ROTARY BLOWERS VS. BLOWING ENGINES FOR LEAD SMELTING
(April 27, 1901)
A note in the communication from S. E. Bretherton on “Pyritic Smelting and Hot Blast,” published in the Engineering and Mining Journal of April 13, 1901, refers to a subject of great interest to lead smelters. Mr. Bretherton remarked that he had been recently informed by August Raht that by actual experiment the loss with the ordinary rotary blowers was 100 per cent. under 10 lb. pressure; that is, it was possible to shut all the gates so that there was no outlet for the blast to escape from the blower and the pressure was only 10 lb., or in other words the blower would deliver no air against 10 lb. pressure. For that reason Mr. Raht expressed himself as being in favor of blowing engines for lead blast furnaces. This is of special interest, inasmuch as it comes from one who is recognized as standing in the first rank of lead-smelting engineers. Mr. Raht is not alone in holding the opinion he does.
The rotary blower did good service in the old days when the air was blown into the lead blast furnace at comparatively moderate pressure. At the present time, when the blast pressure employed is commonly 40 oz. at least, and sometimes as high as 48 oz., the deficiencies of the rotary blower have become more apparent. Notwithstanding the excellent workmanship which is put into them by their manufacturers, the extensive surfaces of contact are inherent to the type, and leakage of air backward is inevitable and important at the pressures now prevailing. The impellers of a rotary blower should not touch each other nor the cylinders in which they revolve, but they are made with as little clearance as possible, the surfaces being coated with grease, which fills the clearance space and forms a packing. This will not, however, entirely prevent leakage, which will naturally increase with the pressure. Even the manufacturers of rotary blowers admit the defects of the type, and concede that for pressures of 5 lb. and upward the cylinder blowing engine is the more economical. Metallurgists are coming generally to the opinion, however, that blowing engines are probably more economical for pressures of 4 lb. or thereabouts, and some go even further. With the blowing engines the air-joints of piston and cylinder are those of actual contact, and the metallurgist may count on his cubic feet of air, whatever be the pressure. Blowing engines were actually introduced several years ago by M. W. Iles at what is now the Globe plant of the American Smelting and Refining Company, and we believe their performance has been found satisfactory.
The fancied drawback to the use of blowing engines is their greater first cost, but H. A. Vezin, a mechanical engineer whose opinions carry great weight, pointed out five years ago in the Transactions of the American Institute of Mining Engineers (Vol. XXVI) that per cubic foot of air delivered the blowing engine was probably no more costly than the rotary blower, but on the contrary cheaper, stating that the first cost of a cylinder blower is only 20 to 25 per cent. more than that of a rotary blower of the same nominal capacity and the engine to drive it. The capacity of a rotary blower is commonly given as the displacement of the impellers per revolution, without allowance for slip or leakage backward. Mr. Vezin expressed the opinion that for the same actual capacity at 2 lb. pressure, that is, the delivery in cubic feet against 2 lb. pressure, the cylinder blower would cost no more than, if as much as, the rotary blower.
In this connection it is worth while making a note of the increasing tendency of lead smelters to provide much more powerful blowers than were formerly considered necessary, due, no doubt, in large measure to the recognition of the greater loss of air by leakage backward at the pressure now worked against. It is considered, for example, that a 42 × 140 in. furnace to be driven under 40 oz. pressure should be provided with a No. 10 blower, which size displaces 300 cu. ft. of air per revolution and is designed to be run at about 100 r.p.m.; its nominal capacity is, therefore, 30,000 cu. ft. of air per minute; although its actual delivery against 40 oz. pressure is much less, as pointed out by Mr. Raht and Mr. Bretherton. The Connersville Blower Company, of Connersville, Ind., lately supplied the Aguas Calientes plant (now of the American Smelting and Refining Company) with a rotary blower of the above capacity, and duplicates of it have been installed at other smelting works. The force required to drive such a huge blower is enormous, being something like 400 h.p., which makes it advisable to provide each blower with a directly connected compound condensing engine.
In view of the favor with which cylindrical blowing engines for driving lead blast furnaces are held by many of the leading lead-smelting engineers, and the likelihood that they will come more and more into use, it will be interesting to observe whether the lead smelters will take another step in the tracks of the iron smelters and adopt the circular form of blast furnace that is employed for the reduction of iron ore. The limit of size for rectangular furnaces appears to have been reached in those of 42 × 145 in., or approximately those dimensions. A furnace of 66 × 160 in., which was built several years ago at the Globe plant at Denver, proved a failure. H. V. Croll at that time advocated the building of a circular furnace instead of the rectangular furnace of those excessive dimensions and considered that the experience with the latter demonstrated their impracticability. In the Engineering and Mining Journal of May 28, 1898, he stated that there was no good reason, however, why a furnace of 300 to 500 tons daily capacity could not be run successfully, but considered that the round furnace was the only form permissible. We are unaware whether Mr. Croll was the first to advocate the use of large circular furnaces for lead smelting, but at all events there are other experienced metallurgists who now agree with him, and the time is, perhaps, not far distant when they may be adopted.
ROTARY BLOWERS VS. BLOWING ENGINES
By J. Parke Channing
(June 8, 1901)
In the issues of the Engineering and Mining Journal for April 13th and 27th reference was made to the relative efficiency of piston-blowing engines and rotary blowers of the impeller type, and in these articles August Raht was quoted as saying that, with an ordinary rotary blower working against 10 lb. pressure, the loss was 100 per cent. I have waited some time with the idea that some of the blower people would call attention to the concealed fallacy in the statement quoted, but so far have failed to notice any reference to the matter. I feel quite sure that Mr. Bretherton failed to quote Mr. Raht in full. The one factor missing in this statement is the speed at which the blower was run when the loss was 100 per cent.
The accepted method of testing the volumetric efficiency of rotary blowers is that of “closed discharge.” The discharge opening of the blower is closed, a pressure gage is connected with the closed delivery pipe, and the blower is gradually speeded up until the gage registers the required pressure. The number of revolutions which the blower makes while holding that pressure, multiplied by the cubic feet per revolution, will give the total slip of that particular blower at that particular pressure. Experience has shown that, within the practical limits of speed at which a blower is run, the slip is a function of the pressure and has nothing to do with the speed. If, therefore, it were found that the particular blower referred to by Mr. Raht were obliged to be revolved at the rate of 30 r.p.m. in order to maintain a constant pressure of 10 lb. with a closed discharge, and if the blower were afterward put in practical service, delivering air, and were run at a speed of 150 r.p.m., it would then follow that its delivery of air would amount to: 150-30 = 120. Its volumetric efficiency would be 120 ÷ 150 = 80 per cent. The above figures must not be relied upon, as I give them simply by way of illustration.
About a year ago I had the pleasure of examining the tabulated results of some extensive experiments in this direction, made by one of the blower companies. I believe they carried their experiments up to 10 lb. pressure, and I regret that I have not the figures before me, so that I could give something definite. I do, however, remember that in the experimental blower, when running at about 150 r.p.m., the volumetric efficiency at 2 lb. pressure was about 85 per cent., and that at 3 lb. pressure the volumetric efficiency was about 81 per cent.
