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By Dr. G. R. Hoey

"Abrasive wear and corrosive wear of carbon-steel grinding balls occurred simultaneously during the wet-grinding of hematite iron ore, and Ni-Cu sulphide ore, in laboratory ball mills. Two corrosion control methods, the use of corrosion inhibitors and pH control, were investigated and found to be effective means of reducing or preventing corrosive wear. Up to 50 % reduction of total ball wear resulted from the use of these methods.Corrosive wear is caused principally by the reaction of oxygen in air with the carbon-steel grinding balls, probably by the following reaction: 2Fe + O2 + 2H20 - 2Fe2 + 40H-.IntroductionA survey of corrosion problems in Canadian sulphide ore mines was completed and the findings were published in 197l(l). The corrosion survey was used to develop the present research program with the objective to find more economical and effective ways and means to reduce the corrosion that occurs in the mining and metallurgical industry. The purpose of this presentation is to review the projects being done by staff members of the Corrosion Section, Extraction Metallurgy Division, on corrosive wear and its control in wet-grind ball mills.Both abrasion and corrosion contribute to metal consumption during wet grinding of ores. The total wear of grinding media in this study is assumed to be equal to the sum of corrosive wear and abrasive wear. A great deal of research and development has been reported in the literature on ways to improve the abrasion resistance of grinding media but only a small number of studies has been reported on the corrosion resistance of conventional grinding media."

Jan 1, 1974

By Freeman C, Dingley W

Addition of sodium nitrite-borax, sodium hydroxide, sodIUm carbonate, lime, or sodium silicate reduced wear of forged steel balls and cast iron slugs in wet grinding a complex zinc-lead-copper sulphide ore in laboratory ball mills. Maximum wear reductions were 27 per cent (forged steel balls) and 33 per cent (cast iron slugs) with sodium hydroxide and nitrite-borax additives respectively.Potential measurements of balls and slugs during grinding, and analysis of the data in terms of corrosion theory demonstrated that corrosive wear had been reduced or prevented by these additives. Actual per cent wear reductions were not equated to the rate of corrosive wear because of the undefined interdependence between abrasion and corrosion in wear processes. The investigation verifies that in laboratory ball mills corrosion is significant in the wear of grinding media and wear can be reduced by corrosion inhibition and pH adjustment.

Jan 1, 1978

By Shane Andrews, Gustavo Vasquez, S. K. Varma

"The effect of solutionizing time on the corrosive wear behavior of 2014 aluminum alloy has been studied using the impact and continuous mode of scratching techniques. The transient currents (TC) evolved during the scratching of the alloy shows a linear dependence between the TC and the solutionizing time. Higher TC values at longer solutionizing time may be associated with the changes in the concentration of quenched-in vacancies, grain sizes and dislocations generated during scratching. These factors can either create (a) anodic sites or (b) pitting corrosion resulting in the generation of transient currents. The results have been compared with the similar study carried out for 6061 aluminum alloys. This 6061 aluminum alloy shows similar behavior but the rate of TC generation per unit solutionizing time is much lower than 2014 aluminum alloy. The morphological changes in the impact scratches and the TEM investigation in the scratch have been presented in this paper.IntroductionThe technique of studying corrosive wear of composites of 6061 aluminum in the monolithic and reinforced forms has been reported in the literature [1]. Two modes of scratching, consisting of impact and continuous, in a corrosive environment indicate that the composites, containing 0.10 and 0.20 volume fractions of alumina particles (VFAP), have higher corrosive wear rates than the monolith. A graph between the transient current (TC) generated during the impact scratching as a function of solutionizing time at 540°C indicates linear relationship between them. A similar slopes for the composites were observed and the TC values were less for the composite containing lesser VFAP."

Jan 1, 1998

By Bryan de Barros Soares

O aço carbono é um material amplamente empregado na indústria petrolífera devido sua elevada resistência mecânica e baixo custo. Entretanto, este material não possui boa resistência à corrosão. Neste contexto a predição da taxa de corrosão é de extrema importância para estimar a durabilidade desses materiais. Atualmente isto é feito com base em dados de ensaios de perda de massa em um único tempo de imersão, estimando-se a taxa de corrosão. Contudo, para converter os dados de perda de massa em taxas de corrosão é imprescindível que os ensaios sejam conduzidos em diferentes tempos de imersão e que haja linearidade entre perda de massa e tempo. Neste trabalho a corrosão do aço carbono em meios saturados com CO2 foi investigada através de ensaios de perda de massa em três tempos de imersão diferentes, com o objetivo de avaliar essa linearidade. Dos quatro meios corrosivos estudados, um não apresentou linearidade entre PM e tempo de imersão, sendo portanto incorreto extrapolar os dados para taxa de corrosão. Os resultados comprovaram que a extrapolação de dados de PM para taxas de corrosão não pode ser aplicada sem uma prévia verificação da linearidade entre perda de massa e tempo de imersão.

