidade das peças, microestrutura e propriedade mecânica. A metodologia da pesquisa se baseou em analisar principalmente pelo processo de fresamento o ferro fundido cinzento GG-25 inoculado diferentemente, denominado de acordo com a Norma DIN EN 1561. As formas de avaliação dos resultados terão como base uma análise quanto ao desgaste da ferramenta, acabamento superficial, análise microestrutural, composição química e propriedades mecânicas do material da peça. Os resultados mostraram que dentre os inoculantes aplicados o com Sulfeto de Estrôncio (SrS) apresentou-se mais estável termodinamicamente que as amostras inoculadas com os outros dois inoculantes, pois ele levou a uma variação da energia livre de Gibbs mais negativa e portanto mais favorável a formação de núcleos com raio crítico (rc) maiores, o que permitirá uma solidificação com nucleação heterogênea. Todavia o respectivo inoculante também foi mais eficiente na formação de uma microestrutura mais estável, com maiores quantidades de células eutéticas e vida útil do inserto mais longa quando usinado.

Aug 16, 2017

By Charlotte Abena Benyarku, Albert Buabeng, Eric Stemn

The abstract highlights the increasing use of mobile equipment in mining and its contribution to accidents. The research analyzes injury narratives from Ghanaian surface gold mines to understand the involvement of mobile mining equipment in injuries and explore the use of injury narrative data as a human factors tool. The study identified priority human factor issues, such as operating mobile mining equipment, part removal/installation, and the operator's cabin. The use of Sankey diagrams for multivariable analysis effectively identified specific operator-centered issues that warrant further investigation for improved safety interventions.

Mar 29, 2022

By R. Hogg, S. Chander

"An overview of various physical and interfacial characterization methods for use in mineral processing is presented in this paper. The availability of a large number of characterization techniques and commercial instruments requires that the reliability and limitation of each be carefully examined. The choice of a method may depend upon the manner in which the information is to be used, for example in development of new processes, plant and process design, plant control, problem solving, etc. IntroductionThe behavior of particulate systems, in mineral processing and other applications, is primarily affected by the physical characteristics (size, shape, surface area, pore structure) of the particles and the nature of the different solid-fluid interfaces present. Mineral processing operations in general are either designed to modify and/or control physical and interfacial properties (grinding, classification, agglomeration, etc.) or are strongly influenced by them (flotation, flocculation, solid/liquid separation, etc.). Characterization of physical and interfacial properties should be an integral part of the development of new processes, plant and process design, and routine plant operation.Our purpose in preparing this article has been to present an overview of the current status of characterization, to point out areas of weakness, and to discuss the reliability and limitations of the various techniques, procedures, and instrumentation available. For more detailed discussion of methods, etc., the reader is referred to standard references.Sampling particulate solids.Sampling is a critical factor in any procedure for characterizing particulate materials — the reliability of characterization data is ultimately limited by the extent to which the sample used was truly represen¬tative of the population from which it was taken.Sampling from particulate solids involves two basic problems: (1) the selection of a random', unbiased sample and (2) choice of the appropriate sample size.The problem of sample selection involves the design of a sample collection procedure that ensures that every kind of particle in the population has the proper chance of being included in the sample without bias. Specific procedures such as the well-known cone-and-quarter technique have been developed and are discussed at length in the literature on particle characterization (Allen, 1981: Herdan, 1960: Gy, 1982)."

Jan 1, 1989

By Guy Bridges

Site formation works requiring excavation of rock at the Queen Mary Hospital redevelopment site was restricted due to extremely low ground vibration limits. Non-Explosive excavation method called Electro Power Impactor (EPI), sometimes called Plasma Blasting, was utilised to help ensure the rock excavation rate could meet the required programme.