It is unnecessary in this connection to call attention to the horse-power efficiency of rotary blowers. This is a matter entirely by itself, and there is considerable difference of opinion among engineers as to the relative horse-power efficiency of rotary blowers and piston blowers. All agree that there is a certain pressure at which the efficiency of the blower becomes less than the efficiency of the blowing engine. This I have heard placed all the way from 2 lb. up to 6 lb.
At the smelting plant of the Tennessee Copper Company we have lately installed blast-furnace piston-blowing engines; the steam cylinders are of the Corliss type and are 13 and 24 in. by 42 in.; the blowing cylinders are two in number, each 57 × 42 in.; the air valves are all Corliss in type. These blowing engines are designed to operate at a maximum air pressure of 2½ lb. per square inch.
At the Santa Fe Gold and Copper Mining Company’s smelter we have recently installed a No. 8 blower directly coupled to a 14 × 32 in. Corliss engine. This blower has been in use about five months and is giving very good results against the comparatively low pressure of 12 oz., or ¾ lb.
During the coming summer it is my intention to make careful volumetric and horse-power tests on these two types of machines under similar conditions of air pressure, and to publish the results; but in the meantime I wish to correct the error that a rotary blower of the impeller type is not a practicable machine at pressure over 5 lb.
BLOWERS AND BLOWING ENGINES FOR LEAD AND COPPER SMELTING
By Hiram W. Hixon
(July 20, 1901)
In the Engineering and Mining Journal for July 6th I note the discussion over the relative merits of blowers and blowing engines for lead and copper smelting, and wish to state that, irrespective of the work to be done, the blast pressure will depend entirely on the charge burden in any kind of blast-furnace work, and that the charge burden governs the reducing action of the furnace altogether. Along these lines the iron industry has raised the charge burden up to 100 ft. to secure the full benefit of the reducing action of the carbon monoxide on the ore.
In direct opposition to this we have what is known as pyritic smelting, wherein the charge burden governs the grade of the matte produced to such an extent that if a charge run with 4 to 6 ft. of burden above the tuyeres, producing 40 per cent. matte, is changed to a charge burden of 10 or 12 ft., the grade of the matte will decrease from 40 per cent. to probably less than 20 per cent. I can state this as a fact from recent experience in operating a blast furnace on heap-roasted ores under the conditions named, with the result as above stated.
I was exceedingly skeptical about pyritic smelting as advocated by some of your correspondents, and still continue to be so; but on making inquiries from some of my co-workers in this line, Mr. Sticht of Tasmania, and Mr. Nutting of Bingham, Utah, I have arrived at the following conclusion, to which some may take exception: That pyritic smelting without fuel, or with less than 5 per cent., with hot blast, is practically impossible; that smelting raw ore with a low charge burden to avoid the reducing action of the carbon monoxide, thereby securing oxidation of the iron and sulphur, is possible and practicable, under favorable conditions; and that a large portion of the sulphur is burned off, and the iron, without reducing action, goes into the slag in combination with silica. These results can be obtained with cold blast.
A blowing engine would certainly be much out of place for operating copper-matting furnaces run with the evident intention of oxidizing sulphur and iron and securing as high a grade of matte as possible, for the reason that to do this it is necessary to run a low charge burden, and with a low charge burden a high pressure of blast cannot be maintained. With a 4 to 6 ft. charge burden the blast pressure at Victoria Mines at present is 3 oz., produced by a No. 6 Green blower run at 120 r.p.m.; and a blowing engine, delivering the same amount of air, would certainly not give more pressure. Under the conditions which we have, a fan would be more effective than a pressure blower, and a blowing engine entirely out of the question as far as economy is concerned.
I installed blowing engines at the East Helena for lead smelting where the charge burden was 21 ft. and the blast pressure at times went up as high as 48 oz. Under these conditions the blowing engines gave satisfaction, but I am of the opinion that the same amount of blast could have been obtained under that pressure with less horse-power by the best type of rotary blower. I do not believe that the field of the blowing engine properly exists below 5 lb., and if this pressure cannot be obtained by charge-burden conditions, their installation is a mistake.
I wish to add the very evident fact that varying the grade of the matte by feeding the furnace at different hights varies the slag composition as to its silica and iron contents and makes the feeder the real metallurgist. The reducing action in the furnace is effected almost entirely by the gases, and when these are allowed to go to waste, reduction ceases.
BLOWING ENGINES AND ROTARY BLOWERS—HOT BLAST FOR PYRITIC SMELTING
By S. E. Bretherton
(August 24, 1901)
I have just read in the Engineering and Mining Journal of July 20th an interesting letter written by Hiram W. Hixon in regard to blowing engines versus the rotary blowers, and also the use of cold blast for pyritic smelting.
The controversy, which I unintentionally started in my letter in the Engineering and Mining Journal of April 13th last, about the advantages of using either blowers or blowing engines for blast furnaces, does not particularly interest me, with the exception that I have about decided, in my own mind, to use blowing engines where there is much back pressure, and the ordinary up-to-date blower for pyritic or matte smelting where much back pressure should not exist. I fully appreciate the fact that so-called pyritic smelting can be done to a limited extent, even with cold blast. Theoretically, enough oxygen can be sent into the blast furnace, contained in the cold blast, to oxidize both the fuel and the sulphur in an ordinary sulphide charge, but I have not yet learned where a high concentration is being made with unroasted ore and cold blast. I experimented on these lines at different times for three years, during 1896, 1897, and 1898, making a fair concentration with refractory ores, most of which had been roasted. I was myself interested in the profits and as anxious as any one for economy. We tried, for fuel in the blast furnace, coke alone, coke and lignite coal, lignite coal alone, lignite coal and dry wood, coal and green wood, and then coke and green wood, all under different hights of ore burden in the furnace.
A description of these experiments would, no doubt, be tiresome to your readers, but I wish to state that the furnace was frozen up several times on account of using too little fuel, when the cold blast would gradually drive nearly all the heat to the top of the furnace, the crucible and between the tuyeres becoming so badly crusted that the furnace had to be cleaned out and blown in again, unless I was called in time to save it by changing the charge and increasing the fuel. We were making high-grade matte under contract, high concentration and small matte fall, which would, no doubt, aggravate matters.
After the introduction of hot blast, heated up to between 200 and 300 deg. F., we made the same grade of matte from the same character of ore, with the exception that we then smelted without roasting, and reduced the percentage of fuel consumption, increased the capacity of the furnace, and almost entirely obviated the trouble of cold crucibles and hot tops. I write the above facts, as they speak for themselves.
I nearly agree with Mr. Hixon, and do not think it practical to smelt with much less than 5 per cent. coke continuously; but there is a great saving between the amount of coke used with a moderately heated blast and cold blast. Regardless of either hot or cold blast, however, the fuel consumption depends very much on the character of the ore to be smelted, the amount of matte-fall and grade of matte made. It is not always advisable or necessary to use hot blast for a matting furnace; that is, where the supply of sulphur is limited. It may then be necessary to use as much fuel in the blast furnace to prevent the sulphur from oxidizing as will be sufficient to furnish the heat for smelting. Such conditions existed at Silver City, N. M. , at times, after our surplus supply of iron and zinc sulphide concentrates was used. I understand that they are now short of sulphur there, on account of getting a surplus amount of oxidized copper ore, and are only utilizing what little heat the slag gives them, without the addition of any fuel on top of the forehearth.
