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By R. B. Anderson

As a part of a Bureau of Mines program on the conversion of coal to liquid and gaseous fuels, the kinetics of the Fischer-Tropsch synthesis on iron catalysts was investigated. Nitrided catalysts were studied at 21.4 atmospheres using feed gases ranging from 2H2+1CO to 0.25H2+1CO, and at 7.8 to 21.4 atmospheres using 1H2+1CO gas. The overall apparent activation energy increased with increasing carbon monoxide content of the feed gas from 19.5 kcal/mole for 2H2+1CO to 23.7 for 0.7H2+1CO. The rate of synthesis with 1H2+1CO feed increased linearly with operating pressure from 7.8 to 21.4 atmospheres. A semifundamental rate equation of the form [r=a'PH2 0.6pco 0.4- ?r0.5PH200.5] was shown to fit the data satisfactorily. The relative usage of hydrogen to carbon monoxide decreased with increasing conversion, passed a minimum, and then increased. Apparently water is the principal primary product and carbon dioxide is produced by subsequent water-gas shift reaction. In another group of experiments water, carbon dioxide, argon, and methane were added to 1H2+1CO feed to a nitrided iron catalyst. Argon and methane were only diluents. Carbon dioxide was a slight inhibitor and water vapor was a strong inhibitor to the rate of synthesis. Water accelerated the water-gas shift and usually increased methane production. Important factors of catalyst geometry in determining rate were found to be particle size and extent of reduction. Apparently the active portion of the catalyst is confined to a layer of 0.01 to 0.02 cm from the external surface of particles. A special solution was made of differential equations relating re¬action at the catalyst surface to diffusion in pores of the catalyst for spherical particles with an active layer surrounding an inert core. The equation derived expressed the experimental data reasonably well. Diffusivities determined by this equation varied from 3 to 24X10-6 cm2/sec; these values are of the same magnitude as observed for the diffusion of gases in liquid. The appendix presents a group of synthesis experiments at pressures of 21.4 to 103 atmospheres. In this range the rate of synthesis on nitrided iron catalysts increased with pressure to the 0.74 to 0.84 power.

Jan 1, 1964

By F. Baerhold, S. Mitterecker

"Acid recovery is an important process step in pickling and leaching processes. Pyrohydrolysis in spray roasters and fluidized beds is a well-established technology to recover hydrochloric acid, but the hydrolysis in the liquid phase is more and more considered as an alternative new process route. The characteristic features of these three processes will be discussed along with select industry trends. For example, hydrochloric acid based leaching processes for ilmenite and nickel laterite ores produce far larger volumes of waste acid streams than the steel industry. Scaling-up of the spray roasting technology to the highest possible capacity is necessary in order to become more cost effective. In addition, process routes based on ferric chloride pyrohydrolysis allow the production of super-azeotropic HCl (up to 30 wt.%). Finally, the new ECOmode concept optimizes the operation of both pickling lines and the spray roast acid recovery plant. Long term results from industrial applications show that fuel savings in the spray roaster of 25% are easily achieved without additional process steps. An attractive feature is also to upgrade existing installations to higher plant capacities.INTRODUCTION Pyrohydrolysis is the conversion of metal halides into metal oxides at elevated temperatures in the presence of water vapor. Reaction products are a hydrogen halide gas and a metal oxide. The process was originally developed and introduced into the market by ANDRITZ Ruthner in the 1960’s. It has found wide technical application in the field of acid recovery and oxide production. The introduction of this process has revolutionized pickling technology. The use of hydrochloric acid has strongly increased over the last decades, since pyrohydrolysis is available for the total recovery of waste pickle liquors from mild steel pickling, as illustrated in Figure 1.Chloride media are also used increasingly in the metallurgical industry. Oxide production processes for Mg, Co, Ni, and Al have been installed. Processes using hydrochloric acid are known for the leaching of ores as well as for the re-extraction of metals from organic solvents. In both cases the recycling of HCl is of substantial importance to achieve highly economic processes and to reduce waste."

