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Jan 1, 1913

Jan 1, 1918

By J. F. Enrietto

It is shown that one cannot approximate the broad internal friction Peaks observed in the Fe-Mn-N system as the sum of a number of subsidiary Peaks and still vetain the concept of a single relaxation time. Furthermore, the approximation of manganese affectirg only nearest neighbor interstitial sites is shown to be incompatible with the experimental results. The results of this experiment show that the influence of manganese in the ferrite lattice extends over 3 to 4 atomic distances. NUMEROUS experiments have shown that the presence of relatively small amounts of substitutional alloying elements can profoundly affect the profile of the damping spectrum in the vicinity of the Snoek peak in bee metals.1-7 Dijkstra and Sladekl observed this phenomenon in Fe-Mn-N alloys where it appeared as a broadening and shifting of the Snoek peak to higher temperatures. The authors interpreted this as being the result of a superposition of two damping maxima—the normal Snoek peak, located at its usual position and having a normal half-width, and an interaction peak having a normal half-width, but occurring at a somewhat higher temperature. They prclposed that the interaction peak was caused by nitrogen atoms jumping from Fe-Mn sites to Fe-Fe sites, the Fe-Mn interstitial sites having one manganese and one iron atom as nearest neighbors, and the Fe-Fe sites having two iron atoms as nearest neighbors. They further suggested that the Fe-Mn sites had a lower energy than the normal Fe-Fe sites because of the presence of the adjacent manganese atoms. From the relative heights of the two postulated damping peaks, the dif-

Jan 1, 1962

By K. C. McCutcheon, John Chipman

To a very large extent the quality of a mild steel ingot is determined by the manner in which it is poured into the ingot mold and its behavior during the first few minutes after it has been poured. The formation of the solid rim of metal around the periphery of the mold, the effervescent action of the liquid steel within this solidified rim and the upward or dounward movement of the metal level are indications of what is going on beneath the surface of the liquid metal, especially at the interface at which solidification occurs. Observation of this "rimming action" enables an experienced observer to predict the thickness of the sound skin of the ingot and the location and extent of cavities or "blowholes" caused by entrapment of gases. The control of rimming action at the hands of a skillful steel pourer is a vital part of the control of quality in mild steel sheet, and the behavior of the liquid steel during the rimming period is a useful criterion of the melting and refining practices employed in the open hearth. The behavior of rimming steel during solidification depends primarily upon the evolution of gases from the molten metal and it is therefore a problem of considerable interest to determine the nature and amount of the gases evolved. A number of attempts to determine either the amount or the analysis of the evolved gases have been reported. Klinger,' in 1925, analyzed the gases extracted from rimming, sernikilled and killed steel ingots and found that the chief constituent of the rimmining gas was carbon monoxide. The relatively large amounts of nitrogen reported in his first few samples may have been due to incomplete removal of air from the gas-collecting mechanism. McKune² reported similar studies of rimming ingots from which surprisingly large proportions of nitrogen and hydrogen were evolved. Ameen and Willners3 determined both the amount and composition of gases extracted from unkilled ingots with the aid of diminished pressure. Their ingots weighed only 80 kg. (176 lb.) and apparently were solid within two or three minutes, so that their

Jan 1, 1938

By John Chipman, K. C. McCutcheon

To a very large extent the quality of a mild steel ingot is determined by the manner in which it is poured into the ingot mold and its behavior during the first few minutes after it has been poured. The formation of the solid rim of metal around the periphery of the mold, the effervescent action of the liquid steel within this solidified rim and the upward or dounward movement of the metal level are indications of what is going on beneath the surface of the liquid metal, especially at the interface at which solidification occurs. Observation of this "rimming action" enables an experienced observer to predict the thickness of the sound skin of the ingot and the location and extent of cavities or "blowholes" caused by entrapment of gases. The control of rimming action at the hands of a skillful steel pourer is a vital part of the control of quality in mild steel sheet, and the behavior of the liquid steel during the rimming period is a useful criterion of the melting and refining practices employed in the open hearth. The behavior of rimming steel during solidification depends primarily upon the evolution of gases from the molten metal and it is therefore a problem of considerable interest to determine the nature and amount of the gases evolved. A number of attempts to determine either the amount or the analysis of the evolved gases have been reported. Klinger,' in 1925, analyzed the gases extracted from rimming, sernikilled and killed steel ingots and found that the chief constituent of the rimmining gas was carbon monoxide. The relatively large amounts of nitrogen reported in his first few samples may have been due to incomplete removal of air from the gas-collecting mechanism. McKune² reported similar studies of rimming ingots from which surprisingly large proportions of nitrogen and hydrogen were evolved. Ameen and Willners3 determined both the amount and composition of gases extracted from unkilled ingots with the aid of diminished pressure. Their ingots weighed only 80 kg. (176 lb.) and apparently were solid within two or three minutes, so that their