Before closing this, which I intended to have been brief, I wish to call your attention to a little experience we had with alumina in the matting furnace at Silverton, Col., where I was acting as consulting metallurgist. The ore we had to smelt contained, on an average, about 20 per cent. Al2O3, 30 per cent. SiO2, with 18 per cent. Fe in the form of an iron pyrite, and no other iron was available except some iron sulphide concentrates containing a small percentage of zinc and lead.
The question naturally arose, could we oxidize and force sufficient of the iron into the slag, and where should we class the alumina, as a base or an acid? My experience in lead smelting led me to believe that Al2O3 could only be classed as an acid in the ordinary lead furnace, and that it would be useless to class it otherwise in a shallow matting furnace; and E. W. Walter, the superintendent and metallurgist in charge, agreed with me.
We then decided to make a bisilicate slag, classing the alumina as silica, and we obtained fairly satisfactory results. The slag made was very clean, but treacherous, which was attributed to two reasons: First, that it required more heat to keep the alumina slag liquid enough to flow than it does a nearly straight silica slag; and, second, that we were running so close to the formula of a bisilicate and aluminate slag (about 31½ per cent. SiO2, 27 per cent. Fe, 20 per cent. CaO, and 18 per cent. Al2O3, or 49½ per cent. acid) that a few charges thrown into the furnace containing more silica or alumina than usual would thicken the slag so that it would then require some extra coke and flux to save the furnace. At times the combined SiO2 and Al2O3 did reach 55 and 56 per cent. in the slag, which did not freeze up the furnace, but caused us trouble.
PART IX
LEAD REFINING
THE REFINING OF LEAD BULLION[49]
By F. L. Piddington
(October 3, 1903)
In presenting this account of the Parkes process of desilverizing and refining lead bullion no originality is claimed, but I hope that a description of the process as carried out at the works of the Smelting Company of Australia may be of service.
Introductory.—The Parkes process may be conveniently summarized as follows:
1. Softening of the base bullion to remove copper, antimony, etc.
2. Removal of precious metals from the softened bullion by means of zinc.
3. Refining the desilverized lead.
4. Liquation of gold and silver crusts obtained from operation No. 2.
5. Retorting the liquated alloy to drive off zinc.
6. Concentrating and refining bullion from No. 5.
Softening.—This is done in reverberatory furnaces. In large works two furnaces are used, copper, antimony, and arsenic being removed in the first and antimony in the second. The size of the furnaces is naturally governed by the quantity to be treated. In these works (refining about 200 tons weekly) a double set of 15-ton furnaces were at work. The sides and ends of these furnaces are protected by a jacket with a 2-in. water space, the jacket extending some 3 in. above the charge level and 6 in. to 9 in. below it. The furnace is built into a wrought-iron pan, and if the brickwork is well laid into the pan there need be no fear of lead breaking through below the jacket.
The bars of bullion (containing, as a rule, 2 to 3 per cent. of impurities) are placed in the furnace carefully, to avoid injuring the hearth, and melted down slowly. The copper dross separates out and floats on top of the charge, which is stirred frequently to expose fresh surfaces. If the furnace is overheated some dross is melted into the lead again and will not separate out until the charge is cooled back. However carefully the work is done some copper remains with the lead, and its effects are to be seen in the later stages. The dross is skimmed into a slag pot with a hole bored in it some 4 in. from the bottom; any lead drained from the pot is returned to the charge. The copper dross is either sent back to the blast furnace direct or may be first liquated. By the latter method some 30 per cent. of the lead contents of the dross is recovered in the refinery.
Base bullion made at a customer’s smelter will often vary greatly in composition, and it is, therefore, difficult to give any hard and fast figures as to percentage of metals in the dross. As a rule our dross showed 65 to 70 per cent. lead, copper 2 to 9 per cent. (average 4 per cent.), gold and silver values varying with the grade of the original bullion, though it was difficult to detect any definite relation between bullion and dross. It was, however, noticed that gold and silver values increased with the percentage of copper.
Immediately the copper dross is skimmed off the heat is raised considerably, and very soon a tin (and arsenic, if present) skimming appears. It is quite “dry” and may be removed in an hour or so. It is a very small skimming, and the tin, not being worth saving, is put with the copper dross.
The temperature is now raised again and antimony soon shows in black, boiling, oily drops, gathering in time into a sheet covering the surface of the lead. When the skimming is about ½-inch thick, slaked lime, ashes, or fine coal is thrown on and stirred in. The dross soon thickens up and may be skimmed off easily. This operation is repeated until all antimony is eliminated. Constant stirring of the charge is necessary. The addition of litharge greatly facilitates the removal of antimony; either steam or air may be blown on the surface of the metal to hasten oxidation, though they have anything but a beneficial effect on the furnace lining. From time to time samples of the dross are taken in a small ladle, and after setting hard the sample is broken in two. A black vitreous appearance indicates plenty of antimony yet in the charge. Later samples will look less black, until finally a few yellowish streaks are seen, being the first appearance of litharge. When all antimony is out the fracture of a sample should be quite yellow and the grain of the litharge long, a short grain indicating impurities still present, in which case another skimming is necessary. The analysis of a representative sample of antimony dross was as follows:
| PbO = | 78.11 per cent. |
| Sb2O4 = | 8.75 ” |
| As2O3 = | 2.18 ” |
| CuO = | 0.36 ” |
| CaO = | 1.10 ” |
| Fe2O3 = | 0.42 ” |
| Al2O3 = | 0.87 ” |
| Insol. = | 4.10 ” |
Antimony dross is usually kept separate and worked up from time to time, yielding hard antimonial lead, used for type metal, Britannia metal, etc.
Desilverization.—The softening being completed the charge is tapped and run to a kettle or pan of cast iron or steel, holding, when conveniently full, some 12 or 13 tons. The lead falling into the kettle forms a considerable amount of dross, which is skimmed off and returned to the softening furnace. By cooling down the charge until it nearly “freezes” an additional copper skimming is obtained, which also is returned to the softener. The kettle is now heated up to the melting point of zinc, and the zinc charge, determined by the gold and silver contents of the kettle, is added and melted. The charge is stirred, either by hand or steam, for about an hour, after which the kettle is allowed to cool down for some three hours and the first zinc crust taken off. When the charge is skimmed clean a sample of the bullion is taken for assay, and while this is being done the kettle is heated again for the second zinc charge, which is worked in the same way as the first; sometimes a third addition of zinc is necessary. The resulting crusts are kept separate, the second and third being added to the next charge as “returns,” allowing 3 lb. of zinc in returns as equal to 1 lb. of fresh zinc. An alternative method is to take out gold and silver in separate crusts, in which case the quantity of zinc first added is calculated on the gold contents of the kettle only. The method of working is the same, though subsequent treatment may differ in that the gold crusts are cupeled direct.