Jan 1, 2016

By Roozbeh Afzali

Designing an appropriate tunnel blasting pattern reduces the cost of rock breakage and subsequent operations such as loading, haulage and support of tunnels. For that reason, we developed a software that designs tunnel blasting patterns. Using this software, three different tunnel blast design models based on parallel hole cut, namely, Swedish, Konya and Energy Balance, were investigated and compared for a virtual tunnel. Accordingly, blasting patterns were designed, for different cross sections. The trend of parameters such as specific charge and specific drilling as functions of tunnel cross-section was determined for each model. Our results show that, in general, The Swedish model gives lower specific charge and specific drilling compared to other models.

Jan 1, 2008

By Juliana Colacioppo

La voladura es un procedimiento que está diseñado para obtener pilas a través de una fragmentación adecuada para el carguío y transporte, aumentando así la productividad de las palas y camiones y asegurando la estabilidad de los taludes y condiciones del piso de los bancos en la mina. Asimismo, el diseño de la voladura debe optimizar las operaciones posteriores tales como chancado, zarandas, clasificación, molienda, flotación y lixiviación. En ese sentido, el presente texto describe la metodología de optimización de procesos de la miina Antamina hasta la oncentradora a través de la recolección de datos de campo, caracterización, modelación matemática y simulación de las operaciones, donde además se utilizan sistemas de rastreamento de minerales y de análisis continuo de imágenes.

Sep 9, 2007

By B. P. Nutter, M. Lebourg, J. A. McAlister

The Multi-Flow Evaluator (MFE) is a new system of tools providing an original approach in drill-stem testing. It improves control during the testing operation and gives a more accurate evaluation of the fluid recovered while providing additional pressure information for reservoir analysis. The tools are operated entirely by up and down motion of the drill stem. This up and down operation provides a positive means of control and offers easily observed surface indications of tool operating position. An unlimited number of shut-in and flow periods may be taken with this tool while in the hole. The key to the success of the MFE system of tools is the safety seal packer. Until this development, operation of the tools by up and down motion only had proven unreliable. The success ratio of the MFE now exceeds the ratio achieved by conventional tools. The equipment includes a 2,750-cc chamber in which a representative sample of the flowing formation fluid is trapped at the end of the last flow period and brought to the surface under pressure. The sample can be evaluated at the wellsite or transferred under pressure for laboratory analysis. The sampling feature of the tool allows a sample to be obtained from a reservoir which has suffered minimum influence from production. A representative sample can be obtained for laboratory or empirical analysis by employing a testing technique to minimize drawdown during flow periods. Interpretation methods to take advantage of this additional information are presented and supported by actual field examples. INTRODUCTlON The Multi-Flow Evaluator (MFE) represents the most recent technical advance in formation testing. Drill-stem testing techniques prior to the MFE have been limited in three areas—(1) operaiion, (2) recovery analysis, and (3) reservoir analysis. The tools in the MFE system allow maximum control during testing and provide test information superior to that obtainable with conventional tools. The system of tools and the operation procedures are described to illustrate the mechanical advantages over present tools. Methods of pressure build-up and fluid

Jan 1, 1966

The Ross Goldfield is a unique gold resource. From its discovery in 1865 there was a growing understanding of the geology of the deposit as mining moved from the surface at Jones Flat in the south to underground extraction at Ross Flat further north. Multiple layers of gold-bearing gravels were penetrated in deep shafts and six rich gold-bearing bottoms were worked. The ultimate bottom on the Old Man Group gravel surface has never been reached either at Jones Flat or at Ross Flat. Mining at Ross Flat was ultimately halted by uncontrollable flooding, by escalating costs and by a shortage of labour during the First World War. Mining at Jones Flat recommenced in the late 1980s and attention has once again been drawn to the potential magnitude of the resource at Ross Flat. Research of historic literature shows that there is potential for a bulk gold grade of 400 mg/m3 in a volume of 82 million cubic metres of gravels which lie in a basin between Jones Flat and north Ross.