Aug 16, 2017

By Lucas Coelho da Silva

Este trabalho tem a pretensão de problematizar a influência da corrosão, sobretudo nos trocadores de calor, e trazer a importância de um estudo prévio detalhado, ressaltando como o desempenho deste equipamento está relacionado ao material utilizado e como esse estudo pode nos abster de grandes dispêndios. Através de ensaios eletroquímicos vamos estudar o comportamento do aço Super Duplex UNS S32750 em diferentes meios corrosivos, com o intuito de avaliar se o nível de corrosão do material é adequado para as condições severas de operação impostas a um permutador de calor.

Aug 16, 2017

By José Helio Sobrinho

O aço inoxidável austenítico alto nitrogênio ISO 5832-9 tem sido promissor na manufatura de próteses ortopédicas com resistência mecânica e de corrosão superiores a do aço tradicional ASTM F-138. Pesquisas recentes têm mostrado que esta liga apresenta resultados peculiares, tais como alta resistência mecânica, resistência à corrosão e presença de partículas de segunda fase (Fase Z) na matriz. Entretanto, não se sabe como se dá a correlação mecânica-microestrutural e de taxa de corrosão em regiões de processabilidade industrial desta liga na conformação a quente de peças. No presente trabalho, ensaios sob torção a quente isotérmico contínuo e interrompidos foram realizados após tratamento térmico de solubilização a 1.200oC, por 300 s com intervalos de temperatura variando entre 1.000°C e 1.200oC com taxa de deformação variando entre 0,05 e 5 s-1, para investigar a trabalhabilidade e a taxa de corrosão sob tensão desta liga com aplicação da técnica de espectroscopia de impedância eletroquímica e caracterização microestrutural, conduzida por microscopia óptica e varredura. Os resultados mostram que a tensão de escoamento é sensível aos parâmetros de deformação. O pico de tensão diminui com o aumento da temperatura e redução da taxa de deformação com alta taxa de recuperação dinâmica com grãos alongados, gerando regiões de acúmulo de tensão. Essas regiões de instabilidade plástica apresentam grau de corrosão superior em relação às amostras como-recebidas. As regiões de alta temperatura e baixa taxa de deformação apresentam boa trabalhabilidade com microestrutura final refinada com grãos recristalizados dinamicamente.

Aug 8, 2018

By Raphael França Assumpção

O aço inoxidável duplex (AID) vem sendo considerado uma excelente alternativa para aplicações que requerem alta resistência à corrosão e alta resistência mecânica. Portanto, a microestrutura e as propriedades dos aços dúplex permanecem como temas de grande interesse. Para este proposito, estudou-se o efeito da laminação a morno em um aço inoxidável duplex. O foco deste trabalho foi o AID 2205 que é um dos tipos mais comuns de AID. O objetivo deste estudo é avaliar o comportamento eletroquímico do AID 2205 laminado a morno no ambiente de íons de cloretos utilizando medidas de polarização potenciodinâmica. O aço inoxidável dúplex foi laminado a morno à 600°C para 60 e 80% de redução na espessura e submetido a testes eletroquímicos. As amostras apresentaram uma excelente resistência à corrosão em solução de NaCl 3,5%. Apesar da existência de certa similaridade entre os valores de potencial de circuito aberto e potenciais de pites obtidos para todas as amostras, o número de pites observado foi maior na amostra CF, seguida pelas amostras com 60 e 80% de redução.

