DISCUSSION OF TESTS WITH FLOTATION CONCENTRATE.

In the tests of flotation concentrate, Table 52 shows that the only satisfactory total extractions of lead were in those roasts that were hot enough to cause considerable volatilization of lead and that rarely was a high extraction of lead obtained by chloridizing roasting followed by leaching with neutral brine or acid brine. Sublimation was the only promising method, as far as the recovery of lead was concerned.

EFFECTS OF TEMPERATURE AND DRAFT.

Tests 1 and 2 (Tables 47 to 52) were designed to be run at a fairly high temperature, hence a rather strong draft suction (2 inch on the water gage) was used. The charge used in test 1 contained 6 per cent of common salt and the one used in test 2 contained 10 per cent salt (see Table 47). Tests 3 and 4 were run to parallel the first two tests, except that the draft was reduced to one-half inch water gage in order that the temperatures might be lower. In each pair of tests the higher percentage of salt gave higher extractions of the lead, and in leaching the calcine a larger proportion of lead was dissolved with acid brine than with neutral brine. The charges used in tests 5 and 6 contained some powdered coal in addition to the natural sulphur content. The higher temperatures resulting did not seem to give any higher extractions of the lead than in the previous tests. Hence test 7 was run with 0.5 inch draft and 6 per cent salt and test 8 with 2 inches draft and 10 per cent salt, to verify the former tests. The results seemed to indicate that with a high draft (high temperature) and a large percentage of salt in the charge more of the lead was volatilized and the total extraction was greater. Therefore, tests 9 and 10 were run with still larger proportions of salt, additional fuel, and high draft in the roaster, with the expectation of obtaining more complete volatilization of the lead. However, the results were not so good as in the former tests.

EFFECTS FROM GRINDING AND MIXING.

In order to determine whether thorough mixing of the salt with the material might improve the reaction, in tests 11, 12, and 13 the material was slightly ground with iron balls in a ball mill with an excess of saturated salt solution, in order to allow intimate mixing of the salt and the mineral particles. The mixture, after being drained dry enough to permit charging into the roaster, was roasted. The excellence of the results shows the importance of thorough mixing. The best way of obtaining a high extraction of lead from material containing mixed sulphides seems to be to grind it fine, mix intimately with about 7.5 per cent salt, and roast in a blast roaster at such a temperature that the lead chloride formed will

be volatilized. A fluffy consistence of the charge in the roaster permits a light draft and good extractions.

DISCUSSION OF TESTS OF ELECTROSTATIC SEPARATOR CONCENTRATES. The results obtained in the tests of material from the electrostatic concentrator, presented in Tables 53 to 58, which was coarser than the flotation concentrate used in the previous tests, confirm these conclusions. Tests 1 and 2, compared with tests 3 and 4, show that the use of 10 per cent NaCl in the charge gives higher extractions of the lead than 6 per cent. The amount of draft used did not seem to make a great deal of difference in these tests. In tests 5 and 6 some of the finely ground flotation concentrate used in the first series of tests (Tables 47 to 52) was mixed with the coarser concentrate, designated as No. 2 material, Table 53, in order to test the effect of enough slimes being present in the charge to cause balling. Considering that less salt was used than in tests 3 and 4, the high extraction in test 6 shows that a better roast was obtained. In test 5 the draft was too low to give a good roast.

ADDITION OF FINELY GROUND MATERIAL.

In tests 7 and 8 the coarse No. 2 concentrate was made to ball by adding some finely ground lead-carbonate ore. In test 8 the draft was too low, but test 7 showed a good extraction of the lead. In test 9 a mixture of coarse No. 2 concentrate, some flotation concentrate, and some finely ground lead carbonate ore was roasted but the draft used was too low.

FINE GRINDING AND THOROUGH MIXING.

Tests 10 to 15 were made to determine whether fine grinding and intimate mixing would improve the roast. In general the results show that some benefit is derived from this treatment. As much as 90 per cent of the lead can be volatilized and leached from the material. In test 11 the charge settled firmly on the grate, owing to improper charging, and a high draft was necessary in order to get any air through. In this instance the high draft does not mean high temperature.

BALLING.

The finely ground material could be balled with water into a mass that would dry to a fluffy consistence during roasting in the blast roaster, leaving room for an excellent penetration of the gases. It is difficult to so mix coarse concentrates containing no slimes that they will go onto the roaster in such a fluffy mass, which permits roasting under a low draft. A part or the whole of the coarse concentrate can be ground finely to provide this slime. Grinding all

of the charge in saturated brine in a ball mill would insure thorough mixing with the necessary salt, but the subsequent filtering and partial drying of the ground material would be rather expensive, although not excessively so. The better practice would probably be to use some flotation concentrate to ball the coarser particles of sulphide and apply the salt as a saturated brine.

RECOVERY OF ZINC.

In the later tests, in which the salt was mixed with the charge by fine grinding, very little of the zinc in the ore was either volatilized or leached, showing that most of the zinc sulphide was unaffected by the roasting of the lead compounds. The highest percentage of the total zinc content volatilized in any of the tests was 7.35 per cent and the highest percentage leached was 25.15 per cent. The average percentage of total zinc volatilized was less than 5 per cent and the average amount leached was about 8 per cent, hence the average loss of zinc from the concentrate during the removal of the lead was 13 per cent. Of course the greater portion of this zinc would later be recovered from the fume and solutions in the form of zinc oxide, so that it need not be lost. It is a remarkable fact that nearly all the lead could be volatilized and leached from a mixed sulphide ore without markedly affecting the sphalerite in any way.