As to the quantity of zinc required:
1. Extracting the gold with as little silver as possible, the following figures were obtained:
| Total Gold in Kettle, oz. | Amount taken out by 1 lb. zinc, oz. |
|---|---|
| 300 | 1.00 |
| 200 | 1.00 |
| 150 | 0.79 |
| 100 | 0.59 |
| 60 | 0.45 |
2. Silver zinking gave the following general results with 11-ton charges:
| Total Silver in Kettle, oz. | Amount taken out by 1 lb. zic, oz. |
|---|---|
| 1,200 | 4.1 |
| 930 | 3.8 |
| 755 | 3.5 |
| 616 | 3.4 |
| 460 | 2.6 |
3. Extracting gold and silver together:
| Total Contents of Kettle | 1 Lb. Zinc Takes Out | ||
|---|---|---|---|
| Au. Oz. | Ag. Oz. | Au. Oz. | Ag. Oz. |
| 494 | 3,110 | 0.59 | 3.60 |
| 443 | 1,883 | 0.64 | 2.80 |
| 330 | 2,417 | 0.45 | 3.34 |
| 204 | 1,638 | 0.36 | 2.86 |
| 143 | 1,330 | 0.28 | 2.65 |
| 123 | 1,320 | 0.23 | 2.54 |
It will be noticed that in each case the richer the bullion the greater the extractive power of zinc. Experiments made on charges of rich bullion showed that the large amount of zinc called for by the table in use was unnecessary, and 250 lb. was fixed on as the first addition of zinc. On this basis an average of 237 charges gave results as follows:
| Total Contents | Zinc Used Lbs. | 1 Lb. Zinc Takes Out | ||
|---|---|---|---|---|
| Au. Oz. | Ag. Oz. | Au. Oz. | Ag. Oz. | |
| 520 | 1,186 | 507.5 | 1.27 | 2.91 |
The zinc used was that necessary to clean the kettle, added as follows: 1st, 250 lb.; 2d (average), 127 lb.; 3d (average), 57 lb. In 112 cases no third addition was required. From these figures it appears that in the earlier work the zinc was by no means saturated.
Refining the Lead.—Gold and silver being removed, the lead is siphoned off into the refining kettle and the fire made up. In about four hours the lead will be red hot, and when hot enough to burn zinc, dry steam, delivered by a ¾ in. pipe reaching nearly to the bottom of the kettle, is turned on. The charge is stirred from time to time and wood is fed on the top to assist dezinking and prevent the formation of too much litharge. In three to four hours the lead will be soft and practically free from zinc. When test strips show the lead to be quite soft and clean, the kettle is cooled down and the scum of lead and zinc oxides skimmed off. In an hour or so the lead will be cool enough for molding; the bar should have a yellow luster on the face when set; if the lead is too cold it will be white, if too hot a deep blue. The refining kettles are subjected to severe strain during the steaming process, and hence their life is uncertain—an average would be about 60 charges; the zinking kettles, on the other hand, last very much longer. Good steel kettles (if they can be obtained) are preferable to cast iron.
Treatment of Zinc Crusts.—Having disposed of the lead, let us return now to the zinc crusts. These are first liquated in a small reverberatory furnace, the hearth of which is formed of a cast-iron plate (the edges of the long sides being turned up some 4 in.) laid on brasque filling, with a fall from bridge to flue of ¾ in. per foot and also sloping from sides to center. The operation is conducted at a low temperature and the charge is turned over at intervals, the liquated lead running out into a small separately fired kettle. This lead rarely contains more than a few ounces of silver per ton; it is baled into bars, and returned to the zinking kettles or worked up in a separate charge. In two to three hours the crust is as “dry” as it is advisable to make it, and the liquated alloy is raked out over a slanting perforated plate to break it up and goes to the retort bin.
Retorting the Alloy.—This is carried on in Faber du Faur tilting furnaces—simply a cast-iron box swinging on trunnions and lined with firebrick. Battersea retorts (class 409) holding 560 lb. each are used; their average life is about 30 charges. The retorts are charged hot, a small shovel of coal being added with the alloy. The condenser is now put in place and luted on; it is made of ⅛ in. iron bent to form a cylinder 12 in. in diameter, open at one end; it is lined with a mixture of lime, clay and cement. It has three holes, one on the upper side close to the furnace and through which a rod can be passed into the retort, a vent-hole on the upper side away from the furnace, and a tap-hole on the bottom for condensed zinc. In an hour or so the flame from the vent-hole should be green, showing that distillation has begun. When condensation ceases (shown by the flame) the condenser is removed and the bullion skimmed and poured into bars for the cupel. The products of retorting are bullion, zinc, zinc powder and dross. Bullion goes to the cupel, zinc is used again in the desilverizing kettles, powder is sieved to take out scraps of zinc and returned to the blast furnace, or it may be, and sometimes is, used as a precipitating agent in cyanide works; dross is either sweated down in a cupel with lead and litharge, together with outside material such as zinc gold slimes from cyanide works, jeweler’s sweep, mint sweep, etc., or in the softening furnace after the antimony has been taken off. In either case the resulting slag goes back to the blast furnace. The total weight of alloy treated is approximately 7 per cent. of the original base bullion. The zinc recovered is about 60 per cent. of that used in desilverizing. The most important source of temporary loss is the retort dross (consisting of lead-zinc-copper alloy with carbon, silica and other impurities), and it is here that the necessity of removing copper in the softening process is seen, since any copper comes out with the zinc crusts and goes on to the retorts, where it enters the dross, carrying gold and silver with it. If much copper is present the dross may contain more gold and silver than the retort bullion itself. In this connection I remember an occasion on which some retort dross yielded gold and silver to the extent of over 800 and 3000 oz. per ton respectively.
Cupellation.—Retort bullion is first concentrated (together with bullion resulting from dross treatment) to 50 to 60 per cent. gold and silver in a water-jacketed cupel. The side lining is protected by an inch water-pipe imbedded in the lining at the litharge level or by a water-jacket, the inner face of which is of copper; the cupel has also a water-jacketed breast so that the front is not cut down. The cupel lining may be composed of limestone, cement, fire-clay and magnesite in various proportions, but a simple lining of sand and cement was found quite satisfactory. When the bullion is concentrated up to 50 to 60 per cent. gold and silver, it is baled out and transferred to the finishing cupel, where it is run up to about 0.995 fine; it is then ready either for the melting-pot or parting plant. The refining test, by the way, is not water-cooled.
Re-melting is done in 200-oz. plumbago crucibles and presents no special features. In the case of doré bullion low in gold, “sprouting” of the silver is guarded against by placing a piece of wood or charcoal on the surface of the metal before pouring, and any slag is kept back. The quantity of slag formed is, of course, very small, so that the bars do not require much cleaning.
The parting plant was not in operation in my time, and I am therefore unable to go into details. The process arranged for was briefly as follows: Solution of the doré bullion in H2SO4; crystallization of silver as monosulphate by dilution and cooling; decomposition of silver sulphate by ferrous sulphate solution giving metallic silver and ferric sulphate, which is reduced to the ferrous salt by contact with scrap iron. The gold and silver are washed thoroughly with hot water and cast into bars.
