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By R. D. Westervelt

IN JULY 1954, the first iron ore from the deposits in the Knob Lake vicinity was mined and shipped to Seven Islands. The beginning of the production may well be the greatest advance in Canadian economy in the past fifty years. In 1955, the Canadian iron ore output was more than twice that of 1954, an in-crease due principally to the iron ore shipments from the Knob Lake area. Previous to this, Canada's iron production was 1.85 per cent of the world's output. Now, this percentage has risen to approximately 5.5 per cent and shows every indication of increasing further as the production rate from the area is increased. In 1955 the Iron Ore Company of Canada shipped over 7 million tons of iron ore and expects to ship over 10 million tons in 1956. Present facilities may be enlarged to handle up to 20 million tons per year. If the expected I 0 million tons per year are exported, Canada's iron exports will be second only to her wheat and newsprint exports. Before production began from the Knob Lake deposits, years of development were necessary in the area. Geologists spent several summers mapping the general area in order to unravel the complicated stratigraphy and petrology. Particular attention was paid to the mapping of the iron formation and all stages of alteration from fresh iron formation to ore were observed.

Jan 1, 1957

By M. J. Ferrante

The preparation and subsequent reduction of vanadium trichloride was investigated as a process for the production of high-purity vanadium metal. Chlorine was reacted with vanadium oxide in the presence of carbon to vanadium tetrachloride. The contaminant vanadium oxytrichloride was also formed and later converted to tetrachloride by reaction with chlorine in the presence of activated charcoal. Vanadium tetrachloride was stirred continuously during thermal dissociation to vanadium trichloride powder from which volatile contaminants were removed by distillation. The vanadium trichloride powder was reduced to vanadium sponge by controlled feeding of the trichloride into molten magnesium. Excess reductant and byproduct salt were removed from the vanadium sponge by high-vacuum distillation. Batches as large as 6.5 kilograms of sponge metal were prepared at an average reaction efficiency of 98 percent. Vanadium sponge after consolidation by arc-melting contained 820 to 11330 ppm of interstitial impurities and had a hardness between Rockwell B 40 and 70, This metal exhibited excellent workability at room temperature. The purification of impure vanadium metal by electron-beam melting resulted in lowering interstitial impurities from 1,890 to 504 ppm and hardness from Rockwell B 87 to 28.

Jan 1, 1968

The beach sand industry is the only Australian mineral industry which, so far, has used electrostatic separation on a full commercial scale. The main application of the process is for the separation of rutile from zircon. Laboratory work has demonstrated that electrostatic methods are effective for ilmenite-monazite and cassiterite-zircon separations, thus facilitating the production of high grade monazite and cassiterite products from top-strip table concentrates, but the commercial application of these processeshas been rather limited so far. Overseas, electrostatic methods are becoming more widely used for the treatment of granular material which is not amenable to flotation. The use of electrostatic separation for the treatment of ores other than beach sands has been investigated in this laboratory 1, 2, 3, and two pilot separators were built during i953 to carry out field tests. Of these, the first was an inclined feed plate separator for the removal of silica from granular lime-sand in Western Australia, the second a conductance roll separator for the removal of diamonds from alluvial gravel in New South Wales. In both cases the whole of the ore passed through the separator, not a gravity concentrate only as in beach sand practice. The two types of separator mentioned are shown diagramatically in Figs. 1 and 2.

Jan 1, 1954

By R. Chryst, R. Kus

In the Integrated IT Systems proposed by Przedsie? biorstwo Robót Geologiczno-Wiertniczych—Geological and Drilling Company (PRGW) a network connects a database with all of the departments of a mining division. A key element thereof is the common graphics environment—Micro Station supplied by the firm, Bentley. PRGW’s SoftMine software and GMSI’s CadsMine that are used to design and schedule mining operate in this environment. The digital form of the design project facilitates the verification of information and the integration of technical solutions with SAP R/3 type financial—accounting systems. The system has been successfully implemented in the Brown Coal Mine in Poland.