As regards improvement of grade of the zinc sulphide concentrate by removal of the lead and the partial roasting of the iron sulphide, it can be seen that the average zinc content in the leached residue of the first four tests is 40 per cent, whereas the material before roasting contained 32.67 per cent zinc. It so happens that material containing 32 per cent zinc is difficult to market as zinc concentrate, whereas in many of the contracts for the complex western zinc sulphide concentrates a 40 per cent zinc concentrate is used as the basis on which the buying price is calculated. Often zinc smelters refuse to smelt zinc sulphide ore carrying less than 35 per cent zinc. Therefore, although the resulting zinc sulphide residue is not of a grade to compare with the concentrates from such districts as the Joplin district, because of the large amount of iron pyrite intergrown with the zinc sulphide, this method applied to the ore will effect practically complete separation of the lead and raise the grade of the zinc concentrate from a product of doubtful commercial value to one which could be easily sold and is more desirable to zinc smelters. As the lead in such zinc concentrates is usually not paid for by the zinc smelter, such a saving is of value in conserving resources as well as adding to the possible profits of the shipper of the zinc concentrates. Roasts 14 and 15 were with zinc concentrates from the electrostatic separator in the Mary Murphy mill,

designated as No. 3 material in the tables. Before treatment these concentrates contained 42.6 per cent zinc and 9.18 per cent lead. The treatment removed 90 per cent of the lead and left a residue containing 94 per cent of the zinc as a concentrate containing 57 per cent zinc. The improvement in the grade of this product is even more noticeable.

RECOVERY OF COPPER AND SILVER.

The extraction of the copper and the silver from such zinc concentrates is of interest, as they are lost along with the lead in ordinary methods of zinc smelting, unless the zinc smelter residue is sent to the lead blast furnaces. An inspection of the tables will show that in most of the tests the extractions of copper are practically negligible. As much as 60 per cent of the silver could be dissolved out by quick leaching, but on standing in contact with the zinc sulphide in the calcine from the volatilizing operation to recover the lead, the brine would lose its dissolved silver. Some tests with the above-mentioned calcines on brines containing dissolved copper and silver chlorides showed that the calcine precipitated all of the silver and copper in solution in 16 hours, and about 50 per cent of the silver and copper in 20 minutes. If the brine was strongly acid, precipitation was not so rapid nor nearly so complete. Roasting evidently changed the properties of the material, as far as precipitation of silver and copper from brine solutions are concerned. The raw material did not act like the calcine in this respect, as the precipitation of silver and of copper was very small. Nevertheless, the roasted material seemed to contain zinc sulphide and very little zinc oxide. It was susceptible of flotation, yielding a dark colored blende concentrate, and acted in every way like zinc sulphide.

Also, the calcine tended to reprecipitate lead from the brine. However, most of the lead could be recovered without much trouble, whereas the silver and copper extractions were always low. The action of zinc oxide on the solutions was practically identical with that of the roasted calcines, which were supposedly zinc sulphide. Hence it does not seem possible to make commercial extractions of the silver and copper in zinc-sulphide concentrates, even when the silver does not accompany the zinc sulphide in intimate crystallization.

LIMITATIONS OF THE METHOD.

This latter fact shows that the process is hardly all that could be desired. Its use insures recovery of the lead from such materials, with a considerable increase of the grade of the zinc concentrate, but yields a low recovery of the silver and copper that may be

present. Copper is not an uncommon constituent of such concentrates, although it is not a usual constituent in the bulk of the present tonnage of leady zinc concentrates being sent to the zinc smelters from the intermountain regions. However, silver is usually present in such materials in proportions varying from small amounts to as much as 20 ounces per ton. Where the usual extraction of the silver may be as much as 25 to 33 per cent, it can be seen that the process is not desirable for the treatment of high silver ores. For the treatment of low silver ores and for ores that do not contain silver, this method has much to recommend it. For instance, the great bulk of zinc concentrates shipped from the western mountain States, excluding those from the Butte mines, contain not over 3 ounces per ton of silver. Usually the lead content of such concentrates falls below 5 per cent, in the improved ore-dressing practice of the present time, although considerable quantities contain 10 per cent or more of lead, which is a loss to the companies shipping the zinc concentrate. For such conditions the process is certainly a step in the right direction.

CONCLUSIONS AS TO EXPERIMENTS WITH LEADY ZINC-SULPHIDE CON

CENTRATES.

The treatment of leady zinc-sulphide concentrates by chloridizing roasting, followed by brine leaching, gives the following results:

1. It is possible to recover 90 per cent or more of the lead in such concentrates by volatilizing most of the lead in a fairly hot roast on a down-draft shallow-bed roaster, such as the Dwight-Lloyd or the Knight-Christensen roaster, followed by a quick leach with saturated brine acidified with sulphuric acid.

2. For treating coarse concentrates it is advisable to add some finely divided flotation concentrate to the roaster charge in order that the mass may be made to ball with a small amount of water and leave the charge porous during roasting.

3. Six to ten per cent of sodium chloride, based on the weight of concentrate, is necessary for good chloridizing with the average

concentrate.

4. The draft necessary to roast a charge bedded 1.5 inches in less than 15 minutes is about 1 inch water gage.

5. The zinc sulphide in the charge is practically unaffected by the combustion of the galena and the pyrite and remains in the residue from leaching. About 90 to 95 per cent of the zinc is recovered in this semisintered residue, which contains a much higher percentage of zinc than the original concentrate treated. Such material should be more acceptable to the zinc smelter because of the absence of lead and its higher grade, assuming that the chloridizing treatment has not injured its roasting properties. In roasting flotation concentrates