Jan 1, 1993

By Kentaro Kuraji

INTRODUCTION Background Tokyo, the capital of Japan, is one of the largest cities of the world. As many as 33 million people live in Tokyo and three neighboring prefectures although the city occupies only 0.6% of the area of the country. The network of Metropolitan Expressways serves as a major artery for daily physical distribution in Tokyo. The Central Circular Route of the Metropolitan Expressway (MECCR) is located innermost among the three circular routes in Tokyo (Figure 1). Once completed, MECCR is expected to drastically mitigate traffic congestion in the central Tokyo area and provide drivers wider varieties of options in route selection. Alleviating congestion is estimated to reduce 400,000 tons of carbon dioxide emissions annually. The eastern and northern sections and part of the western section of MECCR have been completed with the total length reaching 33 km. The western section will be extended southward by approximately 4 km in the spring of 2010. An 11-km-long western section of MECCR that is currently being constructed is composed mostly of tunnels. A large number of power supply, gas supply, telephone, drinking water and sewerage pipes have been placed under the Yamatedori avenue along which MECCR is being constructed. The shield tunneling method has therefore been adopted for constructing tunnels along the route. The method involves the advancement of a shield tunnel boring machine (TBM) through the ground simultaneously assembling segmental rings within the shield. The method not only enables the building of highway tunnels more quickly and efficiently than other methods but also minimizes the effects of construction work on the surrounding area and the environment. Constructing the tunnel using the conventional cut-and-cover method might have required large amounts of time and cost because relocating the numerous objects buried under the Yamatedori avenue might have been necessary during tunnel construction. It was also considered desirable to place the tunnel at a depth of 30 m or more from surface to prevent tunnel construction from adversely affecting railways because the tunnel was expected to cross eleven railways and subways. This is one of the reasons for selecting shield tunneling as an effective method. The shield tunneling method was finally adopted for a length of 7 km in the 11-km western section of MECCR. The shield tunneling method was also employed in the Ohashi Tunnel located southernmost in the 11-km western section of MECCR. This paper describes the construction of the Ohashi Shield Tunnel.

By David C. Oyler, Dennis R. Dolinar, Christopher Mark, Craig S. Compton

Fully-grouted roof bolts comprise more than 80% of the primary roof supports used in U.S. coal mines. However, nearly 1,500 MSHA reportable, non-injury roof falls occur each year, and most of these are attributable to failure of the roof bolt system. Anchorage failure is one failure mechanism for fully-grouted bolts. As roof deformation works its way upward, the bolts can become heavily loaded near their upper ends. If the applied load exceeds the anchorage, the bolts will simply pull out. Research dating back 30 yrs indicates that this type of anchorage failure is most likely when the roof rock is weak, just where roof support is most critical. In soft shale or coal roof, the small amount of data available in the literature indicates that 20-30 in of resin anchorage may be required to achieve the full capacity of the bolt. In other words, the “full resistance zone” of a 60 in bolt may actually be just 30-40 in. Despite its potential importance, there is no widely accepted anchorage test for fully grouted bolts. Standard pull tests have sometimes been employed, but they provide no information on the anchorage near the top of the bolt. An alternative, first described more than 25 yrs ago in the U.S., is the short-encapsulation pull test (SEPT). With this test, the bolt is installed with only a short (1 ft or less) tube of resin. In recent years variations of this test have become international standards. v

By J. Douglas Causey

In 1984, at the request of Bureau of Land Management, the U.S. Bureau of Mines studied the 13,730-acre Burnt Lodge (MT-065-278) Wilderness Study Area (WSA) in order to evaluate the mineral resources. The WSA is located in Phillips and Valley Counties, MT, about 55 miles south of Malta adjoining the Charles M. Russell National Wildlife Refuge. It is underlain by near horizontal sedimentary beds of the Hell Creek Formation, Fox Hills Sandstone, and Bearpaw Shale. No mines or prospects were found in the WSA. Bentonitic beds were examined, but are generally less than 1 ft thick and, except where exposed, are deeply buried. Kaolin-bearing sandstone in the Colgate Member of the Foxhills Sandstone was sampled and may have potential for use in structural clay products. However, remoteness and inaccessibility of the area make it unlikely that the material could be economically extracted even though some of the material is marginally useful.

Jan 1, 1985

By Thomas C. Wilder

Jan 1, 1965

By Stanley Lefond

The health of the industrial minerals industry appears to be improving year by year. In fact, preliminary information suggests that only three commodities (asbestos, bromine, and sulfur-other) failed to show a production gain in 1984 over 1983. In 1983, 10 commodities failed to post gains and six remained constant. The total value of industrial mineral production in 1984 was $16.7 billion, up $1.4 billion from 1983. Cement, clays, gypsum, and crushed stone production increased. Since these are used in housing and construction, this suggests these industries are moving out of their recent doldrums. Barite and clay production also improved in 1984. Local US barite production, however, still has a long way to go to attain the 1981 high of 2.6 Mt (2.9 million st). Barite imports remain high. But they are not as high as the 1982 importation of 2 Mt (2.3 million st). Bentonite production in 1984 increased 20% in volume and 25% in value over the previous year. This suggests improvement in the drilling industry and an increase in demand for iron ore pelletizing grades. All in all, the industrial mineral industry is alive and well. And it shows signs of great improvement in the next few years. Many thanks to those industrial mineral technical committee members for their assistance in assembling this industrial minerals review: M. F. Dibble, H. W. Allen, G. H. Gallup, S. G. Ampian, D. W. L. Spry, H. G. Fleshman, and L. J. Czel.