Jan 1, 2003

By Warder A. Shaver

INTRODUCTION Mine No. 95 of the Consolidation Coal Co., a divi¬sion of Continental Oil Co., is located in Shinnston, WV. Mine No. 95 was opened in January 1967 and has been operating continuously since that time. The mine has 544 hourly and salaried employees. In 1974 this mine was the largest coal producer in West Virginia. Opera¬tions include two longwall sections and eight Marietta Borer continuous miners. Originally there were also two Jeffrey Heliminer continuous miners, but use of these was discontinued in June 1978. Production averages 12 727 t/d (14,000 stpd). Two portals now service the mine with a third portal proposed for 1978. GEOLOGY Mining takes place in the Pittsburgh coal seam which dips 0.04 rad (21h°) northwest. The working seam is steam coal with an average height of 2.2 m (86 in.). Immediately above the seam, the overburden is com¬posed of 6.1 m (20 ft) of shales followed by 6.1 m (20 ft) of weak sandstone and 51.8 (170 ft) of shaley limestone. The shale weathers when exposed to mois¬ture. Surface water has been a problem due to a fire¬clay floor which turns to paste when mixed with water. GENERAL DESCRIPTION The present mining plan for Consol's No. 95 mine has a proposed life of 35 to 40 years. One set of main headings has been driven in the north direction and one set in a western direction. These are the areas in which the majority of the virgin coal lies. The panels are driven off of the mains at 2134 to 3048-m (7,000 to 10,000-ft) intervals. The continuous mining sections are set up 133 x 610 m (600 x 2000 ft). Longwall sections are 259 x 1220 m (850 x 7000 ft). All of these mining sections are retreating uphill. Development is with continuous miners. A Marietta Borer miner is used for the main development work. These miners cut 4.1 m (13 ft 6 in.) wide and 2.1 m (7 ft) high, leaving an arch cut. Along with the arch, a 152.4-mm (6-in.) thick layer of roof coal is left in place to protect the roof from weathering. The combi¬nation of the self-supporting arch and the roof coal al¬lows the entries to be driven 30.5 m (100 ft) without bolting. Full bolting, however, is carried outby the last open crosscut. This bolting is done on 1.5 x 1.5-m (5 x 5-ft) centers using 1.5-m (5-ft) long roof bolts with tandem anchors and 152.4-mm (6-in.) steel washer plates. When needed, crib blocks with a 50.4-mm (2-in.) drill hole are used as supplemental supports. Three-hole steel channel with the ends curved down¬ward to conform to the borer arch may also be used. In the mining sections, a 14 BU loader is used be¬hind the Marietta miner to load two Torkar shuttle cars. These haul to Long-Airdox feeder breakers which load onto a 1.07-m (42-in.) wide conveyor belt. With the exception of the shuttle cars, which use 275-V d-c power, the face equipment operates on 550-V a-c power. The main power line is 7200 V ac. The same loading and hauling equipment is used with the Marietta Borer miner for driving both panel and development entries. In the retreat sections, the same equipment with the exception of the miner is utilized. Because of the advantage of a variable cutting hight, Jeffrey Heliminer ripper-type continuous miners were employed in these retreat sections. For longwall mining the first unit uses an Eickoff EDW 300-L double-drum ranging shearer. Drum di¬ameter is 1.69 m (66 in.) with a drum rotation speed of 45 rpm. The drum is equipped with 48 Carmet 34-L-1 bits. The 300-kW (400-hp) motor operates on 1000 V. Tramming of the shearer head is done through a hydraulically driven 26-mm wide chain. Maximum cutting speed is 7.93 m/min (26 fpm), and maximum tram speed is 12.8 m/min (42 fpm) with an average cutting speed of 3.66 m/min (12 fpm). Web cutting depth is 762 mm (30 in.) and the cutting height is 2.13 m (7 ft). The conveyor width is 686 mm (27 in.). It uses a single chain, 26 mm thick, with flights on 0.91-m (3-ft) spacing. The face conveyor speed is 64.3 m/min (211 fpm) when driven by two 112-kW (150-hp) 1000-V Louis Allis motors through torque convertors. The roof support used is Gullick Dobson chocks [6-legged 77.3 t (85 st) per leg, 463.6 t (510 st) total capacity] with double-acting leg jacks. The hydraulic pressure for the legs is 13.8 MPa (2000 psi), and the hydraulic pressure for rams is 6.8 MPa (1000 psi). The yield valves are set at 57.8 MPa (8500 psi). Each chock weighs 4.09 t (4.5 st). The stage loader is an Eickoff drive, with a 68-kW (90-hp), (00-V Reliance motor. A 26-mm thick single strand chain with flights on 0.6-m (2-ft) centers runs at 75.89 m/min (249 fpm). The stage loader is self-advancing and is connected to the belt tail piece with a special fifth wheel attachment so that the stage loader can advance and guide the tail piece. The hydraulic pressure for the roof support is supplied by two 45-kW (60-hp) Reliance motors op¬erating on 699 V, with 1.26 L/s (20 gpm) constant volume Gullick hydraulic pumps. Each pump services alternate chocks, with the two pressure lines connected at the tailgate by 25.4-mm (1-in.) hydraulic hoses, rated at 27.21 MPa (4000 psi). The 6.8 MPa (1000 psi) pressure rams advance is supplied by a 45-kW (60-hp), 600-V Reliance motor and a 1.26-L/s (20-gpm) con¬stant volume Gullick pump. Return lines are 35.5-mm (1.4-in.) diam hydraulic hoses with rated capacity of 27.21 MPa (4000 psi). The fluid used in the hydraulic system consists of 95% water and 5% Hy-chock fire resistant water soluble oil. Water is stored and mixed in a 6096-L (2000-gal) track-mounted tank and goes to a 655-L (180-gal) reservoir at the pump. The belt conveyor structure is 1.07 m (42 in.) wide with 0.47 rad (27°) idlers, bottom stands located on 3.05-m (10-ft) centers, top stands on 1.5-m (5-ft) centers, and is supported on a 19-mm (3/4-in.) diam