Feb 1, 2020

By S. J. Schofield

INTRODUCTION ONE of the greatest and most interesting problems of economic geology is to discover criteria bearing on the persistence of ore in depth. It is a question which is very much to the fore in connection with many of our important mines in British Columbia, and, as is to be expected, it is a most controversial one. The present paper, summarizing the views of the writers concerning the possibilities, in some of the mining districts of the Province, for the persistence of commercial ore to greater depths than have been, or are at present, mined, is offered as a basis for a general discussion on the subject and to call attention to the need for renewed study and research on the problem. Sooner or later-at some depth or other-individual ore-shoots become impoverished and bottom. In many mines where this has happened, either at the same or at a lower horizon, another ore-shoot, which had no surface outcrop, has been encountered and has continued downward until it, in turn, has pinched out. The Kolar (Mysore, India) gold field, for instance, presents interesting examples of this recurrence of ore-shoots at depth (1). For many years these mines have been under the management of Messrs. John Taylor and Sons. On more than one occasion they have persisted in sinking or drifting on barren quartz or along a narrow fracture for many hundreds of feet before values were again encountered. Their earliest experiment in such bold development took place at a depth comparable to that at which most mines in British Columbia are considered to be bottomed. In the past, is has been concluded that the ores in British Columbian mines do not go below approximately 2,000 feet, and to that depth only in a few instances. Consideration of the depth to which ores have been mined in the Province appears to support this view. The figures shown in the accompaning table have been gathered from various sources and are approximate.

Jan 1, 1935

By John M. Gomes

A two-phase molten halide-silicate technique for extracting tungstic oxide (WO3) from scheelite (CaW04) and wolframite [(Fe, Mn)W04] was investigated. The halide phase extracted over 99 percent of the was, whereas the lower silicate phase retained about 90 percent of the calcium, iron, or manganese oxides. The most efficient extraction and separation of tungsten from scheelite was accomplished at 1,080° C in a molten system containing, in mole-percent, 9 calcium tungstate (CaW04), 42 sodium chloride (NaCl), 28 sodium fluoride (NaF), 17 sodium metasilicate (Na2SiO3), and 4 alumina (Al2O3). Wolframite was also treated at 1,0800°C; the molten system contained, in mole-percent, 11 (Fe, Mn)W04, 74 NaCl, and 15 Na2 SiO3.

Jan 1, 1968

By G. J. Ballmer

Drilling and blasting 12-in. blastholes started about the middle of 1946 and has worked out so well that about one half of the blasting, formerly done with 9-in. holes, is now done with 12-in. holes. In most of this work the harder benches were chosen for the 12-in. size because these were the areas where the greatest difficulty was encountered in thorough fragmentation. At the start of this experimental work a 42-T Bucyrus- Erie churn drill was available whereas all other blasthole drilling was being done with 29-T rigs. The large drill was first equipped with a standard 9-in. string of tools but it could not drill quite as fast as the 29-T drills. How- ever, when the weight of the string was increased about 700 lb, or to a total weight of about 3500 lb, it was found that the 42-T exceeded the drilling speed of the 29-T by about 25 per cent. When a 12-in. bit was substituted for the 9-in. bit the 42-T rig again dropped behind the drill- ing rate of the 29-T rigs. At this stage of the experiment the 42-T was using a 12-in. bit with detachable points and the total weight of the string of tools was about 3800 lb. It took longer to change the drill points than it did the regulation bit and this accounted for some of the slower drilling rate. When a heavier stem and standard construction 12-in. bits were used it was found that the 42-T could drill a 12-in. hole slightly faster than a 29-T could drill a 9-in. hole even in the hardest rock in the pit.

Jan 1, 1949

By Ana Lucia Diegues Skury

Em investigação recente, verificou-se que a associação de zinco no sistema convencional de grafite e de liga metálica como matéria-prima utilizada para produzir diamantes em condições de alta temperatura e alta pressão, promove um aumento da produtividade durante o processo de síntese. A forma como a presença do zinco afeta o mecanismo dea nucleação e crescimento dos diamantes ainda está aberto à discussão. O presente trabalho estudou a influência de Zn sobre os parâmetros estruturais de grafite em 4,7 GPa e 1300°C, utilizando a liga Ni-Mn como catalisador. Os resultados indicam que o Zn melhora a ação catalítica para o crescimento dos cristalitos de grafite. Outro ponto importate é que, provavelmente, o aumento do teor de zinco induza dois mecanismos concorrentes durante a evolução do processo de síntese: nucleação/crescimento dos cristais de diamante e recristalização do grafite.

