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By Myron Rodgers

IN rock- or ore-crushing tests all data, in order to be valuable for study and comparison, should be obtained and tabulated under conditions as uniform as possible. The results of many such tests have no value for comparison, because of incomplete data and the lack of uniform conditions, screens and scale of sizing, etc. The purpose of this paper is to suggest a form of standard specifications which may be developed by the Milling Committee and approved by the A. I. M. E. for rock-crushing tests. It is not intended to discuss any theories of rock crushing. Reports of rock-crushing tests should include the following details: 1. Description of the machine employed (jaw or gyratory crushers, rolls, stamps, tube mills, Chilean mills, etc.). 2. Method and material (timber, concrete, etc.) of foundation. Much power is dissipated in the vibration of poor foundations. 3. Locality from which the rock or ore was obtained, and geological, mineralogical, and physical characters of the material. 4. The power consumption of the machine running with no load and with full load, the unit of power being 1 hp, per 24 hr. 5. The capacity of the machine in tons (of 2000 lb.) per 24 hr. 6. The duty in tons per horsepower-day. 7. The screen analysis of feed and product by the proposed A. I. M. E. standard screen scale. On the foregoing statement, the following remarks are offered.

Jan 9, 1915

By David Hoadley, Kenneth R. Maser, Ashok B. Boghani, James E. Billar, D. Randolph Berry, Mackenzie Burnett, Robert H. Trent

3. Application of Proposed Revised Standards 57.11-50 to Mines Surveyed 3.1 Subparagraph (a) 1 "Each mine shall have at least two separate and properly maintained escapeways from each working place to the surface which are so positioned that physical or environmental damage to one will not lessen the effective¬ness of the other, and which meet the requirements of travelways and man hoisting standards These escape¬ways shall reach to within 1000 feet horizontally or 300 feet vertically of each working place " 3. 1. 1 Discussion This section will deal with the 1000 and 300 feet distance requirements, Section 3. 5 will deal with the "environmental damage" provisions. During the course of the mine surveys, it became obvious that one mine's 1000 foot distance was nowhere near equal to another' s --not in the actual distance but in the time required to traverse that distance. This -Ls true for the 300-foot requirement as well. The most obvious example is that one could traverse down a 300-foot raise much faster than climbing up the same distance. The temptation is strong to recommend that the distance requirement be changed to travel times instead. However, this approach would unfortunately lead to a lot of subjective evaluations. Having a time requirement for an entire mine evacuation (Subparagraph 6) is one thing, time requirements and the requisite inspection proce¬dures for each working place are quite another.

Jan 1, 1976

By Kwang-suk You

This paper describes the characteristics of electric arc furnace slag (EAF slag) as an aggregate of concrete along with the chemical composition, leaching concentration of heavy metals, the soundness of EAF slag and compressive strength of mortar with 20% and 50% replacements of sand. The results showed that leaching of heavy metals and soundness of EAF slag satisfied Korean environmental limit and Korean Industrial Standard for aggregates. The compressive strength of mortar increased as the replacement ratio of EAF slag for sand increased.

Jan 1, 2006

By Mahoney M

For gas content analysis in recent years, it has been standard procedure to report results on a 'dry ash free' (daf) basis, under the assumption that gas only adsorbs onto coaly material. In order to test the relationship between ash and sas content, samples of various rock types were taken from two drillholes in the Newcastle Coalfield. The results of this study confirmed the correlation between gas content and ash, providing an accurate means of relating gas contents for all carbonaceous rock types.