Jan 1, 1982

By S M. Zhang, Y X. Han, J Liu, Y J. Li

"Compared with conventional roasting and magnetic separation processes, higher reduction temperatures and stronger reducing atmospheres are used in reduction technologies to reduce iron oxides to metallic iron, which is often liberated during reduction and amenable to magnetic separation. Reduction and magnetic separation of an Australian refractory iron ore was therefore studied as a function of reduction temperature, time and coal addition. It was found that during reduction, the particle size of the reduced iron increased with reduction temperature and time. Subsequent magnetic separation tests were further conducted on the reduction products. A reduced iron powder with a Fe grade of 85.67 per cent was obtained at a Fe recovery of 87.82 per cent under optimum reduction conditions where a metallisation rate of 85.82 per cent was achieved. The mechanism involved in reduction is discussed based on XRD and SEM examinations.CITATION:Han, Y X, Liu, J, Li, Y J and Zhang, S M, 2013. Upgrading an Australian iron ore by reduction and efficient magnetic separation, in Proceedings Iron Ore 2013, pp 397-402 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Aug 12, 2013

By G. Adderley, A. Wetherelt

The problems of poor drilling in underground operations have been highlighted on many occasions. Incorrect collaring and poor angular deviation can all lead to an increased block size, dilution, over-break adding to both cost and time, while in a tunnelling operation sidewall damage and a poor percentage pull from a blasted round has similar consequences. This paper sets out to further investigate the ability of the laser scanner method of obtaining drill hole deviations and collaring errors at the face as described by Wetherelt and Williams (2006). Tests were carried out to compare measured inter-hole burdens from this method in comparison to that obtained using a total station survey. To investigate the ability of these methods to accurately determine drilling errors, a field test was carried out through an analysis of a series of pre-drilled blast holes of known inclination and direction. A time-motion study was also conducted for a varied number of holes to assess the usability of each method. In line with current research into blast vibration being undertaken at the Camborne School of Mines, 6 cut blasts were carried out to establish the effect of poor drilling on a blasts success to see whether an increased burden at the toe of a hole had an influence on vibration.

Jan 1, 2009

By Camila Soares Fonseca

Os aços inoxidáveis duplex possuem uma microestrutura austeno-ferrítica com fração média de cada fase de cerca 50%. A microestrutura duplex é responsável por melhorar as propriedades mecânicas do material. A soldagem destes aços é frequentemente uma operação crítica. O objetivo desse trabalho é analisar a influência do aporte térmico sobre a microestrutura de juntas soldadas do aço duplex SAF2205. As chapas de aço duplex 2205 foram soldadas pelo processo de soldagem GMAW (Gas Metal Arc Welding), variando apenas o aporte térmico. Após essa etapa, as chapas foram cortadas, preparadas para a análise metalográfica, difração de raio X (DRX) e ensaio de microdureza Vickers. As microestruturas do metal base (MB), da zona termicamente afetada (ZTA) e da zona fundida (ZF) foram caracterizadas nesse estudo. Foi verificado que quanto maior o aporte térmico menor a quantidade de ferrita e maior o crescimento de grão da austenita na região fundida. Os ensaios de microdureza mostraram que quanto maior o aporte térmico maior será a ZTA do material.