In conclusion, some variations in practice may be noted. The use of two furnaces in the softening process has already been mentioned; by this means the drossing and softening are more perfect and subsequent operations thereby facilitated; further, the furnaces, being kept at a more equable temperature, are less subject to wear and tear. Zinc crusts are sometimes skimmed direct into an alloy press in which the excess of lead is squeezed out while still molten; liquation is then unnecessary. Refining of the lead may be effected by a simple scorification in a reverberatory, the soft lead being run into a kettle from which it is molded into market bars.
THE ELECTROLYTIC REFINING OF BASE LEAD BULLION
By Titus Ulke
(October 11, 1902)
Important changes in lead-refining practice are bound to follow, in my opinion, the late demonstration on a large scale of the low working cost and high efficiency of Betts’ electrolytic process of refining lead bullion. It was my good fortune recently to see this highly interesting process in operation at Trail, British Columbia, through the kindness of the inventor, A. G. Betts, and Messrs. Labarthe and Aldridge, of the Trail works.
A plant of about 10 tons daily capacity, which probably cost about $25,000, although it could be duplicated for perhaps $15,000 at the present time, was installed near the Trail smelting works. It has been in operation for about ten months, I am informed, with signal success, and the erection of a larger plant, of approximately 30 tons capacity and provided with improved handling facilities, is now completed.
The depositing-room contains 20 tanks, built of wood, lined with tar, and approximately of the size of copper-refining tanks. Underneath the tank-room floor is a basement permitting inspection of the tank bottoms for possible leakage and removal of the solution and slime. A suction pump is employed in lifting the electrolyte from the receiving tank and circulating the solution. In nearly every respect the arrangement of the plant and its equipment is strikingly like that of a modern copper refinery.
The great success of the process is primarily based upon Betts’ discovery of the easy solubility of lead in an acid solution of lead fluosilicate, which possesses both stability under electrolysis and high conductivity, and from which exceptionally pure lead may be deposited with impure anodes at a very low cost. With such a solution, there is no polarization from formation of lead peroxide on the anode, no evaporation of constituents except water, and no danger in its handling. It is cheaply obtained by diluting hydrofluoric acid of 35 per cent. HF, which is quoted in New York at 3c. per pound, with an equal volume of water and saturating it with pulverized quartz according to the equation:
SiO2 + 6HF = HSiF6 + 2H2O.
According to Mr. Betts, an acid of 20 to 22 per cent. will come to about $1 per cu. ft., or to $1.25 when the solution has been standardized with 6 lb. of lead. One per cent. of lead will neutralize 0.7 per cent. H2SiF6. The electrolyte employed at the time of my inspection of the works contained, I believe, 8 per cent. lead and 11 per cent. excess of fluosilicic acid.
The anodes consist of the lead bullion to be refined, cast into plates about 2 in. thick and approximately of the same size as ordinary two-lugged copper anodes. Before being placed in position in the tanks, they are straightened by hammering over a mold and their lugs squared. No anode sacks are employed as in the old Keith process.
The cathode sheets which receive the regular lead deposits are thin lead plates obtained by electrodeposition upon and stripping from special cathodes of sheet steel. The latter are prepared for use by cleaning, flashing with copper, lightly lead-plating in the tanks, and greasing with a benzine solution of paraffin, dried on, from which the deposited lead is easily stripped.
The anodes and cathodes are separated by a space of 1½ to 2 in. in the tank and are electrically connected in multiple, the tanks being in series circuit. The fall in potential between tanks is only about 0.2 of a volt, which remarkably low voltage is due to the high conducting power of the electrolyte and to some extent to the system of contacts used. These contacts are small wells of mercury in the bus-bars, large enough to accommodate copper pins soldered to the iron cathodes or clamped to the anodes. Only a small amount of mercury is required.
Current strengths of from 10 to 25 amperes per sq. ft. have been used, but at Trail 14 amperes have given the most satisfactory results as regards economy of working and the physical and chemical properties of the refined metal produced.
A current of 1 ampere deposits 3.88 grams of lead per hour, or transports 3¼ times as much lead, in this case, as copper with an ordinary copper-refining solution. A little over 1000 kg., or 2240 lb., requires about 260,000 ampere hours. At 10 amperes per sq. ft. the cathode (or anode) area should be about 1080 sq. ft. per ton of daily output. Taking a layer of electrolyte 1.5 in. thick, 135 cu. ft. will be found to be the amount between the electrodes, and 175 cu. ft. may be taken as the total quantity of solution necessary, according to Mr. Betts’ estimate. The inventor states that he has worked continuously and successfully with a drop of potential of only 0.175 volt per tank, and that therefore 0.25 volt should be an ample allowance in regular refining. Quoting Mr. Betts; “260,000 ampere hours at 0.25 volt works out to 87 electrical h.p. hours of 100 h.p. hours at the engine shaft, in round numbers. Estimating that 1 h.p. hour requires the burning of 1.5 lb. of coal, and allowing say 60 lb. for casting the anodes and refined lead, each ton of lead refined requires the burning of 210 lb. of fuel.” With coal at $6 per ton the total amount of fuel consumed, therefore, should not cost over 60c., which is far below the cost of fire-refining base lead bullion, as we know.
In the Betts electrolytic process, practically all the impurities in the base bullion remain as a more or less adherent coating on the anode, and only the zinc, iron, cobalt and nickel present go into solution. The anode residue consists practically of all the copper, antimony, bismuth, arsenic, silver and gold contained in the bullion, and very nearly 10 per cent. of its weight in lead. Having the analysis of any bullion, it is easy to calculate with these data the composition of the anode residue and the rate of pollution of the electrolyte. Allowing 175 cu. ft. of electrolyte per ton of daily output, it will be found that in the course of a year these impurities will have accumulated to the extent of a very few per cent. Estimating that the electrolyte will have to be purified once a year, the amount to be purified daily is less than 1 cu. ft. for each ton of output. The amount of lead not immediately recovered in pure form is about 0.3 per cent., most of which is finally recovered. As compared with the ordinary fire-refined lead, the electrolytically refined lead is much purer and contains only mere traces of bismuth, when bismuthy base bullion is treated. Furthermore, the present loss of silver in fire refining, amounting, it is claimed, to about 1½ per cent. of the silver present, and covered by the ordinary loss in assay, is to a large extent avoided, as the silver in the electrolytic process is concentrated in the anode residue with a very small loss, and the loss of silver in refining the slimes is much less than in treating the zinc crusts and refining the silver residue after distillation. The silver slimes obtained at Trail, averaging about 8000 oz. of gold and silver per ton, are now treated at the Seattle Smelting and Refining Works. There the slimes are boiled with concentrated sulphuric acid and steam, allowing free access of air, which removes the greater part of the copper. The washed residue is then dried in pans over steam coils, and melted down in a magnesia brick-lined reverberatory, provided with blast tuyeres, and refined. In this reverberatory furnace the remainder of the copper left in the slimes after boiling is removed by the addition of niter as a flux, and the antimony with soda. The doré bars finally obtained are parted in the usual way with sulphuric acid, making silver 0.999 fine and gold bars at least 0.992 fine.
Mr. Betts treated 2000 grams of bullion, analyzing 98.76 per cent. Pb, 0.50 Ag, 0.31 Cu, and 0.43 Sb with a current of 25 amp. per square foot in an experimental way, and obtained products of the following composition:
Refined Lead: 99.9971 per cent. Pb, 0.0003 Ag, 0.0007 Cu, and 0.0019 Sb.