Aug 16, 2017

By Luisa Pérez, Carlos Perea

En el caso de la minería de zinc es común emplear el método de explotación por subniveles, el cual exige rellenar los caserones vacíos, por ejemplo, con una mezcla del relave producto de la planta concentradora de mineral con un 3.5% de cemento, lo que actúa como pilar de sostenimiento de la mina (0.8 MPa). Debido a que al momento de avanzar en los tajos aledaños se producía una dilución con el mineral a explotar y, también, a que era necesario automatizar las operaciones; se decidió evaluar el relleno detrítico cementado, mediante análisis de resistencia, estudios de factibilidad, rendimiento en costos, aumento de productividad y seguridad en la operación. Como resultado, se obtuvieron valores máximos de resistencia de los pilares de 1.6 MPa, lo que evitó la dilución y mejoró el porcentaje de recuperación de mineral.

Jun 30, 2020

Coal Transport Via Pipe Line Claimed Feasible Moving coal via pipe line is a project proposed by the U. S. Bureau of Mines. A pipe line 100 miles long capable of transporting 5000 tons of coal a day with water could be built for about $10 million. The cost of moving the coal 100 miles would be between $1.28 and $2.01 a ton.

Jan 9, 1951

By Yiming Zhang, Yunpeng Li, Chun Feng, Ruyang Cao

In the past, the selection of the number of empty holes between blast holes heavily relied on the subjective experience of practitioners, leading to unsatisfactory blasting results. The number of empty holes between blast holes is particularly important for the perforation of explosive cracks. In order to determine the optimum number of empty holes for inducing the propagation of explosive cracks, a two-dimensional numerical model of two-hole blasting was established based on the continuous-discontinuous element method (CDEM) using polymethyl methacrylate (PMMA) as the medium, and numerical simulation of various working conditions was conducted; the simulation results are basically consistent with the PMMA model test. Through the analysis of blasting parameters, the optimal number of empty holes with a certain hole spacing is fitted. The following conclusions are drawn: (1) There is a threshold for variable x. When x exceeds 50, the presence of empty holes cannot lead to the perforation of explosive cracks between the blast holes. (2) When x is below 50, the optimal number of empty holes inducing the propagation of explosive cracks can be determined through fitting an empirical formula. By adjusting the number of empty holes between blast holes, a significant spacing between adjacent blast holes was achieved, allowing for effective penetration, so as to effectively “induce” the direction of the fracture zone and make full use of the energy of explosives. The research results of this paper are conducive to the research of directional blasting technology and provide certain references for the optimization of precision blasting parameters.

Dec 19, 2023

By D Liu, M Edraki, L Berry

The presence of clay minerals slows down the settling rate and mitigates compact consolidation efficiency in tailings dewatering. The increasing use of saline water in many plants, affects the dewatering performance and presents the mining industry with new concerns. Efficient tailings dewatering is a great challenge due to the complexity character of tailings, especially clay mineralogy, clay content, and water chemistry. In this study, existing studies of the effect of different affecting parameters, such as pH, ionic strength and polymer flocculants, on clay aggregation behaviours are reviewed. The settling behaviours of three types of commonly observed clay minerals in tailings, kaolinite, bentonite, and illite, are also investigated in saline water with different water salinity as a case study. The settling rates of clay minerals are found to be affected by water salinity, clay type, clay content and polymer dosage levels. High salinity water hinders the settling of kaolinite but assists the settling of bentonite and illite through formation of different aggregation structures.The current work has indicated that strategies to promote more rapid dewatering could include adjusting flocculant type or level, water salinity and salt balance, and tailings solid content levels. Implementing dewatering strategies may require a more detailed knowledge of the effect of surface chemistry on the nature and strength of the clay inter-particle forces. This study discusses the research gaps on how to better manage clay-rich tailings in saline water and identifies new research directions. The development of a water management model, supported through in-line monitoring data to predict and manipulate of the surface chemistry of the clay component of the tailings could be a new direction to manage the issues associated with clay minerals in the process of dewatering and tailings management. Mine planning (water use requirements, operating costs) could benefit from more detailed studies on dewatering of clay-rich tailings in saline water. CITATION:Liu, D, Edraki, M and Berry, L, 2018. Improving the dewatering efficiency of clay-rich tailings in saline water, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 476–484 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