Aug 8, 2018

By J. F. Archibald

A research effort has been undertaken, under Workplace Safety and Insurance Board of Ontario (WSIB) sponsorship, to characterize the support capabilities of conventional and innovative spray-on lining materials for mitigating dynamic failure effects created by simulated rockbursts. Conventional spray support media, such as shotcrete and fibrecrete linings, and a variety of new spray-on lining materials [thin, spray-on liners (TSLs)], typically polymer-type thin coatings, were subjected to simulated rockburst influence. For baseline comparison, conventional rockbolt and bolt-and-mesh support products were also evaluated. Observation of lining performance was undertaken through completion of field-scale support and rock detonation trials. Controlled blasting tests were conducted at the Queen’s University Explosives Laboratory to examine the effectiveness of the variety of surface coating agents for modifying dynamic rock spalling processes. All tests were performed upon largely undisturbed and homogeneous rock materials in highly repeatable fashion. All blasts were observed seismically and photographically to provide detailed information concerning rock motion, surface fracturing, ejected fragment motion and support liner survivability characteristics. These material comparison trials were performed to assess the relative effectiveness of a wide range of currently available area support techniques for suppressing hazards associated with dynamic rock ejection. The results of this study validate that thin, spray-on linings may be equally as effective, and often better, than conventional support materials at mitigating rockburst or like damage in highly stressed mine environments.

Jan 1, 2004

By Antonio Lourenço Batista de Souza

Neste trabalho, foi desenvolvido um modelo de elementos finitos para simular o primeiro passe da laminação de tiras a quente de aços C-Mn, durante a etapa de acabamento, via DEFORMTM3D. As dimensões e condições de processo foram obtidas a partir de dados industriais descritos na literatura. Utilizando-se o modelo termomecânico, foram feitas análises de sensibilidade por meio de alterações no fator de atrito e coeficiente de transferência de calor por condução, uma vez que não há um conhecimento preciso dos valores destes parâmetros para o processo simulado. Com isto, foram analisadas as alterações que a mudança destes parâmetros podem causar na carga e temperatura de laminação, perfil de deformação, fração recristalizada e tamanho de grão. O aumento do atrito provoca um aumento na carga de laminação, na homogeneidade do perfil de deformação e menor tamanho de grão. O aumento no coeficiente de transferência de calor, diminui a temperatura superficial da tira, o tamanho de grão e gera um aumento na carga de laminação. Assim, neste estudo foi mostrado a eficácia do modelo em predizer o comportamento do aço no processo de laminação a quente.

Aug 16, 2017

By Felipe Calizaya

This study describes the principles involved in the design of a booster fan installation, the basic requirements, and the provisions needed to guarantee a safe operation of the unit. It also describes how a booster fan should be selected and what measures should be taken so that events such as uncontrolled recirculation of air in a mine can be avoided. A computer program called MFSELECT is presented to assist the ventilation engineer in the design process a For a given mine ventilation network, the program determines the duties of fans and regulators that will fulfill the airflow requirements of a mine and minimize the total power consumption. The application of the program is illustrated by the detailed solution of a sample problem. A Case Study based on the Jim Walter Resources' No. 7 Mine is utilized to describe the basic requirements for the design and operation of booster fans, and illustrate their potential benefits in mine ventilation. Finally, the study discusses a process control system designed to evaluate the safety conditions in the mine and report on the status of both surface fans and underground booster fans.