Oct 30, 2017

By Benjamin A. Allou

La seule production de columbo-tantalite réalisée en Côte d?Ivoire est effectuée dans les éluvions et alluvions du gisement d?Issia, localisé au Centre- Ouest de la Côte d?Ivoire. Même si un lien direct avec les roches granitiques est suspecté, aucune attention particulière n?a été accordée à l?étude de la minéralisation primaire en relation avec l?évolution des granites et des pegmatites associées. Ces granites, qui sont en intrusion dans les roches sédimentaires du Protérozoïque inférieur, comprennent : le granite orienté à deux micas (Lobo), le granite orienté à muscovite dominante (Bitapia) et le granite porphyroïde non orienté à muscovite dominante (Issia). Les résultats de cette étude indiquent que les pegmatites à Nb-Ta-Be-Li sont liées au granite de type Issia. Les observations de terrain et les analyses sur roche totale des granites montrent que: (1) les trois principaux granites de la région sont de type- S et peralumineux dans leur composition, (2) ils ont été formés pendant la période tardi- à post-Éburnéenne, (3) le granite d?Issia fut le dernier à se mettre en place. Les pegmatites liées au granite d?Issia ont été classées en quatre types : Apegmatites stériles; B- pegmatites à Be; C- pegmatites zonées à Nb-Ta-Be; et Dpegmatites zonées à Be-Nb-Ta-Li. Les pegmatites zonées comprennent les zones suivantes: I- zone à aplite; II- zone graphique; III- zone à albite, microcline, béryl et columbo-tantalite; IV - zone à microcline massive; V- zone à quartz. Ces pegmatites présentent une zone d?altération métasomatique externe. La minéralisation de Ta, Nb et Be est liées à la zone III des pegmatites de types C et D, et celle de Li est associée aux greisens qui peuvent être observés n?importe où dans les pegmatites de type D. Les données de terrain et de laboratoire indiquent que le granite d?Issia est la roche parentale des pegmatites à éléments rares de la région. Les pegmatites minéralisées sont distribuées seulement autour du granite d?Issia et les patrons des terres rares des granites et ceux des pegmatites à éléments rares sont très similaires. Les analyses de la composition totale montrent un lien génétique entre le granite d?Issia et les pegmatites à éléments rares. Il est suggéré que les pegmatites ont été formées durant le transport et la cristallisation fractionnée du magma.

May 1, 2003

By Ángel Cabello Victoria

El presente trabajo es un estudio técnico-económico del yacimiento minero Caridad, perteneciente a la compañía del mismo nombre en Lima, cuyas características permiten aplicar el método de corte y relleno en la minería en vetas de plata. Sin embargo, las bajas cotizaciones en el mercado mundial de los metales, minimizaron los costos de los mismos y aumentaron la productividad para mantener la mina en operación. Por tal motivo, se busca utilizar este método en pequeñas minas de plata que tienen características similares y asi recuperar el minado completo de los valores en plata, la eficiencia en la mecanización de los equipos de acarreo y volteo y reducir los accidentes por caídas de rocas, como causa principal en la minería subterránea.

Nov 8, 1982

By Bill Birch, Toby White, Ian Brown

costs/benefits of electronic detonators and measure their effects on primary crusher productivity. 9 blasts (3 shock tube blasts, 6 electronic blasts). Measured electricity consumption & primary crusher throughput. Change to Electronic Detonators resulted in a 7% reduction in primary electricity costs and a 15% increase in primary crusher productivity.

Feb 1, 2020

By J. A. Bullington

Mr. Chairman and Gentlemen: Some of you men may know who are familiar with what we have been doing in the C. F. & I. the last three years, that we have revamped our entire method of cost accounting from the pit bosses' report through to the final report of the president. The basic ideas involved in our cost system are not ours. They were given us by one of the best management and industrial engineering concerns in the country-at least they have that reputation. All I can do now is to try to pass on to you some of the ideas we have from various sources and the methods we have used in teaching those ideas to our coal men on cost accounting. This will not be a technical paper. I tried to put it in language we could all understand, and I hope I will be talking the same language that you gentlemen talk when you think about the cost of your coal in your own mines. Accounting of all kinds is classed as overhead expense, and as overhead expense we must justify its cost by the service it gives the rest of the organization. That brings us up to the problem of making the cost department pay its own way, just the same as the power plant has to pay its own way, and we have gotten away from the old idea of cost accounting in that it is sufficient within itself. The cost sheets are only so good as they enable the operating man to reduce his costs. The cost sheets are only just so much work for the cost department, and if they do not help the operating man to control his costs, we figure that is so much work wasted. The cost department works for the operating department.