Jan 5, 1985

Narrow vein mining, with enough challenge and also where the cutting edge technology can be applied, is still a very important and exciting portion of the mining industry. Based on the deposit characteristics, the most flexible mining method, cut and fill, has been being tested mining this high grade Northern Nevada property, Hollister, since the year of 2008. In the last three years, 254,000 tons of ore were hauled out of the portal and 23,000 feet of development were excavated underground with the winning of several safety and environmental awards. Together with mine geology, the essential segments include mining cycle, ground control, ventilation, grade control and backfill method in cut and fill stopes were covered in this paper.

Jan 1, 2012

Rockbursts continue to be a scourge in the mining industry, being responsible for accidents and damage to mining excavations. Although the problem has been present for more than a century, and although much research has been carried out, a solution is still elusive. Determination of, firstly, the demand on the support system imposed in the rockburst and, secondly, the capacity of the support system, cannot be carried out with any confidence, and therefore rockburst support cannot be designed using a conventional design approach. In contrast with the conventional approach, observations of rockburst damage in a mine have revealed a possible alternative approach to rockburst support ? sacrificial support. It was observed, in rockburst events, that a support system consisting of concrete panels restrained by grouted cables was destroyed, and the concrete panels were ejected. However, the rock behind the panels remained apparently undamaged and in place. This behaviour reignited the concept of sacrificial support, conceived more than 20 years ago, and which is described in this paper. The remedial solution implemented on the mine, involving cables wrapped over the panels and retained by grouting into boreholes, has been subjected to rockburst loading and has confirmed the validity of the sacrificial support concept. The concept of sacrificial support may be controversial, but is deliberately presented here with the aim of generating discussions and contributions, and with the ultimate aim of improving safety and reducing rockburst damage in mines. Keywords rockburst, seismicity, seismic waves, rock support, support design.

Jul 1, 2013

By Sie A

After completing a feasibility study in 1987, Niugini Mining Limited (NML) obtained five Gold Mining Leases within PA 632, covering the Mount Victor and Clarkes Ridge gold deposits. Reserves prior to mining were 260 000t at 3.63g/t Au for Mount Victor and 40 000t at 2.8g/t Au for Clarkes Ridge. Ore was mined from the two deposits by NML from November 1987 to January 1990. A total of 645kg of gold and 346kg of silver were produced from 198 000 tonnes of ore pro- cessed. Mining and milling operations ceased in February 1990. Since then, there has been significant progress in rehabilitating and revegetating the site. Plant and workshop facilities have been dismantled and removed from site, exposed parts of the pit have been landscaped and replanted, tailings water has been neutralised and discharged, and the dam filled. All remaining personnel and equipment left the site in November 1990. It is envisaged that the replanted areas will be rapidly revegetated, leaving little evidence of mining. Niugini Mining did not recoup the initial investment for Mount Victor, but the mine produced a positive cash flow and made a significant' financial contribution to the local communities and Papua New Guinea as a Whole.

Jan 1, 1991

By Paul J. O'brien, George N. Neary

"The base metal mine at Buchans, Newfoundland, has been producing since 1928. For many years, mill tailings were disposed of by dumping into a nearby natural watercourse via which they were transported to Red Indian Lake and deposited. In 1965, environmental concerns prompted the establishment of settling ponds near the mine site. Effluents fed to these ponds included slimes washed from mine backfill sand and waste water pumped from underground as well as mill tailings. Dusting around the perimeter of the ponds was first observed around 1970and within a couple ofyears reached serious proportions. A program to alleviate the condition by developing a vegetative cover over the tailings was commenced in 1974. The work was successfully carried out and demonstrated that the large area of waste could sustain a permanent cover of grasses.This paper outlines the experimental work done at Buchans as well as the procedures followed and the problems encountered in carrying out full-scale revegetation."