Jan 1, 1990

By P. M. Daling

This report, sponsored by the U.S. Bureau of mines, presents the results of an examination of available safety assessment methods and their applicability to the mining industry. Standard safety analysis procedures were evaluated for relevance to the mining industry and mire operators' needs. The major safety concerns and needs of mine operators were identified through telephone contacts and a visit to a mine site. Safety assessment methods recommended for use by the mining Industry were also investigated for potential adaptability to a user-interactive computer system. Several methods were determined to be Immediately transferrable to the mining industry; these techniques are described in a user's handbook produced as part of this project. The handbook gives detailed instructions on appropriate application and utilization of these methods. Other techniques will require further development before they can be effectively transferred to the mining Industry. These methods, their advantages, disadvantages and resource requirements are described in the final project report. Recommended system safety analysis techniques facilitate an evaluation of total mine system safety and a determination of safety-related interactions among systems within a mine. The application of these methods to the mining industry could result In reductions in both the frequency and the consequences of mine accidents.

Jan 1, 1983

By J. C. Gerdeen

It is of upmost importance to keep the size of the program to a minimum for the computer being used. One way is to reduce the size of the data bank by adjusting the large common array A in the main program (page, line 32) and the statement following the assignment of A. The current value of A is 35001 and is optimized for the UNIVAC computer with 5,000 words of useable storage. Another way to reduce the required storage is to overlay the program. That is, to have in the computer only the subroutines which are being used. In large finite element codes, overlays or segmentation of the program is a necessity. The code presented here has three levels of segments and are presented graphically in the tree in Figure D-1.

Jan 1, 1976

By Walter Frank

The purpose of the work covered by this contract is to develop a ground and pilot conductor connector which will significantly improve mine trailing cable splice performance. This work compliments the work completed under Contract 00357100 on the development of crimp connectors for the phase conductors of trailing cables. A brief discussion on splice problems is presented as an introductory means and substantiates the need for the development and test programs outlined in this report. As with the phase development work, the ground and pilot connector work is essentially of an empirical nature. As a result time is devoted to explaining the derivations of the test procedures used. In order to familiarize the reader with connector terminology a section is devoted to a glossary of pertinent terms. The connector design effort includes connectors for splicing ground and pilot conductors. A distinction is made between connectors suitable for splicing round and flat ground configurations. The same also applies to the pilot conductors. During the course of this work, much time has been devoted to the question of suitability of round cable configurations for trailing cable applications. This is discussed in detail along with resulting conclusions and recommendations. A field test program was conducted in order to confirm ground and pilot connector laboratory data. This testing also determined the viability of the connector designs and recommended installation techniques in a mine environment.

Jan 1, 1979

That electricity is an important and valuable agent, the limit of whose capacity for utilisation is as yet unknown, is a truism; although enormous advances have been made, the possible developments are yet greater. Mining, along with other industries, shares already to a large extent in the practical utilisation of this agent. Electrical engineers have been employed for several years past in developing and applying electricity to all classes of machinery, and for every purpose where power is required and in no instance, that the writer is aware of, has its use been curtailed in favour of any other agency now known, while extensions are continually made. The first modern electric motor was shown at the Vienna Exhibition of 1873, and it was used to work a pump through a mile length of cable but it was not before 1878, that an electric motor was put into daily commercial use by Messrs. Chretian and Felix to work a beet sugar hoist. This early plant effected a saving of 40 % as compared with hand work.In 1879, the first Electric Railway was constructed for the Berlin Exhibition, and in the same year, Mr. T. A. Edison was running an experimental line at Menlo Park. The first electric line opened for the public, was constructed by Siemens and Halske, in 1881, at Lichtenfelde, and was 11 miles in length.

By John F. MacDonald

This paper describes the driving of a 400 m long by 25 m diameter highway tunnel through sensitive, overconsolidated silts and clays. The design concept requires the construction of the permanent support system, consisting of 24 concrete filled perimeter drifts, before excavating the mass core. General construction procedures used during the construction of the perimeter drifts are described.