By J Ireland, S Rose

The main outcomes of mining grade control include defining the boundary between ore and waste and then categorising the ore according to which stockpile it should be sent. A typical workflow includes using general mine planning software (GMP) – such as SurpacTM, VulcanTM or CAE StudioTM – to create three-dimensional wireframes of mineralisation domains, which are used to flag the relevant samples of interest and constrain estimation during block modelling. Creating the wireframes is a time-consuming process, requiring reasonable software skills, patience and an eye for detail. The interpretations can be subjective; indeed, even the same geologist will give different shapes if presented with the same data at different times. The process can become particularly tedious when trying to carry out updates due to new data. It can also be difficult to test the effect of alternative interpretations or grade thresholds because of the time constraints in production.There are now several implicit modelling tools available, either as standalone programs or as add-on modules to existing GMP. In this paper, the authors develop a workflow using implicit modelling methods to show how grades and geology can be rapidly interpreted to provide effective domains for grade estimation. The paper includes case studies from open pit gold mines. The implicit modelling approach can save several days of interpretation work for each grade control drilling campaign.CITATION:Rose, S and Ireland, J, 2014. Implicit modelling applied to grade control data, in Proceedings Ninth International Mining Geology Conference 2014 , pp 115–122 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

By J. P. Norrie

The presence of copper-bearing rocks in the Coppermine River area of Arctic Canada has long been known. Word of this was probably circulated along the Arctic and Hudson Bay coasts by the Eskimo, who no doubt traded articles made from native copper among themselves. It is not known whether any of these ever reached the southern shores of Hudson bay. The Indians, also, knew of the occurrence of copper here, as they served as guides to Hearne when he made a successful trip over land from Hudson bay to the Coppermine river. The possible commercial importance of these copper deposits was early appreciated by the Hudson's Bay Company, who on December 7th, 1770, despatched Samuel Hearne from Fort Prince of Wales, now Fort Churchill, to investigate them. He reached the then unnamed river at a point forty miles from its mouth on July 14th, 1771, and after surveying the river to its mouth and naming it the Coppermine, he retraced his steps to Fort Churchill by an overland route. Hearne's expedition was probably the most remarkable and arduous ever attempted in Arctic exploration, but the actual value derived from the trip was entirely out of proportion with the hardships he must have suffered. He verified the occurrence of native copper, but his astronomical and other observations were of little value. Until 1912 no concerted effort was made to obtain any definite information as to the geology and economic value of the Copper-mine deposits.

Jan 1, 1931

By R. White, C. Doyle, B. Martin, C. Díaz, J. Hrepic, G. Cuthbert

"A program initiated to improve the service life of the magnesite-chrome refractory in the top blown rotary converters at INCO’s Copper Cliff Nickel Refinery identified slag corrosion as a major wear mechanism. Target slag compositions were established to produce a non corrosive slag. A plant program to achieve this target slag by the addition of controlled amounts of do lime and MgO was instituted. Addition to rates of the slag modifiers was optimized based on the slag composition assays achieved and the requirement to address corrosion effects on the high alumina content ladle refractory. A60% gain in TBRC refractory performance was eventually achieved and has been maintained yielding a net annual savings of $700,000. A significant gain in available converter operating time has also been realized.IntroductionINCO’s Copper Cliff Nickel Refinery utilizes two top blown rotary converters (Fig. 1) to smelt a blend of semi-refined nickel concentrates, producing a granulated nickel matte suitable for the adjacent carbonyl based nickel refining operation. The converter, or TBRC, operation is a batch process composed of feed charging and melting, oxidation of the excess sulphur in the matte, and reduction of residual nickel oxide. The final matte, containing about 78% nickel, is tapped at 1600°C from the converter and is granulated in a sluice equipped with high pressure water jets."

Jan 1, 1997

By L Nienaber

Mines, after becoming operational, often retain the same support and reinforcement designs developed in the feasibility study. These designs are conceived from limited drilling data and regional experience, thus having an empirical flavour. Empirical designs are perceived to be conservative, catering for a variety of failure mechanisms, which are not always present in mine development drives.This begs the question: If you were to invest in the necessary resources to obtain sufficient data and complete the required analyses for the different development drives in the mine, to engineer a support and reinforcement scheme with the capacity to match the demand of the rock mass, what is the difference in cost when comparing an empirical design to an engineered design?To answer this question, mine tunnels with different geometries and orientations, where detailed structural data was available, were selected for a comparative design exercise. Material cost, development cost, cycle time and professional fees for engineered and empirical designs were estimated. The outcome is summarised in the following paper that discusses the process followed and the results of the design comparison.