Anode Residue: 9.0 per cent. Pb, 36.4 Ag, 25.1 Cu, and 2.95 Sb.
Four hundred and fifty pounds of bullion from the Compania Metalurgica Mexicana, analyzing 0.75 per cent. Cu, 1.22 Bi, 0.94 As, 0.68 Sb, and assaying 358.9 oz. Ag and 1.71 oz. Au per ton, were refined with a current of 10 amp. per square foot, and gave a refined lead of the following analysis: 0.00027 per cent. Cu, 0.0037 Bi, 0.0025 As, 0 Sb, 0.0010 Ag, 0.0022 Fe, 0.0018 Zn and Pb (by difference) 99.9861 per cent.
Although the present method for recovering the precious metals and by-products from the anode residue leaves much room for improvement, the use of the Betts process may be recommended to our lead refiners, because it is a more economical and efficient method than the fire-refining process now in common use. I will state my belief, in conclusion, that the present development of electrolytic lead refining signalizes as great an advance over zinc desilverization and the fire methods of refining lead as electrolytic copper refining does over the old Welsh method of refining that metal.
ELECTROLYTIC LEAD-REFINING[50]
By Anson G. Betts
A solution of lead fluosilicate, containing an excess of fluosilicic acid, has been found to work very satisfactorily as an electrolyte for refining lead. It conducts the current well, is easily handled and stored, non-volatile and stable under electrolysis, may be made to contain a considerable amount of dissolved lead, and is easily prepared from inexpensive materials. It possesses, however, in common with other lead electrolytes, the defect of yielding a deposit of lead lacking in solidity, which grows in crystalline branches toward the anodes, causing short circuits. But if a reducing action (practically accomplished by the addition of gelatine or glue) be given to the solution, a perfectly solid and dense deposit is obtained, having very nearly the same structure as electrolytically deposited copper, and a specific gravity of about 11.36, which is that of cast lead.
Lead fluosilicate may be crystallized in very soluble brilliant crystals, resembling those of lead nitrate and containing four molecules of water of crystallization, with the formula PbSiF6,4H2O. This salt dissolves at 15 deg. C. in 28 per cent. of its weight of water, making a syrupy solution of 2.38 sp. gr. Heated to 60 deg. C., it melts in its water of crystallization. A neutral solution of lead fluosilicate is partially decomposed on heating, with the formation of a basic insoluble salt and free fluosilicic acid, which keeps the rest of the salt in solution. This decomposition ends when the solution contains perhaps 2 per cent. of free acid; and the solution may then be evaporated without further decomposition. The solutions desired for refining are not liable to this decomposition, since they contain much more than 2 per cent. of free acid. The electrical conductivity depends mainly on the acidity of the solution.
My first experiments were carried out without the addition of gelatine to the fluosilicate solution. The lead deposit consisted of more or less separate crystals that grew toward the anode, and, finally, caused short circuits. The cathodes, which were sheet-iron plates, lead-plated and paraffined, had to be removed periodically from the tanks and passed through rolls, to pack down the lead. When gelatine has been added in small quantities, the density of the lead is greater than can be produced by rolling the crystalline deposit, unless great pressure is used.
The Canadian Smelting Works, Trail, B. C. , have installed a refinery, making use of this process. There are 28 refining-tanks, each 86 in. long, 30 in. wide and 42 in. deep, and each receiving 22 anodes of lead bullion with an area of 26 by 33 in. exposed to the electrolyte on each side, and 23 cathodes of sheet lead, about 1/16 in. thick, prepared by deposition on lead-plated and paraffined iron cathodes. The cathodes are suspended from 0.5 by 1 in. copper bars, resting crosswise on the sides of the tanks. The experiment has been thoroughly tried of using iron sheets to receive a deposit thicker than 1/16 in.; that is, suitable for direct melting without the necessity of increasing its weight by further deposition as an independent cathode; but the iron sheets are expensive, and are slowly pitted by the action of the acid solution; and the lead deposits thus obtained are much less smooth and pure than those on lead sheets.
The smoothness and the purity of the deposited lead are proportional. Most of the impurity seems to be introduced mechanically through the attachment of floating particles of slime to irregularities on the cathodes. The effect of roughness is cumulative; it is often observed that particles of slime attract an undue amount of current, resulting in the lumps seen in the cathodes. Samples taken at the same time showed from 1 to 2.5 oz. silver per ton in rough pieces from the iron cathodes, 0.25 oz. as an average for the lead-sheet cathodes, and only 0.04 oz. in samples selected for their smoothness. The variation in the amount of silver (which is determined frequently) in the samples of refined lead is attributed not to the greater or less turbidity of the electrolyte at different times, but to the employment of new men in the refinery, who require some experience before they remove cathodes without detaching some slime from the neighboring anodes.
Each tank is capable of yielding, with a current of 4000 amperes, 750 lb. of refined lead per day. The voltage required to pass this current was higher than expected, as explained below; and for this reason, and also because the losses of solution were very heavy until proper apparatus was put in to wash thoroughly the large volume of slime produced (resulting in a weakened electrolyte), the current used has probably averaged about 3000 amperes. The short circuits were also troublesome, though this difficulty has been greatly reduced by frequent inspection and careful placing of the electrodes. At one time, the solution in use had the following composition in grams per 100 c.c.: Pb, 6.07; Sb, 0.0192; Fe, 0.2490; SiF6, 6.93, and As, a trace. The current passing was 2800 amperes, with an average of about 0.44 volts per tank, including bus-bars and contacts. It is not known what was the loss of efficiency on that date, due to short circuits; and it is, therefore, impossible to say what resistance this electrolyte constituted.
Hydrofluoric acid of 35 per cent., used as a starting material for the preparation of the electrolyte, is run by gravity through a series of tanks for conversion into lead fluosilicate. In the top tank is a layer of quartz 2 ft. thick, in passing through which the hydrofluoric acid dissolves silica, forming fluosilicic acid. White lead (lead carbonate) in the required quantity is added in the next tank, where it dissolves readily and completely with effervescence. All sulphuric acid and any hydrofluoric acid that may not have reacted with silica settle out in combination with lead as lead sulphate and lead fluoride. Lead fluosilicate is one of the most soluble of salts; so there is never any danger of its crystallizing out at any degree of concentration possible under this method. The lead solution is then filtered and run by gravity into the refining-tanks.
The solution originally used at Trail contained about 6 per cent. Pb and 15 per cent. SiF6.
The electrical resistance in the tanks was found to be greater than had been calculated for the same solution, plus an allowance for loss of voltage in the contacts and conductors. This is partly, at least, due to the resistance to free motion of the electrolyte, in the neighborhood of the anode, offered by a layer of slime which may be anything up to ½ in. thick. During electrolysis, the SiF6 ions travel toward the anodes, and there combine with lead. The lead and hydrogen travel in the opposite direction and out of the slime; but there are comparatively few lead ions present, so that the solution in the neighborhood of the anodes must increase in concentration and tend to become neutral. This greater concentration causes an e.m.f. of polarization to act against the e.m.f. of the dynamo. This amounted to about 0.02 volt for each tank. The greater effect comes from the greater resistance of the neutral solution with which the slime is saturated. There is, consequently, an advantage in working with rather thin anodes, when the bullion is impure enough to leave slime sticking to the plates. A compensating advantage is found in the increased ease of removing the slime with the anodes, and wiping it off the scrap in special tanks, instead of emptying the tanks and cleaning out, as is done in copper refineries.