Jan 1, 1938

By W. J. Reagan

The problem of increasing output and decreasing percentage of rejections is a vital one in the manufacture of steel of any kind. The making of basic open-hearth steel for use in rolled steel wheels, tires for locomotive and car wheels, and forgings of various kinds, is a highly specialized industry. The increase in weight of railway rolling stock of all kinds, and the high speed that the public demands of transportation facilities, make it imperative that steel for use in this class of material shall be of the highest quality obtainable. During the past few years it has been the writer's good fortune to have opportunity to study the defects in basic open-hearth steel, and also the results obtained by various changes in open-hearth practice, on such defects. The opportunity for study and the details available are rarely found in the manufacture of steel of any kind. It seems reasonable to assume that many of the results determined here can be applied to the production of steel of other types with a decided improvement in quality, a decrease in the percentage of rejections and an increase in output. All of the material described is basic open-hearth steel with the following range of analysis: carbon, 0.50 to 0.85 per cent.; phosphorus and sulfur, 0.04 max.; silicon, 0.15 to 0.35; manganese, 0.50 to 0.75. All of this material is bottom-cast, therefore tapping and pouring temperatures are higher than in top-cast production. Ingots used in this type of work are of standard length with varying diameters. The ingot is known as duodecagonal, or twelve-sided. The standard body length is 88 in. and the hot top length, 14 in. Fig. 1 shows a typical ingot of this type. The diameter varies from 13 to 30 in. Ingots for special forgings, which cannot be made in molds of standard size, are made in molds of various sizes and designs. These ingots are sliced cold, into blocks of various lengths, depending upon the weight of the forging to be made. Fig. 2 shows the ingot after coming from the slicing lathe. Jusf sufficient metal remains between blocks to hold the ingot in one piece while it is being removed from the lathe; then steel wedges are inserted between the blocks and the blocks are broken apart. This affords abundant opportunity for inspection, as the interior of the

Jan 1, 1930

By J. S. Hardison

In the fall of 1912, the appearance of water in the oil of the Nevada Petroleum Co., Coalinga, Cal., made necessary the installation of a dehydrating plant to reduce the water below the 3 per cent. limit prescribed by the agency. Unlike the mining industry, technical literature of the oil industry is limited and extremely unsatisfactory. Until the recent efforts of the Petroleum and Gas Committee of the American Institute of Mining Engineers, no concerted movement has been instituted to secure publication of papers dealing with the problems of the oil business, and because of the additional fact that dehydrating of oil has not been practiced in California to any extent up to comparatively recent time, it was found necessary to experiment in order to determine the most satisfactory plan for this purpose. It is easy to write regarding successful enterprises, but, while not SO pleasant, it is equally desirable to write of failures so that others may be saved the loss incident to such investments. The water in this oil occurs both free and as an emulsion. Free water easily settles out, but the emulsion requires treatment. So far as can be determined, the emulsion consists of globules of water surrounded and enveloped by a film of oil. Starting with this hypothesis the theory has been evolved that in order to break up this emulsion the water must be heated at least to boiling point, when an explosion takes place destroying the globule. Unfortunately, this condition was not recognized in the installation of the first plant, which was planned on the following lines: A heater was arranged so that all the oil from the wells could be heated before reaching the shipping tanks, these being already fitted with coils for steam heating. Tanks were filled with about 4 ft. of water, live steam was turned into the coils and the water brought nearly to the boiling point, when the heated oil was run in. This for a time was partly successful, but the length of time necessary to apply this heat was too great for practical operating.