Jan 1, 1981

By F. S. Young

Previous laboratory and field experimentation has demonstrated the effect of several variables on drilling rate.1-5 These results have been incorporated into optimization theories6-" for the purpose of reducing drilling cost. Minimum-cost drilling theories rely on a combination of historical data and empirical prediction techniques for selecting optimum bit weight and rotary speed. Past solutions have required extensive use of computing equipment because of both the complexity of the mathematical formulations and the large number of parameters involved. Field attempts to apply these computed results have frequently failed because of the uncertainty of input data requirements. Humble has developed an on-site computer system to control bit weight and rotary speed and thereby implement the concepts of minimum-cost drilling. This system can (1) perform short interval drilling rate tests for formation evaluation, (2) use the results of these tests as input information to solve minimum-cost drilling formulas, and (3) control bit weight and rotary speed in accordance with these computed solutions. The purpose of this paper is to describe the computer control system and present results of initial field testing. Theoretical Considerations Minimum-cost drilling (MCD) requires a quantitative evaluation of the variables involved. Several forms of a basic mathematical model have been sug-gested.'?,' "he relationships reported here form the basis for the solution programmed in this work. Four equations are used, expressing drilling rate, bit bearing wear, bit tooth wear, and cost. A solution for minimum-cost drilling assuming constant bit weight and rotary speed over the entire bit life has been programmed for use in computing MCD schedules. This solution is subject to certain limiting assumptions such as those discussed by Graham et a1.I2: 1. Drilling cost is the summation of bit cost, rotating cost, connection cost and hoisting cost. 2. Diamond bits are excluded. 3. Bit life is limited by either bearing failure or tooth wear, or by a combination of operational factors that make it cheaper to pull an incompletely consumed bit. 4. Circulating hydraulics are adequate and do not limit drilling rate. 5. Bit weight considerations exclude hole deviation. 6. Drilling rate is a function of only bit weight, rotary speed, and degree of tooth dullness; that is, the effects of pressure, lithology, fluid properties hydraulics and drill string dynamics are ignored. In equation form, the four MCD expressions are as follows:

Jan 1, 1970

CrusherMapperTM - A New Tool for Crusher Condition Monitoring The efficiency of mineral processing plants is entirely dependent on the availability of the plant equipment critical path. This typically concerns comminution circuits consisting of primary crushers and primary grinding mills. Crusher and mill availability are determined by the frequency and duration of maintenance shutdowns, which are primarily driven by the need for replacement of sacrificial liners. The change of crusher and mill liner shape over respective life cycles also strongly affects plant efficiency. It is therefore highly desirable to accurately monitor the condition of mills and crushers in order to optimise plant efficiency through maximised availability and targeted liner shape. Traditional comminution liner and other production parameter condition monitoring has been attempted through ineffective methods which prevent satisfactory management of crusher and mill maintenance and reduce operational control. This situation has recently dramatically improved for grinding mills through the development and commercial deployment of MillMapperÖ, a technology based on three-dimensional laser scanning data and dedicated software (Franke, Lichti and Stewart, 2006; Franke and Lichti, 2007; 2008; Franke, 2008; Franke, Redman, Johnson and Michalek, 2009). At the time of writing in late 2009, MillMapperÖ is in successful use at a rate of approximately 170 surveys and associated reports per annum at processing plants in Australasia, South America, North America and Africa. This paper presents the introduction of CrusherMapperÖ, the MillMapperÖ equivalent for crushers. CrusherMapperÖ also utilises high density, high accuracy laser scanning data, and features dedicated algorithms for three-dimensional monitoring of wear on crusher concaves and mantles. This facilitates the detection of wear hot spots or zones or localised liner failures, and accurate forecasting of reline tonnages and dates. It also for the first time provides the means to establish optimised concave and mantle shapes in order to achieve production targets. In contrast to any other currently utilised condition monitoring method, CrusherMapperÖ data collection can occur remotely and does not require any shutdown or crusher isolation. This method therefore does not interrupt crusher production while at the same time eliminating any confined space risk normally associated with taking measurements inside the crusher cavity.