Mar 8, 2011

By L. H. McGuire, L. B. Berger

"During the period October 31 to November 4, 1944, tests were conducted by the Bureau of Mines in the Cascade Tunnel of the Great Northern Railway Cc. to determine the effects upon the tunnel atmosphere of the use of a Diesel freight locomotive through the tunnel. These tests were conducted at the request of the company to ascertain whether undesirable conditions might be caused by the discharge of Diesel engine exhaust gas in the tunnel, which was constructed for an electric haulage system and is not provided with equipment for mechanical ventilation. The use of Diesel locomotives on the division that includes the Cascade Tunnel was under consideration as a means of augmenting haulage facilities for the abnormal freight tonnage that must be handled during wartime.As the Bureau of Mines has received requests for information on the extent of contamination of the air of railroad tunnels by Diesel locomotives, this report was prepared to make available for general information the data obtained during the tests in the Cascade Tunnel. As the test locomotive is of a type used on other railroads, and as the bore and grade of the tunnel and the tonnages moved are comparable to such factors on other roads, the data should be sufficiently representative to be of general interest. The exceptional length of the Cascade Tunnel made it an excellent location for obtaining information on the composition of the tunnel air and on the movement of air induced in the tunnel by the passage of trains.The writers are indebted to C. O. Jenks, vice president of the Great Northern Railway Co., who arranged for the tests, and to the following personnel of the railroad who cooperated in the conduct of the tests: I. E. Clary, superintendent, Spokane Division, J. L. Robson, superintendent of motive power, oil and gas equipment, J. F. N. Gaynor, superintendent of electric operations, William Lowney, master mechanic, Spokane Division, and D. F. Marston and Verne Nelson, assistant superintendents of electric operations.The Cascade Tunnel is on the Spokane Division of the Great Northern Railway and pierces the Cascade Mountains some 49 miles west of Wenatchee, Wash. The west end of this division is electrified from Appleyard Station, 2 miles east of Wenatchee, to Skykomish, Wash., a distance of 73 miles. Until the described tests, only electric haulage had been employed through the tunnel since it was first opened to traffic in January 1929.The Cascade is the longest railroad tunnel in the United States - 7.79 miles. The tunnel is single-track, perfectly straight, concrete-lined, and is on an essentially uniform grade of 1.56 percent, ascending from west to east. The cross-sectional area of the tunnel is 314 square feet, the dimensions being shown in figure 1. Elevation of the east portal is 2,880 feet and of the west portal, 2,240 feet."

Jun 1, 1946

By F. Kongoli

Viscosity is an important parameter of copper smelting and converting processes. It affects the settling of matte or metal droplets in the slag, distribution of elements etc. Despite this importance, confusion exists on the viscosity of copper smelting slags as a result of disagreements found among the measured viscosity data of several authors especially in the slags of high ferric iron content and close to magnetite precipitation. In this work a critical review of the existing data on the viscosity of copper smelting slags has been first carried out and the effect of oxygen potential on the viscosity of copper smelting slags has been clarified through a coupled quantification of viscosity and liquidus temperature. The confusion in the literature has been clarified and the importance of the relation between the viscosity and the liquidus temperature is discussed.

Jan 1, 2003

By L. F. Pett

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

Jan 1, 1952

The topic of the relationships between energy and particle size has had two very active periods -the first in the early thirties, when the relative merits of Rittinger's and Kick's proposals were vigorously debated; the second in the early fifties, when Bond's Third Theory was published. This second stage has produced several contributions that have provided a clearer picture of energy /size correlations in comminution. Derivations of a general relationship have been made from single-fracture experiments and from the rate of comminution, and both methods arrive at an equation of the type E = AK-11? The constant A in the equation may be regarded as a grindability parameter. In some respects, an understanding of the phenomena associated with comminution is still lacking, and research in this field should therefore be expanded. T HE SEARCH for a relationship between the energy consumed in comminution and the size of the broken product has been long and varied, and many

Jan 1, 1964