Jan 1, 1985

Prediction of fragmentation by blasting is most commonly based on theassumption that a single distribution of pre-existing discontinuities is present within a blasted rock volume and that the underlying mechanism of failure is tensile failure. It is assumed that the tensile stress field created by a blast initiates and extends radial cracks around the blastholes. Interaction between radial cracks from different blastholes and the free surface creates rock fragments. In reality, fragmentation of the rock occurs due to two mechanisms.One is related to the compressive-shear failure of the rock (mainly rock matrix) close to the blastholes, while the second mechanism is the previously mentioned tensile failure of the rock mass. The secondmechanism of failure occurs in the form of extension of the larger cracks in the region beyond the crushed rock, and this type of fragmentation occurs after the crushing phase. In the case of hard rock or blasting where the extent of crushing is minimal, the currently used prediction methods give a reasonably good result. However, there are many blasting conditions where the amount of rock crushing is significant. In the case where fragmentation of the given rock volume occurs dueto two different mechanisms, modeling of the rock fragmentation with a single distribution function is not appropriate. To model the fragment size distribution generated by two different mechanisms, it is necessary to use a two-component fragment size distribution function and a scaling parameter that will determine the proportion of the fragmented volume created by each mechanism. The purpose of this paper is to present results of blast fragmentation modeling based on two mechanisms of failure. The two-component model (TCM) utilises experimentally determined parameters from small scale blasting and parameters of the Kuz-Ram model, for more accurate prediction of the complete rock fragment size distribution curve. The methodology is such that it is possible to predict complete fragment size distribution, including fines, of ROM product for a future mine, at the feasibility phase of mine design.

Jan 1, 1999

By Donald E. Raab, Lee W. Saperstein

INTRODUCTION The Premise Not much time has to be spent in a tunneling project to realize the tremendous interdependence that exists among the various unit operations. The research that is to be described attempts to examine one set of these linked operations, namely, the effect of cutters on muck handling properties. All too often research focuses on too minute an aspect of a problem for the results to be of much immediate, practical use. The goal of this research is to avoid this specificity. Apologies are tendered if this goal is missed. It is known that the mode of flow of bulk solids under the influence of gravity -- emptying a bin or hopper -- is dependent on several basic parameters. The size of the particles and the distribution of sizes (consist) determine whether the solids will first flow out of a given aperture; the cohesion of the solids will determine if flow will continue after the bin is opened; and the angle of internal friction determines the size of the flow regime (1,2,3). It is well understood that the latter two parameters affect the behavior of bulk material on a conveyor belt. Cohesion of material to a belt is always an undesired problem. Heaping of material on a belt depends on the material's internal friction. It is also known that varying the particle size, and the packing density of the material can affect the material's angle of internal friction (4,5). The question was posed whether sufficient variation in a mechanical tunneling machine's (mole) cutting function could be arranged to affect materially the handling properties of the muck

Jan 1, 1974

By R H. Sillitoe

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. Geological concepts are employed routinely in porphyry copper exploration at the area selection, prospect appraisal and deposit definition stages. Regional, district-scale, deposit-scale and laboratory-based concepts can all play a role, with the resulting porphyry copper models providing templates for the integration of geochemical and geophysical responses and interpretations. Of the 19 concepts and parameters selected for brief discussion, those relating to subdivision of porphyry intrusions, understanding of diverse and commonly complex alteration-mineralisation patterns and lithocap characterisation are undoubtedly the most influential in exploration. Several of them are of particular importance because of their role in promoting grade development. Additional belt-, district- and deposit-scale features are currently little used but hold promise, whereas others have less practical value but help in underpinning the evolving porphyry copper genetic model.CITATION:Sillitoe, R H, 2015. Recent advances in porphyry copper geology and their exploration utility , in Proceedings PACRIM 2015 Congress, pp 35–44 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By B. Erdem

This article introduces a computer model, which emphasizes panel design for draglines deployed in direct side-casting mode. The study covers flat and inclined coal-seam conditions with the dragline operating along the strike-line. Two casting methods are embedded in the model: material is placed in the open set adjacent to the one on which the dragline is located (the DNS method) and material is placed in the open set adjacent to the one the dragline excavates (the DND method). While in the first method the unproductive walking time is reduced as the set is enlarged, in the second method the cycle time is lessened as the dragline swings along an acute angle. The side-casting model is shaped so as to provide several panel configurations instead of a single one to conform to the characteristics for an opencast mine. This is achieved by assigning the pit width an interval of values ranging from a minimum to a maximum. The model includes solutions to other critical design parameters such as: set length, swing angles at key cut and main cut positions, and required dragline reach at key cut and main cut positions. The model outcomes reveal the following: the stripping operation is largely affected when the coal-seam is inclined. The downhill spoiling mode should be applied wherever possible. The DNS pattern allows wider pit width intervals as well as longer sets and longer reaches.

Jan 1, 2004