Sep 13, 2010

By Dennis Dolinar

Improperly designed support systems have led to unplanned roof falls in longwall tailgates. To prevent such falls requires appropriate support and adequately designed support systems. The National Institute for Occupational Safety and Health (NIOSH) is currently developing a standing support design methodology based on the ground reaction curve. To design the support system with this methodology requires a quantitative assessment of the ground and support interaction, which is developed from instrumented test sites installed in longwall tailgates. To compliment and expand the field data results and to optimize the support design, numerical modeling is also conducted. The models are calibrated from the test site data. In the present study, data was obtained from a longwall tailgate in the Illinois basin. The test site was located in the Herrin No. 6 coal seam at a depth of 500 ft. At the site, the immediate roof was a weak shale overlain by a competent limestone. Above the limestone, sandstone formed the intermediate roof. The floor consisted of a weak underclay. In the study, the performance and interaction with the surrounding rock mass of two different standing supports, an engineered crib and a conventional 4-point wood crib were monitored. The two cribs have similar load-displacement characteristics at least through 6 in of displacement. Beyond 6 in, the engineered crib becomes unstable and begins to strain soften and shed load. The measurements and developed ground reaction curves indicated that the tailgate at the mine was a low convergence environment until well inby the face. Because the convergence was well below their capacity, both supports provide more than adequate levels of support to the tailgate. Significant convergence and support loading did not occur until 90 to 100 ft inby the face. The results of this study were also compared with results from similar studies conducted in the Pittsburgh seam to further advance the development of the ?ground reaction curve? based support system design.

Jan 1, 2009

Pillars in underground limestone mines are required to support the overburden and provide a safe, stable working environment for mining personnel and equipment. Pillar stability was assessed by the National Institute for Occupational Safety and Health in underground limestone mines in the Eastern and Midwestern United States. It was found that current mine layouts have been successful in providing support to the overburden, while a small number of isolated pillar failures were observed. The stable pillar layouts and failed pillars were plotted on a chart that demonstrates the relationship between pillar width-to-height ratio and pillar stress. The results show that pillar failures occurred at the lower range of width-to-height ratios and can occur at relatively low stress levels. Zones indicating the potential risk of pillar instability are shown on the chart, based on hazards associated with the onset of rib spalling, large angular discontinuities and unconfined pillars. The chart can assist limestone mine planners to evaluate potential instability in current or new pillar layouts that are similar to those included on the stability chart.

Jan 1, 2008

By R. Loots

High temperature creep resistant alloys are used in power plants as steam pipes and headers where they have to withstand super heated steam temperatures of up to 600ºC for the lifetime of the plant. Frequently certain portions of pipes may be subjected to excessive temperatures and consequently suffer from creep damage. Damaged portions then have to be replaced by welding new steel to the service exposed steel. Such welds will have to be subjected to a post weld heat treatment (PWHT) to relieve residual stress. During such a PWHT it is then possible for the steel to crack in the heat affected zone (HAZ) by what is known as Reheat Cracking (RHC), also known as stress relief cracking. For the present purpose the RHC susceptibility of service exposed steel was compared with that of new steel by reproducing the temperature cycle to which the HAZ will be subjected during welding in a weld simulator. This was followed by PWHT where the influence of residual stress was studied by applying a constant stress to such a weld-simulated specimen in a creep testing machine. A sharp spiral was machined along the length of the cylindrical specimen to simulate the presence of a possible weld defect. In this way it is possible to subject all portions of the simulated HAZ to the same stress intensity. For the purpose of the investigation 1/2Cr-1/2Mo-1/4V steel known to be susceptible to RHC was chosen. During testing a constant tensile stress was maintained while the temperature was increased to follow the thermal cycle prescribed for PWHT. It was found that the new steel in all instances failed by RHC along the grain boundaries in the high temperature heat affected zone. In the case of the service exposed steel the creep strength of parent metal was lower than that of new steel as a result of exposure to a steam temperature of 540°C for about 15 years. Consequently, all of the failures during testing occurred in a ductile manner by necking outside the HAZ rather than by RHC. When service exposed steel is welded to new material, it is expected that the creep strain associated with stress relaxation will mostly be concentrated in the parent steel outside the HAZ. Accordingly, it should probably not be necessary to make any extra allowance for reheat cracking during repair of steam pipes manufactured from 1/2Cr-1/2Mo-1/4V steel other than that applicable for welding new steel.

Jan 1, 2004