Jan 1, 1995

By J. B. Bassett, E. S. Rowland

The experimental work reported herein firas inspired by the publication of a paper by Grange and stewart, in which it was suggested that at low chromium contents the effect of this element on the Ma point was greater than at higher chromium contents, It was argued that, if a small increase of chromium in solution in austenite caused a sharp lowering of the M, point, it might lead to an explanation of occasional unsolved stress cracking of SAE 52100 during heat treatment. Commercial hardening is carried out at temperatures of incomplete solution, leaving carbides available to enrich the austenite during any inadvertent increase in temperature, such enrichment of the austenite in chromium content would lower the M, point nearer room temperature where cracking is more likely to occur. According, it was decided to investigate the effect of chromium, in quantities between o and 2 pet, on the M, point of commercial steels. Fig I summarizes the data available on chromium steels. For purposes of comparison, the M, temperature values were corrected by means of the Grange and Stewart formulas to (I) Hypoeutectoid compositions: 0.50 pct carbon and 0.80 pct manganese (z) Hsereutectoid compos-sition': 1." pet carbon and O.35 pct manganese. Only the steels having compositions reasonably close to these corrected values were included, because large adjustments inevitably introduce increased error such as are experienced in using the formula. The straight lines of Fig I were drawn by the authors. These data obviously do not settle the question Of the effect Of chromium On the Md Oint, particularly at low chromium contents. While it is fairly well established that chromium causes the M, to be lowered when present in quantities of I pet Or more, Grange and Stewart1 have suggested that in small amounts (less than 1 pet) it is more effective in lowering the Ma point, while Klier and Troiano2 have suggested that in small amounts it is less effective in lowering the Ma point. Payson and Savage3 concur with Klier and Troiano and further suggest that the effect is not linear at these concentrations. Rose and Fischer4 report that small amounts of chromium actually raise the Ma point of a 0.20 carbon steel. Material Two series of 35 Ib induction heats were melted for this work, the analyses of which are given in Table I. One series was aimed at 0.50 pct carbon and 0.80 pct manganese with chromium additions in 0,25 pet increments from o to 1.5 pct with a single heat at 2.0 pct Cr. This series was considered necessary because the commercial heat treatment of SAE 52100 does not generally introduce more than about 0.60 pct carbon in solution in the austenite. The high

Jan 1, 1949

By J. B. Bassett, E. S. Rowland

INTRODUCTION THE experimental work reported herein was inspired by the publication of a paper by Grange and Stewart,1 in which it was suggested that at low chromium contents the effect of this element on the [M,] point was greater than at higher chromium contents. It was argued that, if a small increase of chromium in solution in austenite caused a sharp lowering of the [M,] point, it might lead to an explanation of occasional unsolved stress cracking of SAE 52100 during heat treatment. Commercial hardening is carried out at temperatures of incomplete solution, leaving carbides available to enrich the austenite during any inadvertent increase in temperature. Such enrichment of the austenite in chromium content would lower the [ M,] point nearer room temperature where cracking is more likely to occur. Accordingly, it was decided to investigate the effect of chromium, in quantities between o and 2 pct, on the[ M,] point of commercial steels. Fig I summarizes the data available on chromium steels. For purposes of comparison, the [M, ] temperature values were corrected by means of the Grange and Stewart formula§ to (I) Hypoeutectoid compositions: 0.50 pct carbon and 0.80 pct manganese (2) Hypereutectoid compositions: 1.00 pct carbon and 0.35 pct manganese. Only the steels having compositions reasonably close to these corrected values were included, because large adjustments inevitably introduce increased errors such as are experienced in using the formula. The straight lines of Fig I were drawn by the authors. These data obviously do not settle the question of the effect of chromium on the [M,] point, particularly at low chromium contents. While it is fairly well established that chromium causes the [M,] to be lowered when present in quantities of I pct or more, Grange and Stewart1 have suggested that in small amounts (less than I. pct) it is more effective in lowering the[ M,] point, while Klier and Troiano2 have suggested that in small amounts it is less effective in lowering the []M,] point. Payson and Savage3 concur with Klier and Troiano and further suggest that the effect is not linear at these concentrations. Rose and Fischer4 report that small amounts of chromium actually raise the-[M,] point of a 0.20 carbon steel. MATERIAL Two series of 35 lb induction heats were melted for this work, the analyses of which are given in Table I. One series was aimed at 0.50 pct carbon and 0.80 pct manganese with chromium additions in 0.25 pct increments from o to 1.5 pct with a single heat at 2.0 pct Cr. This series was considered necessary because the commercial heat treatment of SAE 52100 does not generally introduce more than about 0.60 pct carbon in solution in the austenite. The high

Jan 1, 1948

ABSTRACT Based on mineralogical characteristics and processing techniques required, gold ores can be classified into 12 types: gold placers, quartz vein-lode ores, oxidized ores, silver-rich ores, iron oxide copper-gold ores, porphyry copper-gold ores, iron sulphide ores, arsenic sulphide ores, antimony sulphide ores, bismuth sulphide ores, telluride ores, and carbonaceous sulphide ores. The bulk mineralogy and gold deportment of each ore type differs and may have a significant impact on ore processing and metal extraction. This paper summarizes the bulk mineralogy and gold deportment in some major types of gold ore and discusses the potential impact of gold deportment on gold metallurgy.

Jan 1, 2013

By W. Baguley

There are many carbon-in-pulp and resin-in-pulp screening devices which are at present in use in the South African mining industry. Each one of these screens has been an improvement, from the very basic air cleaned and 'EPAC' types which were either circumferential or · diametrical launder screens, which were developed for C I P in the early 1980's, through to the more sophisticated LAP and Kambalda Screens. Each one of these screens not only increased the performance or throughput per square metre than its predecessor, but also addressed maintenance problems, wear, reliability and the ability to handle surge conditions, thereby preventing tank overflow. However, very little theory has been presented concerning, the flow of pulp through internal screens, carbon loss and screen wear rates. Almost all this information is based on empirical data. The NKM screen continued this trend and was able to offer flow rates in excess of 180 cubic metres per square metre of screen per hour and an average screen life of approximately six months. However, with increasing sophistication of the screen and the use of wedge wire as opposed to the normal woven wire, the mass was significantly increased and costs began to mount. Therefore, Mixtec undertook to develop a radically different screen, which incorporated many of the advantages of .the existing equipment and combined a few of its own.

Jan 1, 1992

The achievement of high availability of the Grey River Gold Mining Limited's alluvial gold dredge is a tortuous task which will rely upon the dedication of all staff employed at Grey River Gold Mining Limited (GRGML). The persistent adoption of new technology and better operating procedures, combined with a strong commitment to preventative and predictive maintenance, is necessary. The Grey River Gold Mining Limited alluvial gold dredge is operating in a relatively low grade alluvial gold deposit and hence relies on a high throughput of ore to be profitable. To achieve the required high material throughput the dredge incorporates an overburden removal system which operates simultaneously with a conventional bucketline system.   The overburden system incorporates a cutter wheel and suction pump configuration which is dedicated to removing the low grade (5 mg gold/m3 gravel) overburden. The bucketline conveys the richer grade payzone gravels (up to 890 mg/m3 gravel) to the recovery plant mounted onboard the dredge. The horizon for the overburden system approximates 15 m and the payzone terminates at the interface of the old man rock and the alluvial gravels at a variable depth averaging 22 m.   From a mechanical engineering perspective the dredge is a machine comprised of two materials handling systems which are controlled by Operations Staff and electronic process control systems.   The availability of such a materials handling system is a function of four variables: 1. The environment that the machine operates in. 2. The ability of the people operating the machine. 3. The design of the mechanisms, the configuration that they are grouped in and the effectiveness of the process control systems which control the mechanisms. 4. The procedures utilised to predict the performance of the machines and the skill of staff to analyse that data and carry out effective maintenance.

Jan 1, 1989

By G. Reed Marchant

In 1967 the US Congress passed the "Clean Air Act." This act was amended in 1970. As a result it soon became obvious that American industries were going to have to make significant changes in their methods of operation to be able to comply with quickly developing federal regulations. One of the first regulations developed was on sulfur dioxide. For this reason, the nonferrous smelting industry was immediately affected by these new regulations. This is true because the smelting process produces sulfur dioxide as a natural product of combustion when the sulfur-bearing ore is oxidized to purify the metal. The federal government's approach for setting standards included public hearings at which government, industry, and private citizens were allowed to make comments. For this reason, it was apparent to Kennecott's Metal Mining Div. that it was desirable to make a maximum effort in working with the agency to see that correct input was provided at the hearings to assist the government in setting fair and realistic standards. It was also obvious that more complete monitoring would be required. It soon became apparent that the efforts required input from many levels of the corporation from the chief executive to the plant operators. In order to accomplish both the administrative and the technical needs, an organization was needed that could respond quickly to these needs, obtain information from diverse sources, and bring it to focus in an organized approach to the problem.

Jan 1, 1979

By H. R. Gardner

THE wide variation in tension and compression properties' which exist between the a,ß and ? phases of plutonium indicate that a study of the impact properties would give a greater insight into the variables affecting the mechanical properties and formability of plutonium. In tension the a phase is quite brittle with elongations less than 1 pct. The ß phase at normal test speeds is unusually ductile; elongations up to 621 pct have been obtained. However, the ß phase is quite sensitive to strain rate; elongations decrease rapidly as strain rate is increased. The y and 6 phases are less ductile than the ß phase, with the 5 phase the more ductile of the two. Neither are very sensitive to strain rate. Considering the brittle nature of a Pu in tension and compression and the strain rate sensitivity of ß -phase Pu, both unnotched Charpy specimens and standard V-notched Charpy specimens were fabricated for testing. In addition, since iron in amounts greater than 500 ppm forms a Pu-PlgFe eutectic network which inhibits microcracking in as-cast plutonium,' it was decided to prepare impact specimens with low and high iron contents. Chemical analysis and density data for the impact specimens are presented in Table I. A total of 55 impact specimens, 32 unnotched and 23 with the standard Charpy V-notch were tested at temperatures from —45' to 330°C. The test results for both unnotched and notched specimens are illustrated in Fig. 1.

Jan 1, 1962

To the members of The South African Institute of Mining and Metallurgy I have audited the annual financial statements of The South African Institute of Mining and Metallurgy set out on pages 641 to 644 for the year ended 30 June 2006. These annual financial statements are the responsibility of the Institute?s Council members. My responsibility is to express an opinion on these annual financial statements based on my audit. Scope I conducted my audit in accordance with International Standards on Auditing. Those standards require that I plan and perform the audit to obtain reasonable assurance about whether the annual financial statements are free of material misstatement. An audit includes: ?examining, on a test basis, evidence supporting the amounts and disclosures in the annual financial statements, ?assessing the accounting policies used and significant estimates made by management, and ?evaluating the overall financial statement presentation. I believe that my audit provides a reasonable basis for my opinion. Audit opinion In my opinion, the annual financial statements present fairly, in all material respects, the financial position of the Institute at 30 June 2006 and the results of its operations and cash flows for the year then ended in accordance with generally accepted accounting practice appropriate to the Institute. Supplementary information The supplementary schedule does not form part of the annual financial statements and is presented as additional information. I have not audited this schedule and accordingly do not express an opinion on it.

Jan 1, 2006

By Miguel Araújo Vasconcelos Filho, Ulisses Tetti

In order to determine the influence of the number of increments of the sinter sample on the variability of the chemical composition obtained in Sinter Plant production, a variographic study was carried out. It was performed with two series of samples. The first series was composed of 40 samples with a 6-second interval, with the objective of obtaining the so-called minimum practical error, resulting from the extraction processes of the sample of the conveyor belt and preparation and analysis in the laboratory and the second series composed of 60 samples with a range of 4 minutes, which points to the sample selection error. The results obtained showed that the minimization of the sample selection error occurs for increments spaced in 15 minutes, being in agreement with ISO 3082, but with significantly better results than those currently practiced for intervals between 1hour increments. It was then decided on the suitability of the sampler to allow hourly increments to be collected and on a second opportunity to install a sampling tower with primary and secondary samplers to collect increments every 15 minutes including sample size reduction.

Oct 1, 2019

By B. Das, V. Hucka

In light of the present energy crisis, coal appears now as the only reliable source of energy. The production of coal is expected to regain a rising trend after the latest slump. Aside from the competitive price with natural gas and oil, the other retarder of coal production is its environmental pollution effect. Gasification of coal has been advanced as the partial solution to the energy-ecology crisis. All the Advantages In situ gasification of coal may be considered a further step in this direction. It carries all the advantages of normal coal gasification as a stable source of energy; it is clean to handle and nonpolluting of the environment. It partially and sometimes fully does away with mining, makes uneconomical low grade coal and lignites economical and places some formerly unworkable seams within the scope of modern technology. The seams posing problems for mining due to tectonic disturbances may be converted into gas in situ. Low grade coals, such as lignites, which otherwise would be left in nature as uneconomic wealth, may be exploited by underground or in situ gasification. Thus, in situ gasification is likely to increase the exploitable reserves of coal and may reduce the cost of production. An additional merit of this method is that coal substance never reaches the surface and there is no problem of cinder disposal.

Jan 8, 1973

By M. S. Janssen, J. L. Taylor, M. A. Ford, E. B. Viljoen, H. A. Simonsen

The Stilfontein and Buffelsfontein Gold Mines, near Klerksdorp in the Transvaal, had accumulated a large amount of uranium-containing residue and, when the price of uranium rose in the 1970s, consideration was given to the possible recovery of this uranium. Preliminary tests showed that concentration of the uranium prior to leaching would not be economic. However, the pyrite in the residue could be concentrated by flotation, and the flotation concentrate could be roasted to pro¬vide both enough acid for leaching the uranium and a calcine from which the gold could be recovered. The feasibility study showed that a uranium operation of 270 kt per month would be most economically attrac¬tive, and a plant of that size was accordingly designed and built. In the first six years of its existence, the plant treated over 20 Mt of residue and produced about 3,5 kt of uranium oxide. During that time, the plant was con¬tinually being improved to make it more reliable and cost-efficient. This paper analyses the operation of the plant during its first six years from the viewpoints of its mechanical, process, and economic performance. The criteria on which the selection of the process was based are reviewed and compared with the actual performance of the plant, emphasis being placed on the leaching, solid-liquid separation, recovery, and purification stages.

Jan 1, 1987

By Donald E. Dickie

At Sifto Canada's Goderich salt mine near Goderich, Ontario, the stress control method of mining was introduced to overcome increasing roof failure problems. After successful testing, the conventional room-and-pillar mining method was converted entirely to the stress control method. This conversion has successfully stabilized the ground, increasing the extracton ratio from 38% to 50%. At the same time, it reduced the ground control and rehabilitation cost by more than 80%. The success of the conversion is due to the quantitative technique of mine analysis that was developed to ensure that the conversion was safe and reliable. A finite element model of the ground was constructed to predict and validate the safety of the stress controlled mine opening. A system of in situ stress-property measurement instruments was developed to obtain input data for the computer model as well as to evaluate its output quantitatively with regard to time and location. History of roof problems Severe roof safety problems have been encountered at Goderich Mine since it opened in 1959. These problems continued into the 1980s. Following excavation, the roof strata started to deteriorate by slabbing. Even after roof rehabilitation, deterioration continued to progressively higher roof elevations, requiring repeated rehabilitation. The operational impact of these roof failures was severe, ranging from a considerable safety hazard to major production interruptions.

Jan 1, 1993

By P. S. Rudman

Crystal-structure occurrences in binary phase diagrams of fifth- and sixth-period elements, zirconium through palladium and hafniun through platinum, are surveyed with the conclusion that average .group number GN is the dominant correlative. This is interpreted in terms of curves of enthalpy differences between competitiue structures that are continuous func -tions of GN. It is further reasoned that the absolute formation energies may also be continuous functions of GN and if so can be simply obtained from the regatives of the heats of sublimation of the elements. It is shown that from such a heat of formation-GN plot the heats of mixing of alloy systems can be quantitatively predicted. For example, for sixth-period alloys very large exothermic heats of mixing are predicted for systems that span GN = 6 (i.e., Hf-Pt, Ta-Os, ...). Thermodynamic data with which to test these predictions do not appear to be available. If in the latter transition (A subgroup) elements we eliminate from consideration the fourth period where magnetic terms contribute to crystal-structure stability, as shown in Table I, there is a very regular sequence of structures. Using the group number (GN) assignments of Table I, which differ from the usual assignments of 8 to the elements listed as 8 through 10, it is perhaps trivial to say that the phase occurrence is a unique function of GN. However, the interesting aspect of alloys from among these elements is that the alloy structures also appear to correlate with GN. The GN for an alloy is given as GN = SixiGNi, where xi and GNi = atomic fraction and GN of component i, respectively. It is difficult to credit anyone in particular with having discovered this correlation, so many people have implicitly or explicitly employed it in the last 10 years. However, Bloom and Grant' in their correlation of s-phase occurrence with total number of d + s electrons were certainly one of the first to appreciate the correlation and Hume -Rothery2 appears to be responsible for the use of average GN as defined above. Only in the last few years have the phase diagrams of enough alloy systems been determined for this correlation to be surveyed with any degree of completeness. We shall here, in an abbreviated format, summarize the available phase-diagram literature and discuss some consequences of the GN correlation. BINARY DIAGRAM SURVEY In Figs. 1 to 3 the phase occurrences are summarized. In Fig. 4 the occurrence limits for the more common structures are summarized. That the GN of an alloy falls within the occurrence limits of a particular structure is a necessary but not a sufficient condition for the appearance of that structure. The preceding figures certainly illustrate that GN is a dominant correlative for structures of these elements. Many of the diagrams have complications, with ordering and other phase changes occurring in intermediate phases, that cannot be clearly accommodated in such a schematic survey and have hence been left out. Such omissions cannot, however, negate the GN correlations discussed here. Relative atomic size is also an important crystal-structure determinant and since atomic size varies quite regularly with GN for these elements it is possible that the apparent GN correlation is basically a size correlation. In Fig. 5 relative size occurrence limits are given for intermediate phases and extended terminal solutions. The relative volume, ?v/(v) - 2(vA - vB)/(vA + uB), has been taken as the measure

Jan 1, 1968

By J. B. Gwiazda

In regard to the paper by J.B. Gwiazda, it makes a highly technical approach to show that the .u factor used by designers of lemniscate-guided roof supports has never really been confirmed as a maximum and assumes that convergence is vertical. Also, the paper does not appear to take into account deflection of structures, which occurs when the lemniscate and base members are fully loaded to their maximum stress level, nor the front to back line of the support in relation to differential roof to floor movements caused by strata movements under pressure. It is not unusual for differential movements to be slightly diagonal to the line of the support, particularly in faulted areas and on gradient faces. The paper also does not take into account consolidation of fines immediately above and below the support. Generally speaking, any differential movement is from face to waste and under these conditions the .u of 0.3, which appears to be an international standard, has worked in practice. However, if the face end of the support is lower than the waste end, then the µ of 0.3 can be considerably increased, giving rise to the damage mentioned in the paper. The ideal design should aim for a slightly forward bias in the lemniscate guide so that the last increment of setting is toward the face, tending to close any fissures that may have developed during the support advance cycle. The support should also be fitted with positive set valves to ensure that a high setting load density is attained to minimize bed separation. As far as powered supports are concerned, convergence is irresistible and all powered supports converge at their rated yield load. A similar principle can be applied to the differential roof to floor movements to drastically reduce the very high forces that would otherwise be applied to the lemniscate structures and pins and that, in turn, are transferred to the base arrangement and floor loading. Any differential movements are usually catered for by the 0.3 µ factor or deflection of structures in the lemniscate guide arrangement and consolidation of the floor. The floor loading, due to differential movement, is in addition to the support convergence load and requires additional bearing area to avoid possible floor penetration. Some seven years ago, Fletcher Sutcliffe Wild Ltd. (FSW) introduced a lemniscate-guided shield support where the lemniscate linkage is connected to the roof bar through two horizontally converging rams to allow differential movement to take place above a given rated figure. This is a known force and can be guarded against, whereas with rigid connections the forces, as yet, are unconfirmed. By careful design, a horizontal force in excess of 6 MN (60 tons) opposes differential movements for a total ram loading of only 2.5 MN (25 tons), or 1.25 MN (12% tons) each. This principle can considerably reduce the length and weight of the support in comparison with a rigid pin-type structure; also, the yield load rating can be increased without affecting the lemniscate forces. The graph shows the tensile and compressive forces in a lemniscate linkage of a support with and without hydrostore. These forces react into both the roof beam and base members and, as can be seen from the support height to linkage load graph, a considerable reduction in these reactions is gained by the use of the FSW patented hydrostore system. Floor loading is considerably reduced under maximum µ conditions, and by allowing the roof bar to move with the strata, some degree of improvement to strata control is achieved in line with the assumptions in the paper. In practice, these movements have only been in the region of a few millimeters, which, in turn, reflects on the improvements to strata control by the addition of positive set valves.

Jan 1, 1987

By S. H. Boshlov

Los problemas del diseño estructural de minas se basan en numerosas complejidades como el de una incapacidad para definir explícitamente los parámetros, físicos in situ del medio rocoso y sus materiales asociados, el comportamiento del sistema bajo alguna descripción bosquejada de la aplicación de la carga; otras proceden de la incapacidad para especificar con precisión el estado inicial de la tensión en el medio. En ese sentido, el presente texto señala un modelo de diseño estructural manipulable donde se refleje el comportamiento del prototipo que representa. Así, el análisis por elementos finitos, en el cual la estructura se considera como una reunión de numerosos elementos, cuyas propiedades individuales pueden señalarse a voluntad, indica la posibilidad de un problema de diseño de una estructura minera como un ejercicio de análisis de sistemas. Tal sistema rocoso podría incluir todos los factores naturales y los medios ambientes artificales que afectan la resistencia de las estructuras mineras de superficie interior.

Nov 1, 1971

By Robert S. Moehlman

Is recent years, exploration and development by diamond drilling has become a popular attack on the problem of finding new ore shoots and developing extensions of known shoots. Diamond drilling has been used in recent years in extensive sampling of disseminated copper deposits, in preference to churn drilling. At the Harvard gold mine, near Jamestown, California, Philip R. Bradley, Jr., general superintendent, reports that: "The Harvard work was a singular example of the reliance that can be placed on diamond drilling in evaluating an ore body." Drilling at the Harvard mine will be discussed in a subsequent section of the paper. So popular has diamond drilling become, indeed, that many who have had slight experience with drilling are inclined to regard it as a cheap cure-all. This paper proposes to show some of the limitations on the application of drilling and some of the steps that should be followed in drilling programs to obtain reliable results. Cost oF Drilling Drilling has the advantage, over shaft sinking and drifting, of lower cost and greater speed. As compared with churn drilling, the advantage of bringing up a coherent sample instead of fine cuttings for a sample is obvious. Diamond drilling is generally more expensive, however, and more engineering effort is required in combining the average of core and sludge samples. Exploration of drift walls by hammer drills to a depth of less than 30 ft. is generally cheaper than diamond drilling, but beyond this depth the diamond drill is cheaper and more efficient. The principal factor in determining cost of drilling is the character of the ground drilled. Broken ground with many fissures is most expensive to drill. Hard ground is less of a problem with diamond drills than with ordinary machine drills. Hardness or softness affects only the speed of drilling. Caving, easily washed, sandy or clayey ground requires continual cementing and casing, which is the greatest cost problem. Thus, soft, incompetent ground is more costly to drill than hard, massive ground. Cost of drilling per foot in good to fair ground varies from less than $I per foot, achieved in some districts in Ontario, the Mother Lode, and Missouri, to $5 per foot in many western states. In bad drilling ground, costs may range from $5 to $15 or more per foot, thus approaching the cost of drifting. Other factors that increase the cost per foot are: I. Short total length of hole drilled, so that the cost of moving in crews and equipment must be spread over a relatively few feet. 2. Cost of drilling per foot increases with depth beyond about 500 ft.. as more time is consumed in pulling rods' 3. Cost of preparing drilling stations, getting air underground to stations? etc' 4. Inaccessibility and lack of water on the job make for higher costs.

Jan 1, 1946