Jan 1, 1915

By Clare B. Carlson, C. P. Proctor

THE history of the froth flotation of coal is relatively short. The flotation process was applied to fine-coal cleaning about the time of the end of World War I. Coal flotation finds more widespread use in Europe than in the United States and is used on a commercial scale in nearly all major coal-producing countries in the world. In Fig. 1 the general arrangement of a froth-flotation plant at a British colliery is shown. GENERAL FEATURES OF THE PROCESS Application The flotation process is applicable to the cleaning of the fine sizes, particularly minus 48-mesh coal, although it may be used on sizes as coarse as % in. Since most cleaning processes usually are not effective on sizes smaller than 48-mesh and certainly not effective on minus 100¬mesh, the froth-flotation process offers a solution to the problem of cleaning these sizes. This process consists of agitating the feed pulp with a small amount of reagents, usually 1.0 to 91/ lb. per ton of feed in the presence of water and air; the reagents help to form small bubbles that collect on the particles having a greater affinity for the bubble surface than for the water surface, in this case the fine coal, and the bubbles buoy these particles to the surface to form the froth or concentrate. Slate and shale particles remain wetted and suspended in the water to be drawn off as tailing. The frothers commonly used in coal flotation are cresylic acid, pine oil, and alcohols. More expensive frothers, used in ore flotation, work satisfactorily but do not compare with these cheaper reagents for over-all cost. Collectors used are low-grade petroleum fractions, such as cheap, partly refined fuel oil, or lubricating oil, sometimes mixed with small amounts of kerosene. Other collectors used are the fatty acids and their

Jan 1, 1943

By George F. Comstock

It seems a common opinion among metallographists that all light-gray inclusions seen with the microscope in polished sections of steel are manganese sulphide. Examples of this belief are continually appearing, as for instance in the paper by Dr. Henry Fay on manganese sulphide as a source of danger in steel railsll and in Lieutenant-Commander Cook's paper on Metallography of Steel for U. S. Naval Ordnance.² Slate- colored inclusions are considered to be silicates, and dove-gray inclusions, manganese sulphide. To show the danger in the latter unqualified assumption, it should be sufficient to examine the edge of any piece of steel that is covered with a fairly thick scale, and that has been polished in such a way as to prevent the scale from breaking away entirely below the polished surface. This can be done by protecting the scaled edge of

Jan 1, 1917

By W. V. Vietti

Actual field data from several wells are used to illustrate the application of the method of determining the average density of the fluid column in a flowing oil well in the Yates field, Pecos County, Texas. The depths of the wells in the Yates field vary from 900 to 1700 ft., depending on surface elevation and the location of the wells on structure. Production rates over 8000 bbl. per hour have been determined and many wells are capable of producing 2000 bbl. hourly. Gas-oil ratios are usually below 400 cu. ft. per barrel, but a few wells high on structure have produced over 3000 cu. ft. per barrel. Incidentally, tubing wells has cut these high ratios to as low as 400 cu. ft.' By common agreement of the operators all wells making ratios over 400 cu. ft. will be tubed. Daily production of the field is prorated to 110,000 bbl. from an open-flow potential of 5,000,000 bbl. based on hourly open-flow gages.

Jan 1, 1931

By Henry M. Howe

Preparing the Metals.—Ordinary silver wire was hard-drawn to No. 23 Brown and Sharp gauge; it was cut in two, and one piece was carefully annealed in the ordinary way. The wire was prepared by Messrs. Guild and Delano, jewelers, Boston. The Ansonia Brass and Copper Co., Ansonia, Conn., kindly furnished me with some hard-drawn copper wire No. 31, A. W. G., and some of the same wire after annealing in the ordinary way.

Jan 1, 1885

By R. Hogg, D. W. Fuerstenau

Based on a simplified model for the motion of particulate materials in horizontal, rotating cylinders, approximate expressions have been derived relating the power consumption of a tumbling mill to its dimensions and speed of rotation and to the amount and nature of the charge. The basic assumptions underlying the model are that there is no slip between the charge and the mill shell, that the free surface of the charge remains essentially planar during rotation, and that there is no transfer of momentum from the material flowing down the free surface back to the mill shell. The relationships formulated from this model have been shown to be in excellent agreement with experimental results taken from the literature. While the requirements of the model are most clearly met at low speeds of rotation, the results appear to be valid even for speeds up to 70-80% of the critical speed. Poorer agreement between theory and experiment is found when the charge consists of relatively large particles (balls) and when the filling of the mill exceeds 50% of the available volume. Qualitative explanations of these deviations have been offered.

Jan 1, 1973

By R. G. Mallon, R. L. Braun

Primary copper-sulfide ore deposits that are well below the water table can be chemically mined by in-situ high-pressure leaching. The leaching is accomplished by pumping oxygen gas into the bottom of a large underground flooded column of rubblized ore. Part of the injected oxygen dissolves under the high hydrostatic pressure in the column. With a high enough injection rate, the excess oxygen gas induces a convective circulation of oxygenated solution within the rubble column, which transports dissolved oxygen throughout the system. At elevated temperatures, this results in a relatively rapid dissolution of the copper minerals. A method of calculating the leaching rates is presented. The calculation is done by means of a finite-difference computer code that couples all of the processes relating to copper leaching, convective circulation, and the injection, dissolution, and consumption of oxygen. Results of the calculation for a typical case show that over 70% of the copper can be extracted in less than six years with good efficiency in oxygen utilization.

Jan 1, 1976

By P. de Vitry, Oliver Bowles

THE conventional wire saw, introduced in the slate district of Pennsylvania by the Bureau of Mines in 1927, and used thereafter with remarkable success, consists of a three-strand steel cable having a diameter of 3/16 or 1/4 in. The equipment, methods used, and accomplishments have been described in previous reports.1-7 Similar wire saws have been used extensively in Europe, in quarries and in shaping blocks and slabs. During the period 1932-1937, competition in European marble quarries was keen, prices were declining, and orders were not abundant. When orders were received, there was generally an urgent call for speedy delivery. Because of the diversification of demands, it was impossible to maintain in stock marbles of the color, size, and thickness called for, therefore the material generally had to be fabricated after the order was received. The wire saw enabled stone mills to fill such orders promptly. With skillful handling, 7/8-in. slabs could be cut accurately and rapidly with this equipment. It was found, however, that, as the three-strand wire became worn, the cutting speed dropped rapidly and the cut became less accurate. To overcome these difficulties the junior author of this paper, after much experimentation, developed a new type of wire consisting of a single ribbon-shaped strand, twisted as shown in Fig. 1. To ensure straight sawing, the twist of the ribbon is reversed from right to left and vice versa every 25 ft. The reversal of the [ ] twist is shown in Fig. 2. This is not a new feature; it has been used to advantage with the three-strand wire both in Europe and America for several years. However, some difficulties arise in splicing the three-strand reverse-twist wire, and the claim is made that cutting is less accurate as the wire becomes worn. To make an endless belt, the ends of the single-strand wire are brazed together with

Jan 1, 1941

By G. E. McElroy

THOUGH air jets occupy only a minor role in the ventilation of mines and other enclosures, their importance in this respect is greater than is commonly recognized and may be considerably increased as our conception of their modes of operation becomes clearer. The object of this paper is to discuss some of the ways in which air jets are involved in mine ventilation and some possible applications to the ventilation of mines and other enclosed spaces. The mention of air jets ordinarily suggests the use of some form of an air injector, such as the Venturi blower occasionally used for auxiliary ventilation in mines, or the Saccardo (injector) system used to a limited extent for ventilating railroad tunnels. Actually, the term covers such common phenomena as the discharge of air from fan pipes and coal-mine line brattices, and even from smoke stacks and fans. A knowledge of the general mode of air entrainment by jets also helps to explain certain phenomena-some mine-gas explosions, for instance-that otherwise would remain obscure. Jet action has been involved in three recent series of Bureau of Mines experiments in which the author has participated: experiments on fan-pipe discharge, on the dilution of gases discharged from a stack and on the ventilation of a ship's hold. Available data on the discharge of jets into large openings and on the design and performance of air injectors have been analyzed in considerable detail and supplementary experiments are planned for after the war. The results of the fan-pipe experiments have been combined with the analyses of data on jet discharge into large spaces, and have been published.1 The results of the stack-gas dilution tests, which refer to a jet discharged into a crosswind, have also been published2 and a paper on the design and performance of injectors has been prepared for publication.3 These technical experiments and analyses form the background and the documented evidence for the following nontechnical discussion of the use of air jets in mine ventilation. JET CHARACTERISTICS The jets commonly encountered are those in which the air is in turbulent-as opposed to laminar, or streamline-motion. The characteristics under discussion apply only to such turbulent jets and have their basis in theories of the mixing of turbulent air currents. Such jets have basic characteristics that are present, though variable in form, regardless of whether the jet expands freely in space, whether the expansion is limited by adjacent walls, or whether the jet propels air through a flow circuit. They expand at definite angles and entrain the surrounding air, even though the entrained air must be obtained by recirculation of their own air streams. As the quantity of air in motion in the mixed stream increases, the average velocity of the stream decreases. In the case of entrainment in the free

Jan 1, 1945

By William E. Newman

The smelting of galena in the ore hearth has been practiced in many countries for several hundred years with varying success. In the United States, the water-jacketed American hearths and the Jumbb hearths have found some favor in the Missouri lead belt, where large quantities of non-argentiferous galena concentrates are produced. In form and method of operation the ore hearth has changed but little since its earliest conception. A detailed description of the furnace is lot needed here; it has been fully described in most works on the metallurgy of lead and in many technical papers. Since its field is restricted to non-argentiferous galena, containing over 68 per cent. of lead, its use is limited to a few districts; and, on account of certain disadvantages attending its operation, modern sintering and blast-furnace practice has nearly driven it out of the Missouri field. It is not generally known that the cost of producing pig lead is smaller by the hearth method than by sintering and blast-furnace smelting. Yet it is doubtful if any metallurgist knowing this to be true would recommend an American hearth installation—by reason of the disadvantages which will be briefly enumerated. The furnace has been limited to a length of from 4 to 5 ft., requiring for its operation two men working an 8-hr. shift as a maximum. The product per man is small; the work is hot and laborious; and the difficulty of completely removing the dust and fume has made it a notorious source of lead poisoning, Although suitable for operations on a small scale a large number of hearths would be required for a plant smelting 10,000 tons of 68 to 70 per cent. concentrates per month; and the number of semi-skilled laborers would be six times the number of hearths—a condition which is prohibitive. On account of the high percentage of dust and fume produced, a large flue and bag-house installation is required and, in the past, the handling and re-treatment of this dust and fume has -been a serious problem.

Jan 1, 1916

By Dennis K. Mortensen

Startup of the 9000 tpd Sacaton concentrator is expected to increase Asarco's domestic copper concentrate capacity by 21%. The plant site is located due west of the mining operations, consisting of both crushing and concentrating facilities. Arthur G. McKee & Co. served as the prime contractor for the project. The Sacaton mineralization contains both sulfide and oxide ores. Primary mineral constituents are chalcocite, azurite, malachite, chrysocolla, and minor amounts of silver and occasional gold. Sulfides are distinguished from oxides when more than half the total copper content is in the form of a sulfide copper mineral. While non-sulfide copper minerals are considerably less prevalent throughout the orebody, they tend to occur quite intimately with the sulfide minerals, particularly in the upper zones of the mineralization. For this reason the decision was made to treat all the ore in a flotation circuit using existing sulfidizing methods.

Jan 11, 1974

By A. H. Kline, E. B. Zimmer, Zay Jeffries

It is well known that many properties of a given metal vary with the size of grain or cell. For most industrial purposes, where high ultimate strength and high elastic limit are desired, the manufacturer tries to produce a fine-grained structure. For some purposes, for example in transformer iron, a coarse-grained structure is desirable. The terms fine-grained and coarse-grained are used only in a relative sense. For example, a fine-grained cast copper might have grains 100 times larger than coarse-grained high-speed steel. Similarly, a finegrained steel rail might have grains or cells 100 times larger than coarsegrained wire, which was made of steel from the same heat. In some metals, the change in grain size is more appreciable khan the changes in any of the properties determined by the tensile test. Furthermore, in metals which have been annealed or subjected to high temperatures, grain size may prove to be a better indication of the life of the metal in use than the tensile test. One of the authors was able to supplant the tensile test with grain-size determinations in the control of metal for a specific use, the latter determination being more indicative of the life of the metal part than the tensile test. A grain-size determination with a tensile stest is, of course, more valuable than either singly. At present there is a decided dearth of data on actual grain-size determinations in metals, and consequently a corresponding lack of ability to interpret the determinations after they are made. It would be desirable to have correlation made between grain size and properties of metals, comparing thousands of determinations. In this way, grain-size determinations might be interpreted in much the same manner that tensile tests are now. Several Methods of Measuring Grain Size One of the chief reasons why more grain-size measurements have not been made is the excessive work involved in making a determination.

Jan 1, 1916

By Henry Louis

The work performed by the Royal Commission on Mining Subsidence is likely to prove of permancnt value, less perhaps for the conclusions it has reached and for the recommendations it has based upon these than for the admirable exposition which it affords of the law concerning mining subsidence and of the legal status of the person who has suffered damage by reason of such subsidence, as also of him whose working has caused the damage. The Commission was appointed by Royal Warrant in the middle of 1925 and delivered its final report in the middle of 1927. The terms of reference to the Commission were as follows: "To consider the operation of the law relating to the support of the surface of the land, and of buildings or works on or under the surface, by underlying or adjacent minerals; to enquire into the extent and gravity of the damage caused by subsidence owing to the extraction of minerals and the incidence of the resulting liability; and to report what steps should be taken, by legislation or otherwise, to remedy, equitably to all persons concerned, any defects or hardships that may be found to arise in existing conditions." Report on the Doncaster Area The Commission submitted an interim report at the end of 1925, which referred solely to one area, usually spoken of as the Doncaster area, and recommended that a Commission of a more technical character be appointed to investigate this particular case. This recommendation was acted upon, and the Commission thus appointed reported early in 1928. Here it is necessary only to indicate the nature of the special problem that had to be solved, the proposed solution being mainly of local interest. The Doncaster area consists of about 330 sq. miles of low-lying land in the southeasterly portion of Yorkshire and adjoining parts of Lincolnshire and Nottinghamshire. No less than 78 per cent, of this land lies below the 25-ft. contour line, and a good deal is even below high-water level. It includes a large tract of land known as Hatfield Chase, swampy ground which has always been a source of trouble on that account, so much so that in 1626 a Dutchman, Cornelius Vermuyden, was called in by King Charles I to devise and execute schemes of drainage for its reclamation. The entire Doncaster area is

Jan 1, 1930

By Donald W. Scott, L. A. Roe, B. J. Sweo

INTRODUCTION THIS paper describes the application of ore-dressing methods to the reclamation of milled frit from over-spray, or waste, porcelain enamel. Frit is the name given by enamelers to a granulated glass produced by the smelting and subsequent water quenching of various ingredients capable of forming a glass, usually including a compound to produce opacity. It is the major component of porcelain enamel, and normally commands a price ranging from 7 to 12 cents a pound. In the preparation of a white enamel slip, usually 6 or 7 parts of clay, 2 parts of opacifier, small amounts of electrolytes, in some cases inorganic color stabilizers and organic dyes, and 35 parts of water are added to loo parts of frit; the mixture then is milled to a fineness of the order of 92 to 95 pct through a 200-mesh sieve. The frit forms the base of the enamel; the opacifier imparts opacity or reflectance; the clay suspends the frit while in slip form and serves as a binder' in bisque form (dry and unfired); and the electrolytes control the workability. Spraying is used commercially, in many cases, to apply the finish coat of enamel to the ware, prior to drying and firing. In the spraying operation, only about 50 pct, by dry weight, of the enamel slip adheres to the surface of the ware, the remainder forms the over-spray. From 1/2 to 2/3 of the overspray goes up the flue where it is lost or recovered, depending on whether the plant is equipped with a dust-collecting installation; the coarser portion of the overspray is collected in the spray booth. The former product is termed flue reclaim enamel, and the latter booth reclaim enamel. Although the spraying operation is done in specially constructed booths, which are partially or completely enclosed, it is extremely difficult to keep out contaminating dirt, which joins the reclaim. Some present-day large enameling plants use 20 tons of frit a day, of which 10 tons go into reclaim with a potential value of $1400 to $2400. When two-coat finish enamels were common, the reclaim was used satisfactorily for the undercoat, but the relatively recent advent of one-coat finish enamels prevented use of the reclaim because, even when blended with first-run enamels, it gave dirty finishes with poor gloss. It was thought that the pitting and specking encountered with reclaimed enamels were caused by the contaminating dirt, and that the poor gloss was caused, primarily, by the high clay content; more clay usually was added to give proper suspension properties to the slip when the reclaim was used. A clay content in excess of about 10 pct usually imparted refractoriness to porcelain enamels. Present-day one-coat porcelain enamels must combine high gloss and opacity with freedom from surface defects. Methods of beneficiating the over-spray,

Jan 1, 1947