It is very necessary to have adequate apparatus for washing solution out of the slime. The filter first used consisted of a supported filtering cloth with suction underneath. It was very difficult to get this to do satisfactory work by reason of the large amount of fluosilicate to be washed out with only a limited amount of water. At the present time the slime is first stirred up with the ordinary electrolyte several times, and allowed to settle, before starting to wash with water at all. The Trail plant produces daily 8 or 10 cu. ft. of anode residue, of which over 90 per cent. by volume is solution. The evaporation from the total tank surface of something like 400 sq. ft. is only about 15 cu. ft. daily; so that only a limited amount of wash-water is to be used—namely, enough to replace the evaporated water, plus the volume of the slime taken out.
The tanks are made of 2 in. cedar, bolted together and thoroughly painted with rubber paint. Any leakage is caught underneath on sloping boards. Solution is circulated from one tank to another by gravity, and is pumped from the lowest to the highest by means of a wooden pump. The 22 anodes in each tank together weigh about 3 tons, and dissolve in from 8 to 10 days, two sets of cathodes usually being used with each set of anodes. While 300 lb. cathodes can be made, the short-circuiting gets so troublesome with the spacing used that the loss of capacity is more disadvantageous than the extra work of putting in and taking out more plates. The lead sheets used for cathodes are made by depositing about 1/16 in. metal on paraffined steel sheets in four of the tanks, which are different from the others only in being a little deeper.
The anodes may contain any or all of the elements, gold, silver, copper, tin, antimony, arsenic, bismuth, cadmium, zinc, iron, nickel, cobalt and sulphur. It would be expected that gold, silver, copper, antimony, arsenic and bismuth, being more electronegative than lead, would remain in the slime in the metallic state, with, perhaps, tin, while iron, zinc, nickel and cobalt would dissolve. It appears that tin stands in the same relation to lead that nickel does to iron, that is, they have about the same electromotive forces of solution, with the consequence that they can behave as one metal and dissolve and deposit together. Iron, contrary to expectation, dissolves only slightly, while the slime will carry about 1 per cent. of it. It appears from this that the iron exists in the lead in the form of matte. Arsenic, antimony, bismuth and copper have electromotive forces of solution more than 0.3 volt below that of lead. As there is no chance that any particle of one of these impurities will have an electric potential of 0.3 volt above that of the lead with which it is in metallic contact, there is no chance that they will be dissolved by the action of the current. The same is even more certainly true of silver and gold. The behavior of bismuth is interesting and satisfactory. It is as completely removed by this process of refining as antimony is. No other process of refining lead will remove this objectionable impurity so completely. Tin has been found in the refined lead to the extent of 0.02 to 0.03 per cent. This we had no difficulty in removing from the lead by poling before casting. There is always a certain amount of dross formed in melting down the cathodes; and the lead oxide of this reacts with the tin in the lead at a comparatively low temperature.
The extra amount of dross formed in poling is small, and amounts to less than 1 per cent. of the lead. The dross carries more antimony and arsenic than the lead, as well as all the tin. The total amount of dross formed is about 4 per cent. Table I shows its composition.
The electrolyte takes up no impurities, except, possibly, a small part of the iron and zinc. Estimating that the anodes contain 0.01 per cent. of zinc and soluble iron, and that there are 150 cu. ft. of the solution in the refinery for every ton of lead turned out daily, in one year the 150 cu. ft. will have taken up 93 lb. of iron and zinc, or about 1 per cent. These impurities can accumulate to a much greater extent than this before their presence will become objectionable. It is possible to purify the electrolyte in several ways. For example, the lead can be removed by precipitation with sulphuric acid, and the fluosilicic acid precipitated with salt as sodium fluosilicate. By distillation with sulphuric acid the fluosilicic acid could be recovered, this process, theoretically, requiring but one-third as much sulphuric acid as the decomposition of fluorspar, in which the fluorine was originally contained.
The only danger of lead-poisoning to which the workmen are exposed occurs in melting the lead and casting it. In this respect the electrolytic process presents a distinct sanitary advance.
For the treatment of slime, the only method in general use consists in suspending the slime in a solution capable of dissolving the impurities and supplying, by a jet of steam and air forced into the solution, the air necessary for its reaction with, and solution of, such an inactive metal as copper. After the impurities have been mostly dissolved, the slime is filtered off, dried and melted, under such fluxes as soda, to a doré bullion.
The amount of power required is calculated thus: Five amperes in 24 hours make 1 lb. of lead per tank. One ton of lead equals 10,000 ampere-days, and at 0.35 volts per tank, 3500 watt-days, or 4.7 electric h.p. days. Allowing 10 per cent. loss of efficiency in the tanks (we always get less lead than the current which is passing would indicate), and of 8 per cent. loss in the generator, increases this to about 5.6 h.p. days, and a further allowance for the electric lights and other applications gives from 7 to 8 h.p. days as about the amount per ton of lead. At $30 per year, this item of cost is something like 65c. per ton of lead. So this is an electro-chemical process not especially favored by water-power.
The cost of labor is not greater than in the zinc-desilverization process. A comparison between this process and the Parkes process, on the assumption that the costs for labor, interest and general expenses are about equal, shows that about $1 worth of zinc and a considerable amount of coal and coke have been done away with, at the expense of power, equal to about 175 h.p. hours, of the average value of perhaps 65c., and a small amount of coal for melting the lead in the electrolytic method.
More important, however, is the greater saving of the metal values by reason of increased yields of gold, silver, lead, antimony and bismuth, and the freedom of the refined lead from bismuth.
Tables II, III, and IV show the composition of bullion, slimes and refined lead.
Tables V, VI, VII, and VIII give the results obtained experimentally in the laboratory on lots of a few pounds up to a few hundred pounds. The results in Tables VI and VII were given me by the companies for which the experiments were made.