Jan 1, 1931

By P. K. Chatterjee, G. D. Just

Sublevel open stoping is a low-cost method that has been successfully employed in many large underground mining operations. Where ore body characteristics are suitable, extensive mechanization and the use of large-scale stoping techniques can considerably reduce operating costs. However, careful long-term planning of mine layouts and large capital commitments in development are required prior to production. Implementing precise mine design procedures and carefully controlling operations can result in considerable benefits. This article describes methods to achieve these objectives based largely on mining operations at Mount Isa, where sublevel open stoping has been used since 1932. Ore won by this method accounted for a very large percentage of the total tonnage produced at Mount Isa. Major costs associated with sublevel open stoping can be classified as drilling, explosives and charging, and LHD costs. In this article, analyses of these costs are made with particular reference to stope design. Parameters such as burden, drill hole diameter, explosive density, specific charge, etc., are considered in establishing the interrelationship of unit operating costs, as well as total costs of sublevel open stoping.

Jan 10, 1981

By Esher Hsu

In Taiwan, around 8,000 tons of batteries are consumed per year but only 17% of which were collected and transported to oversea for recovery. The dumping used batteries have become a potential environment problem. How to increase the collection efficiency of used batteries is an important issue for EPA in Taiwan. Besides, some treatment plants are on process to get permission for recovering used batteries. The objective of this study is to estimate the recycling cost of spent consumer battery and explore the feasibility of recovering used consumer batteries in Taiwan based upon the principle of cost-efficiency. The recycling cost including collection cost and recovery is estimated based on a survey. Results would provide some suggestions to Taiwan EPA regarding feasibility of recovering used batteries in Taiwan.

Jan 1, 2006

By J. H. Gasper

A cost analysis for two alternative contour mining methods is performed to compare their results. The methods evaluated are: 1) Front-End Loader/Truck Haulback and 2) Scraper Haulback. Before performing these analyses, output requirements and equipment needs are found based on integrating geometric limitations with material properties. The final geometric design establishes limits on clearances, congestion problems and grades (with associated running surface materials). All calculations are kept practical, with assumptions clearly stated. The cost analysis is a unique, easily utilized compilation of several different approaches incorporating cost engineering principles. The strengths of each are welded into a cohesive entity once again aimed at practical application. The results are presented in tabular form.

Jan 1, 1992

By Ronnie Melker, Joel Webster, Daniel Belardo, Thomas Hyatt

Quality assurance of drilled shafts is highly dependent upon the construction practices used during the excavation and concrete placement process. When performed under slurry, challenges are presented in maintaining borehole stability and placing a uniform and continuous concrete column. An integral part of the construction operation is to evaluate the as-built integrity to verify the drilled shaft is capable of safely transferring the required loads to the subsurface. One such case is exemplified by a project involving the expansion of an existing public transit rail line. The expansion required the construction of a new bridge bent supported by seven 5-ft diameter drilled shafts of various lengths. The specifications for the project required both Cross-Hole Sonic Logging (CSL) and Thermal Integrity Profiling (TIP) for drilled shaft integrity evaluation. Both of these post-construction non-destructive tests (NDT) are designed to detect anomalies or the presence of regions with reduced concrete quality. However, the implementation and methodologies of performing each of these NDT methods are significantly different. Cross-Hole Sonic Logging uses ultrasonic signals emitted from a transmitter in one access tube to a receiver in another access tube. Thermal Integrity Profiling measures elevated concrete temperatures during the curing process using embedded thermal wires. Both test methods require different access details and therefore have different costs associated to the contractor or owner to perform. This paper provides a detailed comparison of these NDT methods as performed on these seven drilled shafts. The associated financial and time costs to perform each method are compared and the results are discussed in detail. INTRODUCTION Drilled shafts are a fundamental deep foundation element that can be designed to resist large bending moments as well as to transfer large compressive axial loads through weaker upper soils to competent soils or rock below. Often times when the upper soils are very soft, lateral support by casing or slurry, or a combination of both, may be used to hold back the softer soils from collapsing into the hole before concrete has been placed. Due to the blind nature of drilled shaft construction and the qualities issues that may arise during the placement process, it is essential for design engineers to specify non-destructive testing (NDT) to assess the quality of the as-built drilled shaft. CSL and TIP testing are two NDT methods designed to detect areas of lower quality concrete or soil inclusions throughout the length of the drilled shaft and determine its overall quality. Each method uses different technologies and methods to achieve the results required to determine the as-built condition of the drilled shaft.

Jan 1, 2019

By M. K. Buder

Over the years, electric utilities have been faced with declining steam-coal quality, rising coal prices, and increasingly rigid and costly environmental standards. As a result, many utilities are beginning to show an interest in coal cleaning to upgrade coal by improving its heating value and reducing its sulfur content. Among the primary questions that must be answered for these electric utilities is whether the cost of coal cleaning can be offset by the potential cost savings in coal transportation and utilization in power generation. In search of at least a preliminary answer to this question, the Electric Power Research Institute (EPRI) engaged Bechtel to perform a case-by-case economic study to examine the cost effects of using cleaned versus raw coal to fuel new electric power plants. The results of two of the seven cases studied which involved western coals - one from Montana and another from Utah - are reported in this paper. Coal cleaning appeared to be a cost effective option for the Utah coal, but only marginally economical for the Montana coal. These cost results are expected to change to favor coal cleaning to a greater extent with the application of the final revised New Source Performance Standards regulating SO2 emissions just issued by the EPA. The study was based on the regulation originally proposed by the EPA under the 1977 Clean Air Act Amendments.

Jan 1, 1979

By Karl-Michael Zett, Marcus Kirschen, Ashraf Hanna

"Bottom purging systems are well established practice in refining metallurgy and an increasingly important technology for the EAF process in steel industry. Direct bottom purging provides efficient mixing of the entire melt volume, increases turbulence intensity in the melt, and improves mass exchange of alloy additions or oxygen. Continuously rising columns of dispersed gas bubbles helps to increase removal of volatile components carbon monoxide and hydrogen avoiding boiling retardation. Direct purging (DPP) systems provide cost benefits especially for EAF processes where metallurgical constraints have to be considered, e.g. for alloyed and stainless steel production, foundries. Case studies are presented.IntroductionCurrent EAF Technology and Production Constraints in Steel IndustryApproxinlately 30 % of steel, 446 Mio. tonnes in 2011, was melted in the electric arc furnace (EAF) with increasing tendency due to its large flexibility in production volume and raw materials. With recent increase of prices for ferrous raw materials the requirements for production of high quality steels from low quality scrap, direct reduced iron (DRI) and blends of ferrous raw material in various qualities increased. Maximum yield of ferrous raw material, oxygen, carbon, and alloys and minimum energy costs are of high importance. At modem high productivity levels even small process improvements provide considerable cost savings. Such improvements are caused, e.g., by efficiency increase of mass and energy transfer in the EAF. Optimization of flow pattern in the steel bath is therefore important for efficient melting of scrap and DRI and high homogeneity of the melt [1-4]."

Jan 1, 2013

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Sublevel long-hole stoping is a large tonnage stoping method that may be used when there is a fairly com¬petent straight vein and the walls will stand with little support. It is a method that lends itself to a high degree of mechanization. Two variations of the method are illustrated in Figs. 1 and 2. Fig. 1 illustrates development typical for a 3-m (10-ft) wide vein. Initial development is by a lower haulageway, which may be either track or trackless. A two-compartment raise is driven in the footwall of the vein. At about 12 m (40 ft) above the main head¬ing, a footwall scram is driven and pockets through the vein are cut on 7.6-m (25-ft) centers. Each of the sub¬level blocks selected in this example is 45.7 m (150 ft) long by 12 m (40 ft) high. Stoping starts at the slot raises at stope endlines and retreats to the timbered raise. This retreat sequence starts on the lowest sublevel and the opening cut is blasted into the slot raise. This retreating system is used in order to provide a good safe back over the miners working in any given sublevel. A crawler-mounted down-the-hole or conventional pneumatic drill puts the holes down from one sublevel to the next on any desired drill pattern. Gravity flow brings the broken ore down to the scram level. The illustration shows a large slusher mounted at the timbered raise and slushing from the pockets back to the raise chute. Development of the ore block in Fig. 2 is typical for a 5.5-m (18-ft) wide vein. Initial development is by a trackless heading driven in the footwall of the vein. Pockets are then put into and through the vein on 7.6-m (25-ft) centers. For ease in using trackless rubber tired mucking equipment, these pockets are crosscut to the vein at approximately 1.04 rad (60°) rather than per¬pendicular. A two-compartment timbered raise is driven down from one level to the next. Endline slot raises are again put through from level to level. Each of the sub¬level blocks selected in this example is 40-m (130 ft) long by 33.5 m (110 ft) high. Stoping starts at the slot raises and retreats back to the timbered raise. Long holes are drilled from one sub¬level down to the next by starting between the lowest pair of sublevels and then working up the ore block. Mucking out of the pocket drawpoints is accomplished by rubber-tired load-haul-dump (LHD) units which then dump into a centrally located orepass.

Jan 1, 1982