TABLE I.—ANALYSES OF DROSS
For analyses of the lead from which this dross was taken, see Table II
| No. | No. in Table II. | Cu. Per Cent. | As. Per Cent. | Sb. Per Cent. | Fe. Per Cent. | Zn. |
|---|---|---|---|---|---|---|
| 1 | 2 | 0.0005 | 0.0003 | 0.0016 | 0.0016 | none |
| 2 | 3 | 0.0010 | 0.0008 | 0.0107 | 0.0011 | " |
TABLE II.—ANALYSES OF BULLION
| No. | Fe. Per Cent. | Cu. Per Cent. | Sb. Per Cent. | Sn. Per Cent. | As. Per Cent. | Ag. Per Cent. | Au. Per Cent. | Pb. Per Cent. | Ag. Per Cent. | Au. Per Cent. |
|---|---|---|---|---|---|---|---|---|---|---|
| 1 | 0.0075 | 0.1700 | 0.5400 | 0.0118 | 0.1460 | 1.0962 | 0.0085 | 98.0200 | 319.7 | 2.49 |
| 2 | 0.0115 | 0.1500 | 0.6100 | 0.0158 | 0.0960 | 1.2014 | 0.0086 | 97.9068 | 350.4 | 2.52 |
| 3 | 0.0070 | 0.1600 | 0.4000 | 0.0474 | 0.1330 | 1.0738 | 0.0123 | 98.1665 | 313.2 | 3.6 |
| 4 | 0.0165 | 0.1400 | 0.7000 | 0.0236 | 0.3120 | 0.8914 | 0.0151 | 97.9014 | 260.0 | 4.42 |
| 5 | 0.0120 | 0.1400 | 0.8700 | 0.0432 | 0.2260 | 0.6082 | 0.0124 | 98.0882 | 177.4 | 3.63 |
| 6 | 0.0055 | 0.1300 | 0.7300 | 0.0316 | 0.1030 | 0.6600 | 0.0106 | 98.2693 | 192.5 | 3.10 |
| 7 | 0.0380 | 0.3600 | 0.4030 | — | tr. | 0.7230 | 0.0180 | 98.4580 | 210.9 | 5.25 |
TABLE III.—ANALYSES OF SLIMES
| Fe. Per Cent. | Cu. Per Cent. | Sb. Per Cent. | Sn. Per Cent. | As. Per Cent. | Pb. Per Cent. | Zn. Per Cent. | Bi. Per Cent. |
|---|---|---|---|---|---|---|---|
| 1.27 | 8.83 | 27.10 | 12.42 | 28.15 | 17.05 | none | none |
| 1.12 | 22.36 | 21.16 | 5.40 | 23.05 | 10.62 | " | " |
TABLE IV.—ANALYSES OF REFINED LEAD
| No. | Cu. Per Cent. | As. Per Cent. | Sb. Per Cent. | Fe. Per Cent. | Zn. Per Cent. | Sn. Per Cent. | Ag. Oz. p. T. | Ni,Co,Cd. Per Cent. | Bi. Per Cent. |
|---|---|---|---|---|---|---|---|---|---|
| 1 | 0.0006 | 0.0008 | 0.0005 | — | — | — | — | — | — |
| 2 | 0.0003 | 0.0002 | 0.0010 | 0.0010 | none | — | — | — | — |
| 3 | 0.0009 | 0.0001 | 0.0009 | 0.0008 | " | — | 0.24 | — | — |
| 4 | 0.0016 | — | 0.0017 | 0.0014 | — | — | 0.47 | none | — |
| 5 | 0.0003 | — | 0.0060 | 0.0003 | — | — | 0.22 | — | — |
| 6 | 0.0020 | — | 0.0010 | 0.0046 | — | — | 0.22 | none | — |
| 7 | 0.0004 | none | 0.0066 | 0.0013 | none | 0.0035 | 0.14 | — | — |
| 8 | 0.0004 | — | 0.0038 | 0.0004 | " | 0.0035 | 0.25 | — | — |
| 9 | 0.0005 | — | 0.0052 | 0.0004 | " | 0.0039 | 0.28 | — | — |
| 10 | 0.0003 | none | 0.0060 | 0.0003 | " | 0.0049 | 0.43 | — | — |
| 11 | 0.0003 | " | 0.0042 | 0.0013 | " | 0.0059 | 0.32 | — | — |
| 12 | 0.0005 | " | 0.0055 | 0.0009 | " | 0.0049 | 0.22 | — | — |
| 13 | 0.0005 | " | 0.0055 | 0.0007 | " | 0.0091 | 0.11 | — | — |
| 14 | 0.0004 | " | 0.0063 | 0.0005 | " | 0.0012 | 0.14 | — | — |
| 15 | 0.0003 | " | 0.0072 | 0.0003 | " | 0.0024 | 0.24 | — | — |
| 16 | 0.0006 | " | 0.0062 | 0.0012 | " | 0.0083 | 0.22 | — | — |
| 17 | 0.0006 | " | 0.0072 | 0.0011 | — | 0.0080 | 0.23 | — | — |
| 18 | 0.0006 | " | 0.0057 | 0.0010 | — | 0.0053 | 0.34 | — | — |
| 19 | 0.0005 | " | 0.0066 | 0.0016 | — | 0.0140 | 0.38 | — | — |
| 19 | 0.0005 | " | 0.0044 | 0.0011 | — | 0.0108 | 0.35 | — | — |
| 20 | 0.0004 | " | 0.0047 | 0.0015 | — | 0.0072 | 0.22 | — | — |
| 20 | 0.0004 | " | 0.0034 | 0.0016 | — | trace | 0.23 | — | — |
| 21 | 0.0022 | " | 0.0010 | 0.0046 | none | 0.0081 | 0.38 | none | none |
TABLE V.—ANALYSES OF BULLION AND REFINED LEAD
| Ag. Per Cent. | Cu. Per Cent. | Sb. Per Cent. | Pb. Per Cent. | |
|---|---|---|---|---|
| Bullion | 0.50 | 0.31 | 0.43 | 98.76 |
| Refined lead | 0.0003 | 0.0007 | 0.0019 | 99.9971 |
TABLE VI.—ANALYSES OF BULLION AND REFINED LEAD
| Cu. Per Ct. | Bi. Per Ct. | As. Per Ct. | Sb. Per Ct. | Ag. Per Ct. | Ag. Per Ct. | Au. Oz. p.T. | Fe. Per Ct. | Zn. Per Ct. | |
|---|---|---|---|---|---|---|---|---|---|
| Bullion | 0.75 | 1.22 | 0.936 | 0.6832 | 358.89 | — | 1.71 | — | — |
| Refined lead | 0.0027 | 0.0037 | 0.0025 | 0.0000 | — | 0.0010 | none | 0.0022 | 0.0018 |
TABLE VII.—ANALYSES OF BULLION, REFINED LEAD AND SLIMES
| Pb. Per Cent. | Cu. Per Cent. | As. Per Cent. | Sb. Per Cent. | Ag. Oz.p.T. | Ag. Per Cent. | Fe,Zn,Ni,Co. Per Cent. | Bi. Per Cent. | |
|---|---|---|---|---|---|---|---|---|
| Bullion | 96.73 | 0.096 | 0.85 | 1.42 | about 275[51] | — | — | — |
| Refined lead | — | 0.0013 | 0.00506 | 0.0028 | — | 0.00068 | 0.0027 | trace |
| Slimes (dry sample) | 9.05 | 1.9 | 9.14 | 29.51 | 9366.9 | — | 0.49 | trace |
TABLE VIII.—ANALYSES OF BULLION, REFINED LEAD AND SLIMES
| Pb. Per Cent. | Cu. Per Cent. | Bi. Per Cent. | Ag. Per Cent. | Sb. Per Cent. | As. Per Cent. | |
|---|---|---|---|---|---|---|
| Bullion | 87.14 | 1.40 | 0.14 | 0.64 | 4.0 | 7.4 |
| Lead | — | 0.0010 | 0.0022 | — | 0.0017 | trace |
| Slimes | 10.3 | 9.3 | 0.52 | 4.7 | 25.32 